A method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching.

Abstract

This invention relates to a method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. The method involves: crushing, roasting, and grinding the vanadium shale to obtain n batches of roasted powder. Water and concentrated sulfuric acid are added to the first batch of roasted powder, followed by the first batch of leaching aid; leaching is performed, followed by solid-liquid separation to obtain a second batch of circulating leaching agent and the first batch of leachate. The second batch of roasted powder is then mixed with the second batch of circulating leaching agent, concentrated sulfuric acid, and water, followed by the second batch of leaching aid; leaching is performed, followed by solid-liquid separation to obtain a third batch of circulating leaching agent and the second batch of leachate. This process continues until the nth batch of circulating leaching agent and the (n-1)th batch of leachate are obtained, where n is a natural number from 3 to 10. The n batches of leachate are mixed, the pH is adjusted, and sodium sulfite is added to obtain the extraction solution. The extraction solution is then subjected to countercurrent extraction and vanadium precipitation to obtain vanadium pentoxide. This invention features low acid consumption, low leaching aid consumption, high vanadium leaching rate, and good separation of vanadium from impurities.

Description

Technical Field

[0001] This invention belongs to the field of vanadium shale roasting and acid leaching technology. Specifically, it relates to a method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. Background Technology

[0002] Vanadium shale is an important vanadium-bearing resource, but because vanadium mainly exists in the mineral in an isomorphous form, lattice release is difficult. To improve recovery rates, high temperatures, high acids, and large amounts of leaching aids are generally introduced during the leaching process, leading to difficulties in subsequent purification and separation.

[0003] The patented technology, "A Method for Decarbonizing and Crystallizing Roasting to Enhance Acid Leaching Vanadium Extraction from Vanadium-Containing Shale Coal" (CN 111304465B), involves suspending and roasting shale coal powder at 700–900°C, followed by decarbonizing roasting at 500–800°C, and then oxidizing at 750–900°C before acid leaching vanadium using a sulfuric acid solution. This method has a simple process flow and produces uniform product properties. However, it uses a sulfuric acid solution with a mass concentration of 25–35%, resulting in high acid consumption and a low vanadium leaching rate of only 80–90%.

[0004] The patented technology "A Comprehensive Utilization Method for Oxidation-Crystallization Roasting of Vanadium Ore from Stone Coal" (CN 111719054B) involves suspending roasting and decarbonizing pulverized stone coal ore, followed by acid-coated leaching or direct leaching. This method has a simple process flow and solves the problem of incomplete vanadium oxidation in traditional processes. However, the acid-coated leaching process consumes 40-45% wt% sulfuric acid, while the direct leaching process consumes 30-40 wt% sulfuric acid, resulting in high acid consumption. Furthermore, the vanadium leaching rate is low, at only 75-80% during direct leaching.

[0005] The patented technology, "A Method for Extracting Valuable Elements from Vanadium Shale" (CN 115807172A), employs a two-stage leaching method to extract valuable elements from vanadium shale. The first stage uses hydrochloric acid leaching to remove impurities such as calcium from the vanadium shale, while the second stage uses a mixed acid solution of sulfuric and hydrochloric acids to extract elements such as vanadium and zinc. However, this method requires a large amount of leaching aid, typically 2-6 wt% of the raw ore powder.

[0006] The patented technology, "A Method for Recovering Vanadium from Primary Shale Coal" (CN 117327930B), uses acids such as sulfuric acid, hydrochloric acid, bromic acid, oxalic acid, sulfurous acid, and phosphoric acid to alter the shale coal, followed by leaching. This method does not require high-temperature, long-term oxidative roasting or sodium treatment. However, this method requires potential control of the solution using an oxidant, followed by the removal of molybdenum and titanium using a molybdenum-titanium extractant or resin to obtain a selectively purified leachate. This method cannot directly separate vanadium and impurities at the source, resulting in poor vanadium-impurity separation.

[0007] In summary, existing vanadium shale extraction processes suffer from problems such as high acid consumption, high leaching aid consumption, low vanadium leaching rate, and poor separation of vanadium from impurities. Summary of the Invention

[0008] The present invention aims to overcome the defects of the prior art and provides a method for enhanced vanadium extraction by roasting and acid leaching of vanadium shale. This method has the advantages of low acid consumption, low leaching aid consumption, high vanadium leaching rate, and good separation effect between vanadium and impurities.

[0009] To achieve the above objectives, the technical solution adopted by the present invention comprises the following steps:

[0010] Step 1, roasting

[0011] The crushed vanadium shale is roasted and ground to obtain roasted powder.

[0012] Step 2: Leaching of roasted powder in batches

[0013] The roasted powder is divided into n equal parts, where n is a natural number from 3 to 10; it is then leached in n batches in sequence.

[0014] The first batch of roasted powder was mixed with water to obtain the first batch of mixed slurry.

[0015] Concentrated sulfuric acid was added to the first batch of mixed pulp to adjust the pulp, thus obtaining the first batch of leach pulp.

[0016] After adding a leaching aid to the first batch of leaching pulp, the pulp was adjusted to obtain a mixed leaching pulp.

[0017] The leaching slurry is stirred and leached at 80-100°C for 4-10 hours, and then solid-liquid separation is performed to obtain clear leaching liquid and secondary leaching residue. 40-60 vol% of the clear leaching liquid is used as the leaching agent for the next batch, and the remaining clear leaching liquid is used as the finished leaching liquid.

[0018] Starting with batch 2, follow these steps:

[0019] The roasted powder is mixed with the recycled leachate from the previous batch and then slurryed to obtain a mixed slurry.

[0020] Concentrated sulfuric acid and water of the same volume as the previous batch of finished leachate were added to the mixed slurry to obtain leachate.

[0021] After adding a leaching aid to the leaching pulp, the pulp is adjusted to obtain a mixed leaching pulp.

[0022] The leaching slurry is stirred and leached at 80-100°C for 4-10 hours, and then solid-liquid separation is performed to obtain clear leaching liquid and secondary leaching residue. 40-60 vol% of the clear leaching liquid is used as the leaching agent for the next batch, and the remaining clear leaching liquid is used as the finished leaching liquid.

[0023] In the nth batch, all the leachate was used as the finished leachate.

[0024] For the first batch of leaching, the mass ratio of concentrated sulfuric acid to calcined powder was 0.16–0.22:1; for the remaining batches, the mass ratio of concentrated sulfuric acid to calcined powder was 0.14–0.20:1.

[0025] For the first batch of leaching, the mass ratio of roasted powder to leaching aid was 1:0.02 to 0.05; for the remaining batches, the mass ratio of roasted powder to leaching aid was 1:0.015 to 0.04.

[0026] The leaching aid is a mixture of sodium nitrate or sodium nitrite and fluoride, wherein the fluoride content in the leaching aid is 70-90 wt%, and the fluoride is one or more of calcium fluoride and sodium fluoride.

[0027] Step 3: Adjust the pH value of the leachate

[0028] Mix all batches of finished product leachate and adjust the pH value to 1.6-2.6 to obtain the adjusted leachate.

[0029] Sodium sulfite was added to the adjusted leachate, stirred, and filtered to obtain the original extract and pH-adjusted residue.

[0030] Step 4: Solvent extraction and vanadium precipitation

[0031] The organic phase is mixed with the extraction solution and subjected to countercurrent extraction at 10–30°C to separate the phases, yielding raffinate and loaded organic phase.

[0032] The loaded organic phase is mixed with the back-extracting agent and subjected to countercurrent back-extraction at 10–30°C to separate the phases, resulting in a vanadium-rich solution and a lean organic phase.

[0033] The vanadium-rich solution was subjected to ammonium salt precipitation of vanadium and ammonia removal to obtain vanadium pentoxide.

[0034] Preferably, in step 3, the liquid-to-solid ratio of sodium sulfite added to the adjusted leachate is 1:(2-6) L / g; stirring is carried out at 60-80°C and at a speed of 1-10 r / min for 30-60 min.

[0035] Preferably, in step 4, the volume ratio of the organic phase to the extraction solution is 1:(2-3), and the volume ratio of the supported organic phase to the back-extraction agent is (4-10):1.

[0036] Preferably, the crushed vanadium shale has a particle size of ≤3mm; the crushed vanadium shale has the following contents: V2O5 content of 0.5-1.5wt%, CaO content of 1-6wt%, SiO2 content of 50-90wt%, Al2O3 content of 2-5wt%, and Fe2O3 content of 1-4wt%.

[0037] In the fractured vanadium shale, mica-like minerals account for 40–95 wt% of the total vanadium content.

[0038] Preferably, the roasting process is as follows: the roasting temperature is 400-700℃; the roasting time is 30-90 min; and the particle size of the roasted powder is less than 0.074 mm, accounting for 20-50 wt%.

[0039] Preferably, the extractant is one or more of di(2-ethylhexyl)phosphonate and 2-ethylhexylphosphonic acid mono-2-ethylhexyl ester.

[0040] Preferably, the organic phase is a mixture of 15-25 vol% extractant, 5-10 vol% octanol and 65-80 vol% sulfonated kerosene.

[0041] Preferably, the stripping agent is a 1-2 mol / L sulfuric acid solution.

[0042] Preferably, the regenerated organic phase is recycled as an organic phase in step 4.

[0043] In this technical solution:

[0044] The chemical composition of each batch of immersion aid is the same;

[0045] The volume of each batch of leachate is equal;

[0046] Each batch of concentrated sulfuric acid has the same chemical composition, and the concentration of the concentrated sulfuric acid is greater than 98 wt%.

[0047] By adopting the above technical solution, the present invention has the following advantages compared with the prior art:

[0048] (1) Low acid dosage and high vanadium leaching rate during the leaching process. Based on existing technologies that add fluorides as leaching aids, this invention further adds nitrates / nitrites as leaching aids. The mechanism by which this invention significantly improves the vanadium leaching rate with low acid dosage is as follows:

[0049] ①Oxidation

[0050] Vanadium in vanadium shale typically exists in mica through lattice substitution, generally in a low valence state (e.g., +3), exhibiting high stability and difficulty in dissolution. During leaching, nitrate ions form nitric acid under acidic conditions. This nitric acid adsorbs at the leaching reaction interface, oxidizing the low-valence vanadium in the vanadium shale mica to a higher valence state (e.g., +4, +5). This oxidation process distorts the mica lattice, increasing the chemical reactivity of vanadium within the lattice and reducing its binding force to the mica lattice, thus facilitating vanadium dissolution under low-acid conditions.

[0051] ②Coordination effect

[0052] Nitrate ions can undergo coordination reactions with vanadium, aluminum, and magnesium ions produced by the decomposition of vanadium-containing mica. During sulfuric acid leaching, nitrate ions act as ligands, combining with the gradually dissolving vanadium, aluminum, and magnesium ions to form relatively stable coordination compounds. This coordination equilibrium promotes the continuous dissolution of mica structural elements, facilitating the continuous transfer of vanadium from vanadium-containing mica into the solution system and improving the vanadium leaching effect.

[0053] (2) Low leaching aid consumption and high vanadium leaching rate. This invention involves batch leaching of vanadium shale roasted powder, with a portion of the acid leaching solution from the previous batch being recycled as a circulating leaching aid to enhance the leaching of the next batch of roasted powder. The mechanism by which this invention improves the vanadium leaching rate with low leaching aid dosage is as follows: the circulating leaching agent contains hydrogen ions, sulfate ions, vanadium ions, and other metal ions. Hydrogen ions and sulfate ions replenish the acid consumed in the leaching reaction, maintaining acidity and promoting the dissolution of vanadium-containing minerals. Furthermore, as the number of cycles increases, the concentration of each metal ion in the leaching solution correspondingly increases, and the ionic strength gradually increases, enhancing the destruction of vanadium-containing minerals and reducing the leaching aid consumption in the leaching process.

[0054] (3) Excellent separation of vanadium and impurities. This invention introduces highly oxidizing nitrate ions into the leaching system of roasted powder. Compared with existing technologies, the mechanism by which this invention achieves selective separation of vanadium and impurities is as follows: nitrate ions can selectively oxidize ferrous ions in the leachate, oxidizing ferrous iron to ferric iron. During pH adjustment, calcium hydroxide or calcium sulfate is generally chosen as the pH adjuster. However, the addition of calcium hydroxide or calcium sulfate can easily lead to local overconcentration. When the local pH value in the leachate is too high, ferric iron easily co-precipitates with calcium sulfate and enters the pH adjustment residue, achieving selective removal of iron ions from the leachate. Measurements show that the iron ion content in the extraction solution can be reduced to below 600 mg / L, and the vanadium ion concentration is greater than 1.6 g / L, achieving the separation of vanadium and iron in the leachate.

[0055] Therefore, the present invention has the characteristics of low acid consumption, low leaching aid consumption, high vanadium leaching rate, and good separation effect between vanadium and impurities during the leaching process. Detailed Implementation

[0056] The present invention will be further described below with reference to specific embodiments, but this is not intended to limit the scope of protection thereof.

[0057] A method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. The steps of the method described in this specific embodiment are as follows:

[0058] Step 1, roasting

[0059] The crushed vanadium shale is roasted and ground to obtain roasted powder.

[0060] Step 2: Leaching of roasted powder in batches

[0061] The roasted powder is divided into n equal parts, where n is a natural number from 3 to 10; it is then leached in n batches in sequence.

[0062] The first batch of roasted powder was mixed with water to obtain the first batch of mixed slurry.

[0063] Concentrated sulfuric acid was added to the first batch of mixed pulp to adjust the pulp, thus obtaining the first batch of leach pulp.

[0064] After adding a leaching aid to the first batch of leaching pulp, the pulp was adjusted to obtain a mixed leaching pulp.

[0065] The leaching slurry is stirred and leached at 80-100°C for 4-10 hours, and then solid-liquid separation is performed to obtain clear leaching liquid and secondary leaching residue. 40-60 vol% of the clear leaching liquid is used as the leaching agent for the next batch, and the remaining clear leaching liquid is used as the finished leaching liquid.

[0066] Starting with batch 2, follow these steps:

[0067] The roasted powder is mixed with the recycled leachate from the previous batch and then slurryed to obtain a mixed slurry.

[0068] Concentrated sulfuric acid and water of the same volume as the previous batch of finished leachate were added to the mixed slurry to obtain leachate.

[0069] After adding a leaching aid to the leaching pulp, the pulp is adjusted to obtain a mixed leaching pulp.

[0070] The leaching slurry is stirred and leached at 80-100°C for 4-10 hours, and then solid-liquid separation is performed to obtain clear leaching liquid and secondary leaching residue. 40-60 vol% of the clear leaching liquid is used as the leaching agent for the next batch, and the remaining clear leaching liquid is used as the finished leaching liquid.

[0071] In the nth batch, all the leachate was used as the finished leachate.

[0072] For the first batch of leaching, the mass ratio of concentrated sulfuric acid to calcined powder was 0.16–0.22:1; for the remaining batches, the mass ratio of concentrated sulfuric acid to calcined powder was 0.14–0.20:1.

[0073] For the first batch of leaching, the mass ratio of roasted powder to leaching aid was 1:0.02 to 0.05; for the remaining batches, the mass ratio of roasted powder to leaching aid was 1:0.015 to 0.04.

[0074] The leaching aid is a mixture of sodium nitrate or sodium nitrite and fluoride, wherein the fluoride content in the leaching aid is 70-90 wt%, and the fluoride is one or more of calcium fluoride and sodium fluoride.

[0075] Step 3: Adjust the pH value of the leachate

[0076] Mix all batches of finished product leachate and adjust the pH value to 1.6-2.6 to obtain the adjusted leachate.

[0077] Sodium sulfite was added to the adjusted leachate, stirred, and filtered to obtain the original extract and pH-adjusted residue.

[0078] Step 4: Solvent extraction and vanadium precipitation

[0079] The organic phase is mixed with the extraction solution and subjected to countercurrent extraction at 10–30°C to separate the phases, yielding raffinate and loaded organic phase.

[0080] The loaded organic phase is mixed with the back-extracting agent and subjected to countercurrent back-extraction at 10–30°C to separate the phases, resulting in a vanadium-rich solution and a lean organic phase.

[0081] The vanadium-rich solution was subjected to ammonium salt precipitation of vanadium and ammonia removal to obtain vanadium pentoxide.

[0082] Preferably, in step 3, the liquid-to-solid ratio of sodium sulfite added to the adjusted leachate is 1:(2-6) L / g; stirring is carried out at 60-80°C and at a speed of 1-10 r / min for 30-60 min.

[0083] Preferably, in step 4, the volume ratio of the organic phase to the extraction solution is 1:(2-3), and the volume ratio of the supported organic phase to the back-extraction agent is (4-10):1.

[0084] Preferably, the crushed vanadium shale has a particle size of ≤3mm; the crushed vanadium shale has the following contents: V2O5 content of 0.5-1.5wt%, CaO content of 1-6wt%, SiO2 content of 50-90wt%, Al2O3 content of 2-5wt%, and Fe2O3 content of 1-4wt%.

[0085] In the fractured vanadium shale, mica-like minerals account for 40–95 wt% of the total vanadium content.

[0086] Preferably, the roasting process is as follows: the roasting temperature is 400-700℃; the roasting time is 30-90 min; and the particle size of the roasted powder is less than 0.074 mm, accounting for 20-50 wt%.

[0087] Preferably, the extractant is one or more of di(2-ethylhexyl)phosphonate and 2-ethylhexylphosphonic acid mono-2-ethylhexyl ester.

[0088] Preferably, the organic phase is a mixture of 15-25 vol% extractant, 5-10 vol% octanol and 65-80 vol% sulfonated kerosene.

[0089] Preferably, the stripping agent is a 1-2 mol / L sulfuric acid solution.

[0090] Preferably, the regenerated organic phase is recycled as an organic phase in step 4.

[0091] In this specific implementation:

[0092] The chemical composition of each batch of immersion aid is the same;

[0093] The volume of each batch of leachate is equal;

[0094] Each batch of concentrated sulfuric acid has the same chemical composition;

[0095] The concentration of the concentrated sulfuric acid is greater than 98 wt%.

[0096] The regenerated organic phase is recycled as an organic phase in step 4.

[0097] The details will not be repeated in the examples.

[0098] Example 1

[0099] A method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. The steps of the method described in this specific embodiment are as follows:

[0100] Step 1, roasting

[0101] The crushed vanadium shale is roasted, and then the roasted vanadium shale is ground into fine powder to obtain roasted powder.

[0102] Step 2: Leaching of n batches of roasted powder

[0103] The roasted powder was divided into three equal parts and leached in three batches in sequence.

[0104] Step 2.1, Leaching of the first batch of roasted powder

[0105] The first batch of roasted powder was mixed with water at a solid-liquid ratio of 1:0.8 kg / L to form a slurry, thus obtaining the first batch of mixed slurry. The first batch of concentrated sulfuric acid was added to the first batch of mixed slurry at a mass ratio of 0.16:1 (first batch of concentrated sulfuric acid:first batch of roasted powder) to form a slurry, thus obtaining the first batch of leachate.

[0106] The first batch of leaching agent was added to the first batch of leaching pulp at a mass ratio of 1:0.02 of the first batch of roasted powder to the first batch of leaching aid, and the pulp was adjusted to obtain the first batch of leaching mixture.

[0107] The first batch of leaching slurry was stirred and leached at 80°C for 4 hours, and the first solid-liquid separation was performed to obtain the first batch of leaching liquid and the first batch of leaching residue; wherein: 40 vol% of the first batch of leaching liquid was used as the second batch of circulating leaching agent, and the remainder of the first batch of leaching liquid was used as the first batch of leaching liquid.

[0108] The first batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0109] The first batch of immersion aid consists of 70 wt% fluoride and 30 wt% sodium nitrate; the fluoride is calcium fluoride.

[0110] Step 2.2, Leaching of the second batch of roasted powder

[0111] The second batch of roasted powder was mixed with the second batch of circulating leachate and slurry was prepared to obtain the second batch of mixed slurry. The second batch of concentrated sulfuric acid was added to the second batch of mixed slurry at a mass ratio of 0.14:1, and then water with the same volume as the first batch of leachate was added and slurry was prepared to obtain the second batch of leachate.

[0112] The second batch of leaching aid was added to the second batch of leaching slurry at a mass ratio of 1:0.015 of the second batch of roasted powder to the second batch of leaching aid to obtain the second batch of leaching mixture.

[0113] The second batch of leaching slurry was stirred and leached at 80°C for 4 hours, followed by a second solid-liquid separation to obtain the second batch of leaching liquid and the second batch of leaching residue. Among them, 40 vol% of the second batch of leaching liquid was used as the third batch of circulating leaching agent, and the remainder of the second batch of leaching liquid was used as the second batch of leaching liquid.

[0114] The second batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0115] Step 2.3, Leaching of the third batch of roasted powder

[0116] The third batch of roasted powder was mixed with the third batch of circulating leachate and slurry was prepared to obtain the third batch of mixed slurry. The third batch of concentrated sulfuric acid was added to the third batch of mixed slurry at a mass ratio of 0.14:1, and then water with the same volume as the second batch of leachate was added and slurry was prepared to obtain the third batch of leachate.

[0117] The third batch of leaching aid was added to the third batch of leaching slurry at a mass ratio of 1:0.015 of the third batch of roasted powder to the third batch of leaching aid to obtain the third batch of leaching slurry.

[0118] The third batch of leaching slurry was stirred and leached at 80°C for 4 hours, followed by a third solid-liquid separation to obtain the third batch of leaching liquid and the third batch of leaching residue.

[0119] The third batch of leachate was entirely from the third batch of leachate.

[0120] The third batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0121] Step 3: Adjust the pH value of the leachate

[0122] The first, second, and third batches of leachate were mixed. A pH adjuster, calcium hydroxide, was added to the mixed leachate to adjust the pH value to 1.6, resulting in an adjusted leachate. Sodium sulfite was then added to the adjusted leachate at a liquid-to-solid ratio of 1:2 L / g. The mixture was then stirred for 30 minutes at 60°C and a rotation speed of 1 r / min, and filtered to obtain the original extract and the pH-adjusted residue.

[0123] Step 4: Solvent extraction and vanadium precipitation

[0124] The organic phase and the extraction solution are mixed at a volume ratio of 1:2, and countercurrent extraction is performed at 10°C to separate the phases, resulting in raffinate and a loaded organic phase. The loaded organic phase and the back-extraction agent are mixed at a volume ratio of 4:1, and countercurrent back-extraction is performed at 10°C to separate the phases, resulting in a vanadium-rich solution and a lean organic phase.

[0125] The vanadium-rich solution was subjected to ammonium salt precipitation of vanadium and ammonia removal to obtain vanadium pentoxide.

[0126] The crushed vanadium shale has the following composition: V2O5 content is 0.5wt%, CaO content is 1wt%, SiO2 content is 50wt%, Al2O3 content is 2wt%, and Fe2O3 content is 1wt%. The vanadium content of mica minerals in the crushed vanadium shale accounts for 40wt% of the total vanadium content.

[0127] The roasting process is as follows: the roasting temperature is 400℃; the roasting time is 30 minutes.

[0128] The roasted powder has a particle size of less than 0.074 mm, accounting for 20 wt%.

[0129] The extractant is di(2-ethylhexyl)phosphonate.

[0130] The organic phase is a mixture of 15 vol% extractant, 5 vol% octanol and 80 vol% sulfonated kerosene.

[0131] The stripping agent is a 1 mol / L sulfuric acid solution.

[0132] Example 2

[0133] A method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. The steps of the method described in this specific embodiment are as follows:

[0134] Step 1, roasting

[0135] The crushed vanadium shale is roasted, and then the roasted vanadium shale is ground into fine powder to obtain roasted powder.

[0136] The roasted powder was divided into four equal parts and leached in four batches in sequence.

[0137] Step 2: Leaching of n batches of roasted powder

[0138] Step 2.1, Leaching of the first batch of roasted powder

[0139] The first batch of roasted powder was mixed with water at a solid-liquid ratio of 1:1.1 kg / L to form a slurry, thus obtaining the first batch of mixed slurry. The first batch of concentrated sulfuric acid was added to the first batch of mixed slurry at a mass ratio of 0.17:1 to the first batch of roasted powder to form a leaching slurry, thus obtaining the first batch of leachate.

[0140] The first batch of leaching agent was added to the first batch of leaching pulp at a mass ratio of 1:0.03 (first batch of roasted powder to first batch of leaching aid), and the pulp was adjusted to obtain the first batch of leaching mixture.

[0141] The first batch of leaching slurry was stirred and leached at 85°C for 5 hours, and the first solid-liquid separation was performed to obtain the first batch of leaching liquid and the first batch of leaching residue; wherein: 45 vol% of the first batch of leaching liquid was used as the second batch of circulating leaching agent, and the remainder of the first batch of leaching liquid was used as the first batch of leaching liquid.

[0142] The first batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0143] The first batch of immersion aid consists of 75 wt% fluoride and 25 wt% sodium nitrite; the fluoride is sodium fluoride.

[0144] Step 2.2, Leaching of the second batch of roasted powder

[0145] The second batch of roasted powder was mixed with the second batch of circulating leachate and slurry was prepared to obtain the second batch of mixed slurry. The second batch of concentrated sulfuric acid was added to the second batch of mixed slurry at a mass ratio of 0.15:1 of the second batch of roasted powder, and then water with the same volume as the first batch of leachate was added and slurry was prepared to obtain the second batch of leachate.

[0146] The second batch of leaching aid was added to the second batch of leaching slurry at a mass ratio of 1:0.02 of the second batch of roasted powder to the second batch of leaching aid to obtain the second batch of leaching mixture.

[0147] The second batch of leaching slurry was stirred and leached at 85°C for 5 hours, followed by a second solid-liquid separation to obtain the second batch of leaching liquid and the second batch of leaching residue. 45 vol% of the second batch of leaching liquid was used as the third batch of circulating leaching agent, and the remainder of the second batch of leaching liquid was used as the second batch of leaching liquid.

[0148] The second batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0149] Step 2.3, Leaching of the third batch of roasted powder

[0150] The third batch of roasted powder was mixed with the third batch of circulating leachate and slurry was prepared to obtain the third batch of mixed slurry. The third batch of concentrated sulfuric acid was added to the third batch of mixed slurry at a mass ratio of 0.15:1 of the third batch of roasted powder. Then, water with the same volume as the second batch of leachate was added and slurry was prepared to obtain the third batch of leachate.

[0151] The third batch of leaching aid was added to the third batch of leaching slurry at a mass ratio of 1:0.02 of the third batch of roasted powder to the third batch of leaching aid to obtain the third batch of leaching mixture.

[0152] The third batch of leaching slurry was stirred and leached at 85°C for 5 hours, followed by a third solid-liquid separation to obtain the third batch of leaching liquid and the third batch of leaching residue. 45 vol% of the third batch of leaching liquid was used as the fourth batch of circulating leaching agent, and the remainder of the third batch of leaching liquid was used as the third batch of leaching liquid.

[0153] The third batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0154] Step 2.4, Leaching of the 4th batch of roasted powder

[0155] The fourth batch of roasted powder was mixed with the fourth batch of circulating leachate and slurry was prepared to obtain the fourth batch of mixed slurry. The fourth batch of concentrated sulfuric acid was added to the fourth batch of mixed slurry at a mass ratio of 0.15:1 of the fourth batch of roasted powder. Then, water with the same volume as the third batch of leachate was added and slurry was prepared to obtain the fourth batch of leachate.

[0156] The fourth batch of leaching aid was added to the fourth batch of leaching slurry at a mass ratio of 1:0.02 of the fourth batch of roasted powder to the fourth batch of leaching aid to obtain the fourth batch of leaching mixture.

[0157] The fourth batch of leaching slurry was stirred and leached at 85°C for 5 hours, followed by a fourth solid-liquid separation to obtain the fourth batch of leaching liquid and the fourth batch of leaching residue.

[0158] The fourth batch of leachate was entirely from the fourth batch of leachate.

[0159] The fourth batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0160] Step 3: Adjust the pH value of the leachate

[0161] The first, second, third, and fourth batches of leachate were mixed. A pH adjuster, calcium oxide, was added to the mixed leachate to adjust the pH value to 1.8, resulting in an adjusted leachate. Sodium sulfite was then added to the adjusted leachate at a liquid-to-solid ratio of 1:3 L / g. The mixture was then stirred at 65°C and 3 r / min for 40 min, filtered, and the original extract and pH-adjusted residue were obtained.

[0162] Step 4: Solvent extraction and vanadium precipitation

[0163] The organic phase and the extraction solution were mixed at a volume ratio of 1:2.3, and the mixture was subjected to countercurrent extraction at 15°C. The phases were separated to obtain raffinate and a loaded organic phase. The loaded organic phase and the back-extraction agent were mixed at a volume ratio of 5:1, and the mixture was subjected to countercurrent back-extraction at 15°C. The phases were separated to obtain vanadium-rich solution and a lean organic phase.

[0164] The vanadium-rich solution was subjected to ammonium salt precipitation of vanadium and ammonia removal to obtain vanadium pentoxide.

[0165] The crushed vanadium shale has the following composition: V2O5 content 0.75wt%, CaO content 2wt%, SiO2 content 60wt%, Al2O3 content 3wt%, and Fe2O3 content 2wt%. The vanadium content of mica minerals in the crushed vanadium shale accounts for 60wt% of the total vanadium content.

[0166] The roasting process is as follows: the roasting temperature is 500℃; the roasting time is 45min.

[0167] The calcined powder has a particle size of less than 0.074 mm, accounting for 30 wt%.

[0168] The extractant is 2-ethylhexylphosphonic acid mono-2-ethylhexyl ester.

[0169] The organic phase is a mixture of 17.5 vol% extractant, 7 vol% octanol and 75.5 vol% sulfonated kerosene.

[0170] The stripping agent is a 1.25 mol / L sulfuric acid solution.

[0171] Example 3

[0172] A method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. The steps of the method described in this specific embodiment are as follows:

[0173] Step 1, roasting

[0174] The crushed vanadium shale is roasted, and then the roasted vanadium shale is ground into fine powder to obtain roasted powder.

[0175] Step 2: Leaching of n batches of roasted powder

[0176] The roasted powder was divided into 6 equal portions and leached in 6 batches in sequence.

[0177] Step 2.1, Leaching of the first batch of roasted powder

[0178] The first batch of roasted powder was mixed with water at a solid-liquid ratio of 1:1.4 kg / L to form a slurry, thus obtaining the first batch of mixed slurry. The first batch of concentrated sulfuric acid was added to the first batch of mixed slurry at a mass ratio of 0.19:1 to the first batch of roasted powder to form a leaching slurry, thus obtaining the first batch of leaching slurry.

[0179] The first batch of leaching agent was added to the first batch of leaching pulp at a mass ratio of 1:0.035 of the first batch of roasted powder to the first batch of leaching aid, and the pulp was adjusted to obtain the first batch of leaching mixture.

[0180] The first batch of leaching slurry was stirred and leached at 90°C for 7 hours, followed by the first solid-liquid separation to obtain the first batch of leaching liquid and the first batch of leaching residue; wherein, 50 vol% of the first batch of leaching liquid was used as the second batch of circulating leaching agent, and the remainder of the first batch of leaching liquid was used as the first batch of leaching liquid.

[0181] The first batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0182] The first batch of immersion aid consists of 80 wt% fluoride and 20 wt% sodium nitrate; the fluoride is a mixture of calcium fluoride and sodium fluoride.

[0183] Step 2.2, Leaching of the second batch of roasted powder

[0184] The second batch of roasted powder was mixed with the second batch of circulating leachate and slurry was prepared to obtain the second batch of mixed slurry. The second batch of concentrated sulfuric acid was added to the second batch of mixed slurry at a mass ratio of 0.17:1, and then water with the same volume as the first batch of leachate was added and slurry was prepared to obtain the second batch of leachate.

[0185] The second batch of leaching aid was added to the second batch of leaching slurry at a mass ratio of 1:0.03 for the second batch of roasted powder to the second batch of leaching aid to obtain the second batch of leaching mixture.

[0186] The second batch of leaching slurry was stirred and leached at 90°C for 7 hours, followed by a second solid-liquid separation to obtain the second batch of leaching liquid and the second batch of leaching residue. Among them, 50 vol% of the second batch of leaching liquid was used as the third batch of circulating leaching agent, and the remainder of the second batch of leaching liquid was used as the second batch of leaching liquid.

[0187] The second batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0188] ...

[0189] The leaching process for the second batch of roasted powder described in step 2.2 can be extrapolated to the leaching process for the fifth batch of roasted powder.

[0190] Step 2.5, Leaching of the 5th batch of roasted powder

[0191] The fifth batch of roasted powder was mixed with the fifth batch of circulating leachate and slurry was prepared to obtain the fifth batch of mixed slurry. The fifth batch of concentrated sulfuric acid was added to the fifth batch of mixed slurry at a mass ratio of the fifth batch of concentrated sulfuric acid to the fifth batch of roasted powder of 0.17:1. Then, water with the same volume as the fourth batch of leachate was added and slurry was prepared to obtain the fifth batch of leachate.

[0192] The fifth batch of leaching aid was added to the fifth batch of leaching slurry at a mass ratio of 1:0.03 of the fifth batch of roasted powder to the fifth batch of leaching aid to obtain the fifth batch of leaching mixture.

[0193] The fifth batch of leaching slurry was stirred and leached at 90°C for 7 hours, followed by a fifth solid-liquid separation to obtain the fifth batch of leaching liquid and the fifth batch of leaching residue. 50 vol% of the fifth batch of leaching liquid was used as the sixth batch of circulating leaching agent, and the remainder of the fifth batch of leaching liquid was used as the fifth batch of leaching liquid.

[0194] The fifth batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0195] Step 2.6, Leaching of the 6th batch of roasted powder

[0196] The 6th batch of roasted powder was mixed with the 6th batch of circulating leaching agent and slurry was prepared to obtain the 6th batch of mixed slurry. The 6th batch of concentrated sulfuric acid was added to the 6th batch of mixed slurry at a mass ratio of 0.17:1, and then water with the same volume as the 5th batch of leaching liquid was added and slurry was prepared to obtain the 6th batch of leaching slurry.

[0197] The sixth batch of leaching aid was added to the sixth batch of leaching slurry at a mass ratio of 1:0.03 of the sixth batch of roasted powder to the sixth batch of leaching aid to obtain the sixth batch of leaching mixture.

[0198] The sixth batch of leaching slurry was stirred and leached at 90°C for 7 hours, followed by a sixth solid-liquid separation to obtain the sixth batch of leaching liquid and the sixth batch of leaching residue.

[0199] The 6th batch of leachate was entirely from the 6th batch of leachate.

[0200] The sixth batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0201] Step 3: Adjust the pH value of the leachate

[0202] The first batch of leachate, the second batch of leachate, ..., the fifth batch of leachate and the sixth batch of leachate were mixed. A pH adjuster, namely calcium hydroxide, was added to the mixed leachate to adjust the pH value to 2.0, thus obtaining the adjusted leachate. Sodium sulfite was then added to the adjusted leachate at a liquid-to-solid ratio of 1:4 L / g. The mixture was then stirred for 45 minutes at 70°C and a rotation speed of 5 r / min, and filtered to obtain the original extract and the pH-adjusted residue.

[0203] Step 4: Solvent extraction and vanadium precipitation

[0204] The organic phase and the extraction solution were mixed at a volume ratio of 1:2.5, and the mixture was subjected to countercurrent extraction at 20°C. The phases were separated to obtain raffinate and a loaded organic phase. The loaded organic phase and the back-extraction agent were mixed at a volume ratio of 7:1, and the mixture was subjected to countercurrent back-extraction at 20°C. The phases were separated to obtain vanadium-rich solution and a lean organic phase.

[0205] The vanadium-rich solution was subjected to ammonium salt precipitation of vanadium and ammonia removal to obtain vanadium pentoxide.

[0206] The crushed vanadium shale has the following composition: V2O5 content 1 wt%, CaO content 3 wt%, SiO2 content 70 wt%, Al2O3 content 3.5 wt%, and Fe2O3 content 2.5 wt%. Mica minerals account for 70 wt% of the total vanadium content in the crushed vanadium shale.

[0207] The roasting process is as follows: the roasting temperature is 550℃; the roasting time is 60min.

[0208] The roasted powder has a particle size of less than 0.074 mm, accounting for 35 wt%.

[0209] The extractant is a mixture of di(2-ethylhexyl)phosphonate and 2-ethylhexylphosphonic acid mono-2-ethylhexyl ester.

[0210] The organic phase is a mixture of 20 vol% extractant, 7.5 vol% octanol and 72.5 vol% sulfonated kerosene.

[0211] The stripping agent is a 1.5 mol / L sulfuric acid solution.

[0212] Example 4

[0213] A method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. The steps of the method described in this specific embodiment are as follows:

[0214] Step 1, roasting

[0215] The crushed vanadium shale is roasted, and then the roasted vanadium shale is ground into fine powder to obtain roasted powder.

[0216] Step 2: Leaching of n batches of roasted powder

[0217] The roasted powder was divided into 8 equal portions and leached in 8 batches in sequence.

[0218] Step 2.1, Leaching of the first batch of roasted powder

[0219] The first batch of roasted powder was mixed with water at a solid-liquid ratio of 1:1.7 kg / L to form a slurry, thus obtaining the first batch of mixed slurry. The first batch of concentrated sulfuric acid was added to the first batch of mixed slurry at a mass ratio of 0.21:1 to the first batch of roasted powder to form a leaching slurry, thus obtaining the first batch of leachate.

[0220] The first batch of leaching agent was added to the first batch of leaching pulp at a mass ratio of 1:0.04 (first batch of roasted powder to first batch of leaching aid), and the pulp was adjusted to obtain the first batch of leaching mixture.

[0221] The first batch of leaching slurry was stirred and leached at 95°C for 9 hours, and the first solid-liquid separation was performed to obtain the first batch of leaching liquid and the first batch of leaching residue; wherein, 55 vol% of the first batch of leaching liquid was used as the second batch of circulating leaching agent, and the remainder of the first batch of leaching liquid was used as the first batch of leaching liquid.

[0222] The first batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0223] The first batch of immersion aid consists of 85 wt% fluoride and 15 wt% sodium nitrite; the fluoride is calcium fluoride.

[0224] Step 2.2, Leaching of the second batch of roasted powder

[0225] The second batch of roasted powder was mixed with the second batch of circulating leachate and slurry was prepared to obtain the second batch of mixed slurry. The second batch of concentrated sulfuric acid was added to the second batch of mixed slurry at a mass ratio of 0.19:1, and then water with the same volume as the first batch of leachate was added and slurry was prepared to obtain the second batch of leachate.

[0226] The second batch of leaching aid was added to the second batch of leaching slurry at a mass ratio of 1:0.035 of the second batch of roasted powder to the second batch of leaching aid to obtain the second batch of leaching mixture.

[0227] The second batch of leaching slurry was stirred and leached at 95°C for 9 hours, followed by a second solid-liquid separation to obtain the second batch of leaching liquid and the second batch of leaching residue. Among them, 55 vol% of the second batch of leaching liquid was used as the third batch of circulating leaching agent, and the remainder of the second batch of leaching liquid was used as the second batch of leaching liquid.

[0228] The second batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0229] ...

[0230] The leaching process for the second batch of roasted powder described in step 2.2 can be extrapolated to the leaching process for the seventh batch of roasted powder.

[0231] Step 2.7, Leaching of the 7th batch of roasted powder

[0232] The 7th batch of roasted powder was mixed with the 7th batch of circulating leaching agent and slurry was prepared to obtain the 7th batch of mixed slurry. The 7th batch of concentrated sulfuric acid was added to the 7th batch of mixed slurry at a mass ratio of 0.19:1, and then water with the same volume as the 6th batch of leaching liquid was added and slurry was prepared to obtain the 7th batch of leaching slurry.

[0233] The seventh batch of leaching aid was added to the seventh batch of leaching pulp at a mass ratio of 1:0.035 of the seventh batch of roasted powder to the seventh batch of leaching aid to obtain the seventh batch of leaching mixture.

[0234] The seventh batch of leaching slurry was stirred and leached at 95°C for 9 hours, followed by a seventh solid-liquid separation to obtain the seventh batch of leaching liquid and the seventh batch of leaching residue. Among them, 55 vol% of the seventh batch of leaching liquid was used as the eighth batch of circulating leaching agent, and the remainder of the seventh batch of leaching liquid was used as the seventh batch of leaching liquid.

[0235] The 7th batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0236] Step 2.8, Leaching of the 8th batch of roasted powder

[0237] The 8th batch of roasted powder was mixed with the 8th batch of circulating leachate and slurry was prepared to obtain the 8th batch of mixed slurry. The 8th batch of concentrated sulfuric acid was added to the 8th batch of mixed slurry at a mass ratio of 0.19:1, and then water with the same volume as the 7th batch of leachate was added and slurry was prepared to obtain the 8th batch of leachate.

[0238] The eighth batch of leaching aid was added to the eighth batch of leaching slurry at a mass ratio of 1:0.035 of the eighth batch of roasted powder to the eighth batch of leaching aid to obtain the eighth batch of leaching slurry.

[0239] The 8th batch of leaching slurry was stirred and leached at 95°C for 9 hours, followed by the 8th solid-liquid separation, to obtain the 8th batch of leaching clear liquid and the 8th batch of leaching residue.

[0240] The 8th batch of leachate was entirely from the 8th batch of leachate.

[0241] The 8th batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0242] Step 3: Adjust the pH value of the leachate

[0243] The first batch of leachate, the second batch of leachate, ..., the seventh batch of leachate and the eighth batch of leachate were mixed. A pH adjuster, namely calcium oxide, was added to the mixed leachate to adjust the pH value to 2.3, thus obtaining the adjusted leachate. Sodium sulfite was then added to the adjusted leachate at a liquid-to-solid ratio of 1:5 L / g. The mixture was then stirred for 50 min at 75°C and a rotation speed of 7 r / min, and filtered to obtain the original extract and the pH-adjusted residue.

[0244] Step 4: Solvent extraction and vanadium precipitation

[0245] The organic phase and the extraction solution were mixed at a volume ratio of 1:2.7, and the mixture was subjected to countercurrent extraction at 25°C. The phases were separated to obtain raffinate and a loaded organic phase. The loaded organic phase and the back-extraction agent were mixed at a volume ratio of 9:1, and the mixture was subjected to countercurrent back-extraction at 25°C. The phases were separated to obtain vanadium-rich solution and a lean organic phase.

[0246] The vanadium-rich solution was subjected to ammonium salt precipitation of vanadium and ammonia removal to obtain vanadium pentoxide.

[0247] The crushed vanadium shale has the following composition: V2O5 content 1.25 wt%, CaO content 4 wt%, SiO2 content 80 wt%, Al2O3 content 4 wt%, and Fe2O3 content 3 wt%. The vanadium content of mica minerals in the crushed vanadium shale accounts for 80 wt% of the total vanadium content.

[0248] The roasting process is as follows: the roasting temperature is 600℃; the roasting time is 75 minutes.

[0249] The calcined powder has a particle size of less than 0.074 mm, accounting for 40 wt%.

[0250] The extractant is di(2-ethylhexyl)phosphonate.

[0251] The organic phase is a mixture of 22.5 vol% extractant, 8 vol% octanol and 69.5 vol% sulfonated kerosene.

[0252] The stripping agent is a 1.75 mol / L sulfuric acid solution.

[0253] Example 5

[0254] A method for enhanced vanadium extraction from vanadium shale through roasting and acid leaching. The steps of the method described in this specific embodiment are as follows:

[0255] Step 1, roasting

[0256] The crushed vanadium shale is roasted, and then the roasted vanadium shale is ground into fine powder to obtain roasted powder.

[0257] Step 2: Leaching of n batches of roasted powder

[0258] The roasted powder was divided into 10 equal portions and leached in 10 batches in sequence.

[0259] Step 2.1, Leaching of the first batch of roasted powder

[0260] The first batch of roasted powder was mixed with water at a solid-liquid ratio of 1:2.0 kg / L to form a slurry, thus obtaining the first batch of mixed slurry. The first batch of concentrated sulfuric acid was added to the first batch of mixed slurry at a mass ratio of 0.22:1 to the first batch of roasted powder to form a leaching slurry, thus obtaining the first batch of leachate.

[0261] The first batch of leaching agent was added to the first batch of leaching pulp at a mass ratio of 1:0.05 of the first batch of roasted powder to the first batch of leaching aid, and the pulp was adjusted to obtain the first batch of leaching mixture.

[0262] The first batch of leaching slurry was stirred and leached at 100°C for 10 hours, and the first solid-liquid separation was performed to obtain the first batch of leaching liquid and the first batch of leaching residue; wherein: 60 vol% of the first batch of leaching liquid was used as the second batch of circulating leaching agent, and the remainder of the first batch of leaching liquid was used as the first batch of leaching liquid.

[0263] The first batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0264] The first batch of immersion aid consists of 90 wt% fluoride and 10 wt% sodium nitrate; the fluoride is sodium fluoride.

[0265] Step 2.2, Leaching of the second batch of roasted powder

[0266] The second batch of roasted powder was mixed with the second batch of circulating leachate and slurry was prepared to obtain the second batch of mixed slurry. The second batch of concentrated sulfuric acid was added to the second batch of mixed slurry at a mass ratio of 0.20:1, and then water with the same volume as the first batch of leachate was added and slurry was prepared to obtain the second batch of leachate.

[0267] The second batch of leaching aid was added to the second batch of leaching slurry at a mass ratio of 1:0.04 of the second batch of roasted powder to the second batch of leaching aid to obtain the second batch of leaching mixture.

[0268] The second batch of leaching slurry was stirred and leached at 100°C for 10 hours, followed by a second solid-liquid separation to obtain the second batch of leaching liquid and the second batch of leaching residue. 60 vol% of the second batch of leaching liquid was used as the third batch of circulating leaching agent, and the remainder of the second batch of leaching liquid was used as the second batch of leaching liquid.

[0269] The second batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0270] ...

[0271] The leaching process for the second batch of roasted powder described in step 2.2 can be extrapolated to the leaching process for the ninth batch of roasted powder.

[0272] Step 2.9, Leaching of the 9th batch of roasted powder

[0273] The 9th batch of roasted powder was mixed with the 9th batch of circulating leaching agent and slurry was prepared to obtain the 9th batch of mixed slurry. The 9th batch of concentrated sulfuric acid was added to the 9th batch of mixed slurry at a mass ratio of 0.20:1, and then water with the same volume as the 8th batch of leaching liquid was added and slurry was prepared to obtain the 9th batch of leaching slurry.

[0274] The 9th batch of leaching aid was added to the 9th batch of leaching slurry at a mass ratio of 1:0.04 (roasted powder to leaching aid) to obtain the 9th batch of leaching mixture.

[0275] The 9th batch of leaching slurry was stirred and leached at 100°C for 10 hours, followed by a 9th solid-liquid separation to obtain the 9th batch of leaching liquid and the 9th batch of leaching residue; wherein, 60 vol% of the 9th batch of leaching liquid was used as the 10th batch of circulating leaching agent, and the remainder of the 9th batch of leaching liquid was used as the 9th batch of leaching liquid.

[0276] The 9th batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0277] Step 2.10, Leaching of the 10th batch of roasted powder

[0278] The 10th batch of roasted powder was mixed with the 10th batch of circulating leachate and slurry was prepared to obtain the 10th batch of mixed slurry. The 10th batch of concentrated sulfuric acid was added to the 10th batch of mixed slurry at a mass ratio of 0.20:1, and then water with the same volume as the 9th batch of leachate was added and slurry was prepared to obtain the 10th batch of leachate.

[0279] The 10th batch of leaching aid was added to the 10th batch of leaching slurry at a mass ratio of 1:0.04 of the 10th batch of roasted powder to the 10th batch of leaching aid to obtain the 10th batch of leaching mixture.

[0280] The 10th batch of leaching slurry was stirred and leached at 100°C for 10 hours, followed by the 10th solid-liquid separation to obtain the 10th batch of leaching clear liquid and the 10th batch of leaching residue.

[0281] All of the 10th batch of leachate was from the 10th batch of leachate.

[0282] The 10th batch of leachate was used as the finished leachate and entered the pH adjustment process.

[0283] Step 3: Adjust the pH value of the leachate

[0284] The first batch of leachate, the second batch of leachate, ..., the ninth batch of leachate and the tenth batch of leachate were mixed. A pH adjuster, namely calcium hydroxide, was added to the mixed leachate to adjust the pH value to 2.6, thus obtaining the adjusted leachate. Then, sodium sulfite was added to the adjusted leachate at a liquid-to-solid ratio of 1:6 L / g. The mixture was then stirred for 60 min at 80°C and a rotation speed of 10 r / min, and filtered to obtain the original extract and the pH-adjusted residue.

[0285] Step 4: Solvent extraction and vanadium precipitation

[0286] The organic phase and the extraction solution are mixed at a volume ratio of 1:3, and countercurrent extraction is performed at 30°C to separate the phases, resulting in raffinate and a loaded organic phase. The loaded organic phase and the back-extraction agent are mixed at a volume ratio of 10:1, and countercurrent back-extraction is performed at 30°C to separate the phases, resulting in a vanadium-rich solution and a lean organic phase.

[0287] The vanadium-rich solution was subjected to ammonium salt precipitation of vanadium and ammonia removal to obtain vanadium pentoxide.

[0288] The crushed vanadium shale has the following composition: V2O5 content 1.5wt%, CaO content 6wt%, SiO2 content 90wt%, Al2O3 content 5wt%, and Fe2O3 content 4wt%. Mica minerals account for 95wt% of the total vanadium content in the crushed vanadium shale.

[0289] The roasting process is as follows: the roasting temperature is 700℃; the roasting time is 90min.

[0290] The roasted powder has a particle size of less than 0.074 mm, accounting for 50 wt%.

[0291] The extractant is 2-ethylhexylphosphonic acid mono-2-ethylhexyl ester.

[0292] The organic phase is a mixture of 25 vol% extractant, 10 vol% octanol and 65 vol% sulfonated kerosene.

[0293] The stripping agent is a 2 mol / L sulfuric acid solution.

[0294] This specific implementation method has the following advantages compared with the prior art:

[0295] (1) The leaching process uses less acid and achieves a high vanadium leaching rate. This specific embodiment adds nitrates / nitrites as leaching aids, building upon existing techniques that use fluorides. The mechanism by which this specific embodiment significantly improves the vanadium leaching rate with low acid dosage is as follows:

[0296] ①Oxidation

[0297] Vanadium in vanadium shale typically exists in mica through lattice substitution, generally in a low valence state (e.g., +3), exhibiting high stability and difficulty in dissolution. During leaching, nitrate ions form nitric acid under acidic conditions. This nitric acid adsorbs at the leaching reaction interface, oxidizing the low-valence vanadium in the vanadium shale mica to a higher valence state (e.g., +4, +5). This oxidation process distorts the mica lattice, increasing the chemical reactivity of vanadium within the lattice and reducing its binding force to the mica lattice, thus facilitating vanadium dissolution under low-acid conditions.

[0298] ②Coordination effect

[0299] Nitrate ions can undergo coordination reactions with vanadium, aluminum, and magnesium ions produced by the decomposition of vanadium-containing mica. During sulfuric acid leaching, nitrate ions act as ligands, combining with the gradually dissolving vanadium, aluminum, and magnesium ions to form relatively stable coordination compounds. This coordination equilibrium promotes the continuous dissolution of mica structural elements, facilitating the continuous transfer of vanadium from vanadium-containing mica into the solution system and improving the vanadium leaching effect.

[0300] (2) Low leaching aid consumption and high vanadium leaching rate. In this specific embodiment, vanadium shale roasted powder is leached in batches, and part of the acid leaching solution from the previous batch is recycled as a circulating leaching aid to enhance the leaching of the next batch of roasted powder. The mechanism by which this specific embodiment improves the vanadium leaching rate with low leaching aid dosage is as follows: the circulating leaching agent contains hydrogen ions, sulfate ions, vanadium ions, and other metal ions. Hydrogen ions and sulfate ions can replenish the acid consumed in the leaching reaction, maintain acidity, and promote the dissolution of vanadium-containing minerals. In addition, as the number of cycles increases, the concentration of each metal ion in the leaching solution increases accordingly, and the ionic strength gradually increases, which enhances the destruction of vanadium-containing minerals and reduces the consumption of leaching aid in the leaching process.

[0301] (3) Good separation effect between vanadium and impurities. In this specific embodiment, a strong oxidizing nitrate ion is introduced into the leaching system of roasted powder. Compared with the prior art, the mechanism by which this specific embodiment achieves selective separation of vanadium and impurities is as follows: nitrate ions can selectively oxidize ferrous ions in the leachate, oxidizing ferrous iron to ferric iron. In the pH adjustment process, calcium hydroxide or calcium sulfate is generally selected as the pH adjuster. However, calcium hydroxide or calcium sulfate is prone to local overconcentration during the addition process. When the local pH value in the leachate is too high, ferric iron is easily co-precipitated with calcium sulfate and enters the pH adjustment residue, achieving selective removal of iron ions in the leachate. It has been determined that the iron ion content in the extraction solution can be reduced to below 600 mg / L, and the vanadium ion concentration is greater than 1.6 g / L, achieving the separation of vanadium and iron in the leachate.

[0302] Therefore, this specific embodiment has the characteristics of low acid consumption, low leaching aid consumption, high vanadium leaching rate, and good separation effect between vanadium and impurities during the leaching process.

Claims

1. A method for strengthening vanadium extraction by roasting and acid leaching of vanadium shale, characterized in that, The method comprises the following steps: Step 1, roasting The crushed vanadium shale is roasted and ground to obtain roasted powder; The particle size of the crushed vanadium shale is ≤3 mm; the crushed vanadium shale has a V2O5 content of 0.5-1.5 wt%, a CaO content of 1-6 wt%, a SiO2 content of 50-90 wt%, an Al2O3 content of 2-5 wt%, and an Fe2O3 content of 1-4 wt%; The vanadium content of the mica mineral in the crushed vanadium shale accounts for 40-95 wt% of the total vanadium content; The roasting process is as follows: the roasting temperature is 400-700 DEG C; the roasting time is 30-90 min; and the particle size of the roasted powder is less than 0.074 mm, accounting for 20-50 wt%; Step 2, batch leaching of the roasted powder The roasted powder is divided into n equal parts, wherein n is a natural number of 3-10; and the n batches are sequentially leached; The first batch of roasted powder is mixed with water to obtain a first batch of mixed slurry; Concentrated sulfuric acid is added to the first batch of mixed slurry to adjust the slurry, and a first batch of leaching slurry is obtained; An immersion aid is added to the first batch of leaching slurry to adjust the slurry, and a leaching mixed slurry is obtained; The leaching mixed slurry is stirred and leached at 80-100 DEG C for 4-10 h, and then solid-liquid separation is performed to obtain leaching clear liquid and secondary leaching residue; 40-60 vol% of the leaching clear liquid is used as a recycled leaching agent for the next batch, and the remaining leaching clear liquid is used as a finished leaching liquid; From the second batch, the following steps are performed: The roasted powder is mixed with the recycled leaching agent produced in the previous batch to obtain a mixed slurry; Concentrated sulfuric acid and water with the same volume as the finished leaching liquid produced in the previous batch are added to the mixed slurry to obtain a leaching slurry; An immersion aid is added to the secondary leaching slurry to adjust the slurry, and a leaching mixed slurry is obtained; The leaching mixed slurry is stirred and leached at 80-100 DEG C for 4-10 h, and then solid-liquid separation is performed to obtain leaching clear liquid and secondary leaching residue; 40-60 vol% of the leaching clear liquid is used as a recycled leaching agent for the next batch, and the remaining leaching clear liquid is used as a finished leaching liquid; In the n th batch, the leaching clear liquid is all used as a finished leaching liquid; In the first batch leaching, the mass ratio of concentrated sulfuric acid to roasted powder is 0.16-0.22:1; in the remaining batches, the mass ratio of concentrated sulfuric acid to roasted powder is 0.14-0.20:1; In the first batch leaching, the mass ratio of roasted powder to immersion aid is 1:0.02-0.05; in the remaining batches, the mass ratio of roasted powder to immersion aid is 1:0.015-0.04; The immersion aid is a mixture of sodium nitrate or sodium nitrite and a fluoride; the content of the fluoride in the immersion aid is 70-90 wt%; and the fluoride is one or more of calcium fluoride and sodium fluoride; Step 3, pH value adjustment of the leaching liquid The finished leaching liquids of all batches are mixed, and the pH value is adjusted to 1.6-2.6 to obtain an adjusted leaching liquid; Sodium sulfite is added to the adjusted leaching liquid, stirred, and filtered to obtain a raw liquid and a pH adjustment residue; Step 4, solvent extraction and vanadium precipitation The organic phase is mixed with the raw liquid, countercurrently extracted at 10-30 DEG C, and separated to obtain a raffinate and a loaded organic phase; The loaded organic phase is mixed with a stripping agent, countercurrent stripping is carried out at 10-30℃, and phase separation is carried out to obtain a vanadium-rich liquid and a lean organic phase; The vanadium-rich liquid is subjected to ammonium salt vanadium precipitation and ammonia removal to obtain vanadium pentoxide.

2. The method for strengthening vanadium extraction by roasting and acid leaching of vanadium shale according to claim 1, characterized in that, In step 3, the adjusted leaching liquid is added with sodium sulfite at a liquid-solid ratio of 1:(2-6) L / g; The stirring is carried out at 60-80℃ and a rotation speed of 1-10 r / min for 30-60 min.

3. The method for strengthening vanadium extraction by roasting and acid leaching of vanadium-containing shale according to claim 1, characterized in that, In step 4, the volume ratio of the organic phase to the original leaching liquid is 1:(2-3), and the volume ratio of the loaded organic phase to the stripping agent is (4-10):

1.

4. The method for strengthening vanadium extraction by roasting and acid leaching of vanadium-containing shale according to claim 1, characterized in that, The organic phase is a mixture of 15-25 vol% of an extractant, 5-10 vol% of sec-octanol and 65-80 vol% of sulfonated kerosene; the extractant is one or more of di(2-ethylhexyl) phosphonate, 2-ethylhexyl phosphonate mono 2-ethylhexyl ester.

5. The method for strengthening vanadium extraction by roasting and acid leaching of vanadium-containing shale according to claim 1, characterized in that, The stripping agent is a 1-2 mol / L sulfuric acid solution.

6. The method for strengthening vanadium extraction by roasting and acid leaching of vanadium-containing shale according to claim 1, characterized in that, The lean organic phase is regenerated and used as the organic phase in step 4. The lean organic phase is regenerated and used as the organic phase in step 4.