Method for extracting vanadium from high-calcium vanadium-silver ore through calcium fixation roasting and acid leaching

Abstract

The invention relates to a high-calcium vanadium-silver ore calcium fixation roasting-acid leaching vanadium extraction method which comprises the following steps: crushing and grinding high-calcium vanadium-silver ore raw ore to obtain raw ore powder; uniformly mixing the raw ore powder with the sulfide, and then carrying out aerobic roasting to obtain a roasted material; mixing the roasted material with a sulfuric acid solution and water to obtain a mixed material, and carrying out heating leaching and solid-liquid separation on the mixed material to obtain a first vanadium leaching solution; wherein the use amount of H2SO4 in the sulfuric acid solution is 15-30wt% of the weight of the roasted material, and the concentration of the sulfuric acid solution is 450-550g / L. According to the method, the raw ore is subjected to aerobic roasting treatment by adopting the sulfide additive, and a calcium carbonate component which can consume a large amount of sulfuric acid in the acid leaching process is consumed in advance, so that the sulfuric acid consumption in the leaching process is reduced, meanwhile, a dolomite crystal structure is damaged in the roasting process, and the dissolution rate of vanadium in subsequent acid leaching is improved; the method has the characteristics of low acid consumption and environmental friendliness, and is suitable for vanadium extraction treatment of vanadium shale with high calcium oxide content.

Description

Technical Field

[0001] This invention belongs to the field of vanadium shale leaching technology, specifically relating to a method for calcium-fixing roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore. Background Technology

[0002] Primary vanadium-bearing ores are an important strategic resource in my country. Vanadium extraction processes mainly include direct acid leaching, blank roasting-acid leaching, and sodium roasting-water leaching. Vanadium shale with a calcium oxide content >5% is generally referred to as high-calcium mica-type vanadium-bearing ore. Vanadium in this ore primarily exists in the trivalent form within the crystal lattice of silicate minerals such as muscovite. These minerals have stable structures, making vanadium extraction difficult. Furthermore, high contents of minerals such as dolomite and calcite significantly increase acid consumption, resulting in a wide variety and high content of leached impurities. Therefore, removing or converting acid-consuming substances such as dolomite into non-acid-consuming substances before leaching is a crucial pretreatment step to improve vanadium leaching efficiency.

[0003] The patented technology, "A Method for Extracting Vanadium from High-Calcium Coal Shale" (CN109136551B), directly transports the ground slurry to a closed tank, mixes it with concentrated sulfuric acid, and performs leaching. The resulting vanadium leachate is then used to directly prepare vanadium pentoxide. This method solves the problems of complex processes and environmental pollution associated with existing sulfuric acid leaching processes. However, this technology consumes a large amount of acid, with a concentrated sulfuric acid consumption of 38–45 wt%.

[0004] The patented technology, "A Comprehensive Utilization Method for Oxidation-Destructive Roasting of Vanadium-Bearing Stone Ore" (CN111719054B), employs suspension roasting to decarbonize and oxidize powdered vanadium stone ore, followed by acid-curing leaching or direct leaching processes to extract vanadium. This method effectively destroys the crystal structure of vanadium-containing minerals; however, the acid-curing process consumes 40–45 wt% sulfuric acid, and the direct leaching process consumes 40–50 wt%, resulting in high acid consumption.

[0005] The patented technology, "A Method for Direct Acid Leaching Enhanced Vanadium Extraction from Vanadium Shale" (CN119753372A), involves crushing and grinding vanadium shale, followed by 3-10 batches of cyclic acid leaching. The process employs batch-by-batch addition of concentrated sulfuric acid and fluoride leaching aids to enhance the leaching process. While this cyclical process reduces acid consumption and achieves a high vanadium recovery rate, the use of strong acids and fluorides exacerbates equipment wear and poses a risk of fluoride contamination. Furthermore, the method has a complex process flow, requiring multiple solid-liquid separation operations and is only suitable for acid leaching vanadium extraction from vanadium shale with a calcium oxide content below 6%.

[0006] The patented technology, "A method for leaching vanadium from vanadium-bearing coal ore using fluorosilicic acid in conjunction with sulfuric acid" (CN101624649A), employs a mixture of sulfuric acid (10-25 wt% of the vanadium-bearing coal ore mass) and fluorosilicic acid (5-15 wt% of the fluorosilicic acid mass) for vanadium extraction. Although this method reduces acid consumption, the presence of fluorosilicic acid results in a large amount of impurity ions such as iron, aluminum, silicon, and fluorine in the resulting leachate. Furthermore, the fluorosilicic acid used is highly corrosive and may generate hydrogen fluoride during the reaction process, polluting the environment.

[0007] In summary, existing vanadium extraction processes from vanadium shale suffer from problems such as high acid consumption and difficulty in processing vanadium shale with high calcium oxide content. Summary of the Invention

[0008] The purpose of this invention is to overcome the above-mentioned technical deficiencies and provide a method for vanadium extraction from high-calcium vanadium-silver ore by calcium fixation roasting and acid leaching, thereby solving the technical problems of high acid consumption and difficulty in processing vanadium shale with high calcium oxide content in the existing vanadium extraction process.

[0009] To achieve the above-mentioned technical objectives, the technical solution provided by this invention is as follows: In a first aspect, the present invention provides a method for calcium-fixed roasting-acid leaching of high-calcium vanadium-silver ore to extract vanadium, comprising the following steps: S1, crushing and grinding the high-calcium vanadium-silver ore to obtain raw ore powder; S2, mixing the raw ore powder with sulfides and then subjecting it to aerobic roasting to obtain roasted material; S3, mixing the roasted material with sulfuric acid solution and water to obtain a mixture, and subjecting the mixture to heating leaching and solid-liquid separation to obtain a first vanadium leaching solution; wherein, the amount of H2SO4 in the sulfuric acid solution is 15-30 wt% of the weight of the roasted material, and the concentration of the sulfuric acid solution is 450-550 g / L.

[0010] Preferably, in step S1, the high-calcium vanadium silver ore has a V2O5 grade of 0.8% to 1.1%, a CaO grade of 6.0% to 12.0%, and a Fe2O3 grade of 6.9% to 10.3%.

[0011] Preferably, in step S1, more than 75% of the raw ore powder has a particle size of less than 0.074 mm.

[0012] Preferably, in step S2, the sulfide includes one or more of pyrite, Na2S, and Na2S2O5.

[0013] Preferably, in step S2, the molar ratio of calcium to sulfide in the raw ore powder is 1:(0.3-0.5).

[0014] Preferably, in step S2, the calcination temperature for aerobic calcination is 700–900°C, and the calcination time is 60–120 min.

[0015] Preferably, in step S3, the liquid-to-solid ratio of the mixture is (1.5-3) L: 1 kg.

[0016] Preferably, in step S3, the leaching temperature for heating is 80–95°C, and the leaching time is 8–12 h.

[0017] Preferably, the method further includes: S4, adding a reducing agent to the first vanadium leaching solution and obtaining a reducing solution through a reduction reaction; adding ammonium sulfate to the reducing solution, dissolving by heating and stirring, evaporating and concentrating, cooling and crystallizing, and separating the solid and liquid to obtain a second vanadium leaching solution and ferrous ammonium sulfate.

[0018] Further preferably, the reducing agent includes at least one of sodium thiosulfate and sodium sulfite; the molar ratio of ferric ions to the reducing agent in the first vanadium leaching solution is 1:(1.5–2.5); the reduction reaction temperature is 25–40°C, and the time is 30–45 min; the Fe in the reducing solution… 2+ The molar ratio of ammonium sulfate to sodium sulfate is 1:(1-1.5); the heating and stirring temperature for dissolution is 40-60℃, and the time is 10-25 min; the evaporation and concentration temperature is 80-90℃, and the time is 15-30 min.

[0019] Compared with the prior art, the beneficial effects of the present invention include: This invention employs aerobic roasting of the raw ore using sulfide additives. At high temperatures, the SO2 or SO3 gases produced by the decomposition of sulfides undergo a solid-phase transformation reaction with calcium carbonate, the main component of dolomite, under sufficient oxygen conditions. This transforms the calcium carbonate into stable calcium sulfate, pre-consuming the calcium carbonate component that would otherwise consume a large amount of sulfuric acid during acid leaching, thus reducing sulfuric acid consumption. Simultaneously, the roasting process disrupts the dolomite crystal structure, exposing and oxidizing the trivalent vanadium within it to higher vanadium, which improves the vanadium dissolution rate during subsequent acid leaching. This invention avoids the use of highly corrosive fluorosilicic acid, effectively reducing the introduction of impurities and is environmentally friendly. Therefore, this invention features low acid consumption and environmental friendliness, making it suitable for vanadium extraction from vanadium shale with high calcium oxide content. Detailed Implementation

[0020] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be further described in detail below with reference to embodiments. It should be understood that the specific embodiments described herein are merely illustrative and not intended to limit the invention.

[0021] To address the shortcomings of current vanadium shale extraction processes, such as high acid consumption and difficulty in processing vanadium shale with high calcium oxide content, this invention provides a method for vanadium extraction from high-calcium vanadium-silver ore through calcium fixation roasting and acid leaching. This method has low acid consumption, can process vanadium shale with a calcium oxide content of over 6%, and is environmentally friendly.

[0022] In a first aspect, the present invention provides a method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore, comprising the following steps: S1, high-calcium vanadium silver ore is crushed and ground to obtain raw ore powder; S2, the raw ore powder and sulfide are mixed evenly and then subjected to aerobic roasting to obtain roasted material; S3, the roasted material is mixed with sulfuric acid solution and water to obtain a mixture, and the mixture is heated and leached and solid-liquid separated to obtain the first vanadium leachate; wherein, the amount of H2SO4 in the sulfuric acid solution is 15-30 wt% of the weight of the roasted material, and the concentration of the sulfuric acid solution is 450-550 g / L.

[0023] Specifically, the amount of H2SO4 used includes, but is not limited to, 15wt%, 16wt%, 18wt%, 20wt%, 22wt%, 25wt%, 27wt%, 30wt% of the weight of the roasted material; the concentration of the sulfuric acid solution includes, but is not limited to, 450 g / L, 460 g / L, 480 g / L, 490 g / L, 500 g / L, 510 g / L, 520 g / L, 530 g / L, 550 g / L.

[0024] In some preferred embodiments, in step S1, the high-calcium vanadium silver ore has a V2O5 grade of 0.8% to 1.1%, a CaO grade of 6.0% to 12.0%, and a Fe2O3 grade of 6.9% to 10.3%.

[0025] In some preferred embodiments, in step S1, the raw ore powder contains more than 75% particles with a diameter of less than 0.074 mm.

[0026] In some preferred embodiments, in step S2, the sulfide includes one or more of pyrite (FeS2), Na2S, and Na2S2O5.

[0027] In some preferred embodiments, in step S2, the molar ratio of calcium to sulfide in the raw ore powder is 1:(0.3-0.5). Specifically, the molar ratio of calcium to sulfide in the raw ore powder includes, but is not limited to, 1:0.3, 1:0.35, 1:0.4, 1:0.45, 1:0.5, etc. If the amount of sulfide is too high, too much sulfur dioxide gas will be generated, polluting the environment; if the amount is too low, the calcium component in the ore sample will have poor solidification, and the sulfuric acid consumption in the subsequent acid leaching process cannot be effectively reduced.

[0028] In some preferred embodiments, in step S2, the aerobic roasting temperature is 700–900°C, and the roasting time is 60–120 min. Specifically, the roasting temperature includes, but is not limited to, 700°C, 710°C, 720°C, 730°C, 750°C, 760°C, 780°C, 800°C, 820°C, 830°C, 840°C, 850°C, 860°C, 880°C, and 900°C, and the roasting time includes, but is not limited to, 60 min, 70 min, 75 min, 80 min, 90 min, 95 min, 100 min, 105 min, 110 min, and 120 min. Among these, a roasting temperature higher than 900°C will cause the powder to sinter into lumps, which is not conducive to the subsequent vanadium leaching; a temperature lower than 700°C will prevent the calcium fixation reaction from proceeding fully and will not effectively reduce acid consumption.

[0029] It should be noted that the aerobic roasting of the present invention only requires the presence of oxygen, such as direct calcination in an air atmosphere.

[0030] In some preferred embodiments, in step S3, the liquid-to-solid ratio of the mixture is (1.5–3) L:1 kg. The liquid-to-solid ratio refers to the ratio of the total volume of sulfuric acid and water to the mass of the roasted material. Specifically, liquid-to-solid ratios include, but are not limited to, 1.5 L:1 kg, 1.8 L:1 kg, 2 L:1 kg, 2.2 L:1 kg, 2.5 L:1 kg, 2.8 L:1 kg, and 3 L:1 kg. A liquid-to-solid ratio greater than 3 L:1 kg results in an excessively large solution volume, leading to a low vanadium concentration in the leachate, which is detrimental to vanadium recovery in subsequent purification and enrichment processes. A liquid-to-solid ratio less than 1.5 L:1 kg results in a small solution volume, increased viscosity of the reaction system, difficulty in stirring, and ultimately a reduced leaching rate.

[0031] In some preferred embodiments, in step S3, the leaching temperature for heating is 80–95°C, and the leaching time is 8–12 h. Specifically, the leaching temperature includes, but is not limited to, 80°C, 82°C, 83°C, 85°C, 88°C, 90°C, 93°C, and 95°C, and the leaching time includes, but is not limited to, 8 h, 8.5 h, 9 h, 9.5 h, 10 h, 10.5 h, 11 h, and 12 h.

[0032] In some preferred embodiments, the method further includes: S4, adding a reducing agent to the first vanadium leaching solution to obtain a reducing solution through a reduction reaction; adding ammonium sulfate to the reducing solution, dissolving by heating and stirring, evaporating and concentrating, cooling and crystallizing, and separating the solid and liquid to obtain a second vanadium leaching solution and ferrous ammonium sulfate. In this invention, ammonium sulfate is added to the reducing solution and heated and stirred until completely dissolved. Then, ferrous ammonium sulfate is precipitated from the solution in crystalline form by evaporation, concentration, and cooling and crystallization. This effectively recovers iron from high-calcium vanadium-silver ore and further purifies the vanadium-containing solution.

[0033] More preferably, the reducing agent includes at least one of sodium thiosulfate and sodium sulfite.

[0034] More preferably, the molar ratio of ferric ions to reducing agent in the first vanadium leachate is 1:(1.5-2.5). Specifically, the molar ratio of ferric ions to reducing agent includes, but is not limited to, 1:1.5, 1:1.6, 1:1.8, 1:2.0, 1:2.2, 1:2.5, etc.

[0035] More preferably, the reduction reaction temperature is 25–40°C and the time is 30–45 min. Specifically, the reduction reaction temperature includes, but is not limited to, 25°C, 28°C, 30°C, 32°C, 35°C, 38°C, and 40°C, and the time includes, but is not limited to, 30 min, 32 min, 35 min, 38 min, 40 min, 43 min, and 45 min.

[0036] Further preferred, Fe in the reducing solution 2+ The molar ratio with ammonium sulfate is 1:(1–1.5). Specifically, Fe 2+ The molar ratio with ammonium sulfate includes, but is not limited to, 1:1, 1:1.1, 1:1.2, 1:1.3, 1:1.4, 1:1.5, etc.

[0037] Further preferably, the heating and stirring dissolution temperature is 40–60°C, and the time is 10–25 min. Specifically, the heating and stirring dissolution temperature includes, but is not limited to, 40°C, 42°C, 45°C, 46°C, 47°C, 50°C, 52°C, 54°C, 56°C, 58°C, and 60°C, and the time includes, but is not limited to, 10 min, 12 min, 15 min, 18 min, 20 min, 22 min, and 25 min.

[0038] Further preferably, the evaporation and concentration temperature is 80–90°C, and the time is 15–30 min. Specifically, the evaporation and concentration temperature includes, but is not limited to, 80°C, 82°C, 83°C, 85°C, 88°C, 90°C, etc., and the time includes, but is not limited to, 15 min, 18 min, 20 min, 22 min, 25 min, 28 min, 30 min, etc.

[0039] The main mechanism of action and advantages of this invention are as follows: (1) This invention uses high-calcium vanadium-silver ore as raw material. Vanadium mainly exists in the trivalent form in the dolomite mineral lattice. Traditional acid leaching processes require a large amount of sulfuric acid to destroy the dolomite structure. This invention innovatively uses sulfide additives to perform aerobic roasting treatment on the raw ore: at high temperature, the SO2 or SO3 gas generated by the decomposition of sulfides undergoes a solid-phase transformation reaction with calcium carbonate, the main component in dolomite, under sufficient oxygen conditions, transforming it into stable calcium sulfate. The key role of this mineral phase transformation is that it pre-consumes the calcium carbonate component, which consumes a large amount of sulfuric acid during acid leaching, thereby reducing the amount of sulfuric acid consumed in the leaching process. At the same time, the roasting process destroys the crystal structure of dolomite, exposing the trivalent vanadium and oxidizing it into high-valent vanadium, creating favorable conditions for the dissolution of vanadium during subsequent acid leaching.

[0040] (2) This invention introduces ammonium sulfate into the first vanadium leaching solution to efficiently recover iron elements added during the aerobic roasting stage and / or those originally present in the raw ore, thereby avoiding the adverse effects of excessive iron content in the leaching solution on subsequent vanadium purification, enrichment, and precipitation processes. If the iron content exceeds the standard, it will not only interfere with the vanadium purification and enrichment process but also cause co-precipitation during the vanadium precipitation stage, seriously affecting the purity of the final vanadium pentoxide product. The core mechanism is that the ammonium ions in ammonium sulfate can combine with the ferrous ions and sulfate ions in the first vanadium leaching solution to form ferrous ammonium sulfate in the solution environment. After evaporating and concentrating the solution and cooling it to room temperature, ferrous ammonium sulfate crystals can precipitate. After solid-liquid separation, efficient separation and recovery of iron can be achieved.

[0041] (3) This invention avoids the use of highly corrosive fluorosilicic acid, effectively reduces the introduction of impurities, and is environmentally friendly.

[0042] Therefore, this invention features low acid consumption and environmental friendliness, and is suitable for vanadium extraction from vanadium shale with high calcium oxide content.

[0043] The present invention will be further described in detail below through specific embodiments. The concentration of the sulfuric acid solution used is 500 g / L.

[0044] Example 1 A method for calcium-fixing roasting and acid leaching of high-calcium vanadium-silver ore to extract vanadium includes the following steps: S1, high-calcium vanadium-silver ore is crushed and ground until 75% of the particles are smaller than 0.074 mm to obtain raw ore powder; S2, the raw ore powder and pyrite (FeS2) are thoroughly mixed at a molar ratio of calcium to FeS2 of 1:0.3, and then calcined at 800℃ for 75 min to obtain the calcined material; S3, the roasted material is mixed with sulfuric acid solution, H2SO4 accounts for 30wt% of the weight of the roasted material, and water is added at a liquid-solid ratio (the liquid-solid ratio is the ratio of the sum of the volumes of sulfuric acid solution and water to the mass of the roasted material) of 1.5L:1kg. The mixture is heated and leached at 85℃ for 9 hours. The solid and liquid are separated to obtain the first vanadium-containing leachate and leachate residue. S4, sodium thiosulfate was added to the first vanadium-containing leachate at a molar ratio of ferric ions to sodium thiosulfate of 1:1.5, and the reaction was carried out at 25°C for 45 min to obtain a reducing solution; ammonium sulfate was added to the reducing solution at a molar ratio of ferrous ions to ammonium sulfate of 1:1.2, and the reaction was carried out at 40°C for 25 min. The resulting solution was evaporated and concentrated at 80°C for 30 min, and then cooled to room temperature. Solid-liquid separation was performed to obtain the second vanadium-containing leachate and ferrous ammonium sulfate precipitate.

[0045] In this embodiment, the raw ore has a V2O5 grade of 0.98%, a CaO grade of 9.91%, and a Fe2O3 grade of 7.57%. The vanadium concentration in the first vanadium-containing leachate is 3520 ppm, and the vanadium leaching rate is 91.2%. The vanadium concentration in the second vanadium-containing leachate is 3740 ppm, and the vanadium recovery rate is 95.63%, while the iron recovery rate is 84.36%.

[0046] Example 2 A method for calcium-fixing roasting and acid leaching of high-calcium vanadium-silver ore to extract vanadium includes the following steps: S1, high-calcium vanadium-silver ore is crushed and ground until 75% of the particles are smaller than 0.074 mm to obtain raw ore powder; S2, the raw ore powder and pyrite (FeS2) are thoroughly mixed at a molar ratio of calcium to FeS2 of 1:0.5, and then calcined at 750℃ for 90 min to obtain the calcined material; S3, the roasted material is mixed with sulfuric acid solution, H2SO4 accounts for 15wt% of the weight of the roasted material, water is added at a liquid-solid ratio of 2L:1kg, and the mixture is heated and leached at 95℃ for 8 hours. The solid and liquid are separated to obtain the first vanadium-containing leachate and leach residue. S4. Sodium thiosulfate was added to the first vanadium-containing leachate at a molar ratio of ferric ions to sodium thiosulfate of 1:2.5. The reaction was carried out at 40°C for 30 min to obtain a reducing solution. Ammonium sulfate was added to the reducing solution at a molar ratio of ferrous ions to ammonium sulfate of 1:1.5. The reaction was carried out at 60°C for 10 min. The resulting solution was evaporated and concentrated at 85°C for 20 min and then cooled to room temperature. Solid-liquid separation was performed to obtain the second vanadium-containing leachate and ferrous ammonium sulfate precipitate.

[0047] In this embodiment, the raw ore has a V2O5 grade of 1.06%, a CaO grade of 10.83%, and a Fe2O3 grade of 7.01%. The vanadium concentration in the first vanadium-containing leachate is 3300 ppm, and the vanadium leaching rate is 89.8%. The vanadium concentration in the second vanadium-containing leachate is 3542 ppm, and the vanadium recovery rate is 96.60%, while the iron recovery rate is 84.71%.

[0048] Example 3 A method for calcium-fixing roasting and acid leaching of high-calcium vanadium-silver ore to extract vanadium includes the following steps: S1, high-calcium vanadium-silver ore is crushed and ground until 75% of the particles are smaller than 0.074 mm to obtain raw ore powder; S2, the raw ore powder and Na2S are thoroughly mixed at a molar ratio of calcium to Na2S of 1:0.4, and then calcined at 700℃ for 120 min to obtain the calcined material. S3, the roasted material is mixed with sulfuric acid solution, H2SO4 accounts for 20wt% of the weight of the roasted material, water is added at a liquid-solid ratio of 2.5L:1kg, and the mixture is heated and leached at 90℃ for 10 h. The solid and liquid are separated to obtain the first vanadium-containing leachate and leach residue. S4. Sodium sulfite was added to the first vanadium-containing leachate at a molar ratio of ferric ions to sodium sulfite of 1:2. The reaction was carried out at 30°C for 40 min to obtain a reducing solution. Ammonium sulfate was added to the reducing solution at a molar ratio of ferrous ions to ammonium sulfate of 1:1.1. The reaction was carried out at 50°C for 15 min. The resulting solution was evaporated and concentrated at 85°C for 25 min and then cooled to room temperature. Solid-liquid separation was performed to obtain the second vanadium-containing leachate and ferrous ammonium sulfate precipitate.

[0049] In this embodiment, the raw ore has a V2O5 grade of 0.96%, a CaO grade of 9.59%, and a Fe2O3 grade of 8.46%. The vanadium concentration in the first vanadium-containing leachate is 3500 ppm, and the vanadium leaching rate is 90.85%. The vanadium concentration in the second vanadium-containing leachate is 3730 ppm, and the vanadium recovery rate is 95.91%, while the iron recovery rate is 84.45%.

[0050] Example 4 A method for calcium-fixing roasting and acid leaching of high-calcium vanadium-silver ore to extract vanadium includes the following steps: S1, high-calcium vanadium-silver ore is crushed and ground until 75% of the particles are smaller than 0.074 mm to obtain raw ore powder; S2, the raw ore powder and Na2S2O5 are thoroughly mixed at a molar ratio of calcium in the raw ore powder to Na2S2O5 of 1:0.45, and then calcined at 850℃ for 60 min to obtain the calcined material. S3, the roasted material is mixed with sulfuric acid solution, H2SO4 accounts for 25wt% of the weight of the roasted material, water is added at a liquid-solid ratio of 3L:1kg, and the mixture is heated and leached at 80℃ for 12 h. The solid and liquid are separated to obtain the first vanadium-containing leachate and leach residue. S4, sodium sulfite was added to the first vanadium-containing leachate at a molar ratio of ferric ions to sodium sulfite of 1:2.5, and the reaction was carried out at 35°C for 40 min to obtain a reducing solution; ammonium sulfate was added to the reducing solution at a molar ratio of ferrous ions to ammonium sulfate of 1:1.35, and the reaction was carried out at 45°C for 20 min. The resulting solution was evaporated and concentrated at 90°C for 15 min, and then cooled to room temperature. Solid-liquid separation was performed to obtain the second vanadium-containing leachate and ferrous ammonium sulfate precipitate.

[0051] In this embodiment, the raw ore has a V2O5 grade of 1.03%, a CaO grade of 10.37%, and a Fe2O3 grade of 9.03%. The vanadium concentration in the first vanadium-containing leachate is 2780 ppm, and the vanadium leaching rate is 87.89%. The vanadium concentration in the second vanadium-containing leachate is 2950 ppm, and the vanadium recovery rate is 95.50%, while the iron recovery rate is 85.13%.

[0052] Comparative Example 1 Compared with Example 1, the only difference is that the calcination temperature in step S2 is adjusted to 650°C, while the other steps and conditions are the same as in Example 1.

[0053] In this comparative example, the vanadium concentration in the first vanadium-containing leachate was 3140 ppm, and the vanadium leaching rate was only 81.61%.

[0054] Comparative Example 2 Compared with Example 1, the only difference is that the roasting temperature in step S2 is adjusted to 950°C, while the other steps and conditions are the same as in Example 1.

[0055] In this comparative example, the vanadium concentration in the first vanadium-containing leachate was 2770 ppm, and the vanadium leaching rate was only 72.29%.

[0056] Comparative Example 3 Compared with Example 1, the only difference is that the aerobic roasting in step S2 is changed to an anaerobic roasting under a nitrogen atmosphere, while the other steps and conditions are the same as in Example 1.

[0057] In this comparative example, the vanadium concentration in the first vanadium-containing leachate was 1470 ppm, and the vanadium leaching rate was only 38.05%.

[0058] Comparative Example 4 Compared with Example 1, the only difference is that the pyrite in step S2 is removed and the raw ore is directly roasted with oxygen. The other steps and conditions are the same as in Example 1.

[0059] In this comparative example, the vanadium concentration in the first vanadium-containing leachate was 2560 ppm, and the vanadium leaching rate was only 66.59%. Only by increasing the amount of H2SO4 used in step S3 to 62 wt% of the weight of the roasted material did the vanadium leaching rate reach the same level as that of Example 1.

[0060] According to Comparative Examples 1-4, suitable roasting temperature, sufficient oxygen atmosphere, and reasonable addition of sulfides are necessary conditions to ensure that the calcium component in dolomite is converted into insoluble calcium sulfate, thereby reducing acid consumption in the subsequent acid leaching process.

[0061] In summary, this invention involves crushing and grinding high-calcium vanadium-silver ore to obtain raw ore powder of suitable particle size, mixing it evenly with sulfides, and then subjecting it to aerobic high-temperature roasting to obtain roasted material. The roasted material is then mixed with sulfuric acid solution and water and heated for leaching. After solid-liquid separation, a first vanadium-containing leachate is obtained. A reducing agent is added to the first vanadium-containing leachate, and after complete reduction, ammonium sulfate is added, heated and stirred until completely dissolved, evaporated and concentrated, and the solution is cooled to room temperature. Solid-liquid separation yields a second vanadium-containing leachate and the byproduct ferrous ammonium sulfate. This invention effectively solves the problem of large-scale calcium dissolution during the acid leaching process of high-calcium vanadium-silver ore, significantly reduces acid consumption, and is environmentally friendly.

[0062] The specific embodiments of the present invention described above do not constitute a limitation on the scope of protection of the present invention. Any other corresponding changes and modifications made in accordance with the technical concept of the present invention should be included within the scope of protection of the claims of the present invention.

Claims

1. A method for calcium-fixing roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore, characterized in that, Includes the following steps: S1, high-calcium vanadium silver ore is crushed and ground to obtain raw ore powder; S2, the raw ore powder and sulfide are mixed evenly and then subjected to aerobic roasting to obtain roasted material; S3, the roasted material is mixed with sulfuric acid solution and water to obtain a mixture, and the mixture is heated and leached and solid-liquid separated to obtain a first vanadium leaching solution; wherein, the amount of H2SO4 in the sulfuric acid solution is 15-30 wt% of the weight of the roasted material, and the concentration of the sulfuric acid solution is 450-550 g / L.

2. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, In step S1, the high-calcium vanadium silver ore has a V2O5 grade of 0.8% to 1.1%, a CaO grade of 6.0% to 12.0%, and a Fe2O3 grade of 6.9% to 10.3%.

3. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, In step S1, more than 75% of the raw ore powder has a particle size of less than 0.074 mm.

4. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, In step S2, the sulfide includes one or more of pyrite, Na2S, and Na2S2O5.

5. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, In step S2, the molar ratio of calcium to sulfide in the raw ore powder is 1:(0.3 to 0.5).

6. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, In step S2, the calcination temperature of the aerobic calcination is 700-900℃, and the calcination time is 60-120 min.

7. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, In step S3, the liquid-to-solid ratio of the mixture is (1.5-3) L: 1 kg.

8. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, In step S3, the leaching temperature for heating is 80–95°C, and the leaching time is 8–12 h.

9. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 1, characterized in that, Also includes: S4, a reducing agent is added to the first vanadium leaching solution, and a reducing solution is obtained through a reduction reaction; Ammonium sulfate was added to the reducing solution, and the solution was dissolved by heating and stirring, then evaporated and concentrated, cooled and crystallized, and the solid and liquid were separated to obtain the second vanadium leachate and ferrous ammonium sulfate.

10. The method for calcium fixation roasting-acid leaching vanadium extraction from high-calcium vanadium-silver ore according to claim 9, characterized in that, The reducing agent includes at least one of sodium thiosulfate and sodium sulfite; The molar ratio of ferric ions to reducing agent in the first vanadium leachate is 1:(1.5~2.5). The reduction reaction is carried out at a temperature of 25–40°C for a time of 30–45 min. Fe in the reducing solution 2+ The molar ratio with ammonium sulfate is 1:(1~1.5); The heating and stirring process is carried out at a temperature of 40–60°C for 10–25 minutes. The evaporation and concentration are carried out at a temperature of 80–90°C for 15–30 minutes.