A method for extracting vanadium by roasting and acid / alkali leaching of vanadium-containing steel slag and vanadium slag in cooperation

Abstract

The present application belongs to the technical field of vanadium chemical engineering and resource recycling, and particularly relates to a method for vanadium extraction by roasting vanadium-containing steel slag and vanadium slag together and then acid / alkali leaching. The steps are as follows: the ball-milled vanadium-containing steel slag and vanadium slag are mixed thoroughly and high-temperature roasted to obtain roasted clinker; the obtained roasted clinker is subjected to acid leaching or alkali leaching; a vanadium precipitation additive is added to the obtained vanadium-containing acid leaching solution to carry out vanadium precipitation reaction, and a precipitate is obtained; the obtained precipitate is heated in an air atmosphere to obtain V2O5. Based on the high calcium and vanadium-containing characteristics of the vanadium-containing steel slag, the high-value utilization of the vanadium-containing steel slag and the vanadium slag is realized by roasting them together. The method has high vanadium recovery rate, can effectively reduce the solid waste storage problem of the vanadium-containing steel slag, and has a simple process flow and is easy to realize industrial application.

Description

Technical Field

[0001] This invention belongs to the field of vanadium chemical technology and resource recycling technology, specifically relating to a method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag followed by acid / alkali leaching. Background Technology

[0002] With the rapid development of the energy storage industry, vanadium, as a core material for advanced energy storage technologies such as flow batteries, is experiencing sustained and rapid market demand growth. Currently, my country's vanadium resources mainly come from vanadium slag produced by smelting vanadium-titanium magnetite and vanadium-containing steel slag, a byproduct of subsequent steelmaking processes. Statistics show that approximately 76% of vanadium supply comes from vanadium slag, with another 10% relying on secondary resources such as vanadium-containing steel slag. However, both of these resources face significant bottlenecks in existing extraction technologies, hindering the green and sustainable development of the vanadium industry.

[0003] In the treatment of vanadium slag, the mainstream process still uses sodium roasting. This method requires a large amount of added sodium salt, which not only generates corrosive gases but also faces problems such as difficulty in separating vanadium and chromium, and high costs for treating high-salt wastewater and slag. Although calcification roasting can avoid vanadium and chromium separation and reduce high-salt wastewater, it often uses calcium carbonate as an additive, releasing a large amount of carbon dioxide during roasting, which is inconsistent with the national "dual-carbon" strategic goal.

[0004] On the other hand, vanadium in vanadium-containing steel slag has a complex occurrence state, often encapsulated in stable phases such as calcium titanium oxide and tricalcium silicate, making it difficult for conventional leaching agents to penetrate effectively, resulting in low vanadium leaching efficiency. Simultaneously, the high calcium content of the steel slag itself causes it to preferentially react with acid during acid leaching, significantly increasing acid consumption. Furthermore, the generated calcium salts form a coating on the slag particle surface, further hindering vanadium leaching. More critically, existing technologies have failed to adequately address the dual impact of phosphate components in vanadium-containing steel slag. Studies have shown that during acid leaching, phosphate anions form soluble, stable vanadium-phosphorus complexes with dissolved vanadium ions. This complex formation can, on the one hand, continuously reduce the concentration of free vanadium ions in the system, promoting the forward dissolution and leaching reaction of vanadium according to the principle of chemical equilibrium shift, thus significantly enhancing the leaching behavior of vanadium. However, on the other hand, this stable vanadium-phosphorus complex strongly inhibits the nucleation and growth of vanadate precipitates during the vanadium precipitation stage, leading to a significant reduction in the vanadium precipitation rate. Existing technologies generally lack systematic control methods for this contradictory effect of "leaching is beneficial, vanadium precipitation is harmful", which makes the phosphate component in vanadium-containing steel slag often the key obstacle restricting the vanadium recovery rate of the whole process.

[0005] For example, Chinese patent CN121228028A discloses a method for the comprehensive utilization of vanadium-containing steel slag and vanadium extraction tailings. This method uses vanadium-containing steel slag and vanadium extraction tailings as raw materials. It should be noted that one of the raw materials used in this patent, the "vanadium extraction tailings," differs significantly from the "vanadium slag" addressed in this application. To convert vanadium in the raw materials into easily soluble vanadates, a composite modifier is added, and calcination is performed to complete phase reconstruction. Subsequently, when the calcined product is leached with a 10-15% sulfuric acid solution, the leaching solution contains high levels of impurities and inevitably forms stable vanadium-phosphorus complexes. Therefore, it is necessary to rely on an extraction process to extract and enrich vanadium, resulting in a complex and cumbersome overall process and a significant increase in production costs. Furthermore, the obtained calcined product is only suitable for acid leaching, and this single leaching path limits the adaptability of the process and restricts the selection of subsequent vanadium precipitation processes.

[0006] Therefore, developing a green vanadium extraction technology that can synergistically and efficiently utilize vanadium slag and vanadium-containing steel slag is of great significance for overcoming the limitations of existing processes and promoting the high-quality and sustainable development of the vanadium chemical industry. Summary of the Invention

[0007] This invention proposes a method for vanadium extraction and V2O5 preparation by co-calcination of vanadium-containing steel slag and vanadium slag followed by acid / alkali leaching, in order to solve the problems of high calcium encapsulation in vanadium-containing steel slag leading to difficult vanadium leaching, single phase limiting the leaching path, and high pollution and high cost.

[0008] Technical solution This invention proposes a method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag followed by acid / alkali leaching, the steps of which are as follows: S1. Calcination: The vanadium-containing steel slag after ball milling is thoroughly mixed with vanadium slag to obtain a mixture. The mixing ratio is calculated based on the ratio of Ca and P in the vanadium-containing steel slag to V in the mixture, with n(Ca / V) = 0.3~2 and n(P / V) = 0.018~0.12. The mixture is then calcined at a high temperature of 600~950℃, preferably 800~950℃, to obtain calcined clinker. S2. Leaching: The roasted clinker obtained in step S1 is subjected to acid leaching or alkali leaching; When acid leaching is used, the pH of the leaching system is adjusted to 0.5-1.5 or 2-3 using an acid solution, the liquid-to-solid ratio is controlled at 3-20 mL / g, the leaching temperature is 10-100℃, and the leaching time is 10-120 min to obtain a vanadium-containing acidic leachate. When alkaline leaching is used, the following steps are taken: leaching is performed using an alkaline solution, with the liquid-to-solid ratio controlled at 3–20 mL / g, the leaching temperature at 10–100℃, and the leaching time at 10–180 min, to obtain a vanadium-containing alkaline leachate. S3, Vanadium Recycling: Add vanadium precipitation additive to the vanadium-containing acidic leachate obtained in step S2, adjust the pH of the system to 2-3 to carry out the vanadium precipitation reaction, filter after the reaction is complete to obtain (NH4)2V6O 16 precipitation; Alternatively, vanadium precipitation additives are added to the vanadium-containing alkaline leachate obtained in step S2, the pH of the system is adjusted to 8-10 to carry out the vanadium precipitation reaction, and after the reaction is completed, the solution is filtered to obtain NH4VO3 precipitate. S4. Preparation of V2O5: The (NH4)2V6O obtained in step S3 16 Alternatively, NH4VO3 can be placed in an air atmosphere and heated at 450–670°C to obtain V2O5.

[0009] Furthermore, in step S1, the vanadium-containing steel slag contains 0.8-3% V2O5 and 35-60% CaO; the vanadium slag is selected from one or more of ordinary vanadium slag, vanadium-chromium slag, and high-chromium vanadium slag, and its particle size is controlled below 200 mesh.

[0010] Preferably, in step S1, the mixing ratio of the roasted material is 3 to 49.6% of the total mass of vanadium-containing steel slag.

[0011] Preferably, in step S1, the holding time for high-temperature calcination is less than 150 min, and the heating rate is 2-10 °C / min.

[0012] Preferably, in step S2, the roasted clinker is ground to a particle size ≤74μm; the acid solution for acid leaching is at least one of sulfuric acid, hydrochloric acid, or nitric acid; and the alkaline solution for alkali leaching is at least one of sodium hydroxide, sodium carbonate, sodium bicarbonate, ammonium carbonate, ammonium bicarbonate, ammonium oxalate, triammonium phosphate, or diammonium hydrogen phosphate.

[0013] Preferably, in step S3, the vanadium precipitation additive is at least one of ammonium chloride, ammonium sulfate, ammonium carbonate, or ammonium bicarbonate.

[0014] Preferably, in step S3, when a vanadium-containing acidic leaching solution is used to add a vanadium precipitation additive, the amount of the vanadium precipitation additive added is based on NH4. + The molar ratio of vanadium to vanadium in the leachate is controlled at 0.5–3; during the vanadium precipitation process, at least one of sulfuric acid, hydrochloric acid, or nitric acid is added to maintain the pH of the system at 2–3; the vanadium precipitation temperature is 60–100°C; and the vanadium precipitation reaction time is 20–180 min.

[0015] Preferably, in step S3, when a vanadium-containing alkaline leaching solution is used, the amount of the vanadium-precipitating additive added is based on NH4. +The molar ratio of vanadium to vanadium in the leachate is controlled at 2–12; the pH of the system is maintained at 8–10 before vanadium precipitation by adding acid or removing carbonate ions, the precipitation temperature is 0–60℃, and the reaction time is 20–180 min.

[0016] Preferably, the acid is at least one of sulfuric acid, hydrochloric acid, or nitric acid; the removal of carbonate ions is achieved by precipitating calcium carbonate with calcium chloride, calcium nitrate, or calcium sulfate.

[0017] Compared with existing technologies, the beneficial effects of this invention are: This invention provides a method for synergistically extracting vanadium from vanadium-containing steel slag and vanadium slag to prepare V₂O₅. Compared to traditional roasting-leaching systems using pure calcium, pure sodium, or pure manganese salts, this invention achieves simultaneous co-leaching and efficient precipitation of vanadium from vanadium slag and vanadium-containing steel slag using a simpler acid / alkali leaching-ammonium salt precipitation system. The method achieves a vanadium leaching rate of 92.57%–94.73% and a V₂O₅ purity of 98.79%–99.79%, both superior to existing vanadium extraction processes and related product quality, ensuring high purity of the product while achieving a high leaching rate.

[0018] This invention utilizes vanadium-containing steel slag, which functions as both a "low-cost / negative-cost calcium source" and a "secondary vanadium source." This significantly reduces auxiliary material consumption and costs, and effectively decreases carbon emissions during the process. By limiting the Ca / V ratio and calcination temperature, this invention reconstructs the phase composition, converting both the "encapsulated vanadium" in the steel slag and the vanadium in the vanadium slag into pyrovanadate (Ca2V2O7), thus transferring vanadium from a stable phase to a soluble phase. Furthermore, the calcium in the steel slag is directed to form pyrovanadate, rather than existing as free calcium. After calcination, the calcium is "locked" in the pyrovanadate or silicate, eliminating preferential acid consumption during acid leaching. In short, this invention transforms the high calcium content in vanadium-containing steel slag from an "obstacle" into a "resource."

[0019] This invention controls the n(P / V) ratio within a specific range in the system by rationally proportioning vanadium-containing steel slag and vanadium slag. During the leaching process, an appropriate amount of phosphate ions forms a soluble vanadium-phosphorus complex with vanadium ions, continuously reducing the concentration of free vanadium ions in the solution. Based on the principle of chemical equilibrium shift, this promotes the forward dissolution and leaching reaction of vanadium, thereby significantly enhancing the vanadium leaching behavior. Simultaneously, because n(P / V) is precisely defined, the form and stability of this complex in the system are effectively constrained, preventing it from significantly inhibiting the nucleation and growth of vanadate precipitates in the subsequent ammonium salt precipitation stage. In other words, this invention, through the design of the phosphorus-vanadium ratio at the raw material end, enables phosphate to "assist" vanadium extraction during leaching and to be "harmless" during vanadium precipitation, without introducing additional complex-breaking steps or increasing process complexity. This resolves the contradiction in traditional technologies where phosphate is "beneficial for leaching but harmful for vanadium precipitation."

[0020] Furthermore, this application, by controlling n(Ca / V) and calcination temperature, selectively generates Ca2V2O7 with amphoteric solubility, overcoming the technical limitations of "single phase → single leaching". Ca2V2O7 can preferentially dissolve under acidic conditions and can also be efficiently leached under alkaline conditions, making the same calcined product suitable for both acid and alkaline leaching processes. Enterprises can flexibly choose the leaching path according to existing equipment conditions, raw material fluctuations, and product requirements.

[0021] The integrated process of "vanadium-containing steel slag-vanadium slag co-roasting-acid / alkali leaching-ammonium salt precipitation-calcination" proposed in this invention eliminates complex extraction steps, resulting in a shorter process flow, simpler operation, and lower overall cost. While achieving efficient and high-value utilization of vanadium slag and vanadium-containing steel slag, it significantly reduces the emission of high-salt wastewater, waste residue, and corrosive gases, providing a reliable technical basis for the industrial-scale promotion of clean vanadium extraction. Attached Figure Description

[0022] Figure 1 XRD patterns of co-calcined clinker phases at different calcination temperatures; Figure 2 XRD patterns of co-calcined clinker phases with different n(Ca / V); Figure 3 The images show SEM images of vanadium-containing steel slag and vanadium slag co-roasted, where A is the SEM image with n(Ca / V)=0 and B is the SEM image with n(Ca / V)=0.5. Figure 4 This is a data analysis chart showing the vanadium leaching rate for different n(Ca / V) values ​​under acid leaching process.

[0023] Figure 5 This is a data analysis chart showing the vanadium precipitation rate for different n(Ca / V) values ​​under the acid leaching process. Detailed Implementation

[0024] The technical solutions in the embodiments of the present invention will be clearly and completely described below.

[0025] This invention provides a method for synergistically utilizing vanadium-containing steel slag and vanadium slag to extract vanadium and prepare V₂O₅. The high calcium content of vanadium-containing steel slag makes it a suitable calcium source for use in calcination roasting processes. Therefore, vanadium-containing steel slag can be used as a natural calcium salt additive, co-processed with vanadium slag, achieving vanadium co-extraction while avoiding the cost and carbon emission problems associated with purchasing calcium salts. This invention fully utilizes the high calcium content of vanadium-containing steel slag, allowing its components, such as calcium oxide, dicalcium silicate, tricalcium silicate, and dicalcium ferrite, to react with vanadium-containing spinel in the vanadium slag under high-temperature conditions, generating acid / alkali-soluble vanadates, thereby achieving phase reconstruction between vanadium slag and vanadium-containing steel slag.

[0026] However, due to the complex composition of vanadium-containing steel slag, under high-temperature roasting conditions, vanadium-iron spinel can undergo oxidation reactions with various calcium-containing phases in the vanadium-containing steel slag, as shown in the following chemical formulas (1) to (17), to generate various calcium vanadates. At the same time, manganese silicate in vanadium slag and magnesium silicate in vanadium-containing steel slag can also react with vanadium-iron spinel to generate manganese vanadate and magnesium vanadate, respectively.

[0027]

[0028] Therefore, direct co-calcination of vanadium-containing steel slag and vanadium slag is difficult to efficiently generate soluble vanadates, resulting in low application prospects. Based on this, this invention proposes a method for vanadium extraction by co-calcination of vanadium-containing steel slag and vanadium slag followed by acid / alkali leaching. This method utilizes the abundant calcium in the steel slag to drive a phase reconstruction reaction with the vanadium slag, efficiently generating soluble vanadates. The specific steps are as follows: S1. Calcination: The vanadium-containing steel slag after ball milling is thoroughly mixed with vanadium slag to obtain a mixture. The mixing ratio is calculated based on the ratio of Ca and P in the vanadium-containing steel slag to V in the mixture, with n(Ca / V) = 0.3-2 and n(P / V) = 0.018-0.12. The mixture is then calcined at a high temperature of 600-950℃, preferably 800-950℃, with a holding time of less than 150 min and a heating rate of 2-10℃ / min to obtain calcined clinker. The vanadium-containing steel slag contains 0-4% V2O5 and 15-50% CaO. The vanadium slag is selected from one or more of ordinary vanadium slag (chromium content <5%), vanadium-chromium slag, and high-chromium vanadium slag (chromium content 5-15%), and its particle size is controlled below 200 mesh.

[0029] Figure 1 The XRD pattern shows the effect of calcination temperature on co-calcined clinker, as shown below. Figure 1 As shown, when n(Ca / V) = 0.5, V₂O₃ in spinel is oxidized to form CaVO₃ in the temperature range of 300°C to 500°C. By controlling the temperature between 600°C and 950°C, vanadium-iron spinel begins to decompose, and the decomposition rate increases with increasing temperature. Under complete oxidation conditions, V(Ⅲ) in the vanadium-containing spinel phase is eventually transformed into easily acid-degraded Ca₂V₂O₇, Mn₂V₂O₇, and Mg₂V₂O₇.

[0030] Figure 2 The effect of different n(Ca / V) on the phase composition of co-calcined clinker is shown. For example... Figure 2As shown, when n(Ca / V) = 0, the vanadium-containing phases in the product are mainly Mn₂V₂O₇ and Mg₂V₂O₇. With further increases in n(Ca / V), Mn₂V₂O₇ and Mg₂V₂O₇ gradually transform into Ca₂V₂O₇. When n(Ca / V) = 2.5, some Ca₂V₂O₇ begins to transform into the insoluble Ca₃V₂O₈, and when n(Ca / V) = 3, Ca₂V₂O₇ has completely transformed into the insoluble Ca₃V₂O₈. Figure 3 As shown, when n(Ca / V) = 0, the vanadium-containing phases in the product are still mainly Mn2V2O7 and Mg2V2O7; when n(Ca / V) = 0.5, easily acid-degraded pyrovanadate has been formed, so n(Ca / V) should be controlled between 0.5 and 2.5. At this time, vanadium-containing steel slag accounts for 4.91 to 49.6% of the total mass of co-roasted clinker.

[0031] In summary, when n(Ca / V) is controlled between 0.5 and 2.5, and the calcination temperature is controlled between 600°C and 950°C, vanadium-containing steel slag reacts with vanadium-containing spinel in vanadium slag to form acid / alkali-soluble vanadates, thereby achieving phase reconstruction between vanadium slag and vanadium-containing steel slag. Vanadium-iron spinel can undergo oxidation reactions with various calcium-containing phases in vanadium-containing steel slag, as shown in the following chemical formula, to generate Ca2V2O7.

[0032] Meanwhile, manganese silicate in vanadium slag and magnesium silicate in vanadium-containing steel slag can also react with vanadium-iron spinel to generate Mn2V2O7 and Mg2V2O7, respectively. Mn2V2O7 and Mg2V2O7 are easily acid-degraded vanadates, while Ca2V2O7 is amphoteric, soluble in both acids and alkalis. Enterprises can flexibly choose between acid leaching and alkali leaching based on their equipment conditions (e.g., whether the existing leaching tank is acid-resistant or alkali-resistant).

[0033] Due to the influence of phosphate components in vanadium-containing steel slag, phosphorus partially dissolves along with vanadium during acid leaching. The dissolved phosphate anions can form soluble vanadium-phosphorus complexes with vanadium ions in the solution, continuously reducing the concentration of free vanadium ions in the system. Based on the principle of chemical equilibrium shift, this promotes the forward dissolution and leaching reaction of vanadium, thus significantly enhancing the vanadium leaching behavior. However, at the same time, this stable vanadium-phosphorus complex inhibits the nucleation and growth of vanadate precipitates, adversely affecting the vanadium precipitation process and leading to a decrease in the vanadium precipitation rate. Within the ratio range of n(Ca / V) = 0.3–2.5 (corresponding to n(P / V) = 0.018–0.15), under optimal conditions, the complexing and solubilizing effect of phosphorus significantly increases the vanadium leaching rate from 87% in the traditional calcification process to 94.73%, but the vanadium precipitation performance decreases. Further adjustment of the ratio revealed that, as... Figure 4As shown, with the increase of the calcium-vanadium ratio (phosphorus-vanadium ratio), the leaching rate first increases and then decreases, while the vanadium precipitation rate continues to decline. When n(Ca / V) is controlled between 0.3 and 2 (corresponding to n(P / V) = 0.018 to 0.12), the vanadium leaching rate is still significantly improved compared to the traditional calcification process, and the vanadium precipitation rate can reach a reasonable level of at least 92.28%. Considering both the vanadium leaching enhancement effect and the acceptability of the vanadium precipitation rate, this study determines the optimal ratio of the acid leaching system to be: n(Ca / V) = 0.3 to 2, n(P / V) = 0.018 to 0.12.

[0034] The reason for this is that the "double-edged sword" effect of phosphate in vanadium-containing steel slag on leaching and vanadium precipitation can be balanced by controlling n(Ca / V) in the raw materials, and thus controlling n(P / V). When n(Ca / V) is controlled between 0.3 and 2, corresponding to n(P / V) = 0.018 to 0.12, the total amount of phosphate in the system is at a low level. During the leaching stage, its complexing and solubilizing effect can still be fully utilized, resulting in a significantly higher vanadium leaching rate than the traditional calcification process. However, during the vanadium precipitation stage, due to the limited absolute concentration of free phosphate, it is insufficient to significantly inhibit the nucleation and growth of vanadate precipitates. Figure 5 As shown, the vanadium precipitation rate can therefore be maintained at a high level of 92.28% to 98.52%. Conversely, when n(Ca / V) increases to 2.5 (corresponding to n(P / V) = 0.15), with the increase of the total amount of phosphate introduced by the vanadium-containing steel slag, the amount of free phosphate in the solution increases. During the vanadium precipitation stage, it forms a stable complex with vanadium ions, which seriously interferes with the precipitation of ammonium salts, resulting in a significant decrease in the vanadium precipitation rate (85.4%). Therefore, this application achieves both "high leaching" and "high vanadium precipitation" by controlling n(Ca / V) within the preferred range of 0.3 to 2 (corresponding to n(P / V) = 0.018 to 0.12), thus avoiding the influence of free phosphate on vanadium precipitation.

[0035] S2, Leaching: The roasted clinker obtained in step S1 is subjected to acid leaching or alkali leaching; the roasted clinker is ground to a particle size ≤74μm.

[0036] When acid leaching is used: the acid solution used for leaching is at least one of sulfuric acid, hydrochloric acid, or nitric acid; as shown in chemical formulas (24)-(26): V2O7 4- +26H⁺=HV 10 O 28 5- +12H2O (24) HV 10 O 28 5- +5H + =5V₂O₅ + 3H₂O (25) V2O5+2H + =2VO2+ +H2O (26) Since vanadates undergo hydrolysis at specific pH levels, acid solutions are used to adjust the pH of the leaching system to 0.5–1.5 or 2–3, control the liquid-to-solid ratio to 3–20 mL / g, and maintain the leaching temperature at 10–100 °C for 10–120 min.

[0037] When alkaline leaching is used: the alkaline solution used for leaching is at least one of sodium hydroxide, sodium carbonate, sodium bicarbonate, ammonium carbonate, ammonium bicarbonate, ammonium oxalate, triammonium phosphate, or diammonium hydrogen phosphate. As shown in chemical formulas (27)-(32), where M represents Ca, Mn, or Mg.

[0038] M2V2O7+4OH - =V2O7 4- +2M(OH)2↓ (27) M2V2O7+2CO3 2- =V2O7 4- +2MCO3↓ (28) M2V2O7 + 4HCO3 2- =V2O7 4- +2MCO3↓+2CO2↑+2H2O (29) M2V2O7+2C2O4 2- =V2O7 4- +2MC2O4↓ (30) 3M2V2O7+4PO4 3- =3V2O7 4- +2M3(PO4)2↓ (31) M2V2O7 + 2HPO4 2- =V2O7 4- +2MHPO4↓ (32) A vanadium-containing alkaline leachate is obtained by leaching with an alkaline solution, controlling the liquid-to-solid ratio at 3–20 mL / g, the leaching temperature at 10–100℃, and the leaching time at 10–180 min. The calcined product Ca₂V₂O₇ of this invention can preferentially dissolve under acidic conditions and can also be efficiently dissolved under alkaline conditions, making the same calcined product suitable for both acid and alkaline leaching processes.

[0039] S3. Vanadium recovery: The vanadium precipitation additive is at least one of ammonium chloride, ammonium sulfate, ammonium carbonate or ammonium bicarbonate.

[0040] For example, chemical formulas (33)-(35): HV 10 O 28 5- +5H + =5V₂O₅ + 3H₂O (33) 3HV 10 O 28 5- +10NH4 + +5H + =5(NH4)2V6O 16 ↓+4H2O (34) 6VO2 + +2NH4 + +4H₂O=(NH₄)₂V₆O 16 ↓+8H + (35) A vanadium precipitation additive is added to the vanadium-containing acidic leachate obtained in step S2, wherein the amount of the vanadium precipitation additive added is based on NH4. + The molar ratio of vanadium to vanadium in the leachate is controlled at 0.5–3; during vanadium precipitation, the pH of the system is maintained at 2–3 by adding at least one of sulfuric acid, hydrochloric acid, or nitric acid. The precipitation temperature is 60–100℃, and the reaction time is 20–180 min. After the reaction is complete, the mixture is filtered to obtain (NH4)2V6O. 16 precipitation.

[0041] Alternatively, a vanadium precipitation additive, such as chemical formula (36), can be added to the vanadium-containing alkaline leachate obtained in step S2: V2O7 4- +2NH4 + +2H + =2NH4VO3↓+H2O (36) When a vanadium-containing alkaline leaching solution is used, a vanadium precipitation additive is added, and the amount of the vanadium precipitation additive added is based on NH4. + The molar ratio of vanadium to vanadium in the leachate is controlled at 2–12; the pH of the system is maintained at 8–10 before vanadium precipitation by adding acid or removing carbonate ions. The acid is at least one of sulfuric acid, hydrochloric acid, or nitric acid; carbonate ions are removed by precipitating calcium carbonate with calcium chloride, calcium nitrate, or calcium sulfate. The precipitation temperature is 0–60℃, and the reaction time is 20–180 min. After the reaction is complete, the mixture is filtered to obtain NH₄VO₃ precipitate. S4. Preparation of V2O5: The (NH4)2V6O obtained in step S3 16 Alternatively, NH4VO3 can be placed in an air atmosphere and heated at 450–670°C to obtain V2O5.

[0042] The leaching rate in this invention is calculated using the formula: the mass of vanadium entering the solution divided by the total mass of vanadium in the co-calcined clinker of vanadium-containing steel slag and vanadium slag. The concentration and content of vanadium in the solution are determined using the standard ferrous ammonium sulfate titration method.

[0043] Example 1 (1) After mixing vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=0.5 (corresponding to n(P / V)=0.03), place them in a box-type muffle furnace, heat them to 850℃ at a heating rate of 5℃ / min, keep them at that temperature for 2 hours, and then let them cool naturally in the furnace to obtain roasted clinker.

[0044] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). Water was added for leaching at a liquid-to-solid volume ratio of 10 mL / g. During the leaching process, sulfuric acid was continuously used to adjust and maintain the pH of the leaching solution at 2.8. The leaching temperature was 60℃ and the leaching time was 60 min. After leaching, solid-liquid separation was performed to obtain a vanadium-containing acidic leaching solution. The calculated vanadium leaching rate was 94.27%.

[0045] (3) Take 40 mL of the acidic vanadium-containing leachate and place it in a three-necked flask. Add ammonium sulfate as a vanadium precipitation additive. The amount added is based on NH4 + The molar ratio of vanadium to vanadium in the leachate was controlled at 2.2. The pH of the system was maintained at 2.5 by adding sulfuric acid dropwise, and the reaction was carried out at 98°C for 60 min to precipitate the vanadium. After the reaction was complete, the mixture was filtered to obtain (NH₄)₂V₆O. 16 precipitation.

[0046] (4) The obtained (NH4)2V6O 16 The product was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5 with a purity of 99.09% and a vanadium precipitation rate of 97.83% in this step.

[0047] Example 2 (1) After mixing vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=0.3 (corresponding to n(P / V)=0.018), place them in a box-type muffle furnace, heat them to 850℃ at a heating rate of 5℃ / min, keep them at the temperature for 2 hours, and cool them naturally with the furnace to obtain roasted clinker.

[0048] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). Water was added for leaching at a liquid-to-solid volume ratio of 10 mL / g. During the leaching process, sulfuric acid was continuously used to adjust and maintain the pH of the leaching solution at 2.8. The leaching temperature was 60℃ and the leaching time was 60 min. After leaching, solid-liquid separation was performed to obtain a vanadium-containing acidic leaching solution. The calculated vanadium leaching rate was 92.78%.

[0049] (3) Take 40 mL of the acidic vanadium-containing leachate and place it in a three-necked flask. Add ammonium sulfate as a vanadium precipitation additive. The amount added is based on NH4 + The molar ratio of vanadium to vanadium in the leachate was controlled at 2.2. The pH of the system was maintained at 2.5 by adding sulfuric acid dropwise, and the reaction was carried out at 98°C for 60 min to precipitate the vanadium. After the reaction was complete, the mixture was filtered to obtain (NH₄)₂V₆O. 16 precipitation.

[0050] (4) The obtained (NH4)2V6O 16 The product was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5 with a purity of 98.89% and a vanadium precipitation rate of 98.52% in this step.

[0051] Example 3 (1) Mix vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=2 (corresponding to n(P / V)=0.12), place them in a box-type muffle furnace, heat to 850℃ at a heating rate of 5℃ / min, keep warm for 2h, and cool naturally with the furnace to obtain roasted clinker.

[0052] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). Water was added for leaching at a liquid-to-solid volume ratio of 10 mL / g. During the leaching process, sulfuric acid was continuously used to adjust and maintain the pH of the leaching solution at 1. The leaching temperature was 60℃ and the leaching time was 60 min. After leaching, solid-liquid separation was performed to obtain a vanadium-containing acidic leaching solution. The calculated vanadium leaching rate was 94.2%.

[0053] (3) Take 40 mL of the acidic vanadium-containing leachate and place it in a three-necked flask. Add ammonium sulfate as a vanadium precipitation additive. The amount added is based on NH4 + The molar ratio of vanadium to vanadium in the leachate was controlled at 2.2. The pH of the system was maintained at 2 by dropwise addition of ammonia, and the reaction was carried out at 98°C for 60 min to precipitate the vanadium. After the reaction was complete, the mixture was filtered to obtain (NH₄)₂V₆O. 16 precipitation.

[0054] (4) The obtained (NH4)2V6O 16 The product was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5 with a purity of 98.79% and a vanadium precipitation rate of 92.28% in this step.

[0055] Example 4 (1) Mix vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=0.5 (corresponding to n(P / V)=0.03), place them in a box-type muffle furnace, heat to 900℃ at a heating rate of 5℃ / min, keep warm for 2h, and cool naturally with the furnace to obtain roasted clinker.

[0056] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). A sodium carbonate solution with a concentration of 100 g / L was used as an alkaline leaching agent, and the liquid-to-solid volume ratio was controlled at 10 mL / g. Leaching was carried out at 95 °C for 120 min. After leaching, the leachate was allowed to cool naturally, and solid-liquid separation was performed to obtain a vanadium-containing alkaline leachate with a pH value of 10.3–10.7. The vanadium leaching rate was calculated to be 94.73%.

[0057] (3) Take 40 mL of the alkaline vanadium-containing leachate and place it in a three-necked flask. Add calcium chloride to precipitate and remove carbonate ions from the solution. The amount of calcium chloride added is controlled according to the molar ratio of calcium ions to carbonate ions in the leachate of 0.8. React at 25℃ for 5 min. After filtration, high-purity calcium carbonate solid and purified vanadium-containing leachate with a pH of 9-9.5 are obtained. The obtained calcium carbonate can be used as an additive in the calcination and roasting of vanadium slag, or as a filler in the plastics, rubber, and paper industries.

[0058] (4) Take 40 mL of the purified vanadium-containing leachate into a three-necked flask, add ammonium chloride as a vanadium precipitation additive, and add it according to the amount of NH4. + The molar ratio of vanadium to vanadium in the leachate was controlled at 8. The reaction was carried out at 5°C for 120 min to precipitate, and the precipitate was obtained by filtration.

[0059] (5) The obtained NH4VO3 was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5. The purity of the product was 99.79% and the vanadium precipitation rate in this step was 90.27%.

[0060] Example 5 (1) Mix vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=0.5 (corresponding to n(P / V)=0.03), place them in a box-type muffle furnace, heat to 900℃ at a heating rate of 5℃ / min, keep warm for 2h, and cool naturally with the furnace to obtain roasted clinker.

[0061] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). A sodium carbonate solution with a concentration of 100 g / L was used as an alkaline leaching agent, and the liquid-to-solid volume ratio was controlled at 10 mL / g. Leaching was carried out at 95 °C for 120 min. After leaching, the leachate was allowed to cool naturally, and solid-liquid separation was performed to obtain a vanadium-containing alkaline leachate with a pH value of 10.3–10.7. The vanadium leaching rate was calculated to be 93.97%.

[0062] (3) Take 40 mL of the alkaline vanadium-containing leachate and place it in a three-necked flask. Add ammonium sulfate as a vanadium precipitation additive. The amount added is based on NH4 + The molar ratio of vanadium to vanadium in the leachate was controlled at 10. The pH of the system was maintained at 9 by adding sulfuric acid dropwise, and the reaction was carried out at 5°C for 120 min to precipitate. After the reaction was complete, the mixture was filtered to obtain NH₄VO₃ precipitate.

[0063] (4) The obtained NH4VO3 was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5. The purity of the product was 99.20%, and the vanadium precipitation rate in this step was 90.73%.

[0064] Example 6 (1) Mix vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=0.5 (corresponding to n(P / V)=0.03), place them in a box-type muffle furnace, heat to 900℃ at a heating rate of 5℃ / min, keep warm for 2h, and cool naturally with the furnace to obtain roasted clinker.

[0065] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). A sodium bicarbonate solution with a concentration of 80 g / L was used as an alkaline leaching agent, and the liquid-to-solid volume ratio was controlled at 10 mL / g. Leaching was carried out at 100 °C for 120 min. After leaching, the leachate was allowed to cool naturally, and solid-liquid separation was performed to obtain a vanadium-containing alkaline leachate with a pH value of about 9.5. The vanadium leaching rate was calculated to be 94.37%.

[0066] (3) Take 40 mL of the alkaline vanadium-containing leachate and place it in a three-necked flask. Add ammonium bicarbonate as a vanadium precipitation additive. The amount added is based on NH4. + The molar ratio of vanadium to vanadium in the leachate was controlled at 8. The reaction was carried out at 5°C for 120 min to precipitate. After the reaction was completed, the mixture was filtered to obtain NH4VO3 precipitate.

[0067] (4) The obtained NH4VO3 was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5. The purity of the product was 99.58%, and the vanadium precipitation rate in this step was 90.78%.

[0068] Example 7 (1) Mix vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=0.5 (corresponding to n(P / V)=0.03), place them in a box-type muffle furnace, heat to 900℃ at a heating rate of 5℃ / min, keep warm for 2h, and cool naturally with the furnace to obtain roasted clinker.

[0069] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). A sodium bicarbonate solution with a concentration of 75 g / L and a sodium hydroxide solution with a concentration of 5 g / L was used as an alkaline leaching agent, and the liquid-to-solid volume ratio was controlled at 10 mL / g. Leaching was carried out at 100 °C for 120 min. After leaching, the leachate was allowed to cool naturally, and solid-liquid separation was performed to obtain a vanadium-containing alkaline leachate with a pH value of about 9.6. The vanadium leaching rate was calculated to be 94.72%.

[0070] (3) Take 40 mL of the alkaline vanadium-containing leachate and place it in a three-necked flask. Add ammonium bicarbonate as a vanadium precipitation additive. The amount added is based on NH4. + The molar ratio of vanadium to vanadium in the leachate was controlled at 8. The reaction was carried out at 5°C for 120 min to precipitate. After the reaction was completed, the mixture was filtered to obtain NH4VO3 precipitate.

[0071] (4) The obtained NH4VO3 was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5. The purity of the product was 99.39%, and the vanadium precipitation rate in this step was 90.28%.

[0072] Example 8 (1) Mix vanadium-containing steel slag and vanadium slag evenly at n(Ca / V)=0.5 (corresponding to n(P / V)=0.03), place them in a box-type muffle furnace, heat to 900℃ at a heating rate of 5℃ / min, keep warm for 2h, and cool naturally with the furnace to obtain roasted clinker.

[0073] (2) The above-mentioned roasted clinker was crushed and ground to about 200 mesh (particle size < 74 μm). A 200 g / L ammonium bicarbonate solution was used as the leaching agent, and leaching was performed at 100 °C for 120 min. After leaching, the leachate was allowed to cool naturally, and solid-liquid separation was performed to obtain a vanadium-containing alkaline leachate with a pH of approximately 9. The calculated vanadium leaching rate was 92.57%.

[0074] (3) Take 40 mL of the alkaline vanadium-containing leachate and place it in a three-necked flask. React at 5 °C for 120 min to precipitate. After the reaction is complete, filter to obtain NH4VO3 precipitate.

[0075] (4) The obtained NH4VO3 was placed in a box furnace and calcined at 500°C for 2 hours in air atmosphere to obtain V2O5. The purity of the product was 99.53% and the vanadium precipitation rate in this step was 92.48%.

[0076] Comparative Example 1 The main difference from Example 1 is that vanadium slag and vanadium-containing steel slag are mixed at a ratio of n(Ca / V)=2.5 (corresponding to n(P / V)=0.09).

[0077] Vanadium slag and vanadium-containing steel slag were mixed at a ratio of n(Ca / V) = 2.5 (corresponding to n(P / V) = 0.09), placed in a box-type muffle furnace, and heated to 850℃ at a heating rate of 5℃ / min. After holding at this temperature for 2 hours, the mixture was allowed to cool naturally to obtain calcified roasted clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74μm), and leached with water at a liquid-to-solid volumetric mass ratio of 10mL / g. During the leaching process, the pH was continuously adjusted with sulfuric acid to maintain the pH of the leachate between 2.5 and 2.8. The leaching temperature was 60℃, and the leaching time was 60min. After leaching, the mixture was filtered to obtain a yellow vanadium-containing leachate. The calculated vanadium leaching rate was 85.33%, and the vanadium precipitation rate was 85.75%. This leachate was subjected to ammonium salt precipitation of vanadium and calcination to obtain vanadium pentoxide with a product purity of 98.17%. Compared with Example 1, the vanadium leaching rate and vanadium precipitation rate decreased by 8.94% and 12.08%, respectively.

[0078] Comparative Example 2 The main difference from Example 4 is that vanadium slag and vanadium-containing steel slag are mixed at a ratio of n(Ca / V)=2.5 (corresponding to n(P / V)=0.09).

[0079] Vanadium slag and vanadium-containing steel slag were mixed at a ratio of n(Ca / V) = 2.5 (corresponding to n(P / V) = 0.09), placed in a box-type muffle furnace, and heated to 850°C at a heating rate of 5°C / min. After holding at this temperature for 2 hours, the mixture was allowed to cool naturally to obtain calcified roasted clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74 μm), and leached with a 100 g / L sodium carbonate solution at a leaching temperature of 95°C for 120 minutes. After leaching, the solution was filtered to obtain a colorless vanadium-containing leachate, and the vanadium leaching rate was calculated to be 82.35%. Compared with Example 4, the vanadium leaching rate decreased by 12.38%.

[0080] Comparative Example 3 The main difference from Example 1 is that calcium carbonate is used instead of vanadium-containing steel slag.

[0081] Vanadium slag and calcium carbonate were mixed at a ratio of n(Ca / V) = 0.5 and placed in a box-type muffle furnace. The mixture was heated to 850°C at a heating rate of 5°C / min, held at that temperature for 2 hours, and then naturally cooled to obtain calcified roasted clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74 μm), and leached with water at a liquid-to-solid volume ratio of 10 mL / g. During the leaching process, the pH was continuously adjusted with sulfuric acid to maintain the pH of the leachate between 2.5 and 2.8. The leaching temperature was 60°C, and the leaching time was 60 min. After leaching, the mixture was filtered to obtain a yellow vanadium-containing leachate, with a calculated vanadium leaching rate of 86.55%. This leachate was then subjected to ammonium salt precipitation of vanadium and calcined to obtain vanadium pentoxide with a product purity of 98.70%. Compared with Example 1, Example 1 used vanadium-containing steel slag to replace the more expensive calcium carbonate additive, which not only reduced carbon dioxide emissions but also increased the vanadium leaching rate by 7.72%, achieving synergistic extraction of vanadium from vanadium-containing steel slag and vanadium slag.

[0082] Comparative Example 4 The main difference from Example 4 is that calcium carbonate is used instead of vanadium-containing steel slag.

[0083] Vanadium slag and calcium carbonate were mixed at a ratio of n(Ca / V) = 0.5 and placed in a box-type muffle furnace. The mixture was heated to 900°C at a heating rate of 5°C / min, held at that temperature for 2 hours, and then naturally cooled to obtain calcified roasted clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74 μm) and leached with a sodium carbonate solution with a concentration of 100 g / L at a leaching temperature of 95°C for 120 minutes. After leaching, the solution was filtered to obtain a colorless vanadium-containing leachate, and the vanadium leaching rate was calculated to be 86.81%. Compared with Examples 4 and 5, using vanadium-containing steel slag can also replace calcium carbonate additives, reducing carbon dioxide emissions, while increasing the vanadium leaching rate by 7.92%, effectively achieving the co-extraction of vanadium from vanadium-containing steel slag and vanadium slag.

[0084] Comparative Example 5 The main difference from Example 1 is that only vanadium slag is used.

[0085] Vanadium slag was placed in a box-type muffle furnace and heated to 850°C at a heating rate of 5°C / min. After holding at this temperature for 2 hours, it was naturally cooled to obtain calcined clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74 μm), and leached with water at a liquid-to-solid volumetric ratio of 10 mL / g. During the leaching process, the pH was continuously adjusted with sulfuric acid to maintain the pH of the leachate between 2.5 and 2.8. The leaching temperature was 60°C, and the leaching time was 60 min. After leaching, the solution was filtered to obtain a yellow vanadium-containing leachate. The calculated vanadium leaching rate was 61.35%. The vanadium leaching efficiency is relatively low, making industrial application difficult.

[0086] Comparative Example 6 The main difference from Example 1 is that only vanadium-containing steel slag is used.

[0087] Vanadium-containing steel slag was placed in a box-type muffle furnace and heated to 850°C at a heating rate of 5°C / min. After holding at this temperature for 2 hours, it was naturally cooled to obtain calcined clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74 μm), and leached with water at a liquid-to-solid volumetric ratio of 10 mL / g. During the leaching process, the pH was continuously adjusted with sulfuric acid to maintain the pH of the leachate between 2.5 and 2.8. The leaching temperature was 60°C, and the leaching time was 60 min. After leaching, the solution was filtered to obtain a yellow vanadium-containing leachate. The calculated vanadium leaching rate was 2.85%. The vanadium leaching efficiency is relatively low, making industrial application difficult.

[0088] Comparative Example 7 The main difference from Example 4 is that only vanadium slag is used.

[0089] Vanadium slag was placed in a box-type muffle furnace and heated to 900℃ at a heating rate of 5℃ / min. After holding at this temperature for 2 hours, it was naturally cooled to obtain calcined clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74μm), and then leached with a 100g / L sodium carbonate solution at a leaching temperature of 95℃ for 120 minutes. After leaching, the solution was filtered to obtain a colorless vanadium-containing leachate. The calculated vanadium leaching rate was 65%. The vanadium leaching efficiency is relatively low, making industrial application difficult.

[0090] Comparative Example 8 The main difference from Example 4 is that only vanadium-containing steel slag is used.

[0091] Vanadium-containing steel slag was placed in a box-type muffle furnace and heated to 900℃ at a heating rate of 5℃ / min. After holding at this temperature for 2 hours, it was naturally cooled to obtain calcined clinker. The clinker was crushed and ground to approximately 200 mesh (particle size < 74μm), and then leached with a 100g / L sodium carbonate solution at a leaching temperature of 95℃ for 120 minutes. After leaching, the solution was filtered to obtain a colorless vanadium-containing leachate. The calculated vanadium leaching rate was 20.76%. The vanadium leaching efficiency is relatively low, making industrial application difficult.

[0092] In summary, the vanadium leaching rate of the method of this invention reaches 92.57%–94.73%, and the purity of the obtained V₂O₅ is 98.79%–99.79%, all of which are significantly superior to traditional roasting processes. This invention effectively overcomes the problems of high additive costs and large carbon emissions in single vanadium slag treatment processes, as well as the difficulty in efficiently utilizing vanadium-containing steel slag. Through the co-roasting of vanadium-containing steel slag and vanadium slag, vanadium is co-extracted from both materials, while significantly improving the overall conversion rate of vanadium. The obtained high-purity V₂O₅ can be directly used in metallurgical alloys, chemical catalysis, and energy storage, while the calcium carbonate byproduct can be reused as a filler or calcination roasting additive, further enhancing the resource utilization value. The short process of "co-roasting-acid / alkali leaching-ammonium salt precipitation-calcination" constructed by this invention is characterized by its simplicity and ease of industrial implementation, reducing auxiliary material consumption and carbon emissions from the source, and providing a reliable path for the clean and high-value comprehensive utilization of vanadium resources.

[0093] The preferred embodiments of the present invention have been described in detail above; however, the present invention is not limited thereto. Within the scope of the inventive concept, various simple modifications can be made to the technical solutions of the present invention, including combinations of various technical features in any other suitable manner. These simple modifications and combinations should also be considered as the content disclosed in the present invention and are all within the protection scope of the present invention.

Claims

1. A method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag followed by acid / alkali leaching, characterized in that, The steps are as follows: S1. Calcination: The vanadium-containing steel slag after ball milling is thoroughly mixed with vanadium slag to obtain a mixture. The mixing ratio is calculated based on the ratio of Ca and P in the vanadium-containing steel slag to V in the mixture, with n(Ca / V) = 0.3~2 and n(P / V) = 0.018~0.

12. The mixture is then calcined at a high temperature of 600~950℃ to obtain calcined clinker. S2. Leaching: The roasted clinker obtained in step S1 is subjected to acid leaching or alkali leaching; When acid leaching is used, the pH of the leaching system is adjusted to 0.5-1.5 or 2-3 using an acid solution, the liquid-to-solid ratio is controlled at 3-20 mL / g, the leaching temperature is 10-100℃, and the time is 10-120 min to obtain a vanadium-containing acidic leachate. When alkaline leaching is used, the following steps are taken: leaching is performed using an alkaline solution, with the liquid-to-solid ratio controlled at 3–20 mL / g, the leaching temperature at 10–100℃, and the leaching time at 10–180 min, to obtain a vanadium-containing alkaline leachate. S3, Vanadium Recycling: Add vanadium precipitation additive to the vanadium-containing acidic leachate obtained in step S2, adjust the pH of the system to 2-3 to carry out the vanadium precipitation reaction, filter after the reaction is complete to obtain (NH4)2V6O 16 precipitation; Alternatively, vanadium precipitation additives are added to the vanadium-containing alkaline leachate obtained in step S2, the pH of the system is adjusted to 8-10 to carry out the vanadium precipitation reaction, and after the reaction is completed, the solution is filtered to obtain NH4VO3 precipitate. S4. Preparation of V2O5: The (NH4)2V6O obtained in step S3 16 Alternatively, NH4VO3 can be placed in an air atmosphere and heated at 450–670°C to obtain V2O5.

2. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag with acid / alkali leaching according to claim 1, characterized in that, In step S1, the vanadium-containing steel slag contains 0.8-3% V2O5 and 35-60% CaO; the particle size of the vanadium slag is controlled below 200 mesh.

3. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag according to claim 1, characterized in that, In step S1, the mixing ratio of the roasted material is 3 to 49.6% of the total mass of vanadium-containing steel slag.

4. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag with acid / alkali leaching according to claim 1, characterized in that, In step S1, the holding time for high-temperature calcination is less than 150 min, and the heating rate is 2-10 °C / min.

5. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag with acid / alkali leaching according to claim 1, characterized in that, In step S2, the roasted clinker is ground to a particle size ≤74μm; the acid solution for acid leaching is at least one of sulfuric acid, hydrochloric acid, or nitric acid; the alkaline solution for alkali leaching is at least one of sodium hydroxide, sodium carbonate, or sodium bicarbonate.

6. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag according to claim 1, characterized in that, In step S3, the vanadium precipitation additive is at least one of ammonium chloride, ammonium sulfate, ammonium carbonate, or ammonium bicarbonate.

7. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag with acid / alkali leaching according to claim 1, characterized in that, In step S3, when a vanadium-containing acidic leaching solution is used, a vanadium precipitation additive is added in an amount equal to NH4. + The molar ratio of vanadium to vanadium in the leachate is controlled at 0.5–3; during the vanadium precipitation process, at least one of sulfuric acid, hydrochloric acid, or nitric acid is added to maintain the pH of the system at 2–3; the vanadium precipitation temperature is 60–100°C; and the vanadium precipitation reaction time is 20–180 min.

8. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag according to claim 1, characterized in that, In step S3, when a vanadium-containing alkaline leaching solution is used, a vanadium precipitation additive is added in an amount calculated according to NH4. + The molar ratio of vanadium to vanadium in the leachate was controlled at 2–12; the pH of the system was maintained at 8–10 by adding acid before vanadium precipitation, the precipitation temperature was 0–60℃, and the reaction time was 20–180 min.

9. The method for vanadium extraction by co-roasting vanadium-containing steel slag and vanadium slag with acid / alkali leaching according to claim 8, characterized in that, The acid is at least one of sulfuric acid, hydrochloric acid, or nitric acid.

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