Method for preparing high-concentration acid leaching vanadium liquid and vanadium pentoxide

Abstract

The invention discloses a method for preparing high-concentration acid leaching vanadium liquid and vanadium pentoxide from vanadium slag calcified clinker, and belongs to the technical field of hydrometallurgy. The method aims to solve the problems of low vanadium concentration of leachate and high wastewater yield of the existing calcification vanadium extraction process. The core technology comprises the steps that multi-stage countercurrent acid leaching is adopted, the concentration of vanadium liquid is enriched stage by stage, and finally high-concentration leachate with the vanadium concentration larger than or equal to 40 g / L is obtained; performing multi-stage countercurrent washing on residues generated by acid leaching, and returning an obtained vanadium-rich washing solution to a leaching system; the washed residues are combined and subjected to deep leaching, residual vanadium is recycled, and the obtained deep leaching liquid serves as washing water; and finally, the high-concentration vanadium liquid is subjected to vanadium precipitation and calcination, and a qualified vanadium pentoxide product is obtained. According to the method, the vanadium concentration and the vanadium recovery rate of the leachate are greatly improved, the discharge of wastewater and waste residues is remarkably reduced, the production cost is reduced, and the method is environment-friendly and has good economic benefits and social benefits.

Description

Technical Field

[0001] This invention belongs to the field of hydrometallurgical technology, specifically relating to a method for efficiently leaching vanadium from vanadium slag calcination and roasting clinker and preparing high-concentration acid leaching vanadium solution and qualified vanadium pentoxide product. Background Technology

[0002] Vanadium, an important strategic metal, is widely used in steel, aerospace, chemical, and energy storage industries. Extracting vanadium from vanadium slag is one of the main methods for obtaining vanadium resources. Currently, the mainstream industrial vanadium extraction processes include sodium roasting-water leaching and calcification roasting-acid leaching.

[0003] The vanadium slag calcination roasting-acid leaching process independently developed by companies such as Panzhihua Iron and Steel Group is a promising clean production process. This process involves mixing calcium additives (such as calcium carbonate and calcium oxide) with vanadium slag and then subjecting it to high-temperature oxidative roasting. This converts the low-valent vanadium (such as V(III)) in the vanadium slag into acid-soluble calcium vanadate (such as Ca(VO3)2, Ca2V2O7, etc.). After acid leaching, the roasted clinker releases vanadium as VO2... + The vanadium pentoxide product is obtained by plasma entering the solution and subsequent precipitation and calcination processes. Compared with the traditional sodium vanadium extraction process, the calcification vanadium extraction process has the following advantages: the calcium additives used (such as limestone) are inexpensive; the vanadium extraction tailings do not contain soluble sodium salts, avoiding environmental pollution caused by sodium salts, and the tailings are easier to utilize as resources (such as returning to steelmaking or using them as building materials); the process wastewater treatment is relatively simple, and it is expected to achieve full-process recycling.

[0004] However, existing vanadium extraction processes using vanadium slag calcification still face significant technical bottlenecks in practical industrial applications. One of the most prominent problems is the low vanadium concentration in the leachate, typically only 20-25 g / L, which is about half the concentration in the leachate from the sodium leaching process. This low vanadium concentration leads to a series of adverse consequences: firstly, the subsequent vanadium precipitation process requires handling a larger volume of solution, resulting in a huge amount of process wastewater; secondly, the large volume of solution treatment increases energy consumption and equipment investment; and thirdly, the large amount of wastewater requires neutralization treatment, generating massive amounts of neutralization slag, increasing solid waste treatment costs and environmental pressure. This problem severely restricts the competitiveness and large-scale application of the vanadium extraction process using calcification.

[0005] Therefore, there is an urgent need in this field for a new method that can significantly increase the vanadium concentration in the acid leaching solution of vanadium slag calcification clinker, in order to fundamentally reduce wastewater production, lower production costs, and improve the overall techno-economic efficiency and environmental friendliness of the calcification vanadium extraction process. Summary of the Invention

[0006] The primary objective of this invention is to overcome the technical defects of low vanadium concentration and large wastewater generation in existing vanadium slag calcification vanadium extraction processes, and to provide a method for preparing high-concentration acidic vanadium leachate.

[0007] Another object of the present invention is to provide a method for preparing qualified vanadium pentoxide products based on the above-mentioned high-concentration vanadium solution.

[0008] Another objective of this invention is to achieve efficient extraction and recovery of vanadium by optimizing the process flow, thereby reducing production costs and environmental pollution.

[0009] To achieve the above objectives, this invention provides a method for preparing high-concentration acid-leached vanadium solution and vanadium pentoxide from vanadium slag calcification clinker. The core of this method lies in the integrated process of multi-stage countercurrent acid leaching combined with deep leaching of residue and multi-stage countercurrent washing. The method specifically includes the following steps: First-stage acid leaching: Vanadium slag calcined and roasted clinker is mixed with leaching mother liquor in a certain proportion. Acid is added under stirring conditions, and the pH and temperature of the leaching system are controlled to carry out the leaching reaction. After the reaction is completed, solid-liquid separation is performed to obtain a first-stage acidic vanadium leaching solution and a first-stage leaching residue.

[0010] Second-stage acid leaching: The first-stage acidic vanadium leaching solution is used as the mother liquor and mixed with fresh vanadium slag calcined clinker. A second-stage acid leaching is then carried out under similar conditions. After the reaction is complete, solid-liquid separation is performed to obtain a second-stage acidic vanadium leaching solution with a higher vanadium concentration and a second-stage leaching residue.

[0011] Third-stage acid leaching: The second-stage acidic vanadium leaching solution is used as the mother liquor and mixed with fresh vanadium slag calcined clinker. Third-stage acid leaching is then carried out under similar conditions. After the reaction is complete, solid-liquid separation is performed to obtain a third-stage acidic vanadium leaching solution with a vanadium concentration reaching the target value (e.g., ≥40 g / L) and third-stage leaching residue.

[0012] Multi-stage countercurrent washing: Vanadium-rich leaching residues are washed with wash water to recover soluble vanadium adhering to the residues. The washing process employs a countercurrent method: fresh wash water is first used to wash the final residue (third-stage leaching residue) with the lowest vanadium content. The resulting wash filtrate is then used to wash the residue from the previous stage (second-stage leaching residue), and so on. Finally, the wash filtrate with a higher vanadium concentration obtained from washing the first-stage leaching residues is returned to the system and used as the mother liquor for the first-stage acid leaching.

[0013] Deep leaching of residue: The first to third stage leaching residues after the above multi-stage countercurrent washing are combined and subjected to short-term deep leaching under lower pH conditions to recover unleached encapsulated vanadium or sparingly soluble vanadium from the residues. The slurry after deep leaching undergoes solid-liquid separation, and the resulting deep leaching solution is rich in a certain amount of vanadium, which is used as the wash water in the multi-stage countercurrent washing process.

[0014] Vanadium precipitation and calcination: The high-concentration acidic vanadium leaching solution obtained from the third-stage acid leaching is subjected to vanadium precipitation treatment (usually using ammonium salts, such as ammonium sulfate or ammonium carbonate). The pH value is adjusted and the solution is heated to precipitate ammonium polyvanadate. After solid-liquid separation, the ammonium polyvanadate is calcined at a suitable temperature to obtain vanadium pentoxide product.

[0015] Preferably, in steps 1, 2, and 3, the mixing mass ratio of the vanadium slag calcined clinker to the leaching mother liquor is 0.5 to 10, more preferably 2.5 to 5.

[0016] Preferably, in steps 1, 2, and 3, the pH of the leaching slurry is controlled within the range of 2.5 to 3.5 during acid leaching, the reaction temperature is ≤65℃, and the reaction time is 40 to 120 minutes. Sulfuric acid is preferably used.

[0017] Preferably, in step 4, the amount of washing water used is consistent with the mass ratio of the calcified roasted clinker added in a single batch, and the ratio of the leaching mother liquor to the clinker.

[0018] Preferably, in step 5, the pH of the leaching slurry is controlled within the range of 0.5 to 2.0 during deep leaching, and the reaction time is 5 to 30 minutes.

[0019] Preferably, in step 6, the NH4 content is controlled during vanadium precipitation. + The molar ratio of vanadium to total vanadium (TV) is 0.8–1.5, the pH for vanadium precipitation is 1.6–2.2, and the reaction is carried out by heating to boiling. The calcination temperature is controlled at 500–600℃, and the calcination time is 60–240 minutes.

[0020] Compared with the prior art, the technical solution provided by the present invention has the following significant advantages and positive effects: Significantly increase the vanadium concentration in the leachate: Through three-stage countercurrent acid leaching, the leachate from the previous stage (which already contains a certain concentration of vanadium) is used as the mother liquor for the next stage, achieving the gradual enrichment of vanadium. This allows the vanadium concentration in the final third-stage leachate to be stably increased to over 40 g / L, or even over 60 g / L, which is far higher than the existing technology level of 20-25 g / L.

[0021] Significantly reduced wastewater production: Because the volume of the solution ultimately entering the vanadium precipitation process is greatly reduced due to the increased vanadium concentration, the total amount of wastewater requiring treatment after vanadium precipitation is also significantly reduced, estimated to be more than 50% less. This not only lowers wastewater treatment costs but also reduces the amount of neutralization residue generated.

[0022] Improving vanadium recovery rate: By combining "multi-stage countercurrent washing" and "deep leaching of residues," vanadium adhering to and encapsulated in the leaching residues is effectively recovered, significantly reducing the vanadium content in the residues and thus improving the overall vanadium recovery rate. Examples show that the overall vanadium leaching rate can reach over 90%.

[0023] Reduced production costs: The reduction in wastewater treatment directly lowers environmental protection costs; the increased concentration of leachate reduces energy consumption in subsequent processes such as vanadium precipitation and evaporation; although the process flow has increased, the efficient utilization of reagents and resources is achieved through the optimization of internal logistics (such as the reuse of washing liquid and deep leachate), thus reducing overall operating costs.

[0024] Environmentally friendly: Reduced emissions of process wastewater and solid waste alleviate environmental pressure and are more in line with the development direction of green metallurgy.

[0025] Excellent product quality: Ammonium polyvanadate obtained by precipitating vanadium from high-concentration vanadium liquid with relatively few impurities can be calcined to obtain qualified vanadium pentoxide products with a purity of over 98.5%, which meet the YB / T 5304-2017 standard. Attached Figure Description

[0026] 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.

[0027] Figure 1 This is a schematic diagram of the process flow for preparing high-concentration acid leaching vanadium solution and vanadium pentoxide from vanadium slag calcification clinker according to the present invention. Detailed Implementation

[0028] The embodiments of this disclosure will be further described in detail below with reference to the accompanying drawings and examples. The detailed description of the embodiments and the accompanying drawings are used to illustrate the principles of this disclosure by way of example, but should not be used to limit the scope of this disclosure. This disclosure can be implemented in many different forms and is not limited to the specific embodiments disclosed herein, but includes all technical solutions falling within the scope of the claims.

[0029] These embodiments are provided to make the disclosure thorough and complete, and to fully express the scope of the disclosure to those skilled in the art. It should be noted that, unless otherwise specifically stated, the relative arrangement of components and steps, material composition, numerical expressions, and values ​​set forth in these embodiments should be interpreted as exemplary only and not as limiting.

[0030] As used in this disclosure, the words “including” or “contains” or similar terms mean that the element preceding the word covers the element listed after the word, and do not exclude the possibility that it may also cover other elements.

[0031] All terms used in this disclosure have the same meaning as understood by one of ordinary skill in the art to which this disclosure pertains, unless otherwise specifically defined. It should also be understood that terms defined in general dictionaries should be interpreted as having meanings consistent with their meanings in the context of the relevant art, and not as idealized or highly formalized, unless expressly defined herein.

[0032] Techniques, methods, and equipment known to those skilled in the art may not be discussed in detail, but where appropriate, they should be considered part of the specification.

[0033] Combination Figure 1 In an embodiment of the present invention, a method 100 for preparing high-concentration acid leaching vanadium solution and vanadium pentoxide from vanadium slag calcification clinker includes the following steps: Step 101, First-stage acid leaching: The vanadium slag calcined and roasted clinker is mixed with the leaching mother liquor in a certain proportion. Acid is added under stirring conditions, and the pH and temperature of the leaching system are controlled to carry out the leaching reaction. After the reaction is completed, solid-liquid separation is performed to obtain the first-stage acidic vanadium leaching solution and the first-stage leaching residue.

[0034] Step 102, Second-stage acid leaching: The first-stage acidic vanadium leaching solution is used as the leaching mother liquor and mixed with fresh vanadium slag calcined clinker. A second-stage acid leaching is then performed under similar conditions. After the reaction is complete, solid-liquid separation is performed to obtain a second-stage acidic vanadium leaching solution with a higher vanadium concentration and a second-stage leaching residue.

[0035] Step 103, Third-stage acid leaching: The second-stage acidic vanadium leaching solution is used as the leaching mother liquor and mixed with fresh vanadium slag calcined clinker. Third-stage acid leaching is then carried out under similar conditions. After the reaction is complete, solid-liquid separation is performed to obtain a third-stage acidic vanadium leaching solution with a vanadium concentration reaching the target value (e.g., ≥40 g / L) and third-stage leaching residue.

[0036] Step 104, Multi-stage Countercurrent Washing: The vanadium-rich leaching residue is washed with wash water to recover soluble vanadium adhering to the residue. The washing process uses a countercurrent method: fresh wash water is first used to wash the final residue (third-stage leaching residue) with the lowest vanadium content. The resulting wash filtrate is used to wash the previous stage residue (second-stage leaching residue), and so on. Finally, the wash filtrate with a higher vanadium concentration obtained from washing the first-stage leaching residue is returned to the system and used as the mother liquor for the first-stage acid leaching.

[0037] Step 105, Deep Leaching of Residue: The first to third stage leaching residues after the above multi-stage countercurrent washing are combined and subjected to short-term deep leaching under lower pH conditions to recover unleached encapsulated vanadium or sparingly soluble vanadium from the residues. The slurry after deep leaching undergoes solid-liquid separation, and the resulting deep leaching solution is rich in a certain amount of vanadium, which is used as the wash water in the multi-stage countercurrent washing process.

[0038] Step 106, Vanadium Precipitation and Calcination: The high-concentration acidic vanadium leaching solution obtained from the third-stage acid leaching is subjected to vanadium precipitation treatment (usually using ammonium salts, such as ammonium sulfate or ammonium carbonate). The pH value is adjusted and the solution is heated to precipitate ammonium polyvanadate. After solid-liquid separation, the ammonium polyvanadate is calcined at a suitable temperature to obtain vanadium pentoxide product.

[0039] The following is combined with Figure 1 The present invention will be further described in detail with reference to specific embodiments, but the implementation of the present invention is not limited thereto. Process parameters not specifically specified in the following embodiments are generally implemented under conventional conditions.

[0040] Example 1 Take 100g of vanadium slag calcined and roasted clinker (with a total vanadium (TV) content of 8.10% and a CaO content of 6.71%) and 350g of leaching mother liquor (initially, it can be water or circulating liquid) and place them in a leaching tank, then start stirring. Slowly add dilute sulfuric acid to adjust and maintain the pH of the slurry at around 3.0. React for 60 minutes at a temperature controlled below 65℃. After the reaction is complete, perform pressure filtration to separate the first-stage acidic vanadium leaching solution and the first-stage leaching residue (wet basis).

[0041] The entire first-stage acidic vanadium leaching solution was used as the leaching mother liquor and mixed with another 100g of vanadium slag calcification roasted clinker. The mixture was reacted for 60 minutes under the same conditions of pH≈3.0 and temperature≤65℃. After solid-liquid separation, the second-stage acidic vanadium leaching solution and the second-stage leaching residue were obtained.

[0042] The entire second-stage acidic vanadium leaching solution was then used as the mother liquor and mixed with a third portion of 100g of vanadium slag calcified roasted clinker. A third-stage leaching was then performed under the same conditions (pH≈3.0, ≤65℃, 60min). After solid-liquid separation, approximately 305ml of the third-stage acidic vanadium leaching solution was obtained, and its TV concentration was found to be as high as 54.2 g / L.

[0043] The residues from the first, second, and third stages of leaching were combined and mixed with process circulating water. Concentrated sulfuric acid was added until the slurry pH reached 1.0, and the mixture was reacted for 15 minutes for deep leaching. After solid-liquid separation, the deep leachate and final tailings were obtained. The TV content of the final tailings was measured to be 0.76%.

[0044] The deep leachate is used as wash water for multi-stage countercurrent washing: first, this wash water is used to wash the third-stage leaching residue to obtain washing filtrate I; washing filtrate I is used to wash the second-stage leaching residue to obtain washing filtrate II; washing filtrate II is used to wash the first-stage leaching residue to obtain washing filtrate III. This washing filtrate III can then be reused as the mother liquor for the next batch of first-stage acid leaching.

[0045] Take 305 ml of the high-concentration third-stage acidic vanadium leaching solution (TV = 54.2 g / L) prepared above, and add 16 g of ammonium sulfate (to control NH4). + The solution was adjusted to pH 1.8 with acid, heated to boiling and held for 60 minutes to precipitate ammonium polyvanadate. The precipitate was filtered and then calcined at 560℃ for a sufficient time to obtain vanadium pentoxide. Analysis showed that the V₂O₅ purity in the product reached 98.45%, and the content of other impurities met the requirements of YB / T 5304-2017 standard.

[0046] The overall vanadium leaching rate in this embodiment is calculated to be 92.57%.

[0047] Example 2 Take 100g of vanadium slag calcified roasted clinker (TV=9.02%, CaO=7.54%) and mix it with 500g of leachate mother liquor. Adjust the pH of the slurry to 2.5 with dilute sulfuric acid and react at ≤65℃ for 40 minutes. Separate the solid and liquid to obtain the first-stage leachate and residue.

[0048] The same procedure was followed for the second and third stages of countercurrent leaching, both under conditions of pH≈2.5, with a reaction time of 40 minutes. After the third stage leaching, approximately 450 ml of acidic vanadium leaching solution was obtained, with a TV concentration of 48.2 g / L.

[0049] The residues from each stage were combined, and the deep leaching conditions were pH=2.0 and reaction time was 30 minutes. The TV content of the resulting tailings was 1.03%.

[0050] During the third stage of high-concentration vanadium liquor precipitation, 32g of ammonium sulfate was added to adjust the pH to 2.2, and the reaction was allowed to boil for 60 minutes. The resulting ammonium pervanadate was calcined at 600℃ to obtain a product with a V₂O₅ purity of 99.12%, which meets the standard.

[0051] The overall vanadium leaching rate in this embodiment is 90.05%.

[0052] Example 3 Take 100g of vanadium slag calcified roasted clinker (TV=7.46%, CaO=7.22%) and mix it with 250g of leaching mother liquor. Adjust the pH of the slurry to 3.5 with dilute sulfuric acid and react at ≤65℃ for 120 minutes to separate the solid and liquid.

[0053] The second and third countercurrent leaching stages were performed as described above (pH≈3.5, 120 min). After the third leaching, approximately 216 ml of acidic vanadium leaching solution was obtained, with a TV concentration as high as 62.8 g / L.

[0054] The combined residues were subjected to deep leaching under the conditions of pH=0.5 and reaction time of 5 minutes, resulting in a TV content of 0.72% in the tailings.

[0055] During the third stage of high-concentration vanadium liquor precipitation, 11g of ammonium sulfate was added to adjust the pH to 1.6, and the reaction was carried out by boiling for 60 minutes. The obtained ammonium pervanadate was calcined at 540℃ to obtain a product with a V2O5 purity of 98.72%, which meets the standard.

[0056] The overall vanadium leaching rate in this embodiment is 91.31%.

[0057] Comparative Example 1 For comparison, 100g of vanadium slag calcined clinker (TV=8.10%) from the same batch as in Example 1 was mixed with 350g of water and leached for 60 minutes at pH=3.0 and ≤65℃. After solid-liquid separation, approximately 310ml of leachate was obtained, and its TV concentration was found to be only 19.5 g / L. After deep leaching of the residue under the same conditions, the TV of the tailings was 0.75%, and the calculated vanadium leaching rate for a single stage was approximately 86%.

[0058] Comparative Analysis of Effects By comparing Examples 1-3 with the comparative examples, the superiority of the present invention can be clearly seen: Leachate concentration: The final vanadium concentration of the leachate obtained by the method of the present invention (48.2-62.8 g / L) is much higher than that of traditional single-stage leaching (19.5 g / L).

[0059] Vanadium recovery rate: The overall vanadium leaching rate of this invention (90.05%-92.57%) is significantly higher than the single-stage leaching rate of the comparative example (approximately 86%).

[0060] Wastewater reduction: Based on the production unit mass of V2O5, the volume of solution entering the vanadium precipitation process is significantly reduced, and the corresponding wastewater generation is expected to be reduced by 50%-65%.

[0061] This invention increases the vanadium concentration in the acid leaching vanadium solution during the production process and reduces production wastewater. However, the vanadium leaching rate decreases due to the increased vanadium concentration in the mother liquor during multi-stage leaching and the subsequent leaching of calcified clinker. Therefore, the vanadium in the residue is recovered through deep leaching of the residue and countercurrent washing to maintain a stable total vanadium yield.

[0062] In summary, this invention, through the synergistic combination of multi-stage countercurrent acid leaching, multi-stage countercurrent washing, and deep leaching of residues, successfully solves the key technical problems of low leachate concentration, high wastewater content, and insufficient recovery rate in existing vanadium extraction processes using calcification. It provides an efficient, economical, and environmentally friendly feasible path for upgrading and optimizing the vanadium extraction process from vanadium slag.

[0063] The embodiments of this disclosure have now been described in detail. To avoid obscuring the concept of this disclosure, some details known in the art have not been described. Those skilled in the art will fully understand how to implement the technical solutions disclosed herein based on the above description.

[0064] While specific embodiments of this disclosure have been described in detail by way of examples, those skilled in the art should understand that the examples are for illustrative purposes only and are not intended to limit the scope of this disclosure. Those skilled in the art should understand that modifications can be made to the above embodiments or equivalent substitutions can be made to some technical features without departing from the scope and spirit of this disclosure.

Claims

1. A method for preparing high-concentration acid leaching vanadium solution and vanadium pentoxide, characterized in that, Includes the following steps: (a) Multi-stage countercurrent acid leaching: The vanadium slag calcified roasted clinker and the leaching mother liquor are subjected to at least two stages of acid leaching reaction in series. After each stage of acid leaching reaction, solid-liquid separation is performed to obtain the acidic vanadium leaching solution and leaching residue of that stage. The acidic vanadium leaching solution obtained in the previous stage is used as the leaching mother liquor of the next stage and mixed with the new vanadium slag calcified roasted clinker for leaching. The last stage of acid leaching yields a high-concentration acidic vanadium leaching solution. (b) Multi-stage countercurrent washing: The last two stages of leaching residue generated in step (a) are washed with wash water in a countercurrent manner, so that the wash water flows in the opposite direction through the leaching residue with decreasing vanadium content, and the final vanadium-rich wash filtrate is returned to step (a) as the leaching mother liquor for the first stage acid leaching. (c) Deep leaching of residue: The leaching residue after washing in step (a) is combined and deep leached under acidic conditions. After solid-liquid separation, deep leachate and final tailings are obtained. (d) Vanadium precipitation and calcination: The high-concentration acidic vanadium leaching solution obtained in step (a) is subjected to vanadium precipitation treatment to obtain ammonium polyvanadate precipitate, which is then calcined to obtain vanadium pentoxide product.

2. The method according to claim 1, characterized in that, In step (a), the multi-stage countercurrent acid leaching is a three-stage countercurrent acid leaching.

3. The method according to claim 1 or 2, characterized in that, In step (a), during each stage of acid leaching, the mass ratio of the vanadium slag calcined clinker to the leaching mother liquor is 0.5 to 10, preferably 2.5 to 5.

4. The method according to claim 1 or 2, characterized in that, In step (a), the acid used for acid leaching is sulfuric acid; the reaction conditions for acid leaching are: controlling the pH of the leaching slurry within the range of 2.5 to 3.5, the reaction temperature ≤ 65℃, and the reaction time 40 to 120 minutes.

5. The method according to claim 1, characterized in that, In step (b), the washing water is the deep leachate obtained in step (c).

6. The method according to claim 1 or 5, characterized in that, In step (b), the amount of washing water used and the mass ratio of the vanadium slag calcification roasted clinker added in a single batch are consistent with the ratio of leaching mother liquor to clinker in step (a).

7. The method according to claim 1, characterized in that, In step (c), the reaction conditions for deep leaching are: controlling the pH of the leaching slurry within the range of 0.5 to 2.0, and the reaction time is 5 to 30 minutes.

8. The method according to claim 1, characterized in that, In step (d), the vanadium precipitation treatment involves precipitation with ammonium salts to control NH4. + The molar ratio of vanadium to total vanadium (TV) in the solution is 0.8–1.5, the pH of vanadium precipitation is 1.6–2.2, and the reaction is carried out by heating to boiling.

9. The method according to claim 1 or 8, characterized in that, In step (d), the calcination conditions are: calcination temperature of 500-600℃ and calcination time of 60-240 minutes.

10. A high-concentration acidic vanadium leaching solution prepared by the method according to any one of claims 1-9, characterized in that, The vanadium concentration TV of the solution is ≥ 40 g / L.

Images