A method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts

Abstract

This application relates to the technical field of waste catalyst recovery, and specifically discloses a method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts. The method includes the following steps: S1: Heating the vanadium-molybdenum-containing waste catalyst for roasting, continuously heating for roasting, and then cooling to obtain a roasted material; S2: Mixing the roasted material with sulfuric acid solution A containing oxalic acid, heating for leaching to obtain leaching solution A and residue A; Mixing residue A with sulfuric acid solution B containing thiourea and ethylenediaminetetraacetic acid, heating for secondary leaching to obtain leaching solution B and residue B, washing residue B with acid solution to obtain a washing solution; Combining the leaching solutions and the washing solution to obtain a combined solution; S3: Adjusting the pH of the combined solution to 2.5 - 3, then adjusting the pH to 4.5 - 5, adding dimethylglyoxime alcohol solution, heating to obtain a purified solution; Mixing the purified solution with an extraction solution evenly to obtain a vanadium-molybdenum-containing organic phase; S4: Back-extraction and product preparation. The ammonium molybdate and vanadium pentoxide obtained in this application have high purity, and the recovery rates of vanadium and molybdenum are relatively high.

Description

Technical Field

[0001] The present application relates to the technical field of waste catalyst recycling, and more specifically, to a method for recycling valuable metals from vanadium-molybdenum-containing waste catalysts. Background Art

[0002] With the gradual reduction of high-quality molybdenum and vanadium ore resources, the recovery of vanadium and molybdenum from secondary resources (such as waste catalysts) has become a key path to ensure resource security. Among them, vanadium-molybdenum-containing waste catalysts, as high-value waste in the petrochemical and metallurgical industries, have a metal content far exceeding that of primary ores and have priority for recovery.

[0003] The main methods for extracting vanadium and molybdenum products from waste vanadium-molybdenum-containing catalysts include acid leaching method, ammonia leaching method and alkali leaching method. The recovery of excessive ammonia water in the ammonia leaching method requires an ammonia distillation and ammonia absorption system, with a long process flow and a low leaching temperature, thus resulting in a long leaching time. The alkali leaching method is further divided into sodium carbonate method and sodium hydroxide method. Sodium hydroxide has strong alkalinity and a fast leaching rate, but a large amount of aluminum, silicon, etc. in the raw materials are leached out, making the solution purification difficult. The acid leaching method has become one of the core methods with the most industrialization potential due to its high leaching rate and applicability to various complex component catalysts.

[0004] The patent application document with the publication number CN103290223A discloses a method for comprehensive recovery of multiple metals from waste catalysts. When the vanadium content in the waste catalyst is above 0.5%, it includes the following steps: (1) Dilute sulfuric acid pre-leaching: Place the waste catalyst in a stirred mill, add dilute sulfuric acid, grind and pre-leach; (2) Ripening water leaching: Ripen the pre-leaching residue in step (1) with concentrated sulfuric acid, add fluoride as an additive during the ripening process to promote the dissociation between elements, and after ripening, perform water leaching to selectively extract vanadium to obtain vanadium pentoxide product; (3) Pyrometallurgical smelting: Perform pyrometallurgical smelting on the vanadium extraction residue in step (2) to recover nickel, cobalt, and molybdenum. Nickel, cobalt, and molybdenum enter the matte, and impurities such as aluminum and silicon enter the slag; (4) Pressure leaching: Subject the matte phase obtained in step (3) to pressure acid leaching to recover nickel and cobalt to obtain nickel sulfate and cobalt sulfate products; (5) Ammonia leaching to extract molybdenum: Use ammonia leaching to recover molybdenum from the leaching residue obtained in step (4) to obtain ammonium molybdate product.

[0005] In this technology, when the pre-leaching residue is ripened with concentrated sulfuric acid, although the addition of fluoride can promote the dissociation of vanadium from impurities and inhibit the leaching of aluminum, the strong acidity of concentrated sulfuric acid will still cause a large amount of valuable metals such as nickel, cobalt, and molybdenum to be non-selectively dissolved and enter the vanadium extraction solution, which will not only affect the purity of the vanadium pentoxide product but also cause a decrease in the content of target metals in the subsequent pyrometallurgical smelting raw material (vanadium extraction residue), indirectly leading to a decrease in the recovery rates of nickel, cobalt, and molybdenum. Summary of the Invention

[0006] In order to improve the yield and purity of vanadium and molybdenum products, the present application provides a method for recycling valuable metals from vanadium-molybdenum-containing waste catalysts.

[0007] This application provides a method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts, adopting the following technical solution:

[0008] A method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts includes the following steps:

[0009] S1: Low-temperature roasting: Roast the vanadium-molybdenum-containing waste catalyst at 180 - 200 °C for 2 - 4 h, then continue to raise the temperature to 450 - 500 °C and roast for 120 - 180 min, and then cool to obtain the roasted material;

[0010] S2: Gradient acid leaching: Mix the roasted material with sulfuric acid solution A containing oxalic acid, and leach at 50 - 60 °C for 90 - 120 min. After solid-liquid separation, obtain leaching solution A and residue A; Mix residue A with sulfuric acid solution B containing thiourea and ethylenediaminetetraacetic acid, and perform secondary leaching at 70 - 80 °C for 90 - 120 min. After solid-liquid separation, obtain leaching solution B and residue B. Wash residue B with acid solution 2 - 3 times to obtain the washing solution; Combine leaching solution A, leaching solution B, and the washing solution to obtain the combined solution;

[0011] S3: Purification and extraction: Adjust the pH of the combined solution to 2.5 - 3, perform solid-liquid separation, then adjust the pH to 4.5 - 5, perform solid-liquid separation again, add dimethylglyoxime alcohol solution, raise the temperature to 35 - 45 °C, keep warm for 20 - 30 min, and perform solid-liquid separation to obtain the purified solution; Mix the purified solution and the extraction solution evenly, let it stand for layering to obtain the vanadium-molybdenum-containing organic phase;

[0012] S4: Back extraction and product preparation: Mix the vanadium-molybdenum-containing organic phase with a weak acid-weak base salt back extraction agent evenly, let it stand for layering to obtain back extraction solution A and vanadium-containing organic phase. Back extraction solution A is alkalized, evaporated, and crystallized to obtain ammonium molybdate; Mix the vanadium-containing organic phase with a strong base back extraction agent evenly, let it stand for layering to obtain back extraction solution B. Back extraction solution B is acidified, ammonium salt is precipitated, then solid-liquid separation is performed, and calcination is carried out to obtain vanadium pentoxide.

[0013] In this technology, two-stage low-temperature roasting is adopted. At 180 - 200 °C, the organic matter, moisture, and volatile impurities adsorbed on the surface of the waste catalyst can be effectively removed, avoiding the pollution of the solution by organic matter during subsequent acid leaching. Subsequently, the temperature is raised to 450 - 500 °C for roasting, which can not only optimize the crystal structure of metal oxides such as vanadium and molybdenum, making them more porous and increasing surface defects, thus significantly increasing the contact area with subsequent acid leaching reagents and enhancing the reaction activity; but also reduce the volatilization loss of vanadium and improve the recovery rate of subsequent vanadium products.

[0014] Afterwards, gradient leaching was carried out using sulfuric acid solution A containing oxalic acid and sulfuric acid solution B containing thiourea. In addition to improving the leaching rates of vanadium and molybdenum and inhibiting the dissolution of impurities, the complexing effect of oxalic acid enabled the preferential dissolution of vanadium and molybdenum in the readily soluble state. The reducing effect of thiourea promoted the reduction of vanadium in the insoluble state to a lower valence state that was more readily soluble and disrupted the binding between molybdenum and the carrier, enabling the leaching of insoluble molybdenum. The synergistic effect of the two further improved the efficiency and selectivity of acid leaching, providing a better basis for subsequent purification and extraction steps. Ethylenediaminetetraacetic acid reduced the decomposition of thiourea by chelating metal ions (such as Fe 3+ , Al 3+ ). Subsequently, stepwise neutralization was carried out to avoid the aggregation of iron / aluminum ions to form large precipitates due to rapid precipitation, which would cause vanadium and molybdenum ions to be co-precipitated and affect the recovery rate of target metals. Dimethylglyoxime specifically chelated and removed nickel, avoiding the competition of nickel ions with vanadium and molybdenum for extraction sites during subsequent extraction, thereby improving the selectivity and efficiency of extraction and achieving the efficient separation of impurities. Finally, through processes such as stripping, crystallization, and calcination, high-purity ammonium molybdate and vanadium pentoxide products could be obtained.

[0015] Preferably, in the sulfuric acid solution A, the concentration of sulfuric acid is 30 g / L to 60 g / L, and the concentration of oxalic acid is 45 to 90 g / L.

[0016] Preferably, in the sulfuric acid solution B, the concentration of sulfuric acid is 150 g / L to 180 g / L, the concentration of thiourea is 80 to 120 g / L, and the concentration of ethylenediaminetetraacetic acid is 2 g / L to 5 g / L.

[0017] Preferably, the mass-to-volume ratio of the vanadium- and molybdenum-containing waste catalyst to the sulfuric acid solution A is 1 g:(3 - 5) mL.

[0018] Preferably, the mass-to-volume ratio of the vanadium- and molybdenum-containing waste catalyst to the sulfuric acid solution B is 1 g:(3 - 5) mL.

[0019] In this technology, first, leaching is carried out with sulfuric acid solution A containing oxalic acid at 50 - 60 °C. Oxalic acid can preferentially form stable complexes with the soluble vanadium ions and molybdenum ions in the catalyst, thereby promoting the preferential dissolution of these soluble vanadium and molybdenum. This selective dissolution method avoids a large amount of co - dissolution with impurities such as aluminum and iron in the initial stage of leaching, reducing the pressure in the subsequent purification process. For residue A, secondary leaching is carried out with sulfuric acid solution B containing thiourea at 70 - 80 °C. Thiourea will form a reducing environment under acidic conditions. In this environment, high - valence vanadium will be reduced to a lower valence state. Vanadium in the lower valence state has better solubility in the solution and is more easily leached. At the same time, thiourea can also break the tight bond between molybdenum and the carrier (such as alumina, silica), enabling the originally insoluble molybdenum to dissociate from the carrier, thus achieving the deep dissolution of insoluble vanadium and molybdenum. Through this "easy - first - difficult - later, step - by - step dissolution" gradient leaching, the efficient activation and selective dissolution of vanadium and molybdenum are achieved.

[0020] Preferably, the dimethylglyoxime alcohol solution includes dimethylglyoxime and ethanol, and the concentration of dimethylglyoxime is 60 - 100 g / L.

[0021] Preferably, the mass - to - volume ratio of the vanadium - and - molybdenum - containing waste catalyst to the dimethylglyoxime alcohol solution is 1 g:(1.5 - 2.5) mL.

[0022] Preferably, in step S2, in the acid - solution washing of residue B, the acid solution used is sulfuric acid solution C, and in sulfuric acid solution C, the concentration of sulfuric acid is 50 g / L - 100 g / L.

[0023] Preferably, the extractant includes 18 g / L - 54 g / L methyltrioctylammonium chloride, 50 g / L - 80 g / L nonylphenol polyoxyethylene ether, and an isoparaffin diluent.

[0024] Preferably, the volume ratio of the purification liquid to the extractant is 1:(1 - 1.5).

[0025] In this technology, methyltrioctylammonium chloride has a high - selectivity coordination ability for molybdate and vanadate. It can preferentially combine with molybdate and vanadate, transferring them from the aqueous phase to the organic phase. Through this selective extraction, the efficient separation of vanadium and molybdenum from impurity ions is achieved.

[0026] Preferably, the weak - acid - weak - base - salt stripping agent is an NH4HCO3 solution, the concentration of the NH4HCO3 solution is 80 g / L - 120 g / L, and the volume ratio of the weak - acid - weak - base - salt stripping agent to the vanadium - and - molybdenum - containing organic phase is 1:(1.5 - 2).

[0027] In this technology, bicarbonate ions have a strong affinity for molybdate ions. Under weakly alkaline conditions, bicarbonate ions can react with molybdate ions, causing the molybdate ions to transfer from the organic phase to the aqueous phase, while vanadium remains in the organic phase, thus achieving efficient separation of vanadium and molybdenum.

[0028] Preferably, the strong base stripping agent includes sodium hydroxide and hydrogen peroxide. The concentration of sodium hydroxide is 120 g / L to 200 g / L, and the concentration of hydrogen peroxide is 6.8 g / L to 10.2 g / L.

[0029] Preferably, in step S4, the specific steps for alkalizing, evaporating, and crystallizing the stripping liquid A are as follows:

[0030] Adjust the pH of the stripping liquid A to 8.0 to 8.5, and then evaporate and crystallize.

[0031] Preferably, in step S4, the specific steps for acidifying, precipitating, and calcining the stripping liquid B are as follows:

[0032] After adjusting the pH of the stripping liquid B to 2.0 to 2.5, precipitate with an ammonium salt, separate the solid and liquid, then transfer it to a sintering furnace, heat up to 580 to 620 °C, calcine for 80 to 100 min, and cool to obtain vanadium pentoxide.

[0033] Preferably, the ammonium salt is ammonium sulfate, and the mass ratio of ammonium sulfate to the vanadium-molybdenum-containing waste catalyst is (0.24 to 0.30):1.

[0034] In this technology, under strongly alkaline conditions, vanadate ions combine with sodium ions and can transfer from the organic phase to the aqueous phase. At the same time, the added H2O2 can oxidize low-valent vanadium, ensuring the stable existence of vanadium in the high-valent state and improving the stripping efficiency of vanadium. This process avoids interference with separation due to vanadium-molybdenum complexation through the method of "stepwise stripping with weak base - strong base".

[0035] Preferably, in the sulfuric acid solution B, ascorbic acid is also included, and the concentration of ascorbic acid is 0.5 to 1.0 g / L.

[0036] In this technology, ascorbic acid, through its synergistic effect with thiourea, plays the "synergistic mechanism of thiourea reducing high-valent vanadium - ascorbic acid stabilizing low-valent state": Thiourea can reduce high-valent vanadium in the solution to low-valent vanadium, realizing the activation and release of insoluble vanadium, and ascorbic acid, as a mild reducing agent, can maintain the stability of low-valent vanadium in the solution after thiourea reduces high-valent vanadium, preventing it from being re-oxidized. At the same time, ascorbic acid can also reduce Fe 3+ to Fe 2+ , inhibiting Fe 3+Hydrolysis generates iron hydroxide precipitate, preventing it from adsorbing and encapsulating vanadium ions, thereby further ensuring the stability of low-valence vanadium in the solution and realizing the synergistic effect of the two in reducing and stabilizing low-valence vanadium.

[0037] Preferably, in step S3, before adding the dimethylglyoxime alcohol solution, first pass through a resin column filled with fluoride-ion-loaded chelating resin at a flow rate of 2-4 BV / h.

[0038] Preferably, the preparation method of the fluoride-ion-loaded chelating resin includes the following steps:

[0039] Immerse the chelating resin in a sodium chloride solution, soak for 22-26 h, then separate the solid and liquid. After washing, immerse it in a sodium hydroxide solution, soak for 3-5 h, separate the solid and liquid, wash, and then load it into an ion exchange column. Pass a sodium fluoride solution through it at a flow rate of 5-10 BV / h until the ratio of the fluoride ion concentration in the effluent to the fluoride ion concentration in the sodium fluoride solution is higher than 95%, and then it is obtained.

[0040] Preferably, the chelating resin is an amino phosphonic acid-based chelating resin.

[0041] In the present technology, an amino phosphonic acid-based chelating resin is used, which is rich in amino and phosphonic acid-based bidentate functional groups on its surface, forming a synergistic impurity removal mechanism of "fluoride competition complexation + resin specific adsorption" with fluoride ions: fluoride ions preferentially complex with Al 3+ / Fe 3+ to form stable low-coordination anions, significantly reducing the concentration of free metal ions, weakening their coordination ability with the extractant, and avoiding the formation of an emulsion layer or consumption of the extractant. The remaining free Al 3+ / Fe 3+ then forms a five-membered ring chelate structure with metal ions through the lone pair electrons of the nitrogen atom and the oxygen atom of the phosphonic acid group of the resin functional group, and realizes deep adsorption by using the high mass transfer efficiency of the macroporous structure of the resin.

[0042] At the same time, the resin adsorption process is not affected by high concentrations of sulfate and oxalate ions in the solution, providing a key guarantee for the efficient stripping of the vanadium- and molybdenum-containing organic phase and the preparation of ammonium molybdate and vanadium pentoxide products with high purity, and realizing the dual optimization of deep impurity removal and high-selectivity separation of target metals.

[0043] Preferably, the concentration of the sodium fluoride solution is 0.5-1.0 mol / L.

[0044] Preferably, the mass fraction of the sodium chloride solution is 4%-6%.

[0045] Preferably, the mass fraction of the sodium hydroxide solution is 8%-10%.

[0046] Preferably, the amino phosphonic acid-based chelating resin is D418 macroporous amino phosphonic acid-based chelating resin.

[0047] In summary, the present application has the following beneficial effects:

[0048] 1. The present application adopts low-temperature roasting activation. While avoiding the densification of the carrier structure, it converts crystalline vanadium molybdenum oxide into an active state that is easily soluble in acid. Gradient acid leaching designs acid solutions with different components step by step, preferentially dissolves soluble vanadium and molybdenum by using the complexing effect of oxalic acid, and uses the reducing property of thiourea to break the bond between insoluble metals and the carrier, realizing the efficient leaching of vanadium and molybdenum and inhibiting the dissolution of impurities; the step-by-step purification process removes aluminum and iron ions in two stages by two-stage neutralization, and combines the selective complexation of nickel by dimethylglyoxime to reduce the interference of impurities on extraction; selective extraction and stripping reverse-extract molybdenum by ammonium bicarbonate and separate and reverse-extract vanadium by strong base sodium hydroxide, realizing the efficient separation and recovery of vanadium and molybdenum by quality.

[0049] 2. In the present application, ascorbic acid is preferably added to sulfuric acid solution B. Ascorbic acid serves as a mild reducing agent to maintain the stability of low-valent vanadium and molybdenum in the solution and prevent their secondary oxidation. It forms a "thiourea-reducing high-valent vanadium - ascorbic acid stabilizing low-valent state synergistic mechanism" with thiourea through the reduction potential difference, significantly improving the leaching efficiency of vanadium and molybdenum.

[0050] 3. After step-by-step neutralization, it is preferably passed through a resin column of chelating resin loaded with fluoride ions to deeply remove aluminum and iron impurities through the "fluoride competition complexation + resin specific adsorption synergistic effect". Specific Embodiments

[0051] The following further elaborates on the present application with reference to examples.

[0052] The raw materials of the examples and comparative examples of the present application are all ordinary commercially available products unless otherwise specified.

[0053] Table 1 XRF detection data of vanadium- and molybdenum-containing waste catalyst before roasting

[0054]

[0055] Example 1

[0056] This example provides a method for recovering valuable metals from vanadium- and molybdenum-containing waste catalysts, including the following steps:

[0057] S1: Low-temperature roasting: Place 1000 g of vanadium- and molybdenum-containing waste catalyst in a box-type resistance furnace, heat it to 180 °C at a rate of 5 °C / min, keep it warm for 4 h, then continue to heat it to 450 °C at a rate of 10 °C / min, keep it warm for 180 min, cool it, and grind it until it completely passes through a 200-mesh sieve to obtain roasted material;

[0058] S2: Gradient acid leaching: Stir and mix the calcined material evenly with 5 L of sulfuric acid solution A, carry out stirring leaching in a water bath at 50 °C for 120 min, after pressure filtration, obtain leaching solution A and residue A; Stir and mix residue A evenly with 5 L of sulfuric acid solution B, carry out stirring leaching in a water bath at 70 °C for 120 min, after pressure filtration, obtain leaching solution B and residue B; Rinse residue B with a sulfuric acid solution with a concentration of 50 g / L, wash 3 times, with a dosage of 200 mL each time, to obtain a water washing solution; Combine leaching solution A, leaching solution B and the water washing solution to obtain a combined solution;

[0059] S3: Purification and extraction: Slowly add ammonia water with a concentration of 50 g / L to the combined solution, adjust the pH to about 3, stir for 20 min, then let it stand for 20 min, after filtration, slowly add ammonia water with a concentration of 50 g / L again, adjust the pH to about 5, stir for 20 min, then let it stand for 20 min, after filtration, slowly add 2.5 L of a dimethylglyoxime ethanol solution with a concentration of 60 g / L, and raise the temperature to 35 °C, keep warm for 20 min, carry out centrifugal separation to obtain a purified solution; Stir and mix the purified solution and the extraction solution in a volume ratio of 1:1.5 for 10 min, let it stand for layering to obtain a vanadium and molybdenum-containing organic phase;

[0060] S4: Stir and mix the vanadium and molybdenum-containing organic phase with an NH4HCO3 solution with a concentration of 80 g / L in a volume ratio of 1:2 for 10 min, let it stand for layering to obtain stripping solution A and a vanadium-containing organic phase; Slowly add ammonia water with a concentration of 50 g / L to stripping solution A, adjust the pH to 8.0, evaporate and concentrate to one-third of the original volume at 80 °C, then carry out ice bath crystallization for 20 h, after centrifugal separation, transfer it to a drying oven at 105 °C and dry to constant weight to obtain ammonium molybdate;

[0061] Stir and mix the vanadium-containing organic phase with a sodium hydroxide solution with a concentration of 120 g / L in a volume ratio of 1:1.5 for 10 min, let it stand for layering to obtain stripping solution B; Slowly add a sulfuric acid solution with a concentration of 50 g / L and 240 g of ammonium sulfate to stripping solution B, adjust the pH to 2.5, raise the temperature to 50 °C, stir and mix for 30 min, let it stand for precipitation for 4 h, after centrifugal separation, wash 3 times with an ammonium sulfate solution with a mass fraction of 2% to obtain vanadium precipitates;

[0062] Transfer the vanadium precipitates into a sintering furnace, heat up to 580 °C at a rate of 10 °C / min, calcine for 100 min, cool with the furnace and then grind to obtain vanadium pentoxide.

[0063] Among them, in step S2, sulfuric acid solution A includes sulfuric acid and oxalic acid, the concentration of sulfuric acid is 30 g / L, and the concentration of oxalic acid is 45 g / L;

[0064] Sulfuric acid solution B includes sulfuric acid, thiourea and ethylenediaminetetraacetic acid, the concentration of sulfuric acid is 150 g / L, the concentration of thiourea is 80 g / L, and the concentration of ethylenediaminetetraacetic acid is 2 g / L.

[0065] In step S3, the extraction solution includes methyltrioctylammonium chloride, nonylphenol polyoxyethylene ether, and Isopar M isoparaffin. The concentration of methyltrioctylammonium chloride is 18 g / L, and the concentration of nonylphenol polyoxyethylene ether is 50 g / L.

[0066] In step S4, the sodium hydroxide solution further includes hydrogen peroxide, and the concentration of hydrogen peroxide is 6.8 g / L.

[0067] In this scheme, the purity of ammonium molybdate > 99%, and the recovery rate of molybdenum is 96.7%; the purity of vanadium pentoxide > 99.5%, and the recovery rate of vanadium is 97.8%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.005%, Al < 0.003%, and Ni < 0.002%.

[0068] Example 2

[0069] This example provides a method for recovering valuable metals from a vanadium-molybdenum-containing waste catalyst, including the following steps:

[0070] S1: Low-temperature roasting: Place 1000 g of the vanadium-molybdenum-containing waste catalyst in a box-type resistance furnace, heat it to 200 °C at a rate of 5 °C / min, keep it warm for 2 h, then continue to heat it to 500 °C at a rate of 10 °C / min, keep it warm for 120 min, cool it, and grind it until it completely passes through a 200-mesh sieve to obtain a roasted material.

[0071] S2: Gradient acid leaching: Stir and mix the roasted material evenly with 3 L of sulfuric acid solution A, stir and leach it in a 60 °C water bath for 90 min, filter it under pressure to obtain leachate A and residue A; stir and mix residue A evenly with 3 L of sulfuric acid solution B, stir and leach it in an 80 °C water bath for 90 min, filter it under pressure to obtain leachate B and residue B; rinse residue B with a sulfuric acid solution with a concentration of 50 g / L, wash it 2 times, with a dosage of 200 mL each time, to obtain a washing solution; combine leachate A, leachate B, and the washing solution to obtain a combined solution.

[0072] S3: Purification and extraction: Slowly add ammonia water with a concentration of 50 g / L to the combined solution, adjust the pH to about 2.5, stir for 25 min, then let it stand for 25 min, filter it, then slowly add ammonia water with a concentration of 50 g / L, adjust the pH to about 4.5, stir for 25 min, let it stand for 25 min, filter it, slowly add 1.5 L of a dimethylglyoxime ethanol solution with a concentration of 100 g / L, and heat it to 45 °C, keep it warm for 30 min, and perform centrifugal separation to obtain a purified solution; stir and mix the purified solution and the extraction solution in a volume ratio of 1:1 for 15 min, let it stand and separate into layers to obtain a vanadium-molybdenum-containing organic phase.

[0073] S4: Mix the vanadium- and molybdenum-containing organic phase with a NH4HCO3 solution at a concentration of 120 g / L in a volume ratio of 1:1.5, stir for 15 min, let it stand for layering, to obtain stripping liquid A and vanadium-containing organic phase; Slowly add ammonia water at a concentration of 50 g / L to the stripping liquid A, adjust the pH to 8.5, evaporate and concentrate to one-third of the original volume at 80 °C, then crystallize in an ice bath for 24 h. After centrifugal separation, transfer it to a drying oven at 105 °C and dry to constant weight to obtain ammonium molybdate;

[0074] Mix the vanadium-containing organic phase with a sodium hydroxide solution at a concentration of 200 g / L in a volume ratio of 1:1, stir for 10 min, let it stand for layering, to obtain stripping liquid B; Slowly add a sulfuric acid solution at a concentration of 50 g / L and 300 g of ammonium sulfate to the stripping liquid B, adjust the pH to 2.0, raise the temperature to 60 °C, stir and mix for 30 min, let it stand for precipitation for 5 h. After centrifugal separation, wash it 3 times with a 2% ammonium sulfate solution by mass to obtain vanadium precipitate;

[0075] Transfer the vanadium precipitate to a sintering furnace, heat it to 620 °C at a rate of 10 °C / min, calcine for 80 min, cool it with the furnace and then grind it to obtain vanadium pentoxide.

[0076] Among them, in step S2, sulfuric acid solution A includes sulfuric acid and oxalic acid, the concentration of sulfuric acid is 60 g / L, and the concentration of oxalic acid is 90 g / L;

[0077] Sulfuric acid solution B includes sulfuric acid, thiourea and ethylenediaminetetraacetic acid, the concentration of sulfuric acid is 180 g / L, the concentration of thiourea is 120 g / L, and the concentration of ethylenediaminetetraacetic acid is 5 g / L.

[0078] In step S3, the extraction liquid includes methyltrioctylammonium chloride, nonylphenol polyoxyethylene ether and Isopar M isoparaffin, the concentration of methyltrioctylammonium chloride is 54 g / L, and the concentration of nonylphenol polyoxyethylene ether is 80 g / L;

[0079] In step S4, the sodium hydroxide solution also includes hydrogen peroxide, and the concentration of hydrogen peroxide is 10.2 g / L.

[0080] In this scheme, the purity of ammonium molybdate > 99%, and the recovery rate of molybdenum is 97.5%; the purity of vanadium pentoxide > 99.5%, and the recovery rate of vanadium is 98.2%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.005%, Al < 0.003%, and Ni < 0.002%.

[0081] Example 3

[0082] This example provides a method for recovering valuable metals from a vanadium- and molybdenum-containing waste catalyst, including the following steps:

[0083] S1: Low-temperature roasting: Place 1000 g of vanadium-molybdenum waste catalyst in a box-type resistance furnace, heat it up to 190 °C at a rate of 5 °C / min, keep it warm for 3 h, then continue to heat it up to 480 °C at a rate of 10 °C / min, keep it warm for 150 min, cool it, and grind it until it completely passes through a 200-mesh sieve to obtain the roasted material;

[0084] S2: Gradient acid leaching: Stir and mix the roasted material with 4 L of sulfuric acid solution A evenly, stir and leach it in a 55 °C water bath for 110 min, after pressure filtration, obtain leachate A and residue A; Stir and mix residue A with 4 L of sulfuric acid solution B evenly, stir and leach it in a 75 °C water bath for 110 min, after pressure filtration, obtain leachate B and residue B; Wash residue B with a sulfuric acid solution with a concentration of 50 g / L, wash it 2 times, with a dosage of 200 mL each time, to obtain the washing solution; Combine leachate A, leachate B, and the washing solution to obtain the combined solution;

[0085] S3: Purification and extraction: Slowly add ammonia water with a concentration of 50 g / L to the combined solution, adjust the pH to about 2.5, stir for 25 min, then let it stand for 25 min, after filtration, slowly add ammonia water with a concentration of 50 g / L again, adjust the pH to about 4.5, stir for 25 min, then let it stand for 25 min, after filtration, slowly add 2 L of a dimethylglyoxime ethanol solution with a concentration of 80 g / L, and heat it up to 40 °C, keep it warm for 25 min, and perform centrifugal separation to obtain the purified solution; Stir and mix the purified solution and the extraction solution at a volume ratio of 1:1.2 for 15 min, let it stand and separate into layers to obtain the vanadium-molybdenum-containing organic phase;

[0086] S4: Mix the vanadium-molybdenum-containing organic phase with an NH4HCO3 solution with a concentration of 100 g / L at a volume ratio of 1:1.8, stir for 15 min, let it stand and separate into layers to obtain the stripping solution A and the vanadium-containing organic phase; Slowly add ammonia water with a concentration of 50 g / L to the stripping solution A, adjust the pH to 8.5, evaporate and concentrate it to one-third of the original volume at 80 °C, then perform ice bath crystallization for 22 h, after centrifugal separation, transfer it to a drying oven at 105 °C and dry it to constant weight to obtain ammonium molybdate;

[0087] Mix the vanadium-containing organic phase with a sodium hydroxide solution with a concentration of 150 g / L at a volume ratio of 1:1.2, stir for 15 min, let it stand and separate into layers to obtain the stripping solution B; Slowly add a sulfuric acid solution with a concentration of 50 g / L and 270 g of ammonium sulfate to the stripping solution B, adjust the pH to 2.0, heat it up to 60 °C, stir and mix for 30 min, let it stand and precipitate for 4.5 h, after centrifugal separation, wash it 3 times with an ammonium sulfate solution with a mass fraction of 2% to obtain vanadium precipitates;

[0088] Transfer the vanadium precipitates to a sintering furnace, heat it up to 600 °C at a rate of 10 °C / min, calcine it for 90 min, cool it with the furnace, and then grind it to obtain vanadium pentoxide.

[0089] Among them, in step S2, sulfuric acid solution A includes sulfuric acid and oxalic acid, the concentration of sulfuric acid is 50 g / L, and the concentration of oxalic acid is 75 g / L;

[0090] Sulfuric acid solution B includes sulfuric acid, thiourea and ethylenediaminetetraacetic acid. The concentration of sulfuric acid is 160 g / L, the concentration of thiourea is 100 g / L, and the concentration of ethylenediaminetetraacetic acid is 4 g / L.

[0091] In step S3, the extractant includes methyltrioctylammonium chloride, nonylphenol polyoxyethylene ether and Isopar M isoparaffin. The concentration of methyltrioctylammonium chloride is 36 g / L, and the concentration of nonylphenol polyoxyethylene ether is 70 g / L;

[0092] In step S4, the sodium hydroxide solution further includes hydrogen peroxide, and the concentration of hydrogen peroxide is 8.5 g / L.

[0093] In this scheme, the purity of ammonium molybdate > 99%, and the recovery rate of molybdenum is 97.2%; the purity of vanadium pentoxide > 99.5%, and the recovery rate of vanadium is 98.0%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.005%, Al < 0.003%, and Ni < 0.002%.

[0094] Example 4

[0095] The difference between this example and Example 3 is:

[0096] In sulfuric acid solution B, ascorbic acid with a concentration of 0.5 g / L is further included.

[0097] Other conditions are the same as those in Example 3.

[0098] In this scheme, the purity of ammonium molybdate > 99%, and the recovery rate of molybdenum is 97.7%; the purity of vanadium pentoxide > 99.5%, and the recovery rate of vanadium is 98.1%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.003%, Al < 0.003%, and Ni < 0.002%.

[0099] Example 5

[0100] The difference between this example and Example 4 is:

[0101] In sulfuric acid solution B, the concentration of ascorbic acid is 1 g / L.

[0102] Other conditions are the same as those in Example 4.

[0103] In this scheme, the purity of ammonium molybdate > 99%, and the recovery rate of molybdenum is 97.9%; the purity of vanadium pentoxide > 99.5%, and the recovery rate of vanadium is 98.2%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.003%, Al < 0.003%, and Ni < 0.002%.

[0104] Example 6

[0105] The difference between this example and Example 5 is as follows:

[0106] In step S3, ammonia water with a concentration of 50 g / L is slowly added to the combined solution, the pH is adjusted to about 2.5, after stirring for 25 min, it is left standing for 25 min, after filtration, ammonia water with a concentration of 50 g / L is slowly added again, the pH is adjusted to about 4.5, after stirring for 25 min, it is left standing for 25 min, after filtration, it passes through a resin column loaded with fluoride ion-chelating resin at a flow rate of 2 BV / h, then 2 L of a dimethylglyoxime ethanol solution with a concentration of 80 g / L is slowly added to the liquid passing through the resin column, the temperature is raised to 45 °C, and it is kept warm for 30 min, and then centrifuged to obtain a purified solution; the purified solution and the extractant are stirred and mixed at a volume ratio of 1:1.2 for 15 min, left standing for liquid separation, to obtain a vanadium- and molybdenum-containing organic phase;

[0107] The preparation method of the fluoride ion-chelating resin includes the following steps:

[0108] The chelating resin is impregnated in a sodium chloride solution, after soaking for 22 h, it is filtered, rinsed 3 times with deionized water, then impregnated in a sodium hydroxide solution, soaked for 3 h, filtered, rinsed 3 times with deionized water, and then loaded into an ion exchange column, and a sodium fluoride solution is passed through at a flow rate of 5 BV / h until the ratio of the fluoride ion concentration in the effluent to the fluoride ion concentration in the sodium fluoride solution is 95.5%, thus obtaining it.

[0109] Among them, the concentration of the sodium fluoride solution is 0.5 mol / L, the mass-volume ratio of the sodium fluoride solution to the chelating resin is 3 mL:1 g; the mass fraction of the sodium chloride solution is 4%, the mass-volume ratio of the sodium chloride solution to the chelating resin is 3 mL:1 g;

[0110] The mass fraction of the sodium hydroxide solution is 8%, the mass-volume ratio of the sodium hydroxide solution to the chelating resin is 3 mL:1 g; the chelating resin is D418 macroporous amino phosphonic acid-based chelating resin.

[0111] Others are the same as in Example 5.

[0112] In this scheme, the purity of ammonium molybdate > 99.5%, the recovery rate of molybdenum is 98.2%; the purity of vanadium pentoxide > 99.8%, the recovery rate of vanadium is 98.5%, and the impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.001%, Al < 0.001%, Ni < 0.001%.

[0113] Example 7

[0114] The difference between this example and Example 6 is as follows:

[0115] In step S3, ammonia water with a concentration of 50 g / L is slowly added to the combined solution to adjust the pH to about 2.5. After stirring for 25 min, it is left standing for 25 min. After filtration, ammonia water with a concentration of 50 g / L is slowly added again to adjust the pH to about 4.5. After stirring for 25 min, it is left standing for 25 min. After filtration, it passes through a resin column filled with chelating resin loaded with fluoride ions at a flow rate of 4 BV / h. Then, 2 L of a dimethylglyoxime ethanol solution with a concentration of 80 g / L is slowly added to the liquid passing through the resin column, and the temperature is raised to 45 °C and kept warm for 30 min. After centrifugal separation, a purified solution is obtained; the purified solution and the extractant are stirred and mixed at a volume ratio of 1:1.2 for 15 min and left standing for liquid separation to obtain a vanadium- and molybdenum-containing organic phase;

[0116] A preparation method of chelating resin loaded with fluoride ions includes the following steps:

[0117] The chelating resin is impregnated in a sodium chloride solution. After soaking for 26 h, it is filtered, rinsed 3 times with deionized water, then impregnated in a sodium hydroxide solution, soaked for 5 h, filtered, rinsed 3 times with deionized water, and then loaded into an ion exchange column. A sodium fluoride solution is passed through at a flow rate of 10 BV / h until the ratio of the fluoride ion concentration in the effluent to the fluoride ion concentration in the sodium fluoride solution is 96.5%, thus obtaining the product.

[0118] Among them, the concentration of the sodium fluoride solution is 1.0 mol / L, and the mass-volume ratio of the sodium fluoride solution to the chelating resin is 5 mL:1 g; the mass fraction of the sodium chloride solution is 6%, and the mass-volume ratio of the sodium chloride solution to the chelating resin is 5 mL:1 g;

[0119] The mass fraction of the sodium hydroxide solution is 10%, and the mass-volume ratio of the sodium hydroxide solution to the chelating resin is 5 mL:1 g; the chelating resin is D418 macroporous aminophosphonic acid-based chelating resin.

[0120] Other conditions are the same as in Example 5.

[0121] In this scheme, the purity of ammonium molybdate > 99.5%, and the recovery rate of molybdenum is 98.5%; the purity of vanadium pentoxide > 99.8%, and the recovery rate of vanadium is 98.8%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.001%, Al < 0.001%, and Ni < 0.001%.

[0122] Comparative Example 1

[0123] The difference between this comparative example and Example 1 is that:

[0124] In step S2, sulfuric acid solution A includes sulfuric acid, and the concentration of sulfuric acid is 30 g / L.

[0125] Other conditions are the same as in Example 1.

[0126] In this solution, the purity of ammonium molybdate is 97.8%, and the recovery rate of molybdenum is 89.7%; the purity of vanadium pentoxide is 98.0%, and the recovery rate of vanadium is 88.1%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe > 0.05%, Al > 0.1%, and Ni > 0.01%.

[0127] Comparative Example 2

[0128] The difference between this comparative example and Example 1 is that:

[0129] In step S2, sulfuric acid solution A contains sulfuric acid with a concentration of 30 g / L; sulfuric acid solution B contains sulfuric acid with a concentration of 150 g / L.

[0130] Others are the same as in Example 1.

[0131] In this solution, the purity of ammonium molybdate is 96.0%, and the recovery rate of molybdenum is 85.5%; the purity of vanadium pentoxide is 96.4%, and the recovery rate of vanadium is 78.0%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe > 0.05%, Al > 0.1%, and Ni > 0.01%.

[0132] Comparative Example 3

[0133] The difference between this comparative example and Example 1 is that:

[0134] In step S3, ammonia water with a concentration of 50 g / L is slowly added to the combined solution to adjust the pH to about 3. After stirring for 20 min, it is left standing for 20 min. After filtration, ammonia water with a concentration of 50 g / L is slowly added again to adjust the pH to about 5. After stirring for 20 min, it is left standing for 20 min. After filtration, a purified solution is obtained; the purified solution and the extraction solution are stirred and mixed at a volume ratio of 1:1.5 for 10 min, left standing for liquid separation, and an organic phase containing vanadium and molybdenum is obtained.

[0135] Others are the same as in Example 1.

[0136] In this solution, the purity of ammonium molybdate is 98.1%, and the recovery rate of molybdenum is 90.1%; the purity of vanadium pentoxide is 98.3%, and the recovery rate of vanadium is 89.4%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe < 0.005%, Al < 0.003%, and Ni > 0.1%.

[0137] Comparative Example 4

[0138] The difference between this comparative example and Example 1 is that:

[0139] In step S3, slowly add ammonia water with a concentration of 50 g / L to the combined solution, adjust the pH to about 5, stir for 20 min, then let it stand for 20 min. After filtration, slowly add 2.5 L of a dimethylglyoxime ethanol solution with a concentration of 60 g / L, and raise the temperature to 35 °C, keep warm for 20 min, and perform centrifugal separation to obtain a purified solution; stir and mix the purified solution and the extraction solution at a volume ratio of 1:1.5 for 10 min, let it stand for layering to obtain a vanadium- and molybdenum-containing organic phase;

[0140] Others are the same as in Example 1.

[0141] In this scheme, the purity of ammonium molybdate is 97.9%, and the recovery rate of molybdenum is 89.5%; the purity of vanadium pentoxide is 98.0%, and the recovery rate of vanadium is 88.7%. The impurity contents in ammonium molybdate and vanadium pentoxide are Fe > 0.03%, Al > 0.05%, and Ni > 0.005%.

[0142] It can be seen from Examples 1 to 3 and Comparative Examples 1 to 4 that during gradient acid leaching, by adding oxalic acid to sulfuric acid solution A, oxalic acid can form stable oxalate complexes with vanadium and molybdenum ions in the easily soluble state in the waste catalyst, reducing the concentration of free vanadium and molybdenum ions in the solution and promoting more vanadium and molybdenum ions to dissolve into the solution. Thiourea can reduce the insoluble metal (such as the high-valent vanadium forming a chemical bond with the carrier) to a low-valent state, break its chemical bond with the carrier, and make the metal ions more easily dissolved by sulfuric acid. At the same time, the reduction effect of thiourea can also inhibit the hydrolysis of impurity ions (such as Fe 3+ ) to avoid the formation of insoluble hydroxide precipitates coating vanadium and molybdenum, thereby improving the leaching rate of vanadium and molybdenum.

[0143] In Comparative Example 3, dimethylglycol oxime was not used, resulting in nickel ions remaining in the solution. During the extraction process, nickel ions compete with vanadium and molybdenum for the extractant, occupying the extraction sites and reducing the extraction efficiency of vanadium and molybdenum.

[0144] Nickel impurities compete with vanadium and molybdenum for extraction, reducing the recovery rates of vanadium and molybdenum. In Comparative Example 4, single-step neutralization (pH directly adjusted to 5) was used, and the aluminum / iron ions in the solution could not be fully hydrolyzed to form precipitates. In contrast, Examples 1 to 3 adopt stepwise neutralization to control the hydrolysis degree of aluminum / iron ions in stages, removing some easily hydrolyzed impurities at a lower pH first to avoid the co-precipitation of vanadium and molybdenum ions being wrapped due to too fast precipitation, which affects the recovery rate of target metals.

[0145] In Examples 4 to 5, by adding ascorbic acid to sulfuric acid solution B, it synergizes with thiourea to further inhibit the oxidation activity of ferric iron, reduce iron impurities, and at the same time reduce the interference of iron on the extraction process, indirectly improving the recovery rates of vanadium and molybdenum. In Examples 6 to 7, D418 resin loaded with fluoride ions is introduced, and aluminum / iron ions are deeply removed through the "fluoride competition complexation + resin specific adsorption synergistic effect", further reducing aluminum / iron impurities.

[0146] This specific embodiment is only an interpretation of the present application and does not limit the present application. After reading this specification, those skilled in the art can make modifications to this embodiment without creative contributions as needed, but as long as it is within the scope of the claims of the present application, it is protected by the patent law.

Claims

1. A method for recovering valuable metals from a vanadium-molybdenum-containing waste catalyst, characterized in that, It includes the following steps: S1: Low-temperature roasting: Roast the vanadium-molybdenum-containing waste catalyst at 180 - 200 °C for 2 - 4 h, then continue to heat up to 450 - 500 °C and roast for 120 - 180 min, and then cool to obtain the roasted material; S2: Gradient acid leaching: Mix the roasted material with sulfuric acid solution A containing oxalic acid, and leach at 50 - 60 °C for 90 - 120 min. After solid-liquid separation, obtain leaching solution A and residue A; Mix residue A with sulfuric acid solution B containing thiourea and ethylenediaminetetraacetic acid, and perform secondary leaching at 70 - 80 °C for 90 - 120 min. After solid-liquid separation, obtain leaching solution B and residue B. Wash residue B with acid solution 2 - 3 times to obtain the washing solution; Combine leaching solution A, leaching solution B and the washing solution to obtain the combined solution; S3: Purification and extraction: Adjust the pH of the combined solution to 2.5 - 3, after solid-liquid separation, then adjust the pH to 4.5 - 5, after solid-liquid separation, add dimethylglyoxime alcohol solution, heat up to 35 - 45 °C, keep warm for 20 - 30 min, and perform solid-liquid separation to obtain the purified solution; And mix the purified solution with the extraction solution evenly, let it stand for layering to obtain the vanadium-molybdenum-containing organic phase; S4: Back-extraction and product preparation: Mix the vanadium-molybdenum-containing organic phase with a weak acid-weak base salt back-extraction agent evenly, let it stand for layering to obtain back-extraction solution A and the vanadium-containing organic phase. Back-extraction solution A is alkalized, evaporated and crystallized to obtain ammonium molybdate; Mix the vanadium-containing organic phase with a strong base back-extraction agent evenly, let it stand for layering to obtain back-extraction solution B. Back-extraction solution B is acidified, ammonium salt precipitated, solid-liquid separated and calcined to obtain vanadium pentoxide.

2. The method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts according to claim 1, wherein In the sulfuric acid solution A, the concentration of sulfuric acid is 30 g / L - 60 g / L, and the concentration of oxalic acid is 45 - 90 g / L.

3. The method for recovering valuable metals from the vanadium-molybdenum-containing waste catalyst according to claim 2, wherein In the sulfuric acid solution B, the concentration of sulfuric acid is 150 g / L - 180 g / L, the concentration of thiourea is 80 - 120 g / L, and the concentration of ethylenediaminetetraacetic acid is 2 g / L - 5 g / L.

4. The method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts according to claim 1, characterized in that, The dimethylglyoxime alcohol solution includes dimethylglyoxime and ethanol, and the concentration of dimethylglyoxime is 60 - 100 g / L.

5. The method for recovering valuable metals from vanadium-molybdenum waste catalysts according to claim 1, wherein The extraction solution includes 18 g / L - 54 g / L methyltrioctylammonium chloride, 50 g / L - 80 g / L nonylphenol polyoxyethylene ether and isoparaffin diluent.

6. The method for recycling valuable metals from vanadium-molybdenum-containing waste catalysts according to claim 1, characterized in that, The weak acid-weak base salt back-extraction agent is an NH4HCO3 solution, the concentration of the NH4HCO3 solution is 80 g / L - 120 g / L, and the volume ratio of the weak acid-weak base salt back-extraction agent to the vanadium-molybdenum-containing organic phase is 1:(1.5 - 2).

7. The method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts according to claim 6, wherein The strong base back-extraction agent includes sodium hydroxide and hydrogen peroxide, the concentration of sodium hydroxide is 120 g / L - 200 g / L, and the concentration of hydrogen peroxide is 6.8 g / L - 10.2 g / L.

8. The method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts according to claim 1, characterized in that, In the sulfuric acid solution B, it also includes ascorbic acid, and the concentration of ascorbic acid is 0.5 - 1.0 g / L.

9. The method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts according to claim 1, wherein In step S3, before adding the dimethylglyoxime alcohol solution, first pass through a resin column loaded with fluoride ion chelating resin at a flow rate of 2 - 4 BV / h.

10. The method for recovering valuable metals from vanadium-molybdenum-containing waste catalysts according to claim 9, wherein The preparation method of the fluoride ion-loaded chelating resin includes the following steps: The chelating resin is impregnated in a sodium chloride solution. After soaking for 22 - 26 h, solid-liquid separation is carried out. After washing, it is impregnated in a sodium hydroxide solution and soaked for 3 - 5 h. Then solid-liquid separation is carried out again. After washing, it is loaded into an ion exchange column, and a sodium fluoride solution is passed through at a flow rate of 5 - 10 BV / h until the ratio of the fluoride ion concentration in the effluent to the fluoride ion concentration in the sodium fluoride solution is higher than 95%, thus obtaining the product.

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