Method for recovering silver and vanadium from silver associated vanadium shale mineral

Abstract

The invention discloses a method for recovering silver and vanadium from a silver associated vanadium shale mineral, which comprises the following steps: grinding the silver associated vanadium shale mineral, mixing with anhydrous sulfuric acid, aging, adding a leaching aid, mixing with water, stirring, leaching to obtain a leaching solution containing vanadium and silver, carrying out neutralization-reduction treatment on the leaching solution, introducing into an extraction process, enriching the extracted vanadium in a vanadium-enriched solution, and recovering the vanadium from the vanadium-enriched solution. And silver is left in the raffinate, the vanadium-rich liquid is introduced into a vanadium precipitation process after being oxidized, a vanadium pentoxide product is obtained after a precipitation product is dried and calcined, zinc powder is added into the raffinate, and pure silver is obtained after replacement. The method can synchronously leach and recover vanadium and silver, has the advantages of high vanadium and silver total recovery rate, high purity, short process flow, low acid consumption and environmental friendliness, and is suitable for industrial application.

Description

TECHNICAL FIELD

[0001] The present application belongs to the technical field of vanadium extraction from shale, and particularly relates to a method for recovering silver and vanadium from silver-associated vanadium shale minerals. BACKGROUND

[0002] It is generally considered that the CaO content in vanadium-containing raw materials is >5%, which is high-calcium vanadium-containing raw materials. The main minerals of such vanadium-containing shale include dolomite, calcite, quartz, illite and kaolin. Vanadium is mostly in the form of trivalent vanadium in the lattice of muscovite minerals, and the structure is stable, so vanadium is not easy to release. Therefore, it is difficult to directly leach vanadium from such vanadium-containing shale. At present, there are many studies on vanadium extraction from vanadium shale, but there are few studies on the simultaneous recovery of associated silver. In order to avoid waste of resources, valuable components in the ore need to be recovered. The traditional silver leaching method is cyanidation. However, cyanide is a highly toxic substance, which has irreversible harm to the environment and human health. Therefore, in recent years, the leaching of silver has been developing towards non-cyanidation.

[0003] Li Jie (Li Jie. Research on the comprehensive recovery of silver and vanadium by silver vanadium ore dressing and smelting [J]. Nonferrous Metals (Smelting Part), 2012, (12): 42-46.) provides a technical method of "decalcification by reverse flotation-high temperature and high acid leaching of vanadium-cyanide leaching of leaching residue" for recovering vanadium and silver from vanadium-silver ore. The leaching rate of vanadium can reach 81%, and the leaching rate of silver can reach 85%. However, this method leaches vanadium and silver under acidic and alkaline conditions respectively, which has the problems of large amount of lime and long process. In addition, cyanide is highly toxic and not environmentally friendly.

[0004] Chinese patent CN103526019A discloses a method for comprehensive recovery of vanadium, selenium and silver from multi-metal associated vanadium ore. The method uses sulfuric acid leaching of vanadium-reduction of selenium-silver leaching to recover multi-metals in vanadium ore. However, the method does not involve the specific recovery method of silver, and the leaching method is stepwise leaching, which has the disadvantage of long process and difficult control.

[0005] Vanadium leaching prefers strong acidic environment (pH <2), while cyanide leaching of silver requires alkaline environment (pH >10) to prevent volatilization. The thiosulfate leaching solution requires weak alkalinity. Therefore, the existing technology cannot satisfy the simultaneous leaching of vanadium and silver in the same leaching system. At the same time, strong oxidants are needed to change the valence state of vanadium, but silver is easy to form insoluble substances in the oxidation environment, while the reduction environment is not conducive to the dissolution of vanadium, so it is difficult to reconcile the conflict between vanadium and silver in the oxidation-reduction environment. In addition, in order to leach silver, Cl - is usually added in the existing technology, but Cl - may form volatile oxychloride (such as VO2Cl) with vanadium in a strong acid oxidation system, causing loss of vanadium. S 2- in sulfide ore +Ag2S is generated, which hinders the leaching of silver. Therefore, how to leach vanadium and silver from shale ore at the same time and solve the problems of environmental unfriendliness, high cost, long process, and low recovery rate of vanadium and silver has become one of the key factors restricting the high value utilization of shale ore. SUMMARY

[0006] In order to solve the above technical problems, the present application provides a method for recovering silver and vanadium from silver-associated vanadium shale ore, which has the advantages of short process and environmental friendliness, can realize the leaching and synchronous recovery of vanadium and silver at the same time, and can prepare vanadium pentoxide and pure silver products with high purity.

[0007] In order to achieve the above purpose, the present application provides a method for recovering silver and vanadium from silver-associated vanadium shale ore, which comprises the following steps: (1) mixing silver-associated vanadium shale ore with anhydrous sulfuric acid and aging for 12-48h to obtain a mixed slurry; (2) mixing the mixed slurry with a leaching aid and adding water to obtain a mixed ore slurry, and leaching to obtain a leaching solution and a leaching residue; (3) adding lime to the leaching solution, mixing and reacting to obtain a feed liquid, adding a reducing agent to the feed liquid, and reducing at 40-60℃ for 1-3h to obtain a crude extract; (4) extracting the crude extract to obtain a vanadium-rich liquid and a raffinate, mixing the vanadium-rich liquid with an oxidizing agent, adding a vanadium precipitating agent to the mixture until the pH is 1.5-2.0, stirring and reacting at 80-95℃ for 30-70min, and then solid-liquid separation and drying to obtain vanadium pentoxide; (5) mixing the raffinate with a silver reducing agent, standing and reacting for 4-6h, and then solid-liquid separation to obtain pure silver.

[0008] Preferably, the amount of anhydrous sulfuric acid is 40-60wt% of the dry weight of the silver-associated vanadium shale ore; and the particle size of the silver-associated vanadium shale ore is 50%-70% less than 0.074mm.

[0009] Preferably, water can be added after the silver-associated vanadium shale ore is mixed with pure sulfuric acid; and the amount of water is 0-15wt% of the dry weight of the silver-associated vanadium shale ore.

[0010] Further preferably, the grade of V2O5 in the silver-associated vanadium shale ore is 0.80%-1.10%, the grade of Ag is 87-110g / t, and the grade of CaO is 6.0%-12.0%.

[0011] Preferably, the ratio of the mixed slurry to the leaching aid in step (2) is 1:(0.20-0.45)L / kg; and the liquid-solid ratio when adding water is (1.20-1.50):1L / kg, wherein the liquid refers to water and the solid refers to the mixed slurry.

[0012] More preferably, the impregnation aid is composed of fluoride, chlorine oxidant and chloride salt in a mass ratio of 1:(8~16):(1~5).

[0013] More preferably, the chlorine oxidant is any one of sodium chlorate, perchloric acid, and sodium hypochlorite; the fluoride is any one of calcium fluoride, sodium fluoride, and sodium fluorosilicate; and the chloride salt is any one of sodium chloride, potassium chloride, and calcium chloride.

[0014] Preferably, the liquid-to-solid ratio of the leachate to lime in step (3) is 1: (0.02~0.03) L / kg; and the liquid-to-solid ratio of the feed liquid to the reducing agent is 1: (0.01~0.03) L / kg.

[0015] More preferably, the reducing agent is one of sodium thiosulfate, sodium sulfite, sodium nitrite, and iron powder.

[0016] Preferably, the extraction in step (4) is a 3-5 stage countercurrent forward extraction and a 2-4 stage countercurrent back extraction; the liquid-solid ratio of the vanadium-rich liquid to the oxidant is 1:(10-15)L / g; and the amount of the vanadium precipitation agent is 3.5-7kg / t (based on the dry ore).

[0017] More preferably, the vanadium precipitation agent has a mass concentration of 15-25 wt%. More preferably, the extraction time during countercurrent extraction is 5-10 min, and the volume ratio of organic phase to aqueous phase is 1:(1.5-2.5); the extraction time during countercurrent extraction is 15-25 min, and the volume ratio of organic phase to aqueous phase is (8-10):1; the oxidant is sodium chlorate or potassium permanganate; and the vanadium precipitant is any one of ammonium chloride, ammonium carbonate, and ammonia water.

[0018] Furthermore, the forward extraction agent is obtained by mixing di(2-ethylhexyl) phosphate (P204): 2-octanol: sulfonated kerosene in a volume ratio of (15~20): 5: (75~80); the reverse extraction agent is 8~12 vol% dilute sulfuric acid.

[0019] Preferably, the liquid-to-solid ratio of the raffinate to the silver reducing agent in step (5) is 1: (0.03~0.06) L / kg.

[0020] More preferably, the silver reducing agent is a metal reducing agent or a chemical reducing agent.

[0021] More preferably, the metal reducing agent is any one of zinc powder, copper powder, and iron powder; and the chemical reducing agent is any one of formaldehyde, hydrazine hydrate, and vitamin C.

[0022] The beneficial effects of this invention are as follows: Anhydrous sulfuric acid aging enhances the destruction of the vanadium mica structure before water leaching, allowing vanadium to break free from its lattice binding and transform into a free state, reducing acid consumption and increasing the vanadium leaching rate. Adding fluorides, chlorine oxidants, and chloride salts during water leaching increases the chloride ion concentration in the system, enabling a complexation reaction between silver ions and chloride ions. This allows silver ions to effectively exist in the solution in the form of AgCl x (x-1)- (x=2,3,4), ensuring simultaneous leaching of vanadium and silver. Simultaneously, it converts low-valence vanadium in the system into a high-valence leaching state, further improving the vanadium leaching rate. Compared with existing cyanide-based silver extraction processes, the aging-water leaching process eliminates the need to adjust the pH to alkaline, shortening the process flow. Furthermore, no highly toxic substances are released during the process, and the total recovery rate of vanadium and silver is above 80%. Therefore, this invention features high total vanadium and silver recovery rate, short process flow, low acid consumption, and environmental friendliness.

[0023] This invention can simultaneously leach vanadium and silver from silver-associated vanadium-bearing shale, and simultaneously recover vanadium and silver, and prepare vanadium pentoxide and pure silver products with high purity. This significantly improves the economic value of vanadium-bearing shale, reduces resource waste, and has the advantages of short recovery process, high efficiency and easy control, and can be applied on a large scale. Attached Figure Description

[0024] Figure 1 This is a schematic diagram of the process of the present invention. Detailed Implementation

[0025] The technical solution of the present invention will be further explained and described below with reference to the accompanying drawings and specific embodiments. It is worth noting that the following embodiments are only preferred embodiments of the present invention and should not be construed as limiting the present invention. The scope of protection of the present invention should be determined by the contents of the claims. Modifications and substitutions made by those skilled in the art to the technical solution of the present invention without creative effort all fall within the scope of protection of the present invention.

[0026] In the following embodiments, the silver-associated vanadium-bearing shale originated from Xingshan, Yichang, with a V2O5 grade of 0.91%, an Ag grade of 103 g / t, and a CaO grade of 11.3%.

[0027] Example 1 (1) Take silver-associated vanadium-bearing shale and ball mill it until the particle size is less than 0.074 mm, accounting for 70%, and use it as the raw ore; (2) The raw ore is mixed evenly with anhydrous sulfuric acid solution and aged for 14 hours to obtain a mixed slurry; wherein the amount of pure sulfuric acid solution used is 50 wt% of the dry mass of the raw ore. (3) The mixed slurry and the leaching aid are mixed at a liquid-solid ratio of 1:0.30 L / kg to obtain a mixed material. Then, water is added to the mixed material at a liquid-solid ratio of 1.50:1 L / kg (where liquid refers to water and solid refers to the mixed slurry) to adjust the slurry and obtain a mixed mineral slurry. The mixed mineral slurry is transported to a leaching tank and leached at 300 r / min and 95℃ for 20 h. The solid and liquid are separated to obtain leaching solution and leaching residue. The leaching aid is prepared by mixing calcium fluoride, sodium chlorate and sodium chloride in a mass ratio of 1:14:1. (4) Add lime to the leachate at a solid-liquid ratio of 1:0.02 L / kg, stir at 400 r / min for 6 h, and then separate the solid and liquid to obtain the feed solution; add sodium thiosulfate to the feed solution at a liquid-solid ratio of 1:0.02 L / kg, reduce at 60 °C for 2 h to obtain the extraction solution; (5) The extraction solution was extracted using a four-stage countercurrent forward extraction and a three-stage countercurrent back extraction process to obtain vanadium-rich solution and raffinate; the forward extraction time was 7 min, the volume ratio of organic phase to aqueous phase was 1:1.5, and the extractant was obtained by mixing di(2-ethylhexyl) phosphate: 2-octanol: sulfonated kerosene in a volume ratio of 20:5:75; the back extraction time was 20 min, the volume ratio of organic phase to aqueous phase was 9:1, and the extractant was a 10 vol% dilute sulfuric acid solution. (6) Mix the vanadium-rich liquid and sodium chlorate at a liquid-solid ratio of 1:15 L / g and stir for 1 hour to obtain the oxidized liquid. Then, gradually add a 20wt% ammonium carbonate solution to the oxidized liquid and stir at 200 r / min for 10 until the pH of the solution is 1.9. Then, stir the reaction at 80℃ and 200 r / min for 50 minutes and then separate the solid and liquid. Calcine the precipitate at 550℃ for 2 hours to obtain vanadium pentoxide product. (7) Mix the raffinate obtained in step (5) with zinc powder at a liquid-solid ratio of 1:0.04 L / kg, let stand for 5 hours, and then separate the solid and liquid. The precipitate is pure silver product.

[0028] Example 2 (1) Take silver-associated vanadium-bearing shale and ball mill it until the particle size is less than 0.074 mm, accounting for 65%, and use it as the raw ore; (2) The raw ore is mixed evenly with anhydrous sulfuric acid solution and aged for 20 hours to obtain a mixed slurry; wherein the amount of pure sulfuric acid solution used is 55 wt% of the dry mass of the raw ore; (3) The mixed slurry and the leaching aid are mixed at a liquid-solid ratio of 1:0.30 L / kg to obtain a mixed material. Water is then added to the mixed material at a liquid-solid ratio of 1.20:1 L / kg to adjust the slurry and obtain a mixed mineral slurry. The mixed mineral slurry is transported to a leaching tank and leached at 300 r / min and 90℃ for 15 h. The solid and liquid are separated to obtain leaching solution and leaching residue. The leaching aid is prepared by mixing calcium fluoride, sodium chlorate and sodium chloride in a mass ratio of 1:10:3. (4) Add lime to the leachate at a solid-liquid ratio of 1:0.0025 L / kg, stir and react at 400 r / min for 6 h, and then separate the solid and liquid to obtain the feed solution; add sodium thiosulfate to the feed solution at a liquid-solid ratio of 1:0.03 L / kg, reduce at 50 °C for 2.5 h to obtain the extraction solution; (5) The extraction solution was extracted using a four-stage countercurrent forward extraction and a three-stage countercurrent back extraction process to obtain vanadium-rich solution and raffinate; the forward extraction time was 7 min, the volume ratio of organic phase to aqueous phase was 1:2, and the extractant was obtained by mixing di(2-ethylhexyl) phosphate, 2-octanol and sulfonated kerosene in a volume ratio of 15:5:80; the back extraction time was 20 min, the volume ratio of organic phase to aqueous phase was 8:1, and the extractant was an 8 ol% dilute sulfuric acid solution. (6) Mix the vanadium-rich solution and potassium permanganate at a liquid-solid ratio of 1:12 L / g and stir for 1 h to obtain the oxidized solution. Then, gradually add a 20wt% ammonium chloride solution to the oxidized solution and stir at 200 r / min for 10 min until the pH of the solution is 1.9. Then, stir the reaction at 90℃ and 200 r / min for 30 min and then separate the solid and liquid. Calcine the precipitate at 550℃ for 2 h to obtain the vanadium pentoxide product. (7) Mix the raffinate obtained in step (5) with zinc powder at a liquid-solid ratio of 1:0.03 L / kg, let stand for 4 hours, and then separate the solid and liquid. The precipitate is the pure silver product.

[0029] Example 3 (1) Take silver-associated vanadium-bearing shale and ball mill it until the particle size is less than 0.074 mm, accounting for 70%, and use it as the raw ore; (2) The raw ore is mixed evenly with anhydrous sulfuric acid solution and then aged for 18 hours to obtain a mixed slurry; wherein the amount of pure sulfuric acid solution used is 60 wt% of the dry mass of the raw ore; (3) The mixed slurry and the leaching aid are mixed at a liquid-solid ratio of 1:0.45 L / kg to obtain a mixed material. Water is then added to the mixed material at a liquid-solid ratio of 1.5:1 L / kg to adjust the slurry and obtain a mixed mineral slurry. The mixed mineral slurry is transported to a leaching tank and leached at 300 r / min and 92℃ for 12 h. The solid and liquid are separated to obtain leaching solution and leaching residue. The leaching aid is prepared by mixing calcium fluoride, sodium chlorate and sodium chloride in a mass ratio of 1:6:4. (4) Add lime to the leachate at a solid-liquid ratio of 1:0.02 L / kg, stir at 400 r / min for 6 h, and then separate the solid and liquid to obtain the feed solution; add sodium sulfite to the feed solution at a liquid-solid ratio of 1:0.02 L / kg, reduce at 55℃ for 2 h to obtain the extraction solution; (5) The extraction solution was extracted using a four-stage countercurrent forward extraction and a three-stage countercurrent back extraction process to obtain vanadium-rich solution and raffinate; the forward extraction time was 7 min, the volume ratio of organic phase to aqueous phase was 1:1.8, and the extractant was obtained by mixing di(2-ethylhexyl) phosphate, 2-octanol and sulfonated kerosene in a volume ratio of 17:5:78; the back extraction time was 20 min, the volume ratio of organic phase to aqueous phase was 10:1, and the extractant was a 9 vol% dilute sulfuric acid solution. (6) Mix vanadium-rich liquid and sodium chlorate at a liquid-solid ratio of 1:12 L / g and stir for 1 hour to obtain an oxidized liquid. Then, gradually add 20wt% ammonia water to the oxidized liquid until the pH of the solution is 1.9. Then, stir the reaction at 85℃ and 200r / min for 70 minutes and then separate the solid and liquid. Calcine the precipitate at 550℃ for 2 hours to obtain vanadium pentoxide product. (7) Mix the raffinate obtained in step (5) with zinc powder at a liquid-solid ratio of 1:0.06 L / kg, let stand for 5 hours, and then separate the solid and liquid. The precipitate is the pure silver product.

[0030] Example 4 (1) Take silver-associated vanadium-bearing shale and ball mill it until the particle size is less than 0.074 mm, accounting for 50%, as the raw ore; (2) After mixing the raw ore with anhydrous sulfuric acid solution evenly, water is added, and the mixture is aged for 36 hours to obtain a mixed slurry; wherein the amount of pure sulfuric acid solution used is 40 wt% of the dry ore mass, and the amount of water used is 5 wt% of the dry ore mass; (3) The mixed slurry and the leaching aid are mixed at a liquid-solid ratio of 1:0.20 L / kg to obtain a mixed material. Water is then added to the mixed material at a liquid-solid ratio of 1.50:1 L / kg to adjust the slurry and obtain a mixed mineral slurry. The mixed mineral slurry is transported to a leaching tank and leached at 300 r / min and 95℃ for 20 h. The solid and liquid are separated to obtain leaching solution and leaching residue. The leaching aid is prepared by mixing sodium fluoride, perchloric acid and potassium chloride in a mass ratio of 1:8:2. (4) Add lime to the leachate at a solid-liquid ratio of 1:0.02 L / kg, stir and react at 400 r / min for 6 h, and then separate the solid and liquid to obtain the feed solution; add sodium nitrite to the feed solution at a liquid-solid ratio of 1:0.01 L / kg, reduce at 60℃ for 2 h to obtain the extraction solution; (5) The extraction solution was extracted using a four-stage countercurrent forward extraction and a three-stage countercurrent back extraction process to obtain vanadium-rich solution and raffinate; the forward extraction time was 7 min, the volume ratio of organic phase to aqueous phase was 1:2.5, and the extractant was a mixture of di(2-ethylhexyl) phosphate, 2-octanol and sulfonated kerosene in a volume ratio of 20:5:75; the back extraction time was 20 min, the volume ratio of organic phase to aqueous phase was 9:1, and the extractant was a 11 vol% dilute sulfuric acid solution. (6) Mix vanadium-rich liquid and sodium chlorate at a liquid-solid ratio of 1:15 L / g to obtain oxidized liquid. Then, gradually add 20wt% ammonium carbonate solution to the oxidized liquid until the pH of the solution is 1.9. Then, stir the reaction at 80℃ and 200r / min for 50min and then separate the solid and liquid. Calcine the precipitate at 550℃ for 2h to obtain vanadium pentoxide product. (7) Mix the raffinate obtained in step (5) with zinc powder at a liquid-solid ratio of 1:0.04 L / kg, let stand for 5 hours, and then separate the solid and liquid. The precipitate is pure silver product.

[0031] Example 5 (1) Take silver-associated vanadium-bearing shale and ball mill it until the particle size is less than 0.074 mm, accounting for 55%, and use it as the raw ore; (2) After mixing the raw ore with anhydrous sulfuric acid solution evenly, water is added and the mixture is aged for 48 hours to obtain a mixed slurry; wherein the amount of pure sulfuric acid solution used is 45 wt% of the dry mass of the raw ore, and the amount of water used is 15 wt% of the dry mass of the raw ore; (3) The mixed slurry and the leaching aid are mixed at a liquid-solid ratio of 1:0.40 L / kg to obtain a mixed material. Water is then added to the mixed material at a liquid-solid ratio of 1.50:1 L / kg to adjust the slurry and obtain a mixed mineral slurry. The mixed mineral slurry is transported to a leaching tank and leached at 300 r / min and 95℃ for 20 h. The solid and liquid are separated to obtain leaching solution and leaching residue. The leaching aid is prepared by mixing sodium fluorosilicate, sodium hypochlorite and calcium chloride in a mass ratio of 1:16:5. (4) Add lime to the leachate at a solid-liquid ratio of 1:0.02 L / kg, stir and react at 400 r / min for 6 h, and then separate the solid and liquid to obtain the feed solution; add iron powder to the feed solution at a liquid-solid ratio of 1:0.03 L / kg, reduce at 60℃ for 2 h to obtain the extraction solution; (5) The extraction solution was extracted using a four-stage countercurrent forward extraction and a three-stage countercurrent back extraction process to obtain vanadium-rich solution and raffinate; the forward extraction time was 7 min, the volume ratio of organic phase to aqueous phase was 1:1.5, and the extractant was a mixture of di(2-ethylhexyl) phosphate, 2-octanol and sulfonated kerosene in a volume ratio of 20:5:75; the back extraction time was 20 min, the volume ratio of organic phase to aqueous phase was 9:1, and the extractant was a 12 vol% dilute sulfuric acid solution. (6) Mix the vanadium-rich liquid and sodium chlorate at a liquid-solid ratio of 1:10 L / g and stir for 1 hour to obtain the oxidized liquid. Then, gradually add a 20wt% ammonium carbonate solution to the oxidized liquid until the pH of the solution is 1.9. Then, stir the reaction at 80℃ and 200r / min for 50 minutes and then separate the solid and liquid. Calcine the precipitate at 550℃ for 2 hours to obtain vanadium pentoxide product. (7) Mix the raffinate obtained in step (5) with zinc powder at a liquid-solid ratio of 1:0.04 L / kg, let stand for 5 hours, and then separate the solid and liquid. The precipitate is pure silver product.

[0032] Comparative Example 1 The methods and steps are the same as in Example 1, except that the anhydrous sulfuric acid solution is replaced with a sulfuric acid solution with a water content of 10%, and the leaching and recovery of vanadium and silver are carried out under the condition of a liquid-solid ratio of 2:1 (mL / g).

[0033] Comparative Example 2 The method and steps are the same as in Example 1, except that the leaching aid is replaced with a mixture of sodium chlorate and calcium fluoride in a mass ratio of 1:14, for the leaching and recovery of vanadium and silver.

[0034] Comparative Example 3 The method and steps are the same as in Example 1, except that the leaching aid is replaced with a mixture of calcium fluoride and sodium chloride in a mass ratio of 1:1, which is used for the leaching and recovery of vanadium and silver.

[0035] Comparative Example 4 The method and steps are the same as in Example 1, except that the leaching aid is replaced with a mixture of sodium chlorate and sodium chloride at a mass ratio of 14:1, which is used for the leaching and recovery of vanadium and silver.

[0036] Comparative Example 5 The method and steps are the same as in Example 1, except that the mass ratio of calcium fluoride:sodium chlorate:sodium chloride in the leaching aid is changed from 1:14:1 to 1:14:8 for the leaching and recovery of vanadium and silver.

[0037] Comparative Example 6 The method and steps are the same as in Example 1, except that the mass ratio of calcium fluoride:sodium chlorate:sodium chloride in the leaching aid is changed from 1:14:1 to 1:20:1 for the leaching and recovery of vanadium and silver.

[0038] Comparative Example 7 The method and steps are the same as in Example 1, except that the mass ratio of calcium fluoride:sodium chlorate:sodium chloride in the leaching aid is changed from 1:14:1 to 3:14:1 for vanadium and silver leaching and recovery.

[0039] Comparative Example 8 The method and steps are the same as in Example 1, except that step (2) is changed to mixing the raw ore with a sulfuric acid solution with a mass concentration of 50wt% and reacting for 12 hours to obtain a mixed slurry.

[0040] Comparative Example 9 The method and steps are the same as in Example 1, except that in step (3), no leaching aid is used. The mixed slurry is directly mixed with water to obtain a mixed mineral slurry, which is then leached.

[0041] Example 6 The leaching rates of vanadium pentoxide and silver in the above embodiments and comparative examples were calculated as follows: leaching rate = ((vanadium and silver concentration in liquid × solution volume)) / (vanadium and silver grade) × raw ore mass), where liquid refers to the leachate obtained in step (3). The results are shown in Table 1. The vanadium pentoxide and pure silver products prepared in the above examples and comparative examples were tested for purity using ICP, and the total recovery rate was calculated using ((product quality) × (vanadium pentoxide and silver purity)) / ((raw ore quality) × (vanadium pentoxide and silver grade)). The results are shown in Table 1.

[0042] Table 1 Leaching and Recovery Results

[0043] The results showed that, compared with Comparative Examples 2-4, the examples all achieved significantly higher vanadium and silver leaching rates. This indicates that the addition of calcium fluoride can enhance the destruction of the mica structure, exposing the vanadium in the mica. The addition of sodium chlorate can increase the potential of the leaching system, change the valence state of vanadium ions, and promote the dissolution of silver. The addition of sodium chloride increases the chloride ion concentration in the leaching system, ensuring that silver exists stably in the solution in ionic form, thus achieving simultaneous leaching of vanadium and silver and significantly improving the vanadium-silver leaching rate.

Claims

1. A method for recovering silver and vanadium from silver-associated vanadium shale minerals, characterized in that: Includes the following steps: (1) A mixed slurry was obtained by mixing and aging silver-associated vanadium shale ore with anhydrous sulfuric acid; (2) After the mixed slurry and the leaching aid are mixed evenly, water is added to adjust the slurry to obtain a mixed slurry. Leaching yields leachate and leaching residue. (3) Add lime to the leachate, mix and react, then separate the solid and liquid to obtain the feed solution. Add reducing agent to the feed solution and reduce for 1-3 hours to obtain the original extract solution. (4) Extraction of the original extract yields vanadium-rich liquid and raffinate. After the vanadium-rich liquid is mixed with the oxidant and reacted, a vanadium precipitant is added. After the reaction, solid-liquid separation, drying and calcination are performed to obtain vanadium pentoxide. (5) Pure silver is obtained by mixing the raffinate with the silver reducing agent and then separating the solid and liquid.

2. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 1, characterized in that: The amount of anhydrous sulfuric acid used is 40-60 wt% of the dry weight of the silver-associated vanadium shale ore; the silver-associated vanadium shale ore contains 50%-70% particles with a diameter less than 0.074 mm.

3. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 1, characterized in that: The ratio of the mixed slurry to the impregnation aid in step (2) is 1:(0.20~0.45)L / kg; the liquid-solid ratio when adding water is (1.20~1.50):1L / kg.

4. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 3, characterized in that: The immersion aid is composed of fluoride, chlorine oxidant and chloride salt in a mass ratio of 1:(8~16):(1~5).

5. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 4, characterized in that: The chlorine oxidant is any one of sodium chlorate, perchloric acid, and sodium hypochlorite; the fluoride is any one of calcium fluoride, sodium fluoride, and sodium fluorosilicate; and the chloride salt is any one of sodium chloride, potassium chloride, and calcium chloride.

6. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 1, characterized in that: The liquid-to-solid ratio of the leachate to lime in step (3) is 1: (0.02~0.03) L / kg; the liquid-to-solid ratio of the feed liquid to the reducing agent is 1: (0.01~0.03) L / kg.

7. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 6, characterized in that: The reducing agent is one of sodium thiosulfate, sodium sulfite, sodium nitrite, and iron powder.

8. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 1, characterized in that: The extraction in step (4) is a 3-5 stage countercurrent forward extraction and a 2-4 stage countercurrent back extraction; the liquid-solid ratio of the vanadium-rich liquid to the oxidant is 1:(10~15)L / g; the amount of the vanadium precipitation agent is 3.5~7kg / t.

9. A method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 8, characterized in that: The extraction time for countercurrent forward extraction is 5-10 min, and the volume ratio of organic phase to aqueous phase is 1:(1.5-2.5); the extraction time for countercurrent back extraction is 15-25 min, and the volume ratio of organic phase to aqueous phase is (8-10):1; the oxidant is any one of sodium chlorate, potassium permanganate, and chlorine; the vanadium precipitant is any one of ammonium chloride, ammonium carbonate, and ammonia water.

10. The method for recovering silver and vanadium from silver-associated vanadium shale minerals according to claim 1, characterized in that: The liquid-to-solid ratio of the raffinate to the silver reducing agent in step (5) is 1: (0.03~0.06) L / kg.

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