A method for strengthening vanadium extraction or selectively separating and recovering vanadium from vanadium-containing minerals

Abstract

The present application relates to vanadium chemical technology and resource recycling technical field, specifically relates to a kind of method for strengthening vanadium extraction or selective separation and recovery of vanadium from vanadium-containing mineral, the method is direct complexation leaching method or separation recomplexation method;Wherein, direct complexation leaching method is: in the low-temperature acid leaching step of vanadium-containing mineral, add H2O2, make it and vanadium in vanadium-containing mineral complexation reaction occurs, obtain vanadium-containing leaching solution;Separation recomplexation method is: after high-temperature acid leaching step of vanadium-containing mineral, add H2O2 in primary leaching residue, make it and vanadium in vanadium-containing mineral complexation reaction occurs, obtain secondary vanadium-containing leaching solution.The present application is leached by hydrogen peroxide complexation simultaneously, on the one hand, vanadium can be efficiently leached, on the other hand, vanadium and impurities can be selectively separated by step leaching.The present application method is high in vanadium recovery rate, realizes selective separation of vanadium and impurities, process is compact, easy to realize industrialization.

Description

Technical Field

[0001] This invention belongs to the field of vanadium chemical technology and resource recycling technology, specifically relating to a method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-containing minerals. Background Technology

[0002] Vanadium, as an important strategic metal, is increasingly widely used in steel, chemical, aerospace, and energy storage (especially vanadium redox flow batteries). In my country, vanadium resources are mainly found in vanadium-titanium magnetite, and the vanadium slag obtained after smelting is the primary raw material for vanadium extraction. Currently, the mainstream processes for extracting vanadium from vanadium slag include sodium roasting-water leaching and calcification roasting-acid leaching. However, existing technologies still face some technical bottlenecks in achieving both efficient vanadium leaching and selective separation.

[0003] During the acid leaching process, the high-valent vanadium ions (mainly VO2) that are leached out + Vanadium pentoxide (V₂O₅) readily undergoes hydrolysis in the acidic pH range of 1.5–2.5, forming insoluble vanadium pentoxide (V₂O₅) precipitate. This side reaction not only significantly reduces the vanadium leaching rate but also leaves some vanadium residue in the leaching residue, increasing the difficulty and cost of subsequent treatment. Due to the hydrolytic characteristics of pentavalent vanadium ions within a specific pH range, traditional acidic leaching processes struggle to achieve both high vanadium leaching efficiency and high leaching solution purity. Furthermore, traditional acid leaching processes produce a large number of impurity ions such as calcium, manganese, magnesium, and iron in the leaching solution, placing a heavy burden on subsequent vanadium separation and purification. Removing these impurities often necessitates ammonium salt precipitation, generating large amounts of ammonia nitrogen wastewater, which is inconsistent with the national green development strategy.

[0004] To address the aforementioned hydrolysis and selectivity issues, existing technologies often employ complex processes involving multi-stage roasting and leaching. For example, patent 202410421010.6 discloses a method for deep vanadium extraction from vanadium slag, achieving a vanadium leaching rate of up to 94.1% and a vanadium recovery rate of up to 92.52%. However, this process is complex, involving a series of steps: vanadium slag and calcium salts undergo primary oxidation roasting, primary acid leaching, secondary oxidation roasting, secondary acid leaching, solution precipitation of vanadium, and further extraction of vanadium from the vanadium-containing leachate before obtaining V₂O₅. While this can improve vanadium recovery to some extent, it results in a lengthy process, high energy and material consumption, and is detrimental to the efficient and economical utilization of resources.

[0005] Previous studies have attempted to improve vanadium leaching efficiency by adding oxidants (such as sodium chlorate and manganese dioxide). These methods mainly utilize the oxidizing properties of the oxidants to remove sparingly soluble low-valence vanadium (V₂O₃) from the minerals. 3+ / V 4+ ) is oxidized to soluble high-valence vanadium (V) 5 + However, the VO generated by the oxidation reaction2+ The ions still face the risk of hydrolysis and precipitation in acidic leachates, failing to fundamentally solve the vanadium loss problem; the Na in the oxidant + Cl - Mn 2+ Plasma can enter the leachate, increasing the difficulty and cost of subsequent separation and purification; these oxidants, while oxidizing vanadium, also oxidize other impurity ions (such as Fe). 2+ This leads to the non-selective consumption of oxidants.

[0006] The methods described above can enhance vanadium leaching, but existing methods cannot solve the problems of efficient vanadium extraction and separation recovery, which are bottleneck issues that urgently need to be addressed in this field. Therefore, it is necessary to develop a highly efficient and selective process to avoid the aforementioned problems. Summary of the Invention

[0007] This invention provides a method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-containing minerals, aiming to solve the technical problems of efficient vanadium extraction and vanadium separation and recovery.

[0008] Technical solution: This invention proposes a method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-containing minerals. The method utilizes hydrogen peroxide to react with vanadium in the vanadium-containing minerals to form a soluble vanadium peroxide complex. This includes a direct complexation leaching method during a low-temperature acid leaching stage, or a separation and re-complexation method following high-temperature acid leaching. The direct complexation leaching method is as follows: In the low-temperature acid leaching step of vanadium-containing minerals, H2O2 is added to cause it to undergo a complexation reaction with vanadium in the vanadium-containing minerals to obtain a vanadium-containing leachate. The separation and re-complexation method is as follows: after the high-temperature acid leaching step of vanadium-containing minerals, H2O2 is added to the primary leaching residue to cause a complexation reaction with the vanadium in the vanadium-containing minerals, thereby obtaining a secondary vanadium-containing leaching solution.

[0009] Furthermore, the steps of the direct complexation leaching method are as follows: S101 raw material pretreatment involves crushing and grinding vanadium-containing minerals; S102 acid leaching involves hydrogen peroxide complexation leaching in a leaching medium, wherein the leaching medium is water, and the liquid-to-solid ratio of the leaching medium to the vanadium-containing mineral is (0.5-50) L:1 kg; the vanadium-containing mineral is calculated as V, and the amount of H2O2 added satisfies n(H2O2 / V) of 0.01-4, resulting in a dark black leaching solution containing vanadium complexed with hydrogen peroxide. The pH of the leaching solution is adjusted to 0-5 using a pH adjuster, and the leaching temperature is 15-35℃. After the S103 leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leachate. The pH of the leachate is adjusted to 2-3 with ammonia water, and ammonium sulfate or ammonium persulfate is added. A precipitation reaction is carried out at 70-100℃ to obtain a solid precipitate. The S104 solid precipitate was repeatedly washed with water and anhydrous ethanol several times and dried in a vacuum oven to obtain ammonium vanadate. S105 heated the obtained ammonium vanadate in a muffle furnace and oxidized and calcined it to obtain the product V2O5.

[0010] Furthermore, the steps of the separation and reconnection method are as follows: S201 raw material pretreatment involves crushing and grinding vanadium-containing minerals; S202 acid leaching, the leaching medium is water, the liquid-solid ratio of the leaching medium to the vanadium-containing mineral is (0.5~20) L:1kg, the pH of the leaching solution is adjusted to 0~3 using a pH adjuster, the leaching temperature is 80~120℃, the leaching time is 30~600min, and a primary leaching solution and a primary leaching residue are obtained; S203 dissolves the primary leaching residue with H2O2 and water, where the primary leaching residue is V, the amount of H2O2 added satisfies n(H2O2 / V) of 0.001 to 10, the leaching temperature is 0 to 100℃, and the leaching time is 1 to 600 min, to obtain the secondary leaching solution; S204 involves heating the secondary leaching solution to precipitate vanadium at a temperature of 30–100°C for 60–100 minutes to obtain the vanadium-precipitated product V5O. 12 • 6H2O; or add ammonium sulfate to the secondary leachate to adjust the pH of the system to 2.0–2.5, and vanadium precipitation is carried out to obtain the vanadium precipitation product (NH4)2V6O. 16 & NH4V4O 10 ; S205 will precipitate vanadium product V5O 12 • 6H₂O or (NH₄)₂V₆O 16 & NH4V4O 10 After calcination, V2O5 is obtained.

[0011] Furthermore, the pH adjuster can be sulfuric acid, hydrochloric acid, nitric acid, citric acid, oxalic acid, or phosphoric acid.

[0012] Furthermore, the vanadium-containing minerals include one or more of the following: vanadium slag roasting products, vanadium-containing steel slag, vanadium extraction tailings, vanadium-containing steel slag roasting products, vanadium-rich slag roasting products obtained by reducing vanadium slag to remove iron and separating vanadium iron, and vanadium-titanium slag roasting products obtained by reducing vanadium-titanium magnetite concentrate to remove iron and separate vanadium titanium slag.

[0013] Furthermore, the vanadium slag is one or more of ordinary vanadium slag, vanadium-chromium slag, or high-calcium and high-phosphorus vanadium slag; the vanadium extraction tailings are one or more of calcification vanadium extraction tailings, sodium vanadium extraction tailings, and manganese vanadium extraction tailings; the vanadium slag roasting product, the vanadium-rich slag roasting product obtained by reducing vanadium slag to remove iron and separating vanadium iron, and the vanadium-titanium slag roasting product obtained by reducing vanadium-titanium magnetite concentrate to remove iron are one or more of blank roasting products, sodium roasting products, calcification roasting products, manganese roasting products, magnesia roasting products, calcium-manganese composite roasting products, calcium-magnesium composite roasting products, sodium-calcium roasting products, and roasting products, as well as roasting products of various additive combinations thereof.

[0014] Furthermore, the H2O2 is a solid powder or a liquid.

[0015] Beneficial effects: This invention provides a method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-containing minerals. By introducing hydrogen peroxide as a complexing agent, it effectively solves the technical problem that pentavalent vanadium ions easily hydrolyze to form insoluble vanadium pentoxide precipitate under acidic conditions during traditional acid leaching. Specifically, it includes two process methods: a direct complexing leaching method and a separation-re-complexing method.

[0016] The direct complexation leaching method significantly improves vanadium leaching efficiency by forming a stable hydrogen peroxide-vanadium complex with hydrogen peroxide during the low-temperature acid leaching step. The separation-re-complexation method, after the high-temperature acid leaching step, adds hydrogen peroxide to re-complex and dissolve the vanadium present in the primary leaching residue as vanadium pentoxide, thus mitigating vanadium loss caused by hydrolysis. Experimental data show that the method of this invention can achieve a vanadium leaching rate of over 96% and a total vanadium recovery rate of over 95%, which is superior to traditional acid leaching processes. Furthermore, the impurities in the leachate are less than 0.5 g / L.

[0017] This invention achieves highly efficient and selective separation of vanadium from impurity elements through rational control of leaching conditions. In the direct complexation leaching method, low-temperature leaching effectively suppresses the co-leaching of impurity ions such as iron, manganese, and magnesium, resulting in a high-purity vanadium-containing leachate. In the separation-re-complexation method, high-temperature acid leaching selectively enriches vanadium in the leaching residue, separating it from impurities. Subsequent heating and vanadium precipitation yields the vanadium product. Notably, the separation-re-complexation method eliminates the need for ammonium salts; vanadium precipitates directly from the peroxyvanadic acid solution upon heating, avoiding the large amounts of ammonia nitrogen wastewater generated by traditional ammonium salt vanadium precipitation processes, thus offering significant environmental advantages.

[0018] The method of this invention features a compact process flow, making it easy to implement in industrial production. Compared to the complex processes of existing technologies that require multiple roasting and leaching stages, the direct complexation leaching method of this invention requires only one complexation leaching step to obtain vanadium-containing leachate, and the separation-re-complexation method requires only two wet processing steps to complete the separation and recovery of vanadium, significantly reducing energy and material consumption. Furthermore, the method of this invention has good adaptability to various vanadium-containing raw materials, including but not limited to vanadium slag roasting products, vanadium-containing steel slag, vanadium extraction tailings, vanadium-rich slag roasting products, and vanadium-titanium slag roasting products, and is applicable to various roasting methods such as blank roasting, sodium roasting, and calcification roasting. The novel process proposed in this invention achieves efficient utilization of vanadium slag and efficient vanadium separation metallurgy, possessing broad industrial application prospects. Attached Figure Description

[0019] Figure 1 This is a flowchart of the present invention; Figure 2 This is a graph showing the relationship between pH and leaching rate in this invention; Figure 3 This is a graph showing the relationship between temperature and leaching rate in this invention; Figure 4 This is a graph showing the relationship between n (H2O2 / V) and leaching rate in this invention; Figure 5 The image shows the detection results of the primary leaching residue in Route 2 of this invention, where A is a SEM (Scanning Electron Microscope) morphology image and B is an EDS (Energy Dispersive Spectroscopy) surface scan image. Figure 6 The image shows the detection results of the secondary leaching residue in Route 2 of this invention. In the image, A is a scanning electron microscope (SEM) morphology image, and B is an EDS energy dispersive spectroscopy (EDS) image. Detailed Implementation

[0020] To further understand the present invention, preferred embodiments of the present invention are described below in conjunction with examples. However, it should be understood that these descriptions are only for further illustrating the features and advantages of the present invention, and not for limiting the scope of the claims of the present invention.

[0021] This invention proposes a method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-containing minerals, utilizing the complexing ability of H2O2, such as... Figure 1 The method is either direct complexation leaching (route 1) or separation and re-complexation (route 2).

[0022] Route 1 of this invention, the direct complexation leaching method, involves adding H₂O₂ during the low-temperature acid leaching step of the vanadium-containing minerals. This H₂O₂ undergoes a complexation reaction with the vanadium in the vanadium-containing minerals to obtain a vanadium-containing leachate. Route 1 directly leaches vanadium from the vanadium-containing minerals via hydrogen peroxide complexation acid leaching, efficiently extracting vanadium and preventing hydrolysis and precipitation of vanadium in an acidic system. The specific steps are as follows:

[0023] S101 raw material pretreatment involves crushing and grinding vanadium-containing minerals, preferably with a particle size of less than 74μm, i.e., about 200 mesh.

[0024] Among them, the vanadium-containing minerals are one or more of the following: vanadium slag roasting products, vanadium-containing steel slag, vanadium extraction tailings, vanadium-containing steel slag roasting products, vanadium-rich slag roasting products obtained by reducing vanadium slag to remove iron and separating vanadium iron, and vanadium-titanium slag roasting products obtained by reducing vanadium-titanium magnetite concentrate to remove iron.

[0025] The vanadium slag may be one or more of the following: 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.

[0026] The vanadium extraction tailings can be calcified vanadium extraction tailings, sodium vanadium extraction tailings, or manganese vanadium extraction tailings.

[0027] The required vanadium slag roasting products, vanadium-containing steel slag roasting products, vanadium-rich slag roasting products obtained by reducing vanadium slag to remove iron and separating vanadium iron, and vanadium-titanium slag roasting products obtained by reducing vanadium-titanium magnetite concentrate to remove iron can be one or more of the following: blank roasting products, sodium-modified roasting products, calcified roasting products, manganese-modified roasting products, magnesian-modified roasting products, calcium-manganese composite roasting products, calcium-magnesium composite roasting products, sodium-calcium roasting products, and roasting products, as well as roasting products with various additive combinations thereof.

[0028] S102 acid leaching involves hydrogen peroxide complexation leaching in a leaching medium, which is water. The pH of the leaching solution is then adjusted to 1-3 using a pH adjuster, and the leaching temperature is 15-35°C. The leaching medium and vanadium-containing minerals are leached at a liquid-to-solid ratio of (0.5-50) L:1 kg, preferably (1-20) L:1 kg. The amount of H2O2 added satisfies n(H2O2 / V) being 0.01-4, preferably n(H2O2 / V) being 1-4, resulting in a dark black leaching solution containing vanadium complexed with hydrogen peroxide.

[0029] During the acid leaching reaction, vanadium in vanadium-containing minerals is initially released as VO2. + It exists in the form of H2O2, and due to the presence of H2O2, the following complexation reaction will occur:

[0030] VO2 + + H2O2 = VO(O2) + + H2O (1)

[0031] VO(O2) + +H₂O₂ = VO(O₂)₂ - +2H + (2)

[0032] VO(O2)2 - + H2O2 = VO(O2)3 3-+ 2H + (3)

[0033] According to Le Chatelier's principle, the presence of H2O2 reduces the amount of product VO. 2+ The concentration of vanadium will significantly promote the acid leaching of vanadium ions, thereby improving the leaching efficiency of vanadium.

[0034] In the absence of hydrogen peroxide, under conditions of pH 1–2.5, the following hydrolysis reaction of vanadium will occur:

[0035] 2VO2 + +H₂O=V₂O₅(s)+2H + (4)

[0036] This inhibits the leaching efficiency of vanadium. Therefore, it is evident that hydrogen peroxide-enhanced leaching in this step can reduce vanadium leaching efficiency. 2+ It complexes to form VO(O2). + VO(O2) 2- and VO(O2)3 3- This promotes the leaching of vanadium while preventing its hydrolysis.

[0037] Based on the above principles, hydrogen peroxide can enhance leaching and prevent vanadium hydrolysis. However, due to the physical properties of hydrogen peroxide, there are still many technical problems in introducing it into the field of acid vanadium leaching. For example, hydrogen peroxide is sensitive to changes in the pH of the system. If the pH is too low, the stability of the vanadium peroxide complex decreases; if the pH is too high, the risk of vanadium hydrolysis and precipitation increases, and hydrogen peroxide itself is also more prone to disproportionation reactions. In the high-temperature environment of conventional acid leaching, hydrogen peroxide is easily hydrolyzed, while low temperatures reduce the leaching reaction rate. It is also important to avoid the formation of a Fenton-like system that consumes H2O2.

[0038] Based on this, the present invention, through in-depth research, has found that pH value, temperature, and hydrogen peroxide addition have significant effects on vanadium leaching under acid leaching conditions. The following analysis focuses on the calcined and roasted products of vanadium slag.

[0039] In the low-temperature acid leaching step, n(H₂O₂ / V) = 3, the leaching temperature was 15℃, and the pH was adjusted to 0–5 to obtain the leaching rates of V, Mn, Mg, and Fe. The relationship between pH and leaching rate is shown in [reference needed]. Figure 2 ,from Figure 2 As can be seen, the leaching efficiency of V, Mn, Mg, and Fe decreases with increasing pH. When the pH is between 0 and 3, the vanadium leaching rate can reach over 94%, especially at pH 1.5, where the vanadium leaching rate reaches 96.49%. Furthermore, Fe in the calcination product is mainly in the form of Fe2+. 3+Since it exists in its original form, it will not produce a Fenton-like system and will not consume H2O2. When the pH is less than 1, the leaching efficiency will further increase, but impurities will also increase further, so increasing the acidity is not very meaningful. When the pH is greater than 3, the acidity is insufficient to efficiently decompose vanadates, so a pH value of 1 to 3 is preferred.

[0040] In the low-temperature acid leaching step, n(H₂O₂ / V) = 3, the leaching pH = 1.5, and the leaching rates of V, Mn, Mg, and Fe were obtained under leaching temperatures ranging from 15 to 55 °C. The relationship between temperature and leaching rate is shown in [reference needed]. Figure 3 ,from Figure 3 As can be seen, with increasing temperature, the leaching efficiency of V and Fe decreases, while the leaching rates of Mn and Mg increase. When the temperature is between 15-35℃, the vanadium leaching rate can reach over 92%, especially at 15℃, where the vanadium leaching rate reaches 96.49%. The reason for this is that increasing the temperature promotes the decomposition of hydrogen peroxide and the hydrolysis reaction of vanadium, thus reducing the vanadium leaching efficiency.

[0041] In the low-temperature acid leaching step, with a leaching pH of 1.5 and a leaching temperature of 15℃, the leaching rates of V, Mn, Mg, and Fe were obtained under the condition that n(H2O2 / V) was adjusted from 0 to 4. The relationship between n(H2O2 / V) and the leaching rate is shown in [reference needed]. Figure 4 ,from Figure 4 As can be seen, the addition of hydrogen peroxide can promote the leaching of vanadium. Preferably, when n(H2O2 / V) is 1-4, the leaching rate of vanadium is at least 92%, especially when n(H2O2 / V) = 3, the leaching rate of vanadium is 96.49%. The leaching rates of Mn, Mg, and Fe are not significantly affected. The addition of H2O2 specifically promotes the leaching of vanadium, with little effect on the leaching rates of Mn, Mg, and Fe (i.e., it has good selectivity). However, when n(H2O2 / V) is greater than 4, further increasing it has limited effect on improving the vanadium leaching rate, and instead increases the cost.

[0042] In summary, under low-temperature acid leaching conditions, this invention achieves a high vanadium leaching rate (96%) and moderate impurity suppression by controlling the pH value; it also controls the temperature to inhibit H2O2 decomposition and prevent vanadium hydrolysis, while simultaneously achieving low-temperature energy saving; and by controlling the amount of hydrogen peroxide added, it provides sufficient complexing agent to protect vanadium and promote efficient vanadium leaching. This fundamentally solves the three major technical challenges of easy vanadium hydrolysis, easy H2O2 decomposition, and severe co-leaching of impurities in traditional acid leaching.

[0043] After the S103 leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leachate. The pH of the leachate is adjusted to 2-3 with ammonia water, and ammonium sulfate or ammonium persulfate is added. The precipitation reaction is carried out at 70-100℃ for 30-120 min to obtain a solid precipitate.

[0044] The S104 solid precipitate was repeatedly rinsed with water and anhydrous ethanol several times, and then dried in a vacuum oven at about 60°C for about 24 hours to obtain ammonium vanadate.

[0045] S105 heated the obtained ammonium vanadate in a muffle furnace at 450-550℃ for 20-120 min to obtain the product V2O5 by oxidation and calcination.

[0046] Route 2 of this invention, the separation and re-complexation method, is as follows: After the high-temperature acid leaching step of the vanadium-containing minerals, H2O2 is added to the primary leaching residue to induce a complexation reaction with the vanadium in the vanadium-containing minerals, resulting in a secondary vanadium-containing leachate. In Route 2, after the acid leaching vanadium extraction process, the complexing agent H2O2 is introduced to dissolve the leaching residue, relieving hydrolysis and reversing the formation reaction of solid V2O5, causing it to redissolve. The specific steps are as follows:

[0047] S201 is the same as step S101;

[0048] S202 acid leaching is performed using water as the leaching medium. The pH is adjusted to 0–3, preferably 1.5–2.4, with acid used during leaching. The leaching temperature is 80–120℃, and the leaching time is 30–600 min, preferably 60–600 min. The liquid-to-solid ratio of the leaching medium to the vanadium-containing mineral is (0.5–20) L:1 kg, preferably (1–5) L:1 kg. This yields a primary leaching solution and a primary leaching residue.

[0049] Depend on Figure 3 It is known that increasing the temperature decreases the leaching efficiency of vanadium, but increases the leaching efficiency of impurities. Therefore, under conditions of pH 1.5, when the leaching temperature is as high as 90℃ or above, the following reaction will occur when vanadate is decomposed with sulfuric acid, due to the hydrolysis reaction of vanadium:

[0050] 2VO2 + +H₂O=V₂O₅(s)+2H + (5)

[0051] This leads to the following overall reaction:

[0052] 1.667Ca 1.4 Mn 0.6 V₂O₇ + 6.667H + + 2.333SO4 2- = 2.333CaSO4 + Mn 2+ + 1.667V2O5+ 3.333H2O(6)

[0053] That is, some vanadium exists as solid V₂O₅ in the primary leaching residue. SEM and EDS analyses were performed on samples of the primary leaching residue; the results are shown below. Figure 5 ,from Figure 5As can be seen, vanadium enters the tailings in the form of V2O5. After solid-liquid separation, vanadium and impurities are separated into solid and liquid components. Vanadium is in the primary leaching residue, while impurities are in the primary leaching solution.

[0054] When dissolving the primary leaching residue with hydrogen peroxide and water in S203, the molar ratio of hydrogen peroxide to vanadium in the primary leaching residue is 0.001–10:1, the leaching temperature is 0–100℃, preferably 0–20℃, and the leaching time is 1–600 min, preferably 10–60 min. A secondary leaching solution is obtained.

[0055] The following chemical reaction occurs when the primary leaching residue is dissolved in hydrogen peroxide and water:

[0056] V₂O₅ + 2H₂O₂ = 2VO(O₂) + +H₂O + 2OH⁻ - (7)

[0057] V₂O₅ + 4H₂O₂ = 2VO(O₂)₂ - ++3H2O+2H + (8)

[0058] V₂O₅ + 6H₂O₂ = 2VO(O₂)₃ 3- +3H₂O+6H + (9)

[0059] The obtained secondary leaching residue was analyzed by SEM and EDS, and the results are shown in the figure. Figure 6 ,from Figure 6 The results show that the vanadium content in the secondary tailings is significantly reduced, which also confirms that the vanadium in the primary leaching residue is effectively dissolved by hydrogen peroxide.

[0060] S204 involves heating the secondary leaching solution to precipitate vanadium at a temperature of 30–100°C. Hydrogen peroxide begins to decompose at temperatures above 30°C, preferably 80–100°C. This temperature is maintained for 60–100 minutes to obtain the vanadium-precipitated product V5O. 12 • 6H2O; or add ammonium sulfate to the secondary leachate to adjust the pH of the system to 2.0–2.5, and vanadium precipitation is carried out to obtain the vanadium precipitation product (NH4)2V6O. 16 & NH4V4O 10 ;

[0061] The secondary leachate is a vanadium peroxide solution. The secondary leachate is directly heated to 80–100°C, causing hydrogen peroxide to decompose. Vanadium then hydrolyzes to form V₂O₅, as shown in the following reaction equation:

[0062] 2H₂O₂ = 2H₂O + O₂(g) (ΔG) T ° = - 189.32–0.057 T kJ / mol; ΔG° 25°C = - 206.31kJ / mol) (10)

[0063] VO(O2)3 3- +4H + =VO(O2) + +O2(g)+2H2O (11)

[0064] VO(O2) + +4H + =VO 2+ +0.75O2(g)+0.5H2O (ΔG° 25°C = -137.74 kJ / mol) (12)

[0065] 2VO(O2) + = 2VO2 + + O2(g) (ΔG° 25°C = -140.92 kJ / mol) (13)

[0066] 4VO2 + (aq) + VO 2+ + 9H2O = V5O 12 • 6H₂O(s) + 6H + (14)

[0067] Increased temperature promotes the decomposition of hydrogen peroxide (see reactions 10 and 13), which in turn weakens its complexing ability with vanadium (see reactions 11 and 12). Due to the increased temperature, vanadium undergoes hydrolysis (see reaction 14), releasing vanadium as V₅O. 12 Vanadium precipitates as a solid phase in the formation of 6H2O.

[0068] S205 vanadium precipitation product V5O 12 • 6H₂O or (NH₄)₂V₆O 16 & NH4V4O 10 After calcining at around 550℃ for about 2 hours, V2O5 is obtained.

[0069] V₂O₅ is obtained after calcination, and the reaction formula is as follows:

[0070] 4V5O 12 • 6H2O+O2(g)=10V2O5+12H2O (15)

[0071] Vanadium precipitated product V5O 12 • When 6H2O is heated in air with the presence of oxygen, it undergoes dehydration and oxidation (see reaction formula 15), eventually forming the V2O5 product.

[0072] In both pathways 1 and 2, the pH adjuster can be any acid, including sulfuric acid, hydrochloric acid, nitric acid, citric acid, oxalic acid, phosphoric acid, etc. The hydrogen peroxide can be a powder or a liquid, preferably a liquid.

[0073] Example 1 S101 crushes and grinds the calcined vanadium slag product to about 200 mesh, that is, to a particle size of less than 74μm.

[0074] S102 was leached with water at a liquid-to-solid ratio of 10 mL / g for the calcined product. During leaching, sulfuric acid was used to adjust the pH, with n(H2O2 / V) set to 3, maintaining the pH of the leachate at 1.5. The leaching temperature was 15℃, and the leaching time was 45 min. After leaching, the solid and liquid were separated by filtration, yielding a leachate containing vanadium complexed with hydrogen peroxide. The leachate was yellow in color, and the calculated vanadium leaching rate was 96.49%.

[0075] S103 vanadium-containing leachate was taken, the pH was adjusted to 2-3 with ammonia water, ammonium sulfate or ammonium persulfate was added, and a precipitation reaction was carried out at 95℃ for 60 min to obtain a solid precipitate with a vanadium precipitation efficiency of 99.87%.

[0076] The S104 solid precipitate was repeatedly washed with water and anhydrous ethanol several times, and dried in a vacuum oven at 60°C for 24 hours to obtain ammonium vanadate.

[0077] S105 involves heating ammonium vanadate in a muffle furnace and oxidizing and calcining it at 450–550 °C for 20–120 min to obtain V2O5, with a total vanadium recovery rate of 96.36%.

[0078] Example 2 S101 crushes and grinds the vanadium-containing steel slag roasting product to about 200 mesh, that is, to a particle size of less than 74μm;

[0079] S102 was leached with water at a liquid-to-solid ratio of 10 mL / g for the calcined product. During leaching, sulfuric acid was used to adjust the pH, with n(H2O2 / V) being 3, to keep the pH of the leaching solution at 0. The leaching temperature was 15℃, and the leaching time was 45 min. After leaching, the solid and liquid were separated by filtration, and a vanadium-containing leaching solution complexed with hydrogen peroxide was obtained. At this time, the vanadium-containing leaching solution was black in color, and the vanadium leaching rate was calculated to be 70.27%.

[0080] Steps S103 to S105 are the same as in Example 1, and the vanadium precipitation efficiency is 99.81%.

[0081] Example 3 S101 crushes and grinds the vanadium extraction tailings to a particle size of less than 74μm.

[0082] S102 was leached with water at a liquid-to-solid ratio of 10 mL / g for vanadium extraction tailings. Sulfuric acid was used to adjust the pH during leaching to maintain the pH at 0. The leaching temperature was 20℃, the leaching time was 30 min, and n(H2O2 / V) was 1. After leaching, the solid and liquid were separated by filtration, yielding a vanadium-containing leachate, which was dark black. The calculated vanadium leaching rate was 80.64%.

[0083] Take 100 mL of vanadium-containing leachate from S103, adjust its pH to 2 with ammonia, add ammonium sulfate or ammonium persulfate, and carry out a precipitation reaction at 95℃ for 60 min.

[0084] S104 separated the reaction product by filtration, washed repeatedly with water and anhydrous ethanol several times, and dried in a vacuum oven at 60°C for 24 hours to obtain ammonium vanadate.

[0085] S105 heated the obtained ammonium vanadate in a muffle furnace at 550℃ for 120 min, and oxidized and calcined it to obtain V2O5 with a purity of 99.11% and a total vanadium recovery rate of 79.24%.

[0086] Example 4 S101 crushes and grinds the calcified roasted product of vanadium-rich slag obtained by reducing vanadium slag to remove iron and separate vanadium iron. The product has a particle size of less than 74μm.

[0087] S102 was leached with water at a liquid-to-solid ratio of 10 mL / g for the calcined product. During leaching, sulfuric acid was used to adjust the pH to maintain the pH of the leachate at 1.5. The leaching temperature was 15℃, the leaching time was 45 min, and n(H2O2 / V) was 4. After leaching, the solid and liquid were separated by filtration, and a vanadium-containing leachate was obtained. At this time, the vanadium-containing leachate was yellow, and the vanadium leaching rate was calculated to be 97.24%.

[0088] Steps S103 to S105 are the same as in Example 1.

[0089] Example 5 S101 crushes and grinds the calcined vanadium slag product to about 200 mesh, that is, to a particle size of less than 74μm.

[0090] S102 was leached with water at a liquid-to-solid ratio of 0.5 mL / g for the calcined product. Sulfuric acid was used to adjust the pH of the leachate to maintain it at 1.5, with n(H2O2 / V) at 0.01. The leaching temperature was 15℃, and the leaching time was 45 min. After leaching, the solid and liquid were separated by filtration, yielding a leachate containing vanadium complexed with hydrogen peroxide. The leachate was yellow at this point, and the calculated vanadium leaching rate was 89.34%.

[0091] Steps S103 to S105 are the same as in Example 1.

[0092] Compared to Example 1, the liquid-to-solid ratio and hydrogen peroxide dosage were reduced, resulting in a lower vanadium leaching rate.

[0093] Example 6 S101 crushes and grinds the calcified roasted product of vanadium-titanium magnetite concentrate obtained by reduction and iron removal to a particle size of less than 74μm.

[0094] S102 was leached with water at a liquid-to-solid ratio of 10 mL / g for the calcined product. During the leaching process, sulfuric acid was used to adjust the pH to 0, n(H2O2 / V) to 1, the leaching temperature to 15℃, and the leaching time to 120 min. The vanadium leaching efficiency was 90.27%.

[0095] Steps S103 to S105 are the same as in Example 1.

[0096] To demonstrate the high conversion efficiency of vanadium by the method of this invention, the vanadium slag was directly roasted without additives, and a direct acid leaching experiment was conducted on the blank roasted product, as shown in Example 7.

[0097] Example 7 S101 directly roasts vanadium slag without additives to obtain blank roasted clinker. The blank roasted vanadium slag product is crushed and ground to a particle size of less than 74μm.

[0098] S102 was leached with water at a liquid-to-solid ratio of 5 mL / g for the calcined product. During leaching, sulfuric acid was used to adjust the pH to maintain the pH of the leachate at 1.5. The leaching temperature was 90℃, the leaching time was 60 min, and n(H2O2 / V) was 3. After leaching, the solid and liquid were separated by filtration, and a vanadium-containing leachate was obtained. The vanadium-containing leachate was initially black and then turned yellow. The vanadium leaching rate was calculated to be 90.35%.

[0099] Control group experiment: Unlike Example 7, no hydrogen peroxide was added during leaching. In this case, the vanadium leaching rate was 10.31% when the leaching pH was 1.5 and 50.24% when the leaching pH was 2.8 (conventional leaching pH).

[0100] In a direct acid leaching experiment on blank roasted products, the leaching rate of vanadium from vanadium without additives was investigated using the method of the present invention. The leaching rate of the method of the present invention (Example 7) was 90.35%, which was about 80% higher than 10.31% (pH 1.5), and about 40% higher than the conventional leaching rate of 50.24% (pH 2.8). This demonstrates that the method of the present invention has a high efficiency in converting vanadium in blank roasted vanadium slag.

[0101] Comparative Example 1 The vanadium slag calcination product was crushed and ground to approximately 200 mesh (particle size less than 74 μm). Water was added at a liquid-to-solid ratio of 10 mL / g for leaching. Sulfuric acid was used to adjust the pH, maintaining the leachate pH between 2.5 and 2.8. The leaching temperature was 60°C, and the leaching time was 60 min. After leaching, the solid and liquid were separated by filtration, yielding a vanadium-containing leachate, which was yellow in color. The calculated vanadium leaching rate was 87.15%. The obtained vanadium-containing leachate was then subjected to ammonium salt precipitation and calcination to obtain V₂O₅, with a total vanadium recovery rate of 85.07%. Compared to Example 1, the hydrogen peroxide complexation leaching process resulted in an approximately 11% higher leaching rate than the traditional non-reducing leaching process. The total vanadium recovery rate was increased by approximately 10%.

[0102] Comparative Example 2 The only difference between Comparative Example 2 and Example 1 is that during the S101 leaching process, only sulfuric acid was added, without hydrogen peroxide. The vanadium leaching efficiency was 88.88%, which is lower than that of Example 1. This highlights the advantage of adding hydrogen peroxide.

[0103] In summary, Route 1 of this invention, by adding hydrogen peroxide under low-temperature acidic conditions, forms a stable vanadium peroxide complex with vanadium, effectively inhibiting the hydrolysis and precipitation of vanadium, thus achieving efficient and selective leaching of vanadium. This method has a short process flow, low energy consumption, a vanadium leaching rate of over 96%, a total recovery rate of over 95%, and is applicable to a variety of vanadium-containing mineral raw materials.

[0104] Example 8 S201 crushes and grinds the calcification roasted product of vanadium slag to a particle size of less than 74μm.

[0105] S202 was leached with water and roasted product liquid-solid ratio of 10 mL / g. During the leaching process, sulfuric acid was used to adjust the pH to keep the pH of the leaching solution at 1.8. The leaching temperature was 95℃ and the leaching time was 120 min. After the leaching was completed, the solid and liquid were separated to obtain a primary leaching residue. At this time, the vanadium leaching efficiency was only 3.10%, and more than 96% of the vanadium was retained in the solution.

[0106] The S203 primary leaching residue was dissolved in a mixture of hydrogen peroxide and water with n(H2O2 / V) of 3 to obtain a hydrogen peroxide and vanadium mixture; the vanadium leaching efficiency was 90.26%.

[0107] Take 100 mL of vanadium-containing leaching solution (S204), heat to 100℃ and maintain for 100 min to obtain vanadium precipitation product, or add ammonium sulfate to adjust the pH of the system to 2.0–2.5 for vanadium precipitation; the vanadium precipitation efficiency is 99.13%.

[0108] The S205 vanadium precipitate was calcined at 550℃ for 2 hours to obtain V2O5.

[0109] Example 9 S201 crushes and grinds the calcified vanadium extraction tailings to a particle size of less than 74μm.

[0110] S202 was leached with water and vanadium extraction tailings at a liquid-to-solid ratio of 5 L:1 kg. Sulfuric acid was used to adjust the pH of the leachate to 2.0. The leaching temperature was 90℃, and the leaching time was 300 min, yielding a primary leachate and a primary leaching residue. At this point, the vanadium leaching efficiency was only 2.56%.

[0111] S203 dissolves the primary leaching residue in water with H2O2. Based on the volume (V) in the residue, n(H2O2 / V) is 5. The leaching temperature is 25℃, and the leaching time is 30 min, yielding a secondary leaching solution (peroxyvanadic acid solution, yellow). At this point, the vanadium leaching efficiency is 91.35%.

[0112] S204 involves heating the secondary leaching solution at 90°C for 100 minutes to precipitate vanadium, yielding the precipitated vanadium product V5O. 12 • 6H₂O; Vanadium precipitation efficiency is 98.67%;

[0113] S205 calcined the vanadium precipitate V2O5•6H2O at 550℃ for 2 hours to obtain V2O5.

[0114] Example 10 S201 crushes and grinds the blank roasted product of vanadium slag (roasted without additives) to a particle size of less than 74μm.

[0115] S202 was leached with water and roasted product at a liquid-to-solid ratio of 3 L:1 kg. The pH was adjusted with hydrochloric acid to maintain the pH of the leachate at 1.5. The leaching temperature was 100℃, and the leaching time was 480 min, yielding a primary leachate and a primary leaching residue. At this point, the vanadium leaching efficiency was only 4.23%.

[0116] S203 dissolves the primary leaching residue in water with H2O2. Based on the volume (V) in the residue, n(H2O2 / V) is 8. The leaching temperature is 10℃ (low temperature inhibits H2O2 decomposition), and the leaching time is 20 min, yielding a secondary leaching solution. At this point, the vanadium leaching efficiency is 92.56%.

[0117] S204 involves adding ammonium sulfate to the secondary leaching solution to adjust the pH of the system to 2.0–2.5, followed by vanadium precipitation at 80°C to obtain the vanadium-precipitated product V2O5•H2O; at this point, the vanadium precipitation efficiency is 98.97%.

[0118] S205 calcined the vanadium precipitate V2O5•H2O at 550℃ for 2 hours to obtain V2O5.

[0119] Comparative Example 3 The vanadium slag calcination product was crushed and ground to approximately 200 mesh, i.e., a particle size of less than 74 μm. Water was added at a liquid-to-solid ratio of 10 mL / g for leaching. Sulfuric acid was used to adjust the pH during leaching, maintaining the pH of the leachate at 1.5 throughout. The leaching temperature was 60°C, and the leaching time was 60 min. After leaching, the solid and liquid were separated by filtration, yielding a vanadium-containing leachate and a primary leaching residue. At this stage, the vanadium-containing leachate was light yellow. Due to the hydrolysis of vanadium, vanadium precipitated as V₂O₅, and the calculated vanadium leaching rate was 16.34%. The primary leaching residue was dissolved in hydrogen peroxide and water, following the same procedure as in Example 8, with an additional vanadium recovery efficiency of 80.12%. The vanadium leaching rate was further increased because hydrogen peroxide complexed with vanadium.

[0120] Comparative Example 3 shows that hydrogen peroxide can inhibit hydrolysis by dissolving the hydrolysis products.

[0121] Comparative Example 4 S201 crushes and grinds the calcification roasted product of vanadium slag to a particle size of less than 74μm.

[0122] S202 was leached with water and roasted product in a liquid-solid ratio of 10 L: 1 kg. The pH was adjusted with sulfuric acid to maintain the pH of the leachate at 1.8. The leaching temperature was 95℃ and the leaching time was 120 min, resulting in a primary leachate and a primary leaching residue.

[0123] S203 involves a secondary leaching of the primary leaching residue with sulfuric acid solution (pH=1.5) at a leaching temperature of 90℃ and a leaching time of 120 min, with a solid-liquid ratio of 5 L:1 kg, resulting in a secondary leaching solution with a vanadium leaching rate of only 10.21%. This is because hydrolyzed V2O5 is difficult to dissolve in acid.

[0124] Take the secondary leachate of S204, adjust the pH to 2-3 with ammonia water, add ammonium sulfate, and precipitate vanadium at 90℃ to obtain ammonium vanadate;

[0125] S205 was obtained by calcining ammonium vanadate at 550℃ for 2 hours to produce V2O5. The total vanadium recovery rate was approximately 10.01%.

[0126] Comparative Example 4 shows that conventional acid leaching cannot effectively dissolve vanadium in the form of V2O5 in the primary leaching residue, and efficient vanadium recovery can only be achieved by relying on the complexation effect of H2O2.

[0127] Comparative Example 5 S201 crushes and grinds the calcified vanadium extraction tailings to a particle size of less than 74μm.

[0128] S202 was leached with water and vanadium extraction tailings in a liquid-solid ratio of 5 L: 1 kg. The pH was adjusted with sulfuric acid to maintain the pH of the leachate at 2.0. The leaching temperature was 90℃ and the leaching time was 300 min, resulting in a primary leachate and a primary leaching residue.

[0129] S203 dissolves the primary leaching residue in water with H2O2, n(H2O2 / V) is 5, the leaching temperature is 25℃, and the leaching time is 30 min, resulting in a secondary leaching solution (peroxyvanadic acid solution). The vanadium leaching rate is only 90.21%.

[0130] S204 heats the secondary leaching solution to 40°C and holds it for only 20 minutes, then filters it to obtain the vanadium-precipitated product and the vanadium-precipitated liquid.

[0131] S205 calcined the vanadium-precipitated product at 550℃ for 2 hours to obtain V2O5, with a vanadium precipitation efficiency of only 5.61%.

[0132] Comparative Example 5 illustrates that the secondary leaching solution must be heated to a sufficient temperature (80-100℃) and maintained for a sufficient time (e.g., 100 min) to ensure complete decomposition of H2O2 and full hydrolysis and precipitation of vanadium; otherwise, vanadium will remain in the solution and cannot be effectively recovered.

[0133] In summary, Route 2 of this invention selectively enriches vanadium in the primary leaching residue (in the form of V2O5) through high-temperature acid leaching, achieving solid-liquid separation from impurities such as calcium, silicon, and iron. The primary leaching residue is then dissolved in water with H2O2, complexing V2O5 into a soluble vanadium peroxide complex. Finally, vanadium is hydrolyzed and precipitated by heating and decomposing H2O2, or by adding ammonium salts to precipitate vanadium, followed by calcination to obtain V2O5. Comparative Examples 3 to 5 demonstrate that H2O2 is crucial for dissolving the V2O5 hydrolysis products and cannot be replaced by conventional acid leaching. Furthermore, sufficient heating (80–100°C, maintained for 100 min) is required during vanadium precipitation to ensure complete decomposition of H2O2; otherwise, vanadium precipitation will be incomplete. This method achieves highly efficient separation of vanadium from impurities, with a total recovery rate exceeding 90%, and allows for the selection of an ammonium-free vanadium precipitation route, avoiding ammonia nitrogen wastewater pollution.

[0134] This invention provides a method for enhanced vanadium extraction or selective separation and recovery from vanadium-containing minerals, achieving the technical objectives of efficient leaching and selective separation through two process routes:

[0135] Route 1 (direct complex leaching method) introduces hydrogen peroxide during the low-temperature acid leaching stage to form a stable vanadium peroxide complex with vanadium ions, effectively suppressing the hydrolysis and precipitation of pentavalent vanadium in acidic media from the source. Experimental data show that this route achieves a vanadium leaching rate of up to 97.24% and a total recovery rate of up to 96.36%, significantly superior to traditional acid leaching processes, thus realizing highly efficient vanadium recovery.

[0136] Route 2 (Separation and Re-complexation Method): High-temperature acid leaching selectively enriches vanadium in the form of V₂O₅ in the primary leaching residue, achieving solid-liquid separation from impurities such as calcium, silicon, and iron. Then, hydrogen peroxide is added to water to dissolve the primary leaching residue, complexing the V₂O₅ into a soluble vanadium peroxide complex, achieving selective vanadium recovery. This route is particularly suitable for complex vanadium-containing raw materials and allows for the selection of an ammonium-free vanadium precipitation process, avoiding the generation of ammonia nitrogen wastewater.

[0137] In summary, this invention addresses the problems of easy hydrolysis of vanadium in vanadium-containing minerals, severe co-leaching of impurities, and the cumbersome nature of traditional processes. It proposes two efficient and selective technical solutions: "direct complexation leaching" and "separation followed by complexation." These solutions achieve both high vanadium leaching and recovery rates, as well as efficient separation of vanadium from impurities, fully demonstrating the core advantages of this method in both efficient recovery and selective separation, and possessing broad prospects for industrial application.

[0138] 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 enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-bearing minerals, characterized in that, The method involves a complexation reaction between hydrogen peroxide and vanadium in vanadium-containing minerals to form a soluble vanadium peroxide complex; it includes a direct complexation leaching method during the low-temperature acid leaching stage, or a separation and re-complexation method after high-temperature acid leaching, wherein... The direct complexation leaching method is as follows: In the low-temperature acid leaching step of vanadium-containing minerals, H2O2 is added to cause it to undergo a complexation reaction with vanadium in the vanadium-containing minerals to obtain a vanadium-containing leachate. The separation and re-complexation method is as follows: after the high-temperature acid leaching step of vanadium-containing minerals, H2O2 is added to the primary leaching residue to cause a complexation reaction with the vanadium in the vanadium-containing minerals, thereby obtaining a secondary vanadium-containing leaching solution.

2. The method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-bearing minerals according to claim 1, characterized in that, The steps of the direct complexation leaching method are as follows: S101 raw material pretreatment involves crushing and grinding vanadium-containing minerals; S102 acid leaching involves hydrogen peroxide complexation leaching in a leaching medium, wherein the leaching medium is water, and the liquid-to-solid ratio of the leaching medium to the vanadium-containing mineral is (0.5-50) L:1 kg; the vanadium-containing mineral is calculated as V, and the amount of H2O2 added satisfies n(H2O2 / V) of 0.01-4, resulting in a dark black leaching solution containing vanadium complexed with hydrogen peroxide. The pH of the leaching solution is adjusted to 0-5 using a pH adjuster, and the leaching temperature is 15-35℃. After the S103 leaching is completed, the solid and liquid are separated by filtration to obtain a vanadium-containing leachate. The pH of the leachate is adjusted to 2-3 with ammonia water, and ammonium sulfate or ammonium persulfate is added. A precipitation reaction is carried out at 70-100℃ to obtain a solid precipitate. The S104 solid precipitate was repeatedly washed with water and anhydrous ethanol several times and dried in a vacuum oven to obtain ammonium vanadate. S105 heated the obtained ammonium vanadate in a muffle furnace and oxidized and calcined it to obtain the product V2O5.

3. The method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-bearing minerals according to claim 1, characterized in that, The steps of the separation and reconnection method are as follows: S201 raw material pretreatment involves crushing and grinding vanadium-containing minerals; S202 acid leaching, the leaching medium is water, the liquid-solid ratio of the leaching medium to the vanadium-containing mineral is (0.5-20) L:1 kg, the pH of the leaching solution is adjusted to 0-3 using a pH adjuster, the leaching temperature is 80-120℃, and the leaching time is 30-600 min, to obtain primary leaching solution and primary leaching residue; S203 dissolves the primary leaching residue with H2O2 and water, where the primary leaching residue is V, the amount of H2O2 added satisfies n(H2O2 / V) of 0.001 to 10, the leaching temperature is 0 to 100℃, and the leaching time is 1 to 600 min, to obtain the secondary leaching solution; S204 involves heating the secondary leaching solution to precipitate vanadium at a temperature of 30–100°C for 60–100 minutes to obtain the vanadium-precipitated product V5O. 12 • 6H2O; or add ammonium sulfate to the secondary leachate to adjust the pH of the system to 2.0–2.5, and vanadium precipitation is carried out to obtain the vanadium precipitation product (NH4)2V6O. 16 & NH4V4O 10 ; S205 will precipitate vanadium product V5O 12 • 6H₂O or (NH₄)₂V₆O 16 & NH4V4O 10 After calcination, V2O5 is obtained.

4. A method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-bearing minerals according to claim 2 or 3, characterized in that, pH adjusters include sulfuric acid, hydrochloric acid, nitric acid, citric acid, oxalic acid, or phosphoric acid.

5. A method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-bearing minerals according to any one of claims 1-3, characterized in that, The vanadium-containing minerals include one or more of the following: vanadium slag roasting products, vanadium-containing steel slag, vanadium extraction tailings, vanadium-containing steel slag roasting products, vanadium-rich slag roasting products obtained by reducing vanadium slag to remove iron and separating vanadium iron, and vanadium-titanium slag roasting products obtained by reducing vanadium-titanium magnetite concentrate to remove iron and separating vanadium titanium slag.

6. A method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-bearing minerals according to claim 5, characterized in that, The vanadium slag is one or more of ordinary vanadium slag, vanadium-chromium slag, or high-calcium and high-phosphorus vanadium slag. The vanadium extraction tailings are one or more of the following: calcified vanadium extraction tailings, sodium vanadium extraction tailings, and manganese vanadium extraction tailings. The vanadium slag roasting product, the vanadium-rich slag roasting product obtained by reducing and removing iron from vanadium slag and separating vanadium-iron, and the vanadium-titanium slag roasting product obtained by reducing and removing iron from vanadium-titanium magnetite concentrate are one or more of the following roasting products: blank roasting product, sodium-modified roasting product, calcified roasting product, manganese-modified roasting product, magnesian-modified roasting product, calcium-manganese composite roasting product, calcium-magnesium composite roasting product, sodium-calcium roasting product, and roasting products, as well as roasting products with various additive combinations thereof.

7. A method for enhanced vanadium extraction or selective separation and recovery of vanadium from vanadium-bearing minerals according to claim 1, characterized in that, The H2O2 is either a solid powder or a liquid.

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