Method for preparing electrolyte from stone coal navajoite

Abstract

The invention discloses a method for preparing an electrolyte by utilizing stone coal vanadium ore. The method comprises the following steps: S1, pretreatment and leaching: carrying out crushing and acid leaching treatment on raw ore; s2, purification and impurity separation: filtering the leachate by adopting a chemical precipitation method to remove precipitates; s3, extraction and back extraction process: adopting a two-stage counter-current extraction process; s4, process optimization and innovation: using digital twin modeling to synchronize actual production data in real time; and S5, finally tempering the electrolyte. In the leaching stage, the ore is heated to 100 DEG C within 10 minutes by virtue of a microwave-assisted acid leaching technology, the reaction time is greatly shortened to 30 minutes from the conventional reaction time, the leaching rate is as high as 95% or above, and meanwhile, the energy consumption is reduced by 45%; and after the leaching aid is added, the vanadium leaching rate is increased from 60% to 90% or above. In the purification step, the molar ratio of calcium to aluminum is accurately controlled to be 1: 0.3, and under the condition that the pH value is 1.5-2.0, the impurity removal rate exceeds 95%, the vanadium loss is lower than 1%, and hydrolysis loss is effectively avoided.

Description

Technical Field

[0001] This invention relates to the field of electrolyte preparation technology, specifically a method for preparing electrolyte using vanadium shale. Background Technology

[0002] With the rapid development of vanadium redox flow battery energy storage technology, the efficient utilization of vanadium shale ore, as an important vanadium resource, is of great significance for electrolyte preparation. Traditional methods for preparing electrolytes from vanadium shale ore face numerous bottlenecks. In the leaching stage, conventional acid leaching reactions are slow, taking several hours to complete, with a vanadium leaching rate of only about 60%. Furthermore, due to the lack of effective means to disrupt the ore structure, a large amount of vanadium oxide is trapped within silicate minerals and difficult to release. During purification, impurity separation is incomplete, making it difficult to simultaneously and effectively remove impurities such as silicon, phosphorus, and heavy metals, and easily leading to vanadium loss. In the extraction stage, low extraction efficiency and poor extractant reuse rates result in high production costs. In addition, traditional processes improperly handle leaching residue, waste acid, and tail gas, causing not only resource waste but also environmental pollution. Moreover, limited means of electrolyte quality control make it difficult to guarantee electrolyte stability and battery performance, failing to meet the demands of large-scale, low-cost vanadium redox flow battery applications. Therefore, innovative technological breakthroughs are urgently needed, requiring a method for preparing electrolytes from vanadium shale ore. Summary of the Invention

[0003] The purpose of this invention is to provide a method for preparing electrolyte using vanadium shale, which improves electrolyte preparation efficiency and reduces costs through a multi-step approach, thereby solving the technical problems mentioned in the background art.

[0004] To achieve the above objectives, the present invention provides the following technical solution: a method for preparing an electrolyte using vanadium shale, comprising the following steps:

[0005] S1. Pretreatment and leaching: including crushing and acid leaching of the raw ore;

[0006] First, the raw ore is crushed to a particle size of less than 1 mm and then sieved through a 200-mesh screen. Larger particles of the raw ore are repeatedly crushed. Then, it is acid-leached with concentrated sulfuric acid (4-5 mol / L) at a liquid-to-solid ratio of 0.8-1.5 L / kg.

[0007] Add leaching aids to the acid leaching process and react at high temperature for 20-40 minutes;

[0008] S2. Purification and impurity separation: Chemical precipitation is used to filter and remove precipitates from the leachate;

[0009] Add a mixed precipitant of calcium salt (CaCO3) and aluminum salt (Al2(SO4)3) to the leachate and adjust the pH to 1.5-2.0;

[0010] S3, extraction and stripping process: two-stage countercurrent extraction process is adopted, anion extractant is used, P204 (di(2-ethylhexyl) phosphate) or N235 (trioctylamine) is used as extractant, and the pH is controlled at 1.8-2.8;

[0011] S4, process optimization and innovation: digital twin modeling is used, a three-dimensional model of the reactor is built based on COMSOL Multiphysics, real-time synchronization of actual production data is realized, impurity concentration distribution and reaction endpoint are predicted, dynamic control is used, ore composition is detected online by X-ray fluorescence spectrometer (XRF), sulfuric acid concentration and microwave power are dynamically adjusted according to ore composition (detected online by XRF), sulfuric acid concentration is dynamically adjusted by ±0.2 mol / L, vanadium concentration is monitored in real time by ultraviolet spectrum (UV-Vis), and two-phase flow rate ratio is adjusted;

[0012] Innovative process converts "waste" such as leaching residue, waste acid, tail gas, etc. into resources: after ammonium fluoride is used to extract silicon and aluminum from acid leaching residue, the residue is made into non-fired building materials (compressive strength ≥ 15 MPa), the concentration of ammonium fluoride solution is 2-4 mol / L, and the reaction temperature is 60-80℃; the raffinate is supplemented with acid and then used for leaching process, and the water recycling rate reaches 95%; hydrogen sulfide waste gas is recovered by catalytic oxidation to recover elemental sulfur, and more than 1,000 tons of SO2 is reduced annually, and through vanadium electrolyte solidification and transportation technology, liquid electrolyte is converted into solid salt (such as NH4VO3), and transportation cost is reduced by 70%;

[0013] S5, final electrolyte conditioning: an electric conductivity meter with an accuracy of ±0.01 mol / L is used, calibrated every 30 minutes, sulfuric acid concentration is adjusted, sulfuric acid concentration is controlled to 1.8-2.0 mol / L, vanadium electrolyte preparation parameters are detected, and vanadium concentration is guaranteed to be ≥1.5 mol / L;

[0014] In-situ electrochemical reduction technology is introduced, high-valence vanadium (V 5+ →V 3+ / V 4+ ) is directly reduced in the electrolyte by electrochemical means, the use of chemical reducing agents is eliminated, and organic-containing wastewater is eliminated.

[0015] Preferably, the acid leaching process in S1 introduces microwave-assisted acid leaching technology, the microwave-assisted acid leaching uses a microwave generator (2.45 GHz) introduced into the acid leaching tank, uses the synchronous action of the thermal effect and non-thermal effect of microwaves, the thermal effect is to directly heat the interior of the ore particles, the temperature is raised to 100℃ within 10 minutes, and the reaction time is shortened to 30 minutes; the non-thermal effect is that the microwave electromagnetic field destroys the silicate mineral lattice and releases the wrapped vanadium oxide, and the leaching rate is increased to more than 95%.

[0016] Preferably, the leaching aid added in the S1 acid leaching process is manganese dioxide or fluorite,

[0017] In the sulfuric acid leaching stage, 3-5% of manganese dioxide powder is added as an oxidizing agent to oxidize low-valence vanadium (V 3+ , V 4+ ) to easily soluble V 5+ (such as VO3 - );

[0018] The selection of fluorite into the reaction can release F - , react with silicate minerals to generate SiF6 2- , and prevent silica gel from wrapping the ore to hinder leaching;

[0019] The high temperature is a temperature environment of 90-95°C, and the reaction is carried out at high temperature for 30 minutes, so that the vanadium leaching rate can be increased from 60% to more than 90%.

[0020] Preferably, the mixed precipitator of calcium salt and aluminum salt is added in S2:

[0021] Under the condition of pH=1.5-2.0,

[0022] Ca 2+ reacts with SiO3 2- to generate CaSiO3(Ksp=2.5×10 -8 );

[0023] Al 3+ reacts with PO4 3- to generate AlPO4(Ksp=6.3×10 -19 );

[0024] At the same time, Al(OH)3 colloid adsorbs heavy metals such as Fe 3+ , As 3+ ;

[0025] In addition, the molar ratio of calcium to aluminum in the mixed precipitator should be precisely controlled, and the ratio is 1:0.3.

[0026] Preferably, the S3 P204 extractant selectively adsorbs vanadate;

[0027] Using 10% P204+90% sulfonated kerosene as the organic phase, under the condition of pH=2.0, the vanadium distribution ratio (D) can reach more than 50, while the D of Fe 3+ , Al 3+ is less than 0.1;

[0028] The P204 has a small D value for Fe 3+ , Al 3+The extraction ability of the P204 is extremely weak, and the hydrolysis thereof can be inhibited at high acidity (pH = 2);

[0029] The P204 can be reused for more than 10 times after stripping (5 mol / L H2SO4), and the cost is reduced by 40%.

[0030] Preferably, the relative flow rates of the organic phase and the aqueous phase in the S4 are two-phase flow rates, a T-shaped microchannel is used, the channel width of the T-shaped microchannel is 0.5 mm, efficient mass transfer is realized in a laminar flow state, the extraction time is shortened from 30 minutes to 5 seconds, and the extraction rate is greater than 99%.

[0031] Preferably, in the S5, in-situ electrochemical reduction is achieved by inserting a graphite electrode in an electrolyte, applying a voltage of 0.8-1.2 V, and directly reducing V 5+ (VO3 - ) to the target valence state (V 3+ / V 4+ ) through an electrochemical reaction.

[0032] Preferably, in the S5, the concentration control method of the sulfuric acid is to supplement concentrated sulfuric acid in the stripping solution, and the H + concentration is monitored in real time by a conductivity meter, and finally adjusted to 2.0 mol / L.

[0033] The H + concentration is monitored in real time by a conductivity meter, and the concentrated sulfuric acid is supplemented to the target range.

[0034] Compared with the prior art, the present application has the following beneficial effects:

[0035] The method for preparing an electrolyte from stone coal vanadium ore has significant advantages in many aspects. In the leaching stage, with the aid of microwave-assisted acid leaching technology, the ore is heated to 100℃ within 10 minutes, the reaction time is greatly shortened from the conventional 30 minutes, the leaching rate is as high as 95% or more, and the energy consumption is reduced by 45%; after adding a leaching aid, the vanadium leaching rate jumps from 60% to more than 90%. In the purification link, the molar ratio of calcium to aluminum is accurately controlled to be 1:0.3, under the condition of pH = 1.5-2.0, the impurity removal rate is more than 95%, and the vanadium loss is less than 1%, effectively avoiding hydrolysis loss;

[0036] In the extraction process, two-stage countercurrent extraction technology is used in combination with P204 or N235 extractant, P204 has high selectivity adsorption to vanadate, the vanadium distribution ratio is more than 50 at pH = 2.0, Fe 3+ , Al 3+The D value is less than 0.1, and P204 can be reused more than 10 times after back-extraction, reducing costs by 40%; T-type microchannel extraction reduces extraction time from 30 minutes to 5 seconds, with an extraction rate of more than 99% and a 60% reduction in extractant usage. In terms of resource utilization, production parameters are dynamically adjusted through digital twin modeling, innovatively converting leaching residue, waste acid, and tail gas into non-fired building materials, recycled liquid, and elemental sulfur, with a water recycling rate of 95% and an annual reduction of over 1,000 tons of SO2 emissions.

[0037] During electrolyte conditioning, a high-precision conductivity meter is used to precisely control the sulfuric acid concentration, and in-situ electrochemical reduction technology is introduced to improve electrolyte stability by 30%. It can also solidify and transport liquid electrolyte, reducing costs by 70%. Calcium-aluminum targeted precipitation controls the calcium-aluminum ratio to 1:0.3, removing multiple impurities in one step and reducing reagent costs by 40%. Attached Figure Description

[0038] Figure 1 This is a schematic diagram of the method flow of the present invention. Detailed Implementation

[0039] To make the objectives, technical solutions, and advantages of the present invention clearer, the technical solutions of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments.

[0040] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0041] This invention provides a method for preparing an electrolyte using vanadium shale ore, such as... Figure 1 As shown, it includes the following steps:

[0042] S1. Pretreatment and leaching: including crushing and acid leaching of the raw ore;

[0043] First, the raw ore is crushed to a particle size of less than 1 mm and then sieved through a 200-mesh screen. Larger particles of the raw ore are repeatedly crushed. Then, it is acid-leached with concentrated sulfuric acid (4-5 mol / L) at a liquid-to-solid ratio of 0.8-1.5 L / kg.

[0044] The acid leaching process incorporates microwave-assisted acid leaching technology. This involves introducing a microwave generator (2.45 GHz) into the acid leaching tank, utilizing the simultaneous thermal and non-thermal effects of microwaves. The thermal effect directly heats the interior of the ore particles, raising the temperature to 100°C within 10 minutes and shortening the reaction time to 30 minutes. The non-thermal effect involves the microwave electromagnetic field disrupting the silicate mineral lattice, releasing the encapsulated vanadium oxides, increasing the leaching rate to over 95%. Through microwave action, the ore is heated to 100°C within 10 minutes, shortening the reaction time; and the microwave electromagnetic field disrupts the silicate lattice, releasing the encapsulated vanadium oxides.

[0045] In the acid leaching process, leaching aids are added, such as manganese dioxide or fluorite. During the sulfuric acid leaching stage, manganese dioxide powder is added at 3-5% of the ore mass as an oxidant to remove low-valent vanadium (V). 3+ V 4+ ) oxidized to easily soluble V 5+ (such as VO3) - ); and by choosing fluorite to participate in the reaction, F can be released. - It reacts with silicate minerals to form SiF6. 2- This prevents silica gel from encapsulating the ore and hindering leaching; the high-temperature environment is 90-95℃, and the reaction is carried out at high temperature for 30 minutes, which can increase the vanadium leaching rate from 60% to over 90%.

[0046] S2. Purification and Impurity Separation: Chemical precipitation is used to filter and remove precipitates from the leachate; a mixed precipitant of calcium salt (CaCO3) and aluminum salt (Al2(SO4)3) is added to the leachate, and under pH conditions of 1.5-2.0, Ca... 2+ With SiO3 2- CaSiO3 is generated (Ksp = 2.5 × 10⁻⁶). -8 );Al 3+ With PO4 3- AlPO4 was generated (Ksp = 6.3 × 10⁻⁶). -19 Simultaneously, Al(OH)3 colloid adsorbs Fe. 3+ As 3+ Heavy metals; in addition, the calcium-aluminum molar ratio of the mixed precipitant should be precisely controlled at 1:0.3; hydrolysis loss should be avoided by controlling the pH value, and when the pH value is less than 2.0, vanadium in the solution will be converted into VOCs. 2+ Stable existence, avoiding hydrolysis loss; furthermore, through the synergistic operation of calcium and aluminum, precise proportioning prevents excessive Ca. 2+ Or Al 3+ Residual residues are minimized, ensuring an impurity removal rate of greater than 95% and vanadium loss of less than 1%.

[0047] S3. Extraction and Back-Extraction Process: A two-stage countercurrent extraction process is employed, utilizing anionic extractants. P204 (di(2-ethylhexyl) phosphate) or N235 (trialkylamine) is used as the extractant, with the pH controlled between 1.8 and 2.8. P204 extractant selectively adsorbs vanadate. Using 10% P204 + 90% sulfonated kerosene as the organic phase, extraction is performed at pH 2.0, achieving a vanadium partition ratio (D) of over 50, while Fe... 3+ Al 3+ The D is less than 0.1; the P2O4 has a high affinity for Fe. 3+ Al 3+ The extraction ability is extremely weak, and its hydrolysis can be inhibited under high acidity (pH=2); the P204 can be reused more than 10 times after back-extraction (5mol / LH2SO4), reducing the cost by 40%.

[0048] S4. Process Optimization and Innovation: Digital twin modeling is used to construct a 3D model of the reactor based on COMSOL Multiphysics. Real-time synchronization of actual production data is used to predict impurity concentration distribution and reaction endpoint. Dynamic control is employed, with parameters dynamically adjusted to sulfuric acid concentration fluctuation ±0.2 mol / L and microwave power ±10%. The ore composition is detected online using X-ray fluorescence spectrometry (XRF). Based on the ore composition (detected online by XRF), the sulfuric acid concentration and microwave power are dynamically adjusted, with the sulfuric acid concentration dynamically adjusted to ±0.2 mol / L. Vanadium concentration is monitored in real time using ultraviolet spectroscopy (UV-Vis). The two-phase flow rate ratio is adjusted, with the two-phase flow rate being the relative flow rate of the organic phase and the water phase. A T-shaped microchannel with a channel width of 0.5 mm is used. The two phases achieve efficient mass transfer under laminar flow conditions, reducing the extraction time from 30 minutes to 5 seconds, with an extraction rate greater than 99%.

[0049] Innovative processes transform leaching residue, waste acid, and tail gas—considered "waste"—into resources: after the acid leaching residue is treated with ammonium fluoride to extract silicon and aluminum, the residue is used to produce non-fired building materials (compressive strength ≥15MPa). The ammonium fluoride solution concentration is 2-4mol / L, and the reaction temperature is 60-80℃. The raffinate is reused in the leaching process after acid replenishment, achieving a water recycling rate of 95%. Hydrogen sulfide waste gas is catalytically oxidized to recover elemental sulfur, reducing SO2 emissions by over 1,000 tons annually. Furthermore, vanadium electrolyte solidification transportation technology converts liquid electrolyte into solid salts (such as NH4VO3), reducing transportation costs by 70%.

[0050] S5. Final electrolyte conditioning: A conductivity meter with an accuracy of ±0.01 mol / L is used to calibrate every 30 minutes, adjust the sulfuric acid concentration to control it to 1.8-2.0 mol / L, and check the vanadium electrolyte preparation parameters to ensure that the vanadium concentration is ≥1.5 mol / L;

[0051] The sulfuric acid concentration is controlled by adding concentrated sulfuric acid to the stripping solution, monitoring it in real time using a conductivity meter, and finally adjusting it to H₂O. + Concentration 2.0 mol / L; H+ was monitored in real time using a conductivity meter. + Concentration, add concentrated sulfuric acid to the target range. If H + If the concentration is too low (<1.8 mol / L), vanadium ions are easily hydrolyzed to form V2O5 precipitate, resulting in a decrease in vanadium concentration; if it is too high (>2.0 mol / L), it will accelerate the corrosion of the battery carbon electrode. If the vanadium concentration is too low, it will reduce the energy density of the all-vanadium redox flow battery; if it is too high, the viscosity will increase, affecting the ion migration rate.

[0052] Furthermore, in-situ electrochemical reduction technology was introduced to directly reduce high-valence vanadium (V2) in the electrolyte using electrochemical means. 5+ →V 3+ / V 4+ This eliminates the need for chemical reducing agents and removes wastewater containing organic matter.

[0053] Preferably, in S5, the in-situ electrochemical reduction is achieved by inserting a graphite electrode into the electrolyte and applying a voltage of 0.8-1.2V.

[0054] V is directly converted through an electrochemical reaction. 5+ (VO3 - ) Restore to the target valence state (V) 3+ / V 4+ );

[0055] Specifically, through microwave and chemical synergistic leaching: the ore structure is destroyed by electromagnetic fields, the leaching efficiency is increased to 95%, and energy consumption is reduced by 45%;

[0056] Calcium-aluminum targeted precipitation: Precisely control the calcium-aluminum ratio (1:0.3) to remove silicon, phosphorus, and heavy metals in one step, reducing reagent costs by 40%.

[0057] Microchannel extraction: T-channels (0.5 mm) achieve second-level mass transfer, reducing extractant usage by 60%;

[0058] Electrochemical intelligent conditioning: Voltage-controlled reduction valence state, no chemical reagents required, electrolyte stability improved by 30%.

[0059] Although embodiments of the invention have been shown and described, it will be understood by those skilled in the art that various changes, modifications, substitutions and alterations can be made to these embodiments without departing from the principles and spirit of the invention, the scope of which is defined by the appended claims and their equivalents.

Claims

1. A method for preparing an electrolyte using vanadium shale, characterized in that: Includes the following steps: S1. Pretreatment and leaching: including crushing and acid leaching of the raw ore; First, the raw ore is crushed to a particle size of less than 1 mm and then sieved through a 200-mesh screen. Larger particles of the raw ore are repeatedly crushed. Then, it is acid-leached with concentrated sulfuric acid (4-5 mol / L) at a liquid-to-solid ratio of 0.8-1.5 L / kg. Add leaching aids to the acid leaching process and react at high temperature for 20-40 minutes; S2. Purification and impurity separation: Chemical precipitation is used to filter and remove precipitates from the leachate; Add a mixed precipitant of calcium salt (CaCO3) and aluminum salt (Al2(SO4)3) to the leachate and adjust the pH to 1.5-2.0; S3. Extraction and back-extraction process: A two-stage countercurrent extraction process is adopted, using anionic extractants, with P204 (di(2-ethylhexyl) phosphate) or N235 (trialkylamine) as the extractant, and the pH is controlled at 1.8-2.

8. S4. Process Optimization and Innovation: Using digital twin modeling, a 3D model of the reactor is constructed based on COMSOL Multiphysics. Real-time synchronization of actual production data is used to predict the distribution of impurity concentration and the reaction endpoint. Dynamic control is applied, and the ore composition is detected online by X-ray fluorescence spectrometry (XRF). Based on the ore composition (detected online by XRF), the sulfuric acid concentration and microwave power are dynamically adjusted. The sulfuric acid concentration is dynamically adjusted by ±0.2 mol / L. The vanadium concentration is monitored in real time by ultraviolet spectroscopy (UV-Vis) to adjust the two-phase flow rate ratio. Innovative processes transform leaching residue, waste acid, and tail gas—"waste"—into resources: after the acid leaching residue is treated with ammonium fluoride to extract silicon and aluminum, the residue is made into non-fired building materials (compressive strength ≥15MPa). The ammonium fluoride solution concentration is 2-4mol / L, and the reaction temperature is 60-80℃. The raffinate is reused in the leaching process after acid replenishment, achieving a water recycling rate of 95%. Hydrogen sulfide waste gas is catalytically oxidized to recover elemental sulfur, reducing SO2 emissions by over 1,000 tons annually. Furthermore, vanadium electrolyte solidification transportation technology converts liquid electrolyte into solid salts (such as NH4VO3), reducing transportation costs by 70%. S5. Final electrolyte conditioning: A conductivity meter with an accuracy of ±0.01 mol / L is used to calibrate every 30 minutes, adjust the sulfuric acid concentration to control it to 1.8-2.0 mol / L, and check the vanadium electrolyte preparation parameters to ensure that the vanadium concentration is ≥1.5 mol / L; Furthermore, in-situ electrochemical reduction technology was introduced to directly reduce high-valence vanadium (V2) in the electrolyte using electrochemical means. 5+ →V 3+ / V 4 + This eliminates the need for chemical reducing agents and removes wastewater containing organic matter.

2. The method for preparing electrolyte using vanadium shale ore according to claim 1, characterized in that: The acid leaching process in S1 incorporates microwave-assisted acid leaching technology. This involves introducing a microwave generator (2.45 GHz) into the acid leaching tank, utilizing the simultaneous thermal and non-thermal effects of microwaves. The thermal effect involves directly heating the interior of the ore particles, raising the temperature to 100°C within 10 minutes and shortening the reaction time to 30 minutes. The non-thermal effect involves the microwave electromagnetic field disrupting the silicate mineral lattice, releasing the encapsulated vanadium oxides, and increasing the leaching rate to over 95%.

3. The method for preparing electrolyte using vanadium shale ore according to claim 1, characterized in that: The leaching aid added in the acid leaching process of S1 is manganese dioxide or fluorite. In the sulfuric acid leaching stage, manganese dioxide powder is added at 3-5% of the ore mass as an oxidant to remove low-valent vanadium (V). 3+ V 4+ ) oxidized to easily soluble V 5+ (such as VO3) - ); Choosing to add fluorite to the reaction can release F - It reacts with silicate minerals to form SiF6. 2- To prevent silica gel from encapsulating the ore and hindering leaching; The high temperature is 90-95℃, and the reaction is carried out at the high temperature for 30 minutes, which can increase the vanadium leaching rate from 60% to over 90%.

4. The method for preparing electrolyte using vanadium shale ore according to claim 1, characterized in that: In step S2, a mixed precipitant of calcium and aluminum salts is added simultaneously: Under pH conditions of 1.5-2.0, Ca 2+ With SiO3 2- CaSiO3 is generated (Ksp = 2.5 × 10⁻⁶). -8 ); Al 3+ With PO4 3- AlPO4 was generated (Ksp = 6.3 × 10⁻⁶). -19 ); Simultaneously, Al(OH)3 colloid adsorbs Fe 3+ As 3+ Heavy metals; In addition, the calcium-aluminum molar ratio of the mixed precipitant should be precisely controlled to be 1:0.

3.

5. The method for preparing electrolyte using vanadium shale ore according to claim 1, characterized in that: The selective adsorption of vanadate by the P204 extractant in S3; Using 10% P2O4 + 90% sulfonated kerosene as the organic phase, extraction at pH 2.0 yielded a vanadium partition ratio (D) exceeding 50, while Fe... 3+ Al 3 + The value of D is less than 0.1; The P2O4 on Fe 3+ Al 3 The extraction ability of + is extremely weak, and its hydrolysis can be inhibited under high acidity (pH=2); The P204 can be reused more than 10 times after back-extraction (5 mol / L H2SO4), reducing costs by 40%.

6. The method for preparing electrolyte using vanadium shale ore according to claim 1, characterized in that: In S4, the two-phase flow rate is the relative flow rate between the organic phase and the aqueous phase. A T-shaped microchannel with a channel width of 0.5 mm is used. The two phases achieve efficient mass transfer in laminar flow, reducing the extraction time from 30 minutes to 5 seconds and achieving an extraction rate of over 99%.

7. The method for preparing electrolyte using vanadium shale ore according to claim 1, characterized in that: The in-situ electrochemical reduction in S5 involves inserting a graphite electrode into the electrolyte, applying a voltage of 0.8-1.2V, and directly reducing V through an electrochemical reaction. 5+ (VO3 - ) Restore to the target valence state (V) 3+ / V 4+ ).

8. The method for preparing electrolyte using vanadium shale ore according to claim 1, characterized in that: The sulfuric acid concentration control method in S5 involves adding concentrated sulfuric acid to the back-extraction solution, monitoring it in real time using a conductivity meter, and finally adjusting it to H₂O. + Concentration 2.0 mol / L; H was monitored in real time using a conductivity meter. + Concentration, add concentrated sulfuric acid to the target range.

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