Method for preparing vanadium electrolyte from vanadium-containing shale by short process

Abstract

The application discloses a method for preparing vanadium electrolyte from vanadium-containing shale through a short process, and comprises the following steps: crushing, grinding and slurry preparation of vanadium-containing shale to obtain ore slurry; flotation of the ore slurry to obtain flotation concentrate; concentration of the flotation concentrate, sulfuric acid aging and water leaching to obtain vanadium-containing leaching solution; removal of aluminum from the vanadium-containing leaching solution to obtain post-aluminum-removal solution; neutralization and reduction pretreatment of the post-aluminum-removal solution to obtain post-pretreatment solution; one-stage extraction of the post-pretreatment solution to obtain vanadium-rich solution I; two-stage extraction of the vanadium-rich solution I to obtain vanadium-rich solution II; and oil removal treatment of the vanadium-rich solution II to obtain vanadium electrolyte. The method realizes efficient extraction of vanadium from multi-metal vanadium-containing shale through flotation and sulfuric acid aging and water leaching, directly uses the vanadium-containing leaching solution as raw material, and prepares vanadium electrolyte through neutralization and reduction pretreatment, one-stage extraction, two-stage extraction and oil removal process. The method has the characteristics of short process, low cost, environmental friendliness, strong adaptability of raw materials, less reagents and energy consumption, etc.

Description

Technical Field

[0001] This invention relates to the field of vanadium electrolyte technology, and more particularly to a short-process method for preparing vanadium electrolyte from vanadium-containing shale. Background Technology

[0002] Vanadium redox flow batteries, as a highly promising large-scale energy storage technology, offer advantages such as cleanliness, high efficiency, safety, and long lifespan, leading to their increasing application. The vanadium electrolyte determines the energy density and performance of vanadium batteries, but the high purity requirements result in complex and costly preparation processes, limiting the widespread adoption of vanadium batteries. Currently, the mainstream preparation process for vanadium electrolytes uses high-purity solid vanadium pentoxide as raw material, employing chemical reducing agents such as SO2, oxalic acid, and V2O3, or using electrolysis to reduce V2O5 and dissolve it in sulfuric acid. However, using high-purity vanadium pentoxide as a raw material results in high preparation costs.

[0003] The conventional vanadium extraction process from vanadium shale involves an acid leaching-extraction-acidic vanadium precipitation-alkali dissolution-alkali vanadium precipitation flow. However, this process is lengthy, requires multiple oxidation-reduction steps during vanadium precipitation, consumes large amounts of energy and reagents, causes severe pollution, and incurs high production costs. In conventional vanadium shale extraction, the vanadium-containing acid leaching solution often contains various impurity ions such as iron, aluminum, magnesium, potassium, and phosphorus. These impurity ions not only severely affect the purification and enrichment process of vanadium but also reduce the quality of the final vanadium product. High concentrations of iron ions, in particular, easily lead to vanadium loss and hinder the selective enrichment of vanadium ions by organic compounds during extraction, thus affecting the purity of vanadium pentoxide. The traditional process for preparing high-purity V₂O₅ from acidic vanadium-rich solutions typically involves multiple steps, including oxidation, vanadium precipitation, alkali dissolution for impurity removal, re-precipitation of vanadium, and calcination. This lengthy process is accompanied by wastewater and exhaust gas emissions, making it difficult to meet the efficiency and environmental requirements for large-scale vanadium battery applications.

[0004] Therefore, there is an urgent need to provide a short-process method for preparing vanadium electrolyte from vanadium-containing shale to solve the problems of the traditional process for preparing high-purity V2O5 from acidic vanadium-rich liquid being lengthy, energy-intensive, and unable to meet electrolyte standard requirements. Summary of the Invention

[0005] The purpose of this invention is to overcome the above-mentioned technical deficiencies and propose a short-process method for preparing vanadium electrolyte from vanadium-containing shale, which solves the technical problems of the traditional process for preparing high-purity V2O5 from acidic vanadium-rich liquid being lengthy, energy-intensive, and unable to meet electrolyte standard requirements.

[0006] This invention provides a short-process method for preparing vanadium electrolyte from vanadium-containing shale, comprising the following steps: S1. Vanadium-bearing shale is crushed, ground, and mixed to produce slurry; S2. The slurry is subjected to flotation to obtain flotation concentrate and flotation tailings; S3. After concentration of flotation concentrate, it is subjected to sulfuric acid aging and water leaching to obtain vanadium-containing leachate; S4. The vanadium-containing leaching solution is subjected to aluminum removal to obtain the aluminum-removed solution; S5. After the aluminum removal liquid undergoes neutralization-reduction pretreatment, the pretreated liquid is obtained. S6. After pretreatment, the liquid is extracted in one stage to obtain vanadium-rich solution I; S7. Vanadium-rich solution I is subjected to two-stage extraction to obtain vanadium-rich solution II; S8 and vanadium-rich solution II are deoiled to obtain vanadium electrolyte.

[0007] Compared with the prior art, the beneficial effects of the present invention include: This invention achieves efficient extraction of vanadium from polymetallic vanadium-bearing shale by employing a "flotation pre-tailings removal-sulfuric acid ripening-water leaching" process. After obtaining the vanadium-bearing leachate, it can be directly used as raw material, eliminating the complex steps of "oxidation-vanadium precipitation-alkali dissolution-vanadium precipitation-calcination-reduction dissolution in sulfuric acid" in traditional processes. The vanadium electrolyte can be prepared through a single process of "neutralization-reduction pretreatment-first-stage extraction-second-stage extraction-deoiling". This invention can effectively prepare high-concentration vanadium electrolyte from vanadium-bearing shale, and has the characteristics of short process, low cost, environmental friendliness, strong raw material adaptability, and low reagent and energy consumption, combining the advantages of cleanliness, high efficiency and environmental friendliness. Attached Figure Description

[0008] Figure 1 This is a process flow diagram of one embodiment of the method for preparing vanadium electrolyte from vanadium-containing shale using a short process provided by the present invention; Figure 2 This is a process flow diagram of one embodiment of the flotation process in the short-process method for preparing vanadium electrolyte from vanadium-containing shale provided by the present invention. Detailed Implementation

[0009] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be further described in detail below with reference to the accompanying drawings and embodiments. It should be understood that the specific embodiments described herein are merely illustrative and not intended to limit the invention.

[0010] This invention provides a short-process method for preparing vanadium electrolyte from vanadium-containing shale, comprising the following steps: S1. Vanadium-bearing shale is crushed, ground, and mixed to produce slurry; S2. The slurry is subjected to flotation to obtain flotation concentrate and flotation tailings; S3. After concentration of flotation concentrate, it is subjected to sulfuric acid aging and water leaching to obtain vanadium-containing leachate; S4. The vanadium-containing leaching solution is subjected to aluminum removal to obtain the aluminum-removed solution; S5. After the aluminum removal liquid undergoes neutralization-reduction pretreatment, the pretreated liquid is obtained. S6. After pretreatment, the liquid is extracted in one stage to obtain vanadium-rich solution I; S7. Vanadium-rich solution I is subjected to two-stage extraction to obtain vanadium-rich solution II; S8 and vanadium-rich solution II are deoiled to obtain vanadium electrolyte.

[0011] In this invention, a "flotation pre-tailing-sulfuric acid ripening-water leaching" process is employed in the leaching stage to achieve efficient extraction of vanadium from polymetallic vanadium-bearing shale. This invention pre-enriches vanadium-bearing mica through flotation while simultaneously removing impurities such as dolomite and pyrite, reducing the amount of impurity ions leached at the source and significantly reducing sulfuric acid consumption. The sulfuric acid ripening-water leaching effectively disrupts the structure of the vanadium-bearing mica, further saving sulfuric acid usage, inhibiting impurity leaching, and eliminating the need for leaching aids compared to direct acid leaching. This fundamentally avoids the formation of fluoride-aluminum complexes, ensuring the smooth implementation of subsequent crystallization and aluminum removal steps, thereby obtaining a leachate with low impurity content. This invention, after obtaining vanadium-containing leachate, directly uses it as raw material, eliminating the complex processes of "oxidation-vanadium precipitation-alkali dissolution-vanadium precipitation-calcination-reduction dissolution in sulfuric acid" in traditional processes. It can prepare vanadium electrolyte through a single process of "neutralization-reduction pretreatment-first-stage extraction (extraction + back-extraction)-second-stage extraction (extraction + washing + back-extraction)-deoiling," reducing preparation costs by more than 20%. This invention can effectively prepare high-concentration vanadium electrolyte from vanadium-containing shale, featuring a short process, low cost, environmental friendliness, strong raw material adaptability, and low reagent and energy consumption, combining the advantages of cleanliness, high efficiency, and environmental friendliness.

[0012] Existing acidic vanadium extraction technologies are only suitable for low-concentration acidic vanadium-containing solutions (V₂O₅ content 1~10 g / L), exhibiting poor raw material adaptability and failing to meet the requirements for efficient vanadium recovery and impurity separation in vanadium-rich solutions. This invention achieves vanadium enrichment in vanadium-containing solutions through a two-stage extraction process. The first stage extraction enriches vanadium from the vanadium-containing leachate to obtain vanadium-rich solution I. The second stage extraction uses vanadium-rich solution I as raw material, saponifying the extractant (to improve vanadium extraction capacity) to further enrich vanadium in vanadium-rich solution I, yielding vanadium-rich solution II, which meets electrolyte standards. Furthermore, during the extraction process, the vanadium concentration in the electrolyte can be flexibly controlled within the range of 80~150 g / L by adjusting the ratio of the organic phase to the aqueous phase. This method offers convenient control and yields an electrolyte product with high concentration and excellent charge / discharge capacity.

[0013] In this embodiment, in step S1, the vanadium-bearing shale is high-calcium vanadium-bearing shale. Conventional vanadium extraction processes for high-calcium vanadium-bearing shale face the problem of excessively high sulfuric acid consumption. To ensure vanadium leaching rate, excessive sulfuric acid needs to be added, which not only directly increases production costs but also leads to the dissolution of more calcium, magnesium, and other impurity ions into the leachate, increasing the burden and difficulty of subsequent solution purification. This invention, by employing a "flotation pre-tailing-sulfuric acid ripening-water leaching" process, can achieve efficient vanadium extraction from high-calcium vanadium-bearing shale while reducing the leaching amount of impurity ions and decreasing sulfuric acid consumption.

[0014] In some specific embodiments of the present invention, the V2O5 grade in the high-calcium vanadium-bearing shale is 0.70%~1.10%, and the CaO grade is 9%~16%.

[0015] In this embodiment, in step S1, the mass concentration of the slurry is 15%~30%, and the proportion of particles with a size of -0.074 mm is 65%~75%.

[0016] In this embodiment, step S2 includes flotation: reverse flotation and forward flotation.

[0017] Preferably, reverse flotation includes roughing and scavenging.

[0018] More preferably, the coarse selection is performed once.

[0019] More preferably, the number of scans is 1 to 3.

[0020] Preferably, positive flotation includes: roughing, cleaning and scavenging.

[0021] More preferably, the coarse selection is performed once.

[0022] More preferably, the number of selections is 1 to 3.

[0023] More preferably, the number of scans is 1 to 3.

[0024] Specifically, please refer to Figure 2 Reverse flotation includes: After roughing, the slurry yields rougher concentrate I and rougher concentrate I; After roughing, ore I is scavenged to obtain scavenged ore I and final tailings I.

[0025] Among them, the scavenging of ore I is returned to the roughing step of reverse flotation.

[0026] Specifically, please refer to Figure 2 Positive flotation includes: Rough concentrate I is further roughed to obtain rough concentrate II and rough concentrate II; After further refining, rough concentrate II yields flotation concentrate and middlings. After roughing, ore II is scavenged to obtain scavenged ore II and final tailings II.

[0027] Among them, the middlings are returned to the finer flotation step.

[0028] Among them, the roughing step of the scavenging ore II is returned to the positive flotation.

[0029] Preferably, the reagents for reverse flotation include: a pH adjuster and a reverse flotation collector.

[0030] More preferably, the pH adjuster is sulfuric acid, and its dosage is 15~20 kg / t based on the initial dry weight of the slurry.

[0031] More preferably, the reverse flotation collector is an anionic fatty acid collector, including but not limited to sodium oleate, and its dosage is 500~1000 g / t based on the initial dry weight of the pulp.

[0032] More preferably, during the reverse flotation process, after adding the reverse flotation reagent, the pH of the pulp is 4 to 4.5.

[0033] Preferably, the reagents for positive flotation include: inhibitors and positive flotation collectors.

[0034] More preferably, the inhibitor is sodium fluorosilicate, and its dosage is 500~2100 g / t based on the initial dry weight of the slurry, and more preferably 700~2100 g / t.

[0035] More preferably, the positive flotation collector is a highly selective cationic collector, including but not limited to dodecylamine, and its dosage is 200~750 g / t based on the initial dry weight of the pulp, and more preferably 300~750 g / t.

[0036] In this embodiment, the flotation concentrate is concentrated to a pulp mass concentration of 55%~60%.

[0037] In this embodiment, step S3, sulfuric acid aging-water immersion, includes: S31. Mix sulfuric acid with concentrated flotation concentrate and mature it to obtain matured feed; S32. The hydrated material is subjected to water leaching and solid-liquid separation to obtain vanadium-containing leachate and leachate residue.

[0038] Preferably, the mass fraction of sulfuric acid is 93% to 99%.

[0039] Preferably, the amount of sulfuric acid added is 20% to 50% of the dry weight of the concentrated flotation concentrate.

[0040] Preferably, the mixing and aging time is 4-10 hours. In existing sulfuric acid aging processes, external heating is usually required to regulate the temperature of the aging process, thereby ensuring that the aging reaction proceeds fully and improving the vanadium leaching rate. In this invention, vanadium-containing mica is first pre-enriched by flotation while simultaneously removing impurities such as dolomite and pyrite. Then, by controlling the slurry concentration and the ratio of slurry to sulfuric acid, the aging process can be self-heated without the need for additional heating. The temperature of the self-exothermic process is maintained at 120-150°C for a long time to ensure sufficient aging. Finally, efficient vanadium leaching can be achieved through a simple water leaching process.

[0041] Preferably, the ratio of the amount of calcining material to water is 1g:(1~3)mL.

[0042] Preferably, the water immersion temperature is 20~30℃ (i.e., room temperature), the water immersion time is 4~8 h, and the water immersion is carried out under stirring conditions.

[0043] In this embodiment, step S4, aluminum removal includes: mixing and reacting a vanadium-containing leaching solution and an aluminum removal agent, followed by crystallization and solid-liquid separation to obtain an aluminum-removed liquid.

[0044] Preferably, the aluminum removal agent is potassium sulfate.

[0045] Preferably, the molar ratio of aluminum in the vanadium-containing leachate to potassium in the aluminum removal agent is 1:(0.8~1.2).

[0046] Preferably, the crystallization temperature is 0~10℃ and the crystallization time is 2~3 h.

[0047] In this embodiment, the neutralization-reduction pretreatment includes: adjusting the pH of the solution after aluminum removal to 1.9~2.0 with a neutralizing agent, and adding a reducing agent to a potential of -260~-280 mV.

[0048] Preferably, the neutralizing agent is at least one of lime milk, sodium hydroxide, potassium hydroxide, and ammonia water.

[0049] Preferably, the reducing agent is at least one of sodium sulfite, iron powder, sodium thiosulfate, and sulfur dioxide.

[0050] In this embodiment, step S6, the extraction step includes: S61. The pretreated liquid is subjected to normal extraction with the first extractant to obtain the first loaded organic phase and the first raffinate. S62. The first loaded organic phase is back-extracted by the first back-extracting agent to obtain the first lean organic phase and the first back-extraction solution (i.e., vanadium-rich solution I).

[0051] Preferably, the first extractant comprises: di(2-ethylhexyl) phosphate (P204), 2-octanol and sulfonated kerosene, and the volume ratio of di(2-ethylhexyl) phosphate (P204), 2-octanol and sulfonated kerosene is (15~20):5:(75~80).

[0052] Preferably, during the extraction process of the pretreated liquid with the first extractant, the volume ratio of the first extractant to the pretreated liquid (i.e., the aqueous phase) is 1:(1.5~2.5).

[0053] Preferably, during the extraction process of the pretreated liquid with the first extractant, the extraction temperature is 20~30℃ and the extraction time for each stage is 5~10 min.

[0054] Preferably, during the extraction process of the pretreated liquid with the first extractant, the number of extraction stages is 3 to 5.

[0055] Preferably, countercurrent extraction is used during the extraction process of the pretreated liquid with the first extractant.

[0056] Preferably, the first stripping agent is 8 vol% to 12 vol% dilute sulfuric acid.

[0057] Preferably, the volume ratio of the first loaded organic phase to the first stripping agent is (8~10):1.

[0058] Preferably, during the back-extraction process of the first loaded organic phase by the first back-extracting agent, the back-extraction temperature is 20~30℃ and the back-extraction time for each stage is 15~25 min.

[0059] Preferably, during the back-extraction process of the first loaded organic phase by the first back-extractant, the number of back-extraction stages is 2 to 4.

[0060] Preferably, countercurrent back-extraction is used during the back-extraction process of the first loaded organic phase by the first back-extractant.

[0061] In this embodiment, the two-stage extraction includes: S61. Vanadium-rich solution I is subjected to normal extraction with a second extractant to obtain a second loaded organic phase and a second raffinate. S62. The second loaded organic phase is washed with detergent to obtain a washed organic phase and a washed liquid; S63. The washed organic phase is back-extracted by the second back-extracting agent to obtain the second organic-lean phase and the second back-extracting solution (i.e., vanadium-rich solution II).

[0062] Preferably, the second extractant comprises: a mixture of 15 vol% to 30 vol% di(2-ethylhexyl) phosphate and 70 vol% to 85 vol% sulfonated kerosene, which is then saponified by a saponifying agent, with a saponification degree of 40% to 80%.

[0063] The saponifying agent is at least one of sodium hydroxide and sodium ethoxide.

[0064] Preferably, the volume ratio of the second extractant to the vanadium-rich liquid I (i.e., the aqueous phase) is 1:(1~4).

[0065] Preferably, during the extraction process of the vanadium-rich liquid I with the second extractant, the extraction temperature is 20~30℃ and the extraction time for each stage is 8~10 min.

[0066] Preferably, during the extraction process of the vanadium-rich liquid I with the second extractant, the number of extraction stages is 4 to 6.

[0067] Preferably, during the forward extraction process of the vanadium-rich solution I with the second extractant, countercurrent forward extraction is used.

[0068] Preferably, the detergent is 5 vol% to 10 vol% dilute hydrochloric acid.

[0069] Preferably, during the washing process, the volume ratio of detergent to the second loaded organic phase is 1:(1~5).

[0070] Preferably, during the washing process, the washing temperature is 25~40℃, and the washing time for each stage is 5~20 minutes.

[0071] Preferably, the washing process involves 1 to 3 washing stages.

[0072] Preferably, countercurrent washing is used during the washing process.

[0073] Preferably, the second stripping agent is 10 vol% to 15 vol% dilute sulfuric acid.

[0074] Preferably, during the back-extraction of the washing organic phase by the second back-extractant, the volume ratio of the second back-extractant to the washing organic phase is 1:(4~10).

[0075] Preferably, during the back-extraction process of the washing organic phase by the second back-extracting agent, the back-extraction temperature is 20~30℃ and the back-extraction time for each stage is 10~40 min.

[0076] Preferably, during the back-extraction process of the washing organic phase by the second back-extracting agent, the number of back-extraction stages is 3 to 5.

[0077] Preferably, countercurrent back-extraction is used during the back-extraction process of the washing organic phase with the second back-extractant.

[0078] In this embodiment, a stripping process is used for oil removal. Stripping is a prior art technique, and this invention does not limit its application.

[0079] To avoid redundancy, in the following embodiments and comparative examples of the present invention, the high-calcium vanadium-bearing shale is sourced from Yichang, Hubei Province, wherein the V2O5 grade is 0.98% and the CaO grade is 12.5%.

[0080] Example 1 Step 1, Flotation: The slurry with 65% -0.074 mm particles and a mass concentration of 15% is subjected to a reverse flotation decalcification process of "one roughing and one scavenging" and a direct flotation process of "one roughing, one cleaning, and one scavenging" to obtain flotation concentrate and flotation tailings. The reagent dosage is based on the initial dry slurry. The reverse flotation decalcification process is as follows: first, add pH adjuster (15 kg / t sulfuric acid) to the slurry and stir for 2 minutes, then add collector (500 g / t sodium oleate) and stir for 2 minutes. The slurry pH is 4.5. Skim off the bubbles for 3 minutes to obtain rougher concentrate I and rougher medium ore I. Rougher medium ore I is then scavenged and skimmed off for 2 minutes to obtain scavenger medium ore I and final tailings I. The direct flotation process is as follows: first, add inhibitor (500 g / t sodium fluorosilicate) to rougher concentrate I and stir for 2 minutes, then add collector (200 g / t sodium fluorosilicate)... Stir the mixture with 200 g / t dodecylamine for 3 min and skim it for 5 min to obtain rougher concentrate II and middlings II; first add inhibitor (200 g / t sodium fluorosilicate) to rougher concentrate II and stir for 2 min, then add collector (100 g / t dodecylamine) and stir for 2 min and skim it for 5 min to obtain flotation concentrate and middlings; middlings II is then scavenged and skimmed for 3 min to obtain scavenged middlings II and final tailings II; Step 2, Sulfuric Acid Maturation-Water Leaching: The flotation concentrate obtained in Step 1 is concentrated to a mass concentration of 55% using a thickener. Sulfuric acid (mass fraction 98%) is added at 20% of the dry basis mass of the concentrated flotation concentrate. The mixture is matured for 10 hours. The maturation process generates heat and raises the temperature spontaneously. The matured material is mixed with water at a ratio of 1g:1mL and stirred at 20℃ for 8 hours. After solid-liquid separation, vanadium-containing leachate and leaching residue are obtained. Step 3, aluminum removal: The vanadium-containing leachate obtained in Step 2 and the aluminum removal agent (potassium sulfate) are mixed at a molar ratio of 1:1 between aluminum in the vanadium-containing leachate and potassium in the aluminum removal agent. The mixture is then subjected to low-temperature crystallization at 2°C for 2 hours to remove aluminum, resulting in the aluminum-removed solution. Step 4, Neutralization-Reduction Pretreatment: Add a neutralizing agent (lime milk) to the aluminum-removed solution obtained in Step 3 to adjust the solution pH to 1.9, and add a reducing agent (sodium sulfite) until the solution potential is -260 mV to obtain the pretreated solution; Step 5, First-stage extraction: The pretreated liquid obtained in Step 4 is subjected to three-stage countercurrent forward extraction at a volume ratio of organic phase to aqueous phase of 1:1.5, at room temperature, and with each extraction stage lasting 5 min. It is then subjected to two-stage countercurrent back-extraction at a volume ratio of loaded organic phase to back-extraction agent of 8:1, at room temperature, and with each back-extraction stage lasting 15 min, to obtain vanadium-rich solution I. The organic phase is a mixture of di(2-ethylhexyl) phosphate (P204), 2-octanol, and sulfonated kerosene at a volume ratio of 15:5:80. The back-extraction agent is 8 vol% dilute sulfuric acid. Step Six, Two-Stage Extraction: The vanadium-rich solution I obtained in Step Five was subjected to five stages of countercurrent forward extraction at a volume ratio of 1:2 between the extractant and the aqueous phase, at room temperature, and with each extraction stage lasting 8 minutes. Then, it was washed for 20 minutes at 25°C with a volume ratio of 1:2 between the washing solution (5 vol% dilute hydrochloric acid) and the supported organic phase. Finally, it underwent four stages of countercurrent back-extraction at a volume ratio of 1:5 between the back-extraction agent and the washed supported organic phase, at room temperature, and with each back-extraction stage lasting 10 minutes, to obtain vanadium-rich solution II. The extractant was obtained by saponification with sodium hydroxide after mixing 15 vol% di(2-ethylhexyl) phosphate and 85 vol% sulfonated kerosene, with a saponification degree of 50%. The back-extraction agent was 10 vol% dilute sulfuric acid. Step 7: Vanadium-rich liquid II is deoiled by a stripping process to obtain vanadium electrolyte.

[0081] Example 2 Step 1, Flotation: The slurry with 70% -0.074 mm particles and a mass concentration of 23% is subjected to a reverse flotation decalcification process of "one roughing and one scavenging" and a direct flotation process of "one roughing, one cleaning, and one scavenging" to obtain flotation concentrate and flotation tailings. The reagent dosage is based on the initial dry slurry. The reverse flotation decalcification process is as follows: First, add a pH adjuster (17 kg / t sulfuric acid) to the slurry and stir for 2 minutes, then add a collector (750 g / t sodium oleate), stir for 2 minutes until the slurry pH is 4.0, and skim for 3 minutes to obtain rougher concentrate I and rougher medium ore I. Rougher medium ore I is then scavenged and skimmed for 2 minutes to obtain scavenged medium ore I and final tailings I. The direct flotation process is as follows: First, add an inhibitor (1000 g / t sodium fluorosilicate) to the rougher concentrate I and stir for 2 minutes, then add a collector (350 g / t sodium oleate). Stir the mixture with 400 g / t dodecylamine for 3 min and skim it for 5 min to obtain rougher concentrate II and middlings II. First, add inhibitor (400 g / t sodium fluorosilicate) to rougher concentrate II and stir for 2 min, then add collector (180 g / t dodecylamine) and stir for 2 min and skim it for 5 min to obtain flotation concentrate and middlings. After scavenging and skimming the middlings II, obtain scavenged middlings II and final tailings II. Step 2, Sulfuric Acid Maturation-Water Leaching: The flotation concentrate obtained in Step 1 is concentrated to a mass concentration of 57% using a thickener. Sulfuric acid (mass fraction 98%) is added at 35% of the dry basis mass of the concentrated flotation concentrate. The mixture is matured for 7 hours. The maturation process generates heat spontaneously. The matured material is mixed with water at a ratio of 1g:1.5mL and stirred at 25℃ for 6 hours. After solid-liquid separation, vanadium-containing leachate and leaching residue are obtained. Step 3, aluminum removal: The vanadium-containing leachate obtained in Step 2 and the aluminum removal agent (potassium sulfate) are mixed at a molar ratio of aluminum in the vanadium-containing leachate to potassium in the aluminum removal agent of 1:1.2. Then, the mixture is subjected to low-temperature crystallization at 5°C for 2.5 h to remove aluminum, resulting in the aluminum-removed solution. Step 4, Neutralization-reduction pretreatment: Add a neutralizing agent (lime milk) to the aluminum-removed solution obtained in Step 3 to adjust the solution pH to 1.95, and add a reducing agent (sodium sulfite) until the solution potential is -270 mV to obtain the pretreated solution; Step 5, First-stage extraction: The pretreated liquid obtained in Step 4 is subjected to three-stage countercurrent forward extraction at a volume ratio of organic phase to aqueous phase of 1:2, at room temperature, and at each extraction time of 7 min. It is also subjected to two-stage countercurrent back-extraction at a volume ratio of loaded organic phase to back-extraction agent of 9:1, at room temperature, and at each back-extraction time of 20 min, to obtain vanadium-rich solution I. The organic phase is a mixture of di(2-ethylhexyl) phosphate (P204): 2-octanol: sulfonated kerosene at a volume ratio of 17.5:5:77.5, and the back-extraction agent is 10 vol% dilute sulfuric acid. Step Six, Two-Stage Extraction: The vanadium-rich solution I obtained in Step Five was subjected to five stages of countercurrent positive extraction under the conditions of an extractant to aqueous phase volume ratio of 1:3, room temperature, and extraction time of 9 min for each stage. It was then subjected to four stages of countercurrent back-extraction under the conditions of a washing solution (5 vol% dilute hydrochloric acid) to the supported organic phase volume ratio of 1:3, a temperature of 30℃, and a washing time of 15 min. Finally, it was subjected to four stages of countercurrent back-extraction under the conditions of a back-extraction agent to the washed supported organic phase volume ratio of 1:7, room temperature, and back-extraction time of 20 min for each stage, to obtain vanadium-rich solution II. The extractant was obtained by saponification of a mixture of 20 vol% di(2-ethylhexyl) phosphate and 80 vol% sulfonated kerosene with sodium ethoxide, with a saponification degree of 55%. The back-extraction agent was 12 vol% dilute sulfuric acid. Step 7: Vanadium-rich liquid II is deoiled by a stripping process to obtain vanadium electrolyte.

[0082] Example 3 Step 1, Flotation: The slurry with 75% -0.074 mm particles and a mass concentration of 30% is subjected to a reverse flotation decalcification process of "one roughing and one scavenging" and a direct flotation process of "one roughing, one cleaning, and one scavenging" to obtain flotation concentrate and flotation tailings. The reagent dosage is based on the initial dry slurry. The reverse flotation decalcification process is as follows: first, add a pH adjuster (20 kg / t sulfuric acid) to the slurry and stir for 2 minutes, then add a collector (1000 g / t sodium oleate) and stir for 2 minutes. The slurry pH is 4.2. Skim off the bubbles for 3 minutes to obtain rougher concentrate I and rougher medium ore I. Rougher medium ore I is then scavenged and skimmed off for 2 minutes to obtain scavenger medium ore I and final tailings I. The direct flotation process is as follows: first, add an inhibitor (1500 g / t sodium fluorosilicate) to rougher concentrate I and stir for 2 minutes, then add a collector (500 g / t sodium fluorosilicate)... Stir the mixture with 600 g / t dodecylamine for 3 min and skim it for 5 min to obtain rougher concentrate II and middlings II; first add inhibitor (600 g / t sodium fluorosilicate) to rougher concentrate II and stir for 2 min, then add collector (250 g / t dodecylamine) and stir for 2 min and skim it for 5 min to obtain flotation concentrate and middlings; scaveng the middlings II and skim it for 3 min to obtain scavenged middlings II and final tailings II; Step 2, Sulfuric Acid Maturation-Water Leaching: The flotation concentrate obtained in Step 1 is concentrated to a mass concentration of 60% using a thickener. Sulfuric acid (mass fraction 98%) is added at 50% of the dry basis mass of the concentrated flotation concentrate. The mixture is matured for 4 hours. The maturation process generates heat spontaneously. The matured material is mixed with water at a ratio of 1g:2mL and stirred at 30℃ for 4 hours. After solid-liquid separation, vanadium-containing leachate and leaching residue are obtained. Step 3, aluminum removal: The vanadium-containing leachate obtained in Step 2 and the aluminum removal agent (potassium sulfate) are mixed at a molar ratio of aluminum in the vanadium-containing leachate to potassium in the aluminum removal agent of 1:1.1. Then, the mixture is crystallized at a low temperature of 2°C for 3 hours to remove aluminum, and the aluminum-removed solution is obtained. Step 4, Neutralization-Reduction Pretreatment: Add a neutralizing agent (lime milk) to the aluminum-removed solution obtained in Step 3 to adjust the pH of the solution to 2.0, and add a reducing agent (sodium sulfite) until the solution potential is -280 mV to obtain the pretreated solution; Step 5, First-stage extraction: The pretreated liquid obtained in Step 4 is subjected to three-stage countercurrent forward extraction at a volume ratio of organic phase to aqueous phase of 1:2.5, at room temperature, and at each extraction time of 10 min. It is also subjected to two-stage countercurrent back-extraction at a volume ratio of loaded organic phase to back-extraction agent of 10:1, at room temperature, and at each back-extraction time of 25 min, to obtain vanadium-rich solution I. The organic phase is a mixture of di(2-ethylhexyl) phosphate (P204), 2-octanol, and sulfonated kerosene at a volume ratio of 20:5:75. The back-extraction agent is 10 vol% dilute sulfuric acid. Step Six, Two-Stage Extraction: The vanadium-rich solution I obtained in Step Five was subjected to five stages of countercurrent positive extraction under the conditions of an extractant to aqueous phase volume ratio of 1:4, room temperature, and extraction time of 10 min for each stage. It was then washed for 8 min at 40°C with a washing solution (5 vol% dilute hydrochloric acid) to the supported organic phase volume ratio of 1:4, and followed by four stages of countercurrent back-extraction under the conditions of a back-extraction agent to the washed supported organic phase volume ratio of 1:9, room temperature, and back-extraction time of 30 min for each stage, to obtain vanadium-rich solution II. The extractant was obtained by saponification of a mixture of 25 vol% di(2-ethylhexyl) phosphate and 75 vol% sulfonated kerosene with sodium hydroxide, achieving a saponification degree of 60%. The back-extraction agent was 15 vol% dilute sulfuric acid. Step 7: Vanadium-rich liquid II is deoiled by a stripping process to obtain vanadium electrolyte.

[0083] Comparative Example 1 All other conditions are the same as in Example 3, except that in step one, there is no forward flotation after reverse flotation.

[0084] Comparative Example 2 All other conditions are the same as in Example 3, except that in step one, forward flotation is performed directly without reverse flotation.

[0085] Comparative Example 3 All other conditions are the same as in Example 3, except that in step two, the flotation concentrate is concentrated to a mass concentration of 40% by a thickener.

[0086] Comparative Example 4 All other conditions are the same as in Example 3, except that in step two, the flotation concentrate is concentrated to a mass concentration of 80% by a thickener.

[0087] Comparative Example 5 All other conditions are the same as in Example 3, except that step three was not performed.

[0088] Comparative Example 6 All other conditions are the same as in Example 3, except that step six is ​​performed directly without step five.

[0089] Comparative Example 7 All other conditions are the same as in Example 3, except that in step five, the positive extraction agent is obtained by saponification with sodium hydroxide after mixing 25 vol% di(2-ethylhexyl) phosphate and 75 vol% sulfonated kerosene, with a saponification degree of 60%.

[0090] Comparative Example 8 All other conditions are the same as in Example 3, except that in step six, the positive extraction agent is obtained by mixing di(2-ethylhexyl) phosphate (P204): 2-octanol: sulfonated kerosene in a volume ratio of 20:5:75.

[0091] Comparative Example 9 All other conditions are the same as in Example 3, except that in step six, the loaded valence phase is not washed with dilute hydrochloric acid.

[0092] Table 1 Composition of vanadium electrolyte

[0093] Please refer to Table 1. As can be seen from Table 1, the electrolyte components prepared in Examples 1 to 3 of this invention all meet the requirements of Grade I tetravalent electrolyte in the national standard (GB / T-37204-2018).

[0094] Compared with Example 3, Comparative Examples 1 and 2 only performed reverse flotation and direct flotation, respectively. The concentrations of aluminum, iron, and magnesium in the electrolyte were higher, indicating that the combination of direct flotation and reverse flotation can more effectively remove gangue minerals such as dolomite and pyrite from the ore, thereby reducing their concentration in the electrolyte.

[0095] Compared with Example 3, the concentration of flotation concentrate in Comparative Examples 3 and 4 was too low or too high, and the vanadium concentration in the electrolyte was too low. This indicates that both too low and too high pulp concentration (i.e., water content) will have an adverse effect on the leaching rate of the leaching process. Too high water content leads to incomplete maturation, while too low water content leads to a short self-heating maturation time, both of which result in a low vanadium concentration in the electrolyte.

[0096] Compared with Example 3, Comparative Example 5 did not undergo aluminum removal, and the aluminum content in the electrolyte was higher, indicating that the aluminum removal process in step three can effectively remove impurities Al from the solution and obtain a vanadium electrolyte that meets the requirements.

[0097] Compared with Example 3, Comparative Example 6 skipped step five and went directly to step six. The vanadium concentration in the electrolyte was lower and the impurity ion content was significantly higher, indicating that using a single-stage soap extraction alone could not enrich vanadium to a qualified concentration and the impurity ion content was higher.

[0098] Compared with Example 3, in Comparative Example 7, the positive extraction agent in steps five and six was obtained by saponification of a mixture of 25 vol% di(2-ethylhexyl) phosphate and 75 vol% sulfonated kerosene. Although the vanadium concentration was increased to a certain extent, the content of impurity ions in the electrolyte was significantly increased, indicating that the two-stage saponification extraction method could not separate vanadium and impurity ions, resulting in a high content of impurity ions.

[0099] Compared with Example 3, in Comparative Example 8, the positive extraction agent in steps five and six was obtained by mixing di(2-ethylhexyl) phosphate (P204): 2-octanol: sulfonated kerosene in a volume ratio of 20:5:75. The vanadium concentration in the electrolyte was extremely low, indicating that the two-stage extraction alone could not enrich vanadium to a qualified concentration.

[0100] Compared with Example 3, Comparative Example 9 did not undergo washing with dilute hydrochloric acid, resulting in a significant increase in the iron ion concentration of the electrolyte, which caused the electrolyte to fail to meet the standards.

[0101] 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 preparing vanadium electrolyte from vanadium-bearing shale using a short-process method, characterized in that, Includes the following steps: Vanadium-bearing shale is crushed, ground, and mixed to produce slurry; The slurry is subjected to flotation to obtain flotation concentrate and flotation tailings; The flotation concentrate is concentrated and then subjected to sulfuric acid aging and water leaching to obtain a vanadium-containing leachate; The vanadium-containing leachate is subjected to aluminum removal to obtain an aluminum-removed solution; The aluminum-removed liquid is subjected to neutralization-reduction pretreatment to obtain the pretreated liquid; The pretreated liquid is then subjected to a first-stage extraction to obtain vanadium-rich liquid I; The vanadium-rich solution I is subjected to two-stage extraction to obtain vanadium-rich solution II; The vanadium-rich liquid II is deoiled to obtain vanadium electrolyte.

2. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The vanadium-bearing shale is a high-calcium vanadium-bearing shale; and / or, The vanadium-bearing shale is a high-calcium vanadium-bearing shale, and in the high-calcium vanadium-bearing shale, the V₂O₅ grade is 0.70%~1.10%, and the CaO grade is 9%~16%; and / or, The slurry has a mass concentration of 15% to 30% and a particle size of -0.074 mm accounts for 65% to 75%.

3. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The flotation process includes: reverse flotation and forward flotation; wherein... The reverse flotation includes: roughing and scavenging; the roughing is performed once; the scavenging is performed 1 to 3 times; and / or, The positive flotation includes: roughing, cleaning, and scavenging; the roughing is performed once; the cleaning is performed 1 to 3 times; the scavenging is performed 1 to 3 times; and / or, The reverse flotation reagents include: a pH adjuster and a reverse flotation collector; the pH adjuster is sulfuric acid, and its dosage is 15-20 kg / t based on the initial dry weight of the pulp; the reverse flotation collector is an anionic fatty acid collector, and its dosage is 500-1000 g / t based on the initial dry weight of the pulp; and / or, During the reverse flotation process, after adding the reverse flotation reagents, the pH of the pulp is 4-4.5; and / or, The reagents for the positive flotation include: a depressant and a positive flotation collector; the depressant is sodium fluorosilicate, and its dosage is 500~2100 g / t based on the initial dry weight of the pulp; the positive flotation collector is a highly selective cationic collector, and its dosage is 200~750 g / t based on the initial dry weight of the pulp.

4. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The flotation concentrate is concentrated to a pulp concentration of 55% to 60%.

5. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The sulfuric acid aging-water immersion includes: Sulfuric acid is mixed with concentrated flotation concentrate and matured to obtain matured feedstock; The matured material is subjected to water leaching and solid-liquid separation to obtain vanadium-containing leachate and leaching residue; wherein, The sulfuric acid has a mass fraction of 93% to 99%; and / or, The amount of sulfuric acid added is 20% to 50% of the dry basis mass of the concentrated flotation concentrate; and / or, The mixing and maturation time is 4-10 hours; and / or, The ratio of the cured material to water is 1g:(1~3)mL; and / or, The water immersion temperature is 20~30℃, the water immersion time is 4~8 hours, and the water immersion is carried out under stirring conditions.

6. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The aluminum removal process includes: mixing and reacting the vanadium-containing leaching solution and the aluminum-removing agent, followed by crystallization and solid-liquid separation to obtain the aluminum-removed solution; wherein, The aluminum removal agent is potassium sulfate; and / or... The molar ratio of aluminum in the vanadium-containing leachate to potassium in the aluminum removal agent is 1:(0.8~1.2); and / or, The crystallization temperature is 0~10℃, and the crystallization time is 2~3 h.

7. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The neutralization-reduction pretreatment includes: adjusting the pH of the dealuminized solution to 1.9~2.0 with a neutralizing agent, and adding a reducing agent until the potential is -260~-280 mV; wherein, The neutralizing agent is at least one of lime milk, sodium hydroxide, potassium hydroxide, and ammonia water; and / or, The reducing agent is at least one of sodium sulfite, iron powder, sodium thiosulfate, and sulfur dioxide.

8. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The extraction segment includes: The pretreated liquid is subjected to normal extraction with the first extractant to obtain the first loaded organic phase and the first raffinate; The first loaded organic phase is back-extracted using the first back-extracting agent to obtain a first organic-lean phase and a vanadium-rich solution I; wherein, The first extractant comprises: di(2-ethylhexyl) phosphate (P204), 2-octanol, and sulfonated kerosene, wherein the volume ratio of di(2-ethylhexyl) phosphate (P204), 2-octanol, and sulfonated kerosene is (15~20):5:(75~80); and / or, During the extraction process of the pretreated liquid with the first extractant, the volume ratio of the first extractant to the pretreated liquid is 1:(1.5~2.5); and / or, During the extraction process of the pretreated liquid with the first extractant, the extraction temperature is 20-30℃, and the extraction time for each stage is 5-10 min; and / or, During the extraction process of the pretreated liquid with the first extractant, the number of extraction stages is 3 to 5; and / or, During the pretreated liquid's extraction with the first extractant, countercurrent extraction is employed; and / or, The first stripping agent is 8 vol% to 12 vol% dilute sulfuric acid; and / or, The volume ratio of the first supported organic phase to the first stripping agent is (8~10):1; and / or, During the back-extraction of the first loaded organic phase with the first back-extractant, the back-extraction temperature is 20-30°C, and the back-extraction time for each stage is 15-25 min; and / or, During the back-extraction process of the first loaded organic phase using the first back-extractant, the number of back-extraction stages is 2 to 4; and / or, In the process of back-extracting the first loaded organic phase with the first back-extractant, countercurrent back-extraction is used.

9. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The two-stage extraction includes: The vanadium-rich liquid I is subjected to positive extraction with a second extractant to obtain a second loaded organic phase and a second raffinate. The second loaded organic phase was washed with a detergent to obtain a washed organic phase and a washing liquid; The washed organic phase is back-extracted by a second back-extracting agent to obtain a second organic-lean phase and a vanadium-rich solution II; wherein, The second extractant comprises: a mixture of 15 vol% to 30 vol% di(2-ethylhexyl) phosphate and 70 vol% to 85 vol% sulfonated kerosene, saponified by a saponifying agent, with a degree of saponification of 40% to 80%; wherein the saponifying agent is at least one of sodium hydroxide and sodium ethoxide; and / or, The volume ratio of the second extractant to vanadium-rich solution I is 1:(1~4); and / or, During the extraction process of the vanadium-rich solution I using the second extractant, the extraction temperature is 20-30℃, and the extraction time for each stage is 8-10 min; and / or, During the extraction process of the vanadium-rich solution I using the second extractant, the number of extraction stages is 4 to 6; and / or, During the extraction of the vanadium-rich solution I using the second extractant, countercurrent extraction is employed; and / or, The detergent is 5 vol% to 10 vol% dilute hydrochloric acid; and / or, During the washing process, the volume ratio of detergent to the second loaded organic phase is 1:(1~5); and / or, During the washing process, the washing temperature is 25~40℃, and the washing time for each stage is 5~20 min; and / or, During the washing process, the washing level is 1 to 3; and / or, During the washing process, countercurrent washing is employed; and / or, The second stripping agent is 10 vol%~15 vol% dilute sulfuric acid; and / or, During the back-extraction of the washed organic phase using the second back-extractant, the volume ratio of the second back-extractant to the washed organic phase is 1:(4~10); and / or, During the back-extraction of the washed organic phase using the second back-extractant, the back-extraction temperature is 20-30°C, and the back-extraction time for each stage is 10-40 min; and / or, During the back-extraction process of the washed organic phase by the second back-extracting agent, the number of back-extraction stages is 3 to 5; and / or, The washing organic phase is back-extracted using a second back-extractant in a countercurrent back-extraction process.

10. The method for preparing vanadium electrolyte from vanadium-bearing shale using a short-process method according to claim 1, characterized in that, The oil removal process is carried out using a stripping process.

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