Method for extracting vanadium and rubidium from high-silicon shale vanadium ore and preparing high-purity quartz

Abstract

The invention discloses a method for extracting vanadium and rubidium from high-silicon shale vanadium ore and synchronously preparing high-purity quartz, and belongs to the technical field of comprehensive utilization of resources and hydrometallurgy. According to the method, high-silicon shale vanadium ore is used as a raw material, and high-purity vanadium, rubidium and silicon products are efficiently separated and prepared through a series of synergistic processes such as superconducting high-intensity magnetic separation, hot-pressing acid leaching and multi-stage counter-current extraction. In the finally obtained product, the purity of V2O5 is greater than or equal to 99%, the purity of RbCl is greater than or equal to 99%, and the content of SiO2 in high-purity quartz is greater than or equal to According to the method, gradient separation and full-quantification resource recovery of multiple components such as silicon, vanadium, rubidium and iron in the high-silicon shale vanadium ore are successfully achieved, the key technical problems of silica gel interference, rubidium element loss, low added value of quartz and the like are effectively solved, and a new way is provided for efficient development and utilization of the ore difficult to treat; the method has important industrial application value and environmental benefit.

Description

TECHNICAL FIELD

[0001] The present application relates to the field of comprehensive utilization of resources and hydrometallurgy, and in particular to a method for extracting vanadium and rubidium from high-silicon shale vanadium ore and preparing high-purity quartz. BACKGROUND

[0002] Shale vanadium ore is a key component of China's vanadium resource system, and plays a crucial strategic role in ensuring national resource security, supporting the development of a steel power and new energy industry. Among them, high-silicon shale vanadium ore has huge reserves, but its mineral composition is complex, the crystal lattice structure is stable, the vanadium grade is low, and it is often associated with rubidium, molybdenum, cobalt and other dispersed metals. Traditional processes such as "sodium roasting-water leaching" or "calcium roasting-acid leaching" mainly focus on the extraction of vanadium, but fail to achieve efficient recovery of valuable components such as silicon and rubidium, resulting in a large amount of silicon slag that can easily cause environmental accumulation and resource waste, thereby restricting the green and sustainable development of the industry.

[0003] Although the prior art has improved in terms of single vanadium extraction or quartz purification, there are still obvious bottlenecks: (1) Vanadium extraction: Conventional roasting has poor leaching effect on vanadium in high-silicon ore; a large amount of silicon is dissolved in the acid leaching process to form a colloid, which seriously interferes with the subsequent liquid-solid separation and extraction of vanadium and rubidium.

[0004] (2) High-purity quartz preparation: The preparation of high-purity quartz requires high purity of the raw material itself, and it is extremely difficult to obtain high-purity quartz from complex ores through beneficiation, especially the deep removal of impurities such as iron and aluminum.

[0005] (3) Rubidium recovery: As an associated element, rubidium is not recovered in the original process and is dispersed in various slag phases or solutions, causing the loss of strategic resources.

[0006] (4) Process sustainability: The existing preparation process is long, high in energy consumption, and produces a large amount of waste, which is poor in economy and environmental protection.

[0007] Therefore, it is of great significance to develop a collaborative processing method that can simultaneously solve the efficient extraction of vanadium and rubidium and the high-value utilization of high-silicon slag, and realize the resource utilization of all components. SUMMARY

[0008] The main purpose of the present application is to provide a method for extracting vanadium and rubidium from high-silicon shale vanadium ore and preparing high-purity quartz, which realizes the step-by-step separation and full recovery of silicon, vanadium, rubidium and iron components in the ore through multi-technology coupling, and solves the key technical problems of silicon colloid interference, rubidium resource loss and low added value of quartz.

[0009] To achieve the above purpose, the present application provides the following technical solutions: A method for extracting vanadium, rubidium from high-silicon shale vanadium ore and preparing high-purity quartz, comprising the following steps: S1: mixing high-silicon shale vanadium ore with additives, mechanically grinding and activating the mixture to obtain activated powder; S2: microwave roasting, water quenching, and wet grinding of the activated powder in an oxygen-rich environment to obtain ore slurry A; superconducting magnetic separation of the ore slurry A to obtain coarse silicon powder, filtrate, and iron, vanadium, and rubidium-rich ore, wherein the filtrate is returned to the high-temperature water quenching system; S3: floating the coarse silicon powder using a composite collector to obtain quartz sand and flotation tailings, and then hot-pressing and acid leaching the quartz sand to obtain high-purity quartz with SiO2 purity > 99.95% and acid leaching filtrate; S4: magnetic separation of the iron, vanadium, and rubidium-rich ore to obtain vanadium and rubidium concentrates and iron-rich ore; then mixing the vanadium and rubidium concentrates with the flotation tailings and acid leaching filtrate from S3 to obtain ore slurry B; and then supplementing acid to the ore slurry B and performing ultrasonic enhanced acid leaching to obtain tailings and vanadium and rubidium acid solution; S5: selectively reducing iron ions from the vanadium and rubidium solution to obtain a purified vanadium and rubidium solution, and then performing three-stage countercurrent extraction I to obtain a vanadium-containing loaded organic phase and an extracted solution; S6: stripping I of the vanadium-containing loaded organic phase to obtain a vanadium-rich solution, and then precipitating and calcining the vanadium-rich solution to obtain ≥99% V2O5 product; S7: using an ion exchange-electric field strengthening system to separate and enrich rubidium from the extracted solution in step S5 to obtain a rubidium-rich solution, and then performing three-stage countercurrent extraction II, stripping II, and evaporation crystallization to obtain ≥99% pure RbCl product.

[0010] Preferably, the SiO2 content in the high-silicon shale vanadium ore in S1 is ≥70%, the V content is ≥0.3%, and the Rb content is ≥0.03%; the additives are one of sodium carbonate, calcium fluoride, sodium hydroxide, or waste alkali slag from the petroleum, paper, and textile industries; the mass ratio of high-silicon shale vanadium ore to additives is 1: (0.02-0.05); and the mechanical grinding particle size is 75-150 µm.

[0011] Preferably, the microwave roasting temperature in S2 is 500-900°C, the roasting time is 50-120 min, the air flow rate is 9-12 m 3 / h, the water quenching time is 3-10 min, the wet grinding particle size is 25-58 µm, the superconducting magnetic separation strength is 4-7 T, the ore slurry A concentration is 8-15%, and the flow rate is 50-150 m 3 / t.

[0012] Preferably, the composite collector in S3 is prepared according to the mass ratio of anionic collector to nonionic surfactant (3-1):1; the anionic collector is one of plant oil sulfonate or petroleum sulfonate; the nonionic surfactant is one of fatty alcohol polyoxyethylene ether or alkylphenol polyoxyethylene ether; the total amount of the composite collector is 150-600 g / t; the hot pressure acid leaching temperature is 180-220 DEG C, the reaction time is 5-10 h, the liquid-solid ratio is (3-5):1 m 3 / t, the mixed acid for acid leaching is hydrochloric acid and sulfuric acid with a volume ratio of (1.5-2):1.

[0013] Preferably, the magnetic field strength of the magnetic separation in S4 is 0.5-1.5 T, the concentration of the ore pulp B is 6-10%, and the flow rate is 80-100 m 3 / t; the volume ratio of the ore pulp B to the mixed acid is (3-5):1, the ultrasonic power is 400-1200 W, the reaction time is 2-5 h, and the reaction temperature is 60-80 DEG C.

[0014] Preferably, in S5, ascorbic acid is added to the vanadium and rubidium solution to reduce the iron ions, and the molar ratio of ascorbic acid to Fe 3+的 is (1.2-1.5):1; after reduction for 1-2 h, calcium carbonate is added to adjust the pH of the solution to 3.5-4, and the solution is reacted for 0.5-1 h, so that Fe 2+ forms Fe(OH)2 precipitate or FeCO3 precipitate, and the vanadium and rubidium purified solution is obtained after filtration; the pH of the vanadium and rubidium purified solution is adjusted to 1.8-1.9, NaClO3 is added according to m(V):m(NaClO3)=(0.6-1):1, and low-valence vanadium is oxidized to V 5+ ; 0.5-1.0 mol / L methyltrioctylammonium carbonate and 0.2-0.5 mol / L tributyl phosphate are used for three-stage countercurrent extraction (O / A=1:(2-3)) with sulfonated kerosene as a diluent.

[0015] Preferably, in S6, the stripping agent is 2.5-3.5 mol / L NaOH, and the calcination temperature is 350-500 DEG C.

[0016] Preferably, in the ion exchange-electric field strengthening system in S7, the resin is one of D418 amino phosphoric acid chelating resin or macroporous strong basic anion exchange resin, the operating voltage is 50-150 V, the eluent is 4% NaOH+4% NaCl solution, the three-stage countercurrent extraction agent is t-BAMBP with a concentration of 0.5-1.0 mol / L, the extraction phase ratio (O / A) is (4-6):1, and the stripping agent is hydrochloric acid with a concentration of 1-6 mol / L.

[0017] The beneficial effects of the present application are: 1. The silicon-breaking and vanadium-extracting idea of the present application efficiently destroys the structure of vanadium minerals while avoiding the activation of a large amount of silicon through mechanical activation-microwave roasting coupling, and directly separates most of the silicon through subsequent magnetic separation, thereby eliminating the problem of silica gel from the source and creating conditions for subsequent wet process.

[0018] 2. The multi-component synergistic recovery method of the present application realizes the step-by-step separation and high-value utilization of Si, V, Rb, and Fe through quartz flotation purification, vanadium-rubidium ultrasonic acid leaching, rubidium ion exchange-electric field strengthening recovery, and converts waste slag into products, thereby achieving significant economic benefits.

[0019] 3. The present application reduces energy consumption by more than 50% through the use of microwave heating, realizes waste treatment with waste through the use of waste alkali slag as an additive, and maximizes waste emission reduction and environmental protection through multiple cycles such as filtrate, tailings, and acid liquid within the process.

[0020] 4. The high-selectivity separation method of the present application has extremely high selectivity through the use of methyltrioctylammonium carbonate-tri-n-butyl phosphate to synergistically extract vanadium, and solves the problem of low-grade rubidium recovery through the use of ion exchange-electric field strengthening technology to adsorb and separate trace rubidium.

[0021] 5. The present application first realizes the synergistic and efficient extraction of vanadium, rubidium, and quartz from high-silicon shale vanadium ore, and has advanced technology, strong innovation, and high industrial application value. BRIEF DESCRIPTION OF DRAWINGS

[0022] Figure 1 The process flow diagram for extracting vanadium and rubidium from high-silicon shale vanadium ore and preparing high-purity quartz is described in the present application. DETAILED DESCRIPTION

[0023] The technical solutions in the embodiments of the present application will be described clearly and completely below with reference to the drawings in the embodiments of the present application. Obviously, the described embodiments are only a part of the embodiments of the present application, rather than all the embodiments. Based on the embodiments in the present application, all other embodiments obtained by those skilled in the art without creative labor fall within the scope of protection of the present application.

[0024] Example 1 Raw material: high-silicon shale vanadium ore from a certain place, composition analyzed as follows: SiO2 70%, V 0.4%, Rb 0.03%, and the rest being Fe, Al, K, Na, Ca, and other impurities. The preparation method is as follows: S1: Mix the high-silicon shale vanadium ore with petroleum refining waste alkali slag (main components are NaOH and Na2CO3) at a mass ratio of 1:0.02, and mechanically grind and activate to a particle size of 150 µm to obtain an activated powder.

[0025] S2: The activated powder is placed in a microwave calcination furnace and calcined at an air flow rate of 10 m 3 / h and a temperature of 750℃ for 120 min. After the calcined clinker is quenched in high-temperature water for 5 min, wet grinding is performed to a particle size of about 40 µm. Subsequently, superconducting high-intensity magnetic separation is performed at a magnetic field strength of 5.5 T, a slurry concentration of 12%, and a flow rate of 100 m 3 / t. Coarse silicon powder, filtrate, and iron, vanadium, and rubidium-rich ore are obtained; the filtrate is returned to the water quenching system for recycling.

[0026] S3: The coarse silicon powder is subjected to flotation using a composite collector composed of petroleum sulfonate and alkyl phenol polyoxyethylene ether at a mass ratio of 2:1, with a total dosage of 150 g / t. Quartz sand and flotation tailings are obtained. The quartz sand is subjected to hot-pressure acid leaching using mixed acid (hydrochloric acid:sulfuric acid = 1.5:1) at 180℃ and a liquid-solid ratio of 5:1 m 3 / t for 7 h. High-purity quartz sand with a SiO2 purity of 99.95% and acid leaching filtrate are obtained.

[0027] S4: The iron, vanadium, and rubidium-rich ore obtained in S2 is subjected to magnetic separation scavenging (magnetic field strength 0.5 T, slurry concentration 10%) to obtain vanadium and rubidium concentrates and iron-rich ore; the vanadium and rubidium concentrates are mixed with the flotation tailings and acid leaching filtrate from S3, and after adding acid, acid leaching is performed at 60℃ for 2 h under ultrasonic intensification of 1200 W (slurry to mixed acid volume ratio 4:1), finally obtaining tailings and vanadium and rubidium acid liquor.

[0028] S5: Ascorbic acid (molar ratio to Fe 3+ 2+ of 1.3:1) is added to the acid liquor and reduced for 1 h. Calcium carbonate is then added to adjust the pH to 3.8, and the reaction is carried out for 0.8 h to precipitate Fe 2+ 2+, and the purified vanadium and rubidium liquor is obtained by filtration; the pH of the purified liquor is adjusted to 1.85, and NaClO3 is added in a m(V):m(NaClO3) ratio of 0.6:1 for oxidation. Three-stage countercurrent extraction (O / A = 1:3) is performed using 0.5 mol / L methyltrioctylammonium carbonate and 0.5 mol / L tributyl phosphate as the diluent, and sulfonated kerosene as the diluent, to obtain vanadium-containing loaded organic phase and rubidium-containing post-extraction liquid.

[0029] S6: The vanadium-containing loaded organic phase is stripped using 3.5 mol / L NaOH solution to obtain a vanadium-rich solution; after ammonium salt precipitation and calcination at 350℃, high-purity V2O5 product with a purity of 99.5% is obtained.

[0030] S7: The rubidium-containing extract solution was treated using an ion exchange-electric field enhancement integrated system with D418 resin and an operating voltage of 150V. After adsorption, it was desorbed with 4% NaOH + 4% NaCl solution to obtain a rubidium-rich solution. The rubidium-rich solution was subjected to three-stage countercurrent extraction, back-extraction, and evaporation crystallization to finally obtain a high-purity RbCl product with a purity of 99.2%. The three-stage countercurrent extractant was t-BAMBP with a concentration of 0.5 mol / L and an extraction ratio (O / A) of 5:1. The back-extraction agent was hydrochloric acid with a concentration of 6 mol / L. Example 2

[0031] Raw material: High-silica shale vanadium ore, with the following composition after analysis: SiO2 85%, V 0.3%, Rb 0.04%, and the remainder being impurities such as Fe, Al, K, Na, and Ca.

[0032] S1: Mix high-silica shale vanadium ore with sodium carbonate at a mass ratio of 1:0.05, and mechanically grind and activate it to a particle size of 75µm to obtain activated powder.

[0033] S2: Place the activated powder in a microwave baking oven with an airflow of 9m³ / h. 3 The clinker was calcined at 900℃ for 80 minutes per hour. After being quenched in high-temperature water for 10 minutes, it was wet-milled to a particle size of approximately 25µm. Subsequently, it underwent superconducting magnetic separation at a magnetic field strength of 4T, a slurry concentration of 8%, and a flow rate of 150m / h. 3 / t. The process yields coarse silica powder, filtrate, and rich iron, vanadium, and rubidium ores; the filtrate is returned to the water quenching system for recycling.

[0034] S3: Crude silica powder is floated using a composite collector, which is a mixture of vegetable oil sulfonate and fatty alcohol polyoxyethylene ether at a mass ratio of 3:1, with a total dosage of 400 g / t. Quartz sand and flotation tailings are obtained. The quartz sand is then floated at 220℃ with a liquid-to-solid ratio of 3:1. 3 Under the given conditions, hot-pressing acid leaching was performed for 5 hours using a mixed acid solution (hydrochloric acid: sulfuric acid = 1.8:1). High-purity quartz sand with SiO2 purity of 99.96% and acid leaching filtrate were obtained.

[0035] S4: The iron, vanadium, and rubidium rich ore obtained from S2 is subjected to magnetic separation (magnetic field strength 1.5T, slurry concentration 8%) to obtain vanadium, rubidium concentrate and rich iron ore; the vanadium and rubidium concentrate is mixed with the flotation tailings and acid leaching filtrate generated from S3, and after acid replenishment, it is acid leached at 80℃ for 5 hours under ultrasonic enhancement of 400 (slurry to mixed acid volume ratio 5:1) to finally obtain tailings and vanadium and rubidium acid solution.

[0036] S5: Add ascorbic acid (with Fe) to the acid solution. 3+ The molar ratio was 1.5:1) for reduction for 1 hour. Then, calcium carbonate was added to adjust the pH to 4.0, and the reaction was continued for 1 hour to allow Fe... 2+Precipitation and filtration yielded a vanadium- and rubidium-containing purified solution. The pH of the purified solution was adjusted to 1.8, and NaClO3 was added at a ratio of m(V):m(NaClO3) = 0.8:1 for oxidation. Three-stage countercurrent extraction (O / A = 1:2) was performed using 0.8 mol / L methyltrioctylammonium carbonate and 0.2 mol / L tributyl phosphate, diluted with sulfonated kerosene, to obtain a vanadium-supported organic phase and a rubidium-containing extract.

[0037] S6: The vanadium-supported organic phase was back-extracted with 2.5 mol / L NaOH solution to obtain a vanadium-rich solution; after precipitation with ammonium salt and calcination at 500℃, a high-purity V2O5 product with a purity of 99.1% was obtained.

[0038] S7: The rubidium-containing extract solution was treated using an integrated ion exchange-electric field enhancement system. A macroporous strong-base anion exchange resin was used, with an operating voltage of 50V. After adsorption, the solution was desorbed with 4% NaOH + 4% NaCl solution to obtain a rubidium-rich solution. The rubidium-rich solution was subjected to three-stage countercurrent extraction, back-extraction, and evaporation crystallization to finally obtain a high-purity RbCl product with a purity of 99.0%. The three-stage countercurrent extractant was t-BAMBP with a concentration of 0.8 mol / L and an extraction ratio (O / A) of 6:1. The back-extraction agent was hydrochloric acid with a concentration of 3 mol / L.

[0039] Example 3 Raw material: High-silica shale vanadium ore, with the following composition after analysis: SiO2 78%, V 0.4%, Rb 0.03%, and the remainder being impurities such as Fe, Al, K, Na, and Ca.

[0040] S1: Mix high-silica shale vanadium ore with calcium fluoride at a mass ratio of 1:0.04, and mechanically grind and activate it until the particle size is 120µm to obtain activated powder.

[0041] S2: Place the activated powder in a microwave calcination oven with an airflow of 12m³ / h. 3 The clinker was roasted at 500℃ for 50 minutes per hour. After being quenched in high-temperature water for 3 minutes, it was wet-milled to a particle size of approximately 58µm. Subsequently, it underwent superconducting magnetic separation at a magnetic field strength of 7T, a slurry concentration of 15%, and a flow rate of 50m / h. 3 / t. The process yields coarse silica powder, filtrate, and rich iron, vanadium, and rubidium ores; the filtrate is returned to the water quenching system for recycling.

[0042] S3: Crude silica powder is floated using a composite collector, which is a mixture of vegetable oil sulfonate and alkylphenol polyoxyethylene ether at a mass ratio of 1:1, with a total dosage of 600 g / t. Quartz sand and flotation tailings are obtained. The quartz sand is then floated at 200℃ with a liquid-to-solid ratio of 4:1. 3Under the given conditions, hot-pressing acid leaching was performed for 10 hours using a mixed acid solution (hydrochloric acid: sulfuric acid = 2:1). High-purity quartz sand with SiO2 purity of 99.95% and acid leaching filtrate were obtained.

[0043] S4: The iron, vanadium, and rubidium rich ore obtained from S2 is subjected to magnetic separation (magnetic field strength 1.2T, slurry concentration 6%) to obtain vanadium and rubidium concentrate and rich iron ore; the vanadium and rubidium concentrate is mixed with the flotation tailings and acid leaching filtrate generated from S3, and after acid replenishment, it is acid leached at 70℃ for 3 hours under ultrasonic enhancement of 800W (slurry to mixed acid volume ratio 5:1) to finally obtain tailings and vanadium and rubidium acid solution.

[0044] S5: Add ascorbic acid (with Fe) to the acid solution. 3+ The reduction was carried out at a molar ratio of 1.2:1 for 1 hour. Then, calcium carbonate was added to adjust the pH to 3.5, and the reaction was continued for 0.5 hours to allow Fe to... 2+ Precipitation and filtration yielded a vanadium- and rubidium-purified solution. The pH of the purified solution was adjusted to 1.9, and NaClO3 was added at a ratio of m(V):m(NaClO3) = 1:1 for oxidation. A three-stage countercurrent extraction (O / A = 1:2.5) was performed using 1 mol / L methyltrioctylammonium carbonate and 0.3 mol / L tributyl phosphate, diluted with sulfonated kerosene, to obtain a vanadium-supported organic phase and a rubidium-containing extract.

[0045] S6: The vanadium-supported organic phase was back-extracted with 3.0 mol / L NaOH solution to obtain a vanadium-rich solution; after precipitation with ammonium salt and calcination at 400℃, a high-purity V2O5 product with a purity of 99.5% was obtained.

[0046] S7: The rubidium-containing extract solution was treated using an ion exchange-electric field enhancement system. A macroporous strong-base anion exchange resin was used, with an operating voltage of 100V. After adsorption, the solution was desorbed with 4% NaOH + 4% NaCl solution to obtain a rubidium-rich solution. The rubidium-rich solution was subjected to three-stage countercurrent extraction, back-extraction, and evaporation crystallization to finally obtain a high-purity RbCl product with a purity of 99.1%. The three-stage countercurrent extractant was t-BAMBP with a concentration of 1.0 mol / L and an extraction ratio (O / A) of 4:1. The back-extraction agent was hydrochloric acid with a concentration of 1 mol / L.

[0047] The present invention's silicon-vanadium extraction approach, through mechanical activation coupled with microwave roasting, efficiently destroys the vanadium mineral structure while avoiding excessive silicon activation. Subsequent magnetic separation directly separates most of the silicon, eliminating the silica gel problem at its source and creating conditions for subsequent wet processing. The multi-component synergistic recovery method of the present invention, through quartz flotation purification, vanadium-rubidium ultrasonic acid leaching, and rubidium ion exchange-electric field enhanced recovery, achieves the stepwise separation and high-value utilization of the four main components: Si, V, Rb, and Fe, transforming waste residue into products with significant economic benefits. The present invention reduces energy consumption by more than 50% by using microwave heating; utilizes waste alkali residue as an additive to achieve waste-to-waste treatment; and incorporates multiple internal processes such as filtrate, tailings, and acid to minimize waste emissions, making it environmentally friendly. The highly selective separation method of the present invention employs methyltrioctylammonium carbonate-tributyl phosphate synergistic extraction of vanadium, achieving extremely high selectivity; and utilizes ion exchange-electric field enhanced technology to specifically adsorb and separate trace amounts of rubidium, solving the problem of rubidium recovery at low grades. This invention is the first to achieve the synergistic and efficient extraction of vanadium, rubidium, and quartz from high-silica shale vanadium ore. The process is advanced, highly innovative, and has extremely high industrial application value.

[0048] In summary, this invention successfully achieves the cascade separation and full-scale resource recovery of multiple components such as silicon, vanadium, rubidium, and iron in high-silica shale vanadium ore, effectively solving key technical problems such as silica interference, rubidium loss, and low added value of quartz. It provides a new approach for the efficient development and utilization of such difficult-to-process ores and has significant industrial application value and environmental benefits.

[0049] The specific embodiments of the invention have been described in detail above, but these are merely examples, and the invention is not limited to the specific embodiments described above. For those skilled in the art, any equivalent modifications or substitutions to the invention are also within the scope of this invention. Therefore, all equivalent transformations, modifications, and improvements made without departing from the spirit and principles of this invention should be included within the scope of this invention.

Claims

1. A method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz, characterized in that, Includes the following steps: S1: Mix high-silica shale vanadium ore with additives, and mechanically grind and activate the mixture to obtain activated powder; S2: The activated powder is microwave roasted, water quenched, and wet ground in an oxygen-rich environment to obtain slurry A; slurry A is subjected to superconducting strong magnetic separation to obtain coarse silicon powder, filtrate, and iron, vanadium, and rubidium rich ore, of which the filtrate is returned to the high-temperature water quenching system; S3: The coarse silica powder is floated using a composite collector to obtain quartz sand and flotation tailings. The quartz sand is then subjected to hot pressing and acid leaching to obtain high-purity quartz with SiO2 purity >99.95% and acid leaching filtrate. S4: Magnetic separation is performed on iron, vanadium, and rubidium rich ore to obtain vanadium and rubidium concentrate and rich iron ore; then, the vanadium and rubidium concentrate is mixed with the flotation tailings and acid leaching filtrate from S3 to obtain slurry B; slurry B is then subjected to ultrasonic enhanced acid leaching after acid replenishment to finally obtain tailings and vanadium and rubidium acid solution; S5: After selectively reducing vanadium and rubidium solutions to remove iron ions, a vanadium and rubidium purified solution is obtained. This solution is then subjected to three-stage countercurrent extraction I to obtain a vanadium-loaded organic phase and an extract solution. S6: The vanadium-supported organic phase is back-extracted to obtain a vanadium-rich solution, which is then precipitated and calcined to obtain a V2O5 product with ≥99% purity. S7: Rubidium is separated and enriched from the extraction liquid in step S5 using an ion exchange-electric field enhancement system to obtain a rubidium-rich solution, which is then subjected to three-stage countercurrent extraction II, back-extraction II, and evaporation crystallization to finally obtain an RbCl product with a purity ≥99%.

2. The method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz as described in claim 1, characterized in that: The high-silica shale vanadium ore in S1 contains SiO2 content ≥70%, V content ≥0.3%, and Rb content ≥0.03%; the additive is one of sodium carbonate, calcium fluoride, sodium hydroxide, or waste alkali residue generated from petroleum, papermaking, or textile industries; the mass ratio of high-silica shale vanadium ore to additive is 1:(0.02~0.05); the mechanically ground particle size is 75~150µm.

3. The method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz as described in claim 1, characterized in that, In S2, the microwave calcination temperature is 500~900℃, the calcination time is 50~120min, and the air flow rate is 9~12m. 3 / h; water quenching time is 3~10min; wet grinding particle size is 25~58µm; superconducting magnetic separation intensity is 4~7T; slurry A concentration is 8~15%; flow rate is 50~150m / h 3 / t.

4. The method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz as described in claim 1, characterized in that, The S3 flotation composite collector is prepared according to the mass ratio of anionic collector to nonionic surfactant (3~1:1); the anionic collector is one of vegetable oil sulfonate or petroleum sulfonate; the nonionic surfactant is one of fatty alcohol polyoxyethylene ether or alkylphenol polyoxyethylene ether; the total amount of composite collector is 150~600g / t; the hot-pressing acid leaching temperature is 180~220℃, the reaction time is 5~10h, and the liquid-solid ratio is (3~5):1m. 3 / t, the acid leaching mixture is hydrochloric acid and sulfuric acid in a volume ratio of (1.5~2):

1.

5. The method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz as described in claim 1, characterized in that, In step S4, the magnetic field strength for magnetic separation is 0.5~1.5T, the pulp B concentration is 6~10%, and the flow rate is 80~100m. 3 / t; the volume ratio of slurry B to mixed acid is (3~5):1, the ultrasonic power is 400~1200W, the reaction time is 2~5h, and the reaction temperature is 60~80℃.

6. The method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz as described in claim 1, characterized in that, In step S5, ascorbic acid is added to the vanadium and rubidium solution to reduce and remove iron ions. The ascorbic acid reacts with Fe... 3+ The molar ratio of Fe is (1.2~1.5):1; after reduction for 1~2 hours, calcium carbonate is added to adjust the pH of the solution to 3.5~4, and the reaction is allowed to proceed for 0.5~1 hours, allowing Fe to... 2+ Fe(OH)₂ or FeCO₃ precipitate is formed, and after filtration, vanadium and rubidium purification solution is obtained. The pH of the vanadium and rubidium purification solution is adjusted to 1.8~1.9, and NaClO₃ is added at a ratio of m(V):m(NaClO₃) = (0.6~1):1 to oxidize the low-valent vanadium to V. 5+ Three-stage countercurrent extraction was performed using 0.5-1.0 mol / L methyltrioctylammonium carbonate and 0.2-0.5 mol / L tributyl phosphate, with sulfonated kerosene as a diluent (O / A=1:(2-3)).

7. The method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz as described in claim 1, characterized in that, The stripping agent in step S6 is 2.5~3.5 mol / L NaOH; the calcination temperature is 350~500℃.

8. The method for extracting vanadium and rubidium from high-silica shale vanadium ore and preparing high-purity quartz as described in claim 1, characterized in that, In the ion exchange-electric field enhancement system in step S7, the resin is either D418 aminophosphate chelating resin or macroporous strong basic anion exchange resin, the operating voltage is 50~150V, the desorbent is 4% NaOH + 4% NaCl solution; the three-stage countercurrent extractant is t-BAMBP with a concentration of 0.5~1.0 mol / L and an extraction ratio (O / A) of (4~6):1; the back-extraction agent is hydrochloric acid with a concentration of 1~6 mol / L.

Images