Method for recovering valuable metals from waste hydrogenation catalyst

Abstract

The invention discloses a method for recovering valuable metals from a waste hydrogenation catalyst, and relates to the technical field of waste catalyst recovery, and the method comprises the following steps: carrying out oxidizing roasting and oxygen pressure alkaline leaching on the waste hydrogenation catalyst to obtain an alkaline leaching solution and alkaline leaching residues; aluminum, vanadium and molybdenum are recycled from the alkaline leaching solution, wherein carbon dioxide is introduced into the alkaline leaching solution for aluminum precipitation; ammonium sulfate and magnesium sulfate are added into the filtrate subjected to aluminum precipitation to remove phosphorus, silicon and arsenic, molybdenum and vanadium are adsorbed through resin, and vanadium precipitation and molybdenum precipitation are carried out; and aluminum and nickel are recycled from alkaline leaching residues, specifically, aluminum is precipitated through oxygen pressure acid leaching, then alkali is added into filtrate to further recycle aluminum, and then sodium carbonate is added for nickel precipitation. By means of the method, molybdenum, vanadium, nickel, aluminum and other elements in the waste catalyst can be effectively recycled, the recycling rate is high, the molybdenum recycling rate can reach 97%, the vanadium recycling rate can reach 95%, the nickel recycling rate can reach 92%, comprehensive utilization of the waste catalyst is achieved, and good social value and economic value are achieved.

Description

Technical Field

[0001] The invention relates to the technical field of waste catalyst recovery, and in particular to a method for recovering valuable metals from waste hydrogenation catalysts. Background Art

[0002] A large amount of waste hydrogenation catalysts are produced during the petroleum refining process. These catalysts contain a large amount of useful metals such as Mo, V, Ni, Al, etc., and are important secondary resources. Recovering metals and other useful components from them can not only reduce the pollution caused by direct emissions of waste hydrogenation catalysts to the environment, but also have certain economic benefits.

[0003] At present, the mainstream method for recycling waste hydrogenation catalysts is to use the sodium calcination-water leaching process, which is to use a high-temperature sodium calcination at 800-1000°C in a rotary kiln or a boiling furnace, with an organic matter ignition reduction rate of more than 99.99%. After high-temperature sodium calcination, water leaching is used to extract Mo, V, Ni, etc. in a wet process. In this method, Ni and Al cannot be comprehensively utilized, and there are problems such as high material consumption, serious equipment corrosion, large amount of high-salt wastewater, and heavy environmental burden. Summary of the invention

[0004] Based on the technical problems existing in the background technology, the present invention proposes a method for recovering valuable metals from waste hydrogenation catalysts, and the recovery rate of molybdenum, vanadium, nickel and aluminum elements is high.

[0005] The present invention provides a method for recovering valuable metals from a waste hydrogenation catalyst, comprising the following steps: 1) Oxidative roasting: The spent hydrogenation catalyst is subjected to oxidative roasting and crushed to obtain a roasting material; 2) Oxygen pressure alkali leaching: Place the roasted material in a pressure autoclave, add a mixed alkali consisting of sodium carbonate and sodium hydroxide, perform oxygen pressure alkali leaching, filter, and obtain alkali leaching solution and alkali leaching residue; 3) Recovery of aluminum, vanadium and molybdenum in alkaline leaching solution: 3.1) Aluminum precipitation 1: Add carbon dioxide to the alkaline leaching solution to adjust the pH of the system to 9.5-10.5, age, filter, and obtain pseudo-boehmite and filtrate; calcine the pseudo-boehmite to obtain activated alumina; 3.2) Impurity removal: adjust the pH of the filtrate after aluminum precipitation 1 to 10-11, add ammonium sulfate and magnesium sulfate to purify and remove impurities, stir the reaction, filter, and obtain a purified liquid after phosphorus, silicon and arsenic are removed; The pH of the purified liquid is adjusted to 3.5-4.5, and the molybdenum and vanadium are adsorbed by an anion exchange resin, and the liquid after the resin is collected; the resin adsorbed with molybdenum and vanadium is analyzed to obtain an analysis liquid rich in molybdenum and vanadium; 3.3) Vanadium precipitation: adjust the pH of the analytical solution to 7.5-8.5, add ammonium sulfate, stir to react, filter, and obtain ammonium metavanadate precipitate and vanadium precipitation filtrate; 3.4) Molybdenum precipitation: add sulfuric acid to the vanadium precipitation filtrate to adjust the pH value to 2.0-3.0, stir to react, filter, and obtain molybdic acid precipitate and molybdenum precipitation filtrate; 4) Recovery of aluminum and nickel from alkaline leaching residue: 4.1) Oxygen pressure acid leaching: Place the alkali leaching residue in a pressure autoclave for oxygen pressure acid leaching, add sulfuric acid to adjust the pH to 1.0-1.5, perform oxygen pressure acid leaching, filter, and obtain bauxite and acid leaching solution; 4.2) Aluminum precipitation 2: add alkali solution to the acid leaching solution to adjust the pH to 5, stir the reaction, filter, and obtain aluminum hydroxide and the filtrate after aluminum precipitation; 4.3) Nickel precipitation: Add sodium carbonate to the filtrate after aluminum precipitation, stir to react, filter, and obtain basic nickel carbonate precipitate.

[0006] Furthermore, in step 1), when the oil content in the waste hydrogenation catalyst is ≥15%, it also includes performing thermal desorption pretreatment on the waste hydrogenation catalyst before oxidative roasting, specifically, heating the waste hydrogenation catalyst in a closed oxygen-proof environment to perform thermal desorption to remove oil; Preferably, the oil content in the spent hydrogenation catalyst after thermal desorption treatment is controlled to be ≤8%; preferably, the thermal desorption and oil removal is performed by heating to ≤550°C.

[0007] Further, in step 1), the calcination temperature is 600-700°C and the calcination time is 2-5h; Preferably, after the oxidative calcination, the calcined material is wet ball-milled and sieved.

[0008] Further, in step 2), the mass percentage of sodium carbonate and sodium hydroxide in the mixed alkali is 65-75:25-35; Preferably, the conditions for oxygen pressure alkali leaching are: liquid-solid ratio 3-4:1, oxygen pressure 0.5-0.6 MPa, temperature 170-190° C., leaching time 60-120 min; and the endpoint pH after oxygen pressure alkali leaching is controlled to be 11.5-12.5.

[0009] Furthermore, in step 3.1), the alkaline leaching solution is heated to 80-90°C, carbon dioxide is introduced thereinto to adjust the pH value of the system to 9.5-10.5, and the temperature is lowered and aged; preferably, the aging time is 3-4 hours.

[0010] Further, in step 3.2), ammonium sulfate and magnesium sulfate are added, and the mixture is stirred at 50-70°C for 20-40 minutes; Preferably, the process further comprises recycling the post-resin liquid, that is, the post-resin liquid is concentrated, frozen and crystallized, and centrifuged to obtain sodium sulfate crystals and a centrifuge mother liquor rich in ammonium sulfate; wherein the centrifuge mother liquor is recycled to the vanadium precipitation process in step 3.3).

[0011] Furthermore, in step 3.3), the temperature of vanadium precipitation is 50-70°C.

[0012] Further, in step 3.4), the stirring reaction temperature is 70-80°C and the reaction time is 40-80min; Preferably, the method further comprises enriching or concentrating the molybdenum precipitation filtrate and repeating the resin adsorption, vanadium precipitation and molybdenum precipitation operations in the above steps 3.2) to 3.4).

[0013] Furthermore, in step 4.1), the conditions of oxygen pressure acid leaching are: liquid-solid ratio 4-6:1, oxygen pressure 0.4-0.6 MPa, temperature 160-190°C, and leaching time 2-2.5 h.

[0014] Furthermore, in step 4.3), sodium carbonate is added to the filtrate after precipitation of aluminum 2 to control the pH at the reaction end point to be 8.0-8.5; Preferably, the method further comprises adjusting the pH of the filtrate after nickel precipitation to 3.0-3.5, adsorbing nickel through a cation exchange resin, analyzing to obtain a nickel-containing analytical solution, and returning the solution to the oxygen pressure acid leaching process for treatment.

[0015] In the present invention, in step 1), metals such as molybdenum, vanadium and nickel in the waste catalyst are oxidized by oxidative roasting to promote the subsequent alkaline leaching process.

[0016] Furthermore, in addition to molybdenum, vanadium, nickel and carrier aluminum as active ingredients, the waste hydrogenation catalysts generated in the oil refining process often also contain a high content of petroleum hydrocarbons with a calorific value of more than 4,000 kcal, which affects the subsequent roasting and leaching efficiency. Therefore, when the oil content in the waste hydrogenation catalyst is ≥15%, it needs to be pretreated by thermal desorption before oxidative roasting. This pretreatment process not only recovers the oil, but also reduces the subsequent roasting temperature.

[0017] Furthermore, by oxidative roasting at 600-700°C, metals such as molybdenum, vanadium and nickel are well oxidized, and the aluminum oxide at this temperature is γ-alumina, the subsequent aluminum leaching effect is good and it is more energy-saving and efficient than the traditional roasting process of 800-1000°C.

[0018] Furthermore, after roasting, the material is ground to a certain particle size by wet ball milling, which can ensure the subsequent alkali leaching efficiency and the leaching rate of molybdenum and vanadium; preferably, the material is wet ball milled and then sieved through a 60-100 mesh sieve.

[0019] In the present invention, oxygen pressure alkali leaching is used in step 2). Compared with the traditional sodium soda ash roasting, the present application directly roasts and then leaches under high pressure, which has strong uniformity and reduces the alkali preparation and aging processes; the process is more compact, reduces the open-air stacking of materials, reduces the emission of a certain amount of unorganized waste gas, and is more economical and environmentally friendly. In addition, the mixed alkali oxygen pressure leaching composed of soda ash and flake alkali has good leaching properties for molybdenum and vanadium, and the obtained alkali leaching solution is rich in substances such as sodium molybdate, sodium vanadate and sodium metaaluminate.

[0020] The valuable metals in the alkali leaching solution and the alkali leaching residue in the present invention are recovered.

[0021] Recovery of aluminum, vanadium and molybdenum from alkaline leaching solution: In the present invention, in step 3.1), aluminum is first precipitated in the alkaline leaching solution, and the solution is heated and stirred and carbon dioxide is introduced to adjust the pH value to precipitate aluminum to obtain pseudo-boehmite, which is aluminum hydroxide with higher purity, and activated alumina can be obtained after roasting.

[0022] In step 3.2), auxiliary materials such as ammonium sulfate and magnesium sulfate are added to the filtrate after aluminum precipitation to remove impurities such as phosphorus, arsenic, silicon, etc. in the solution to obtain a purified solution.

[0023] The main chemical reactions involved are: 2Na3PO4+ 2MgSO4 + (NH4)2SO4+ 12H2O = 2MgNH4PO4·6H2O↓+ 3Na2SO4 2Na3PO4 + 3MgSO4 = Mg3(PO4)2↓+ 3Na2SO4 Na2SiO3 + MgSO4 = MgSiO3↓+ Na2SO4 2Na3AsO4+ 3MgSO4= Mg3(AsO4)2↓ + 3Na2SO4 The amount of ammonium sulfate and magnesium sulfate added can be based on the theoretical amount required for precipitating phosphorus, arsenic, and silicon.

[0024] Since the content of molybdenum and vanadium in the purified liquid is low (generally 10-20 g / L), it is not suitable for direct precipitation of vanadium and molybdenum. You can first use anion exchange resin for adsorption and then analyze, and then precipitate molybdenum and vanadium in the analyzed liquid. The anion exchange resin here can be D314 anion exchange resin.

[0025] Furthermore, the post-resin liquid is rich in sodium sulfate and ammonium sulfate, which can be recycled. Sodium sulfate in the post-resin liquid is frozen and crystallized using freezing technology to obtain sodium sulfate decahydrate, which is further double-effect evaporated to obtain the glauber powder product (sodium sulfate). The crystallization mother liquor is rich in ammonium sulfate, which is concentrated and reused in the vanadium precipitation process.

[0026] In step 3.3), the pH value of the analytical solution is adjusted to 7.5-8.5. The main components of the solution are sodium metavanadate and sodium molybdate. Sodium metavanadate often polymerizes into Na3(VO3)3 or Na4(VO3)4 in aqueous solution, and the polymer further dissociates into Na + 、V3O9 3- 、V4O 12 4- ions, when NH 4+ A double decomposition reaction will occur to generate a white or yellow precipitate of ammonium metavanadate which is insoluble in water. The amount of ammonium sulfate added depends on the content of metavanadate in the solution.

[0027] The main reaction involved is: 2NaVO3+(NH4)2SO4→2NH4VO3↓+Na2 SO4 In step 3.4), slowly add an appropriate amount of dilute sulfuric acid (30% sulfuric acid can be used) to the filtrate after vanadium precipitation to precipitate molybdenum, and the precipitate is the molybdic acid product.

[0028] Furthermore, the molybdenum precipitation filtrate still contains a small amount of vanadium and molybdenum, which can be enriched or concentrated by adding alkali to adjust the pH to 3.5-4.5, and then adsorbing the residual molybdenum and vanadium through anion exchange resin, repeating the above steps 3.2)-3.4) to perform resin adsorption, vanadium precipitation, and molybdenum precipitation operations.

[0029] Recovery of aluminum and nickel from alkaline leaching residue: In the present invention, in step 4.1), oxygen-pressurized sulfuric acid is used for leaching. Specifically, the alkali leaching residue can be pulped and then added to a high-pressure reactor (specifically, a horizontal titanium high-pressure reactor can be selected), oxygen is introduced to control the partial pressure within a certain range, and the nickel and a part of the aluminum in the alkali leaching residue are leached into the solution (acid leaching solution) by controlling the amount of sulfuric acid, temperature, pressure, and time, and most of the aluminum forms sodium aluminum sulfate-type bauxite.

[0030] In step 4.2), since the acid leaching solution still contains some aluminum, alkali solution is added thereto to convert it into aluminum hydroxide precipitate.

[0031] In step 4.3), sodium carbonate is added to the filtrate after aluminum precipitation to carry out nickel precipitation reaction. The amount of sodium carbonate added is based on the theoretical amount of precipitated nickel carbonate.

[0032] The nickel ion content in the filtrate after nickel precipitation is determined, and then the pH value is adjusted to 3.0-3.5. The nickel is then adsorbed by a nickel-type cation exchange resin and analyzed by sulfuric acid. The obtained nickel-containing analysis solution can be returned to the oxygen pressure acid leaching process and treated together with the alkaline leaching residue; specifically, the nickel-type cation exchange resin can be selected from D861 resin.

[0033] Compared with the prior art, the beneficial effects of the present invention are embodied in that the present invention effectively recovers molybdenum, vanadium, nickel, aluminum and other elements in waste catalysts through processes such as thermal desorption, selective temperature-controlled calcination, high-pressure alkali leaching, wet extraction of molybdenum and vanadium, pressurized acid leaching of nickel-aluminum alkali leaching residues, and high-temperature calcination of acid leaching residues. The recovery rate is high, and the content of harmful elements such as sulfur, phosphorus and arsenic in the recovered products is low. Among them, the molybdenum recovery rate can reach 97%, the vanadium recovery rate can reach 95%, and the nickel recovery can reach 92%, thereby realizing the comprehensive utilization of waste catalysts, and the entire recovery process is relatively simple, with good social value and economic value. BRIEF DESCRIPTION OF THE DRAWINGS

[0034] Figure 1 is a process flow chart of an embodiment of the present invention; Figure 2 This is the XRD pattern of bauxite. DETAILED DESCRIPTION

[0035] The oil content of the waste hydrogenation catalyst in the present invention is about 30% after testing, and the main components are: 4.28% molybdenum, 8.92% vanadium in the form of vanadium pentoxide, 2.38% nickel, and 51.2% aluminum in the form of aluminum trioxide.

[0036] The technical solution of the present invention is described in detail below through specific embodiments. Example 1

[0037] A method for recovering valuable metals from spent hydrogenation catalysts, the process flow is as follows: Figure 1 As shown, the specific steps include: 1) Oxidation roasting The oil-containing molybdenum-vanadium waste hydrogenation catalyst was deoiled at 380°C (the oil content after deoiling was 6%), and then calcined at 650°C for 3 hours to obtain an oxidized calcined material; the oxidized calcined material was then ground into 60-100 mesh.

[0038] 2) Oxygen pressure alkali leaching Weigh 200g of ground oxidized roasted material, add 60g of soda ash, 30g of flake caustic soda, 700g of water, introduce oxygen partial pressure of 0.55Mpa, heat to 180℃, stir at 250rpm, and leaching at high temperature for 1.5h; release the pressure when the temperature is below 60℃, take out the material and filter it to obtain 800mL of alkaline leaching solution (filtrate) (pH=12.5) and 150g of alkaline leaching residue.

[0039] The alkali leached residue was washed with water and its composition was analyzed by chemical analysis method. The main components were as follows: molybdenum 0.22%, vanadium pentoxide 0.68%, nickel 5.02%, aluminum oxide 61.07%; 3) Recovery of aluminum, vanadium and molybdenum in alkaline leaching solution: 3.1 Aluminum deposition 1 The alkaline leaching solution is heated to 90°C in a water bath, and carbon dioxide gas is introduced until the pH value of the system drops to 10, resulting in pseudo-boehmite precipitation. The solution is cooled and aged for 3-4 hours, and filtered to obtain 92g of pseudo-boehmite wet weight. After drying, 40g of aluminum hydroxide is obtained; the filtrate after aluminum precipitation enters the next step of processing.

[0040] 3.2 Impurity Removal The filtrate after aluminum precipitation still contains trace impurities such as P and Si. Add appropriate amounts of magnesium sulfate and ammonium sulfate to the filtrate, adjust the pH to 10.5, and stir the reaction at 60°C for 30 minutes to make the phosphate and silicate in the filtrate form magnesium ammonium phosphate, magnesium phosphate and magnesium silicate precipitation; the main chemical reactions involved are: 2Na3PO4 + 2MgSO4 + (NH4)2SO4 + 12H2O = 2MgNH4PO4·6H2O↓+ 3Na2SO4 2Na3PO4 + 3MgSO4 = Mg3(PO4)2↓+ 3Na2SO4 Na2SiO3 + MgSO4 = MgSiO3↓+ Na2SO4 The material after the reaction was filtered to obtain a purified liquid (filtrate). The purified liquid was adjusted to pH 4, passed through D314 anion exchange resin (Ningbo Zhengguang Resin Co., Ltd.) to adsorb molybdenum and vanadium, and eluted with 8% sodium hydroxide solution after saturation to obtain a molybdenum-vanadium-rich analytical solution.

[0041] 3.3 Vanadium precipitation The molybdenum-vanadium-rich analytical solution was pumped into the reactor, the pH was adjusted to 8, and then an excess of 8% ammonium sulfate was added, and the reaction was carried out at 60°C for 2 hours to generate ammonium metavanadate precipitation; the chemical reaction involved is: 2NaVO3+(NH4)2SO4= 2NH4VO3↓+Na2SO4 The material after vanadium precipitation was filtered to obtain vanadium precipitation filtrate and ammonium metavanadate filter cake; the filter cake was washed, 27.20 grams, the ammonium metavanadate content obtained was 98.80%, and the vanadium recovery rate reached 95%.

[0042] 3.4 Molybdenum deposition The pH of the vanadium precipitation filtrate was adjusted to 2.5 with 30% sulfuric acid and reacted at 75°C for 1 hour to obtain molybdic acid precipitate; the chemical reaction involved is: Na2MoO4+H2SO4 = H2MoO4↓+Na2SO4 The material after molybdenum precipitation was filtered to obtain molybdenum precipitation filtrate and molybdic acid filter cake; the molybdic acid filter cake was washed and dried, and 14.83 grams of molybdic acid was obtained. The molybdenum content was 56.0%, and the molybdenum recovery rate reached 97%.

[0043] 3.5 Ion Exchange The molybdenum precipitation filtrate is enriched and concentrated and then returned to the resin adsorption process of step 3.2. The pH of the molybdenum precipitation filtrate is adjusted to 3.5-4.5 with sodium hydroxide solution, and then the above vanadium precipitation and molybdenum precipitation operations are repeated.

[0044] The main components of the liquid after ion exchange are ammonium sulfate and sodium sulfate, which enter the subsequent sulfate recovery process; the main component of the solution after analysis is molybdate, which returns to the molybdenum precipitation process.

[0045] 3.6 Sulfate Recovery In steps 3.2 and 3.5, the resin liquid after the anion exchange resin adsorbs molybdenum and vanadium is rich in sodium sulfate and ammonium sulfate. The resin liquid is concentrated, then cryogenically frozen and crystallized (freezing temperature is -5°C, crystallization time is 2h), and centrifuged to obtain sodium sulfate decahydrate and centrifuged mother liquor. Among them, sodium sulfate decahydrate is dried to obtain anhydrous sodium sulfate (sodium sulfate); the centrifuged mother liquor contains a large amount of ammonium sulfate, which can be reused in the vanadium precipitation process as an ammonium sulfate auxiliary material.

[0046] 4) Recovery of aluminum and nickel from alkaline leaching residue: 4.1 Sulfuric acid oxygen pressure acid leaching Weigh 180 g of alkaline leaching residue, add 1000 mL of 18% sulfuric acid, and introduce oxygen for pressurized leaching: the leaching temperature is 180°C, the oxygen is introduced at 0.50 MPa, the leaching time is 2 h, the stirring speed is 250 rpm, and the acid leaching residue precipitate and 700 mL of acid leaching solution are obtained; wherein the concentrations of the main components in the acid leaching solution are: nickel 10.1 g / L, and the aluminum content is 15.0 g / L calculated as alumina.

[0047] The acid leaching residue is bauxite (sodium aluminum sulfate trihydrate), which is characterized by XRD (such as Figure 2 As shown), its structural formula satisfies: NaAl3(SO4)2(OH)6, which is a hexagonal bauxite containing Al2(SO4)3 and Al2O3 with crystal water, wherein the content of Al2O3 is>33%; specifically, the main components are: molybdenum 0.16%, vanadium oxide 0.23%, aluminum oxide 33.49%, sodium 2.71%, potassium 0.32%, nickel 0.31%, calcium 1.05%, and magnesium 0.78%.

[0048] 4.2 Dealumination The acid leaching solution still contains 15g / L of aluminum that needs to be removed. Sodium hydroxide solution is added to the acid leaching solution to adjust the pH to 5.0. The aluminum ions in the solution form aluminum hydroxide precipitation, which is incorporated into bauxite for subsequent treatment; the chemical reactions involved are: Al2(SO4)3 + 6NaOH = 2Al(OH)3↓+ 3Na2SO4 After the reaction is completed, suction filtration is performed to obtain aluminum hydroxide residue and filtrate after aluminum precipitation, and the filtrate after aluminum precipitation enters the nickel extraction process.

[0049] 4.3 Nickel deposition Sodium carbonate is added to the filtrate after aluminum precipitation to adjust the pH to 8.5 to form nickel carbonate precipitation. The main reactions that occur are: 10NiSO4+ 7Na2CO3+ 6NaOH+ H2O=7NiCO3·3Ni(OH)2·H2O +10Na2SO4 After the reaction is completed, suction filtration is performed to obtain a basic nickel carbonate product and a filtrate after nickel precipitation, and the filtrate after nickel precipitation enters a resin filtration process.

[0050] The basic nickel carbonate was washed with water and dried, yielding 11.31 g with a nickel content of 38.72% and a nickel recovery rate of 92%.

[0051] 4.4 Resin Filtration The filtrate after nickel precipitation still contains a small amount of nickel, which is adsorbed using nickel-type cation exchange resin D861 (Ningbo Zhengguang Resin Co., Ltd.). First, the pH of the filtrate after nickel precipitation is adjusted to 3.0-3.5, adsorbed by D861 resin, and analyzed with 4% sulfuric acid to obtain a nickel-containing analytical solution, which is reused in the oxygen pressure acid leaching process. Example 2

[0052] A method for recovering valuable metals from spent hydrogenation catalysts, the process flow is as follows: Figure 1 As shown, the specific steps include: 1) Oxidation roasting The oil-containing molybdenum-vanadium waste hydrogenation catalyst is deoiled at 400°C (the oil content after deoiling is 5%), and then calcined at 600°C for 4 hours to obtain an oxidized calcined material; then the oxidized calcined material is ground into 60-100 meshes.

[0053] 2) Oxygen pressure alkali leaching Weigh 200 g of ground oxidized roasted material, add 65 g of soda ash, 25 g of flake caustic soda, 600 g of water, introduce an oxygen partial pressure of 0.50 MPa, heat to 170 ° C, stir at 200 rpm, and leaching at high temperature for 1.5 h; release the pressure when the temperature is below 60 ° C, take out the material and filter it to obtain 500 mL of alkaline leaching solution (filtrate) (pH = 12.0) and 140 g of alkaline leaching residue.

[0054] 3) Recovery of aluminum, vanadium and molybdenum in alkaline leaching solution: 3.1 Aluminum deposition 1 The alkaline leaching solution was heated to 80°C in a water bath, and carbon dioxide gas was introduced until the pH value of the system dropped to 9.5, resulting in pseudo-boehmite precipitation. The solution was cooled and aged for 3-4 hours, and filtered to obtain 55 g of pseudo-boehmite wet weight. After drying, 30 g of aluminum hydroxide was obtained. The filtrate after aluminum precipitation was processed into the next step.

[0055] 3.2 Impurity Removal The filtrate after aluminum precipitation still contains trace impurities such as P and Si. Add appropriate amounts of magnesium sulfate and ammonium sulfate to the filtrate, adjust the pH to 10, and stir the reaction at 50°C for 40 minutes to make the phosphate and silicate in the filtrate form magnesium ammonium phosphate, magnesium phosphate and magnesium silicate precipitation; the main chemical reactions involved are: 2Na3PO4 + 2MgSO4 + (NH4)2SO4 + 12H2O = 2MgNH4PO4·6H2O↓+ 3Na2SO4 2Na3PO4 + 3MgSO4 = Mg3(PO4)2↓+ 3Na2SO4 Na2SiO3 + MgSO4 = MgSiO3↓+ Na2SO4 The material after the above reaction is filtered to obtain a purified liquid (filtrate).

[0056] The pH value of the purified liquid was adjusted to 3.5, and passed through D314 anion exchange resin to adsorb molybdenum and vanadium. After saturated adsorption, it was eluted with 8% sodium hydroxide solution to obtain a molybdenum- and vanadium-rich analytical solution.

[0057] 3.3 Vanadium precipitation The molybdenum-vanadium-rich analytical solution was pumped into the reactor, the pH was adjusted to 7.5, and then an excess of 6% ammonium sulfate was added, and the reaction was carried out at 50°C for 3 hours to generate ammonium metavanadate precipitation; the chemical reaction involved is: 2NaVO3+(NH4)2SO4= 2NH4VO3↓+Na2SO4 The material after vanadium precipitation was filtered to obtain vanadium precipitation filtrate and ammonium metavanadate filter cake; the filter cake was washed to obtain 26.45 grams of ammonium metavanadate with a content of 98.5%, and the vanadium recovery rate reached 92%. 3.4 Molybdenum deposition The pH of the vanadium precipitation filtrate was adjusted to 2 using 30% sulfuric acid, and the reaction was carried out at 70°C for 80 minutes to obtain molybdic acid precipitate; the chemical reaction involved is: Na2MoO4+H2SO4 = H2MoO4↓+Na2SO4 The material after molybdenum precipitation was filtered to obtain molybdenum precipitation filtrate and molybdic acid filter cake; the molybdic acid filter cake was washed and dried, and 14.2 grams of molybdic acid was obtained. The molybdenum content was 56.00%, and the molybdenum recovery rate reached 93%.

[0058] 3.5 Ion Exchange The molybdenum precipitation filtrate is enriched and concentrated and then returned to the resin adsorption process of step 3.2. The pH of the molybdenum precipitation filtrate is adjusted to 3.5-4.5 with sodium hydroxide solution, and then the above vanadium precipitation and molybdenum precipitation operations are repeated.

[0059] The main components of the liquid after ion exchange are ammonium sulfate and sodium sulfate, which enter the subsequent sulfate recovery process; the main component of the solution after analysis is molybdate, which returns to the molybdenum precipitation process.

[0060] 3.6 Sulfate Recovery In steps 3.2 and 3.5, the resin liquid after the anion exchange resin adsorbs molybdenum and vanadium is rich in sodium sulfate and ammonium sulfate. The resin liquid is concentrated, then cryogenically frozen and crystallized (freezing temperature is -5°C, time is 2h), and centrifuged to obtain sodium sulfate decahydrate and centrifuged mother liquor. Among them, sodium sulfate decahydrate is dried to obtain anhydrous sodium sulfate (sodium sulfate); the centrifuged mother liquor contains a large amount of ammonium sulfate, which can be reused in the vanadium precipitation process as an ammonium sulfate auxiliary material.

[0061] 4) Recovery of aluminum and nickel from alkaline leaching residue: 4.1 Sulfuric acid oxygen pressure acid leaching Weigh 180 g of alkaline leaching residue, add 800 mL of 18% sulfuric acid, and introduce oxygen for pressurized leaching: the leaching temperature is 170°C, the oxygen is introduced at 0.40 MPa, the leaching time is 2.5 h, the stirring speed is 250 rpm, and the acid leaching residue precipitate and 600 mL of acid leaching solution are obtained; wherein, the concentration of each main component in the acid leaching solution is: nickel 12.6 g / L, and the aluminum content is calculated as alumina as alumina 8.5 g / L.

[0062] The acid leaching residue is bauxite (sodium aluminum sulfate trihydrate), the main component of which is sodium aluminum sulfate.

[0063] 4.2 Dealumination of acid leaching solution The acid leaching solution still has 8.5 g / L of aluminum that needs to be removed. Sodium hydroxide solution is added to the acid leaching solution to adjust the pH to 4.5. The aluminum ions in the solution form aluminum hydroxide precipitation, which is incorporated into bauxite for subsequent treatment; the chemical reactions involved are: Al2(SO4)3 + 6NaOH = 2Al(OH)3↓+ 3Na2SO4 After the reaction is completed, suction filtration is performed to obtain aluminum hydroxide residue and filtrate after aluminum precipitation, and the filtrate after aluminum precipitation enters the nickel extraction process.

[0064] 4.3 Nickel deposition Sodium carbonate is added to the filtrate after aluminum precipitation to adjust the pH to 8.0 to form nickel carbonate precipitation. The main reactions that occur are: 10NiSO4+ 7Na2CO3+ 6NaOH+ H2O=7NiCO3·3Ni(OH)2·H2O +10Na2SO4 After the reaction is completed, suction filtration is performed to obtain a basic nickel carbonate product and a filtrate after nickel precipitation, and the filtrate after nickel precipitation enters a resin filtration process.

[0065] The basic nickel carbonate was washed with water and dried to obtain 10.95 g, with a nickel content of 40% and a nickel recovery rate of 92%.

[0066] 4.4 Resin Filtration The filtrate after nickel precipitation still contains a small amount of nickel, which is adsorbed by nickel-type cation exchange resin D861. First, the pH of the filtrate after nickel precipitation is adjusted to 3.0-3.5, adsorbed by D861 resin, and analyzed by 4% sulfuric acid to obtain a nickel-containing analytical solution, which is reused in the oxygen pressure acid leaching process. Example 3

[0067] A method for recovering valuable metals from spent hydrogenation catalysts, the process flow is as follows: Figure 1 As shown, the specific steps include: 1) Oxidation roasting The oil-containing molybdenum-vanadium waste hydrogenation catalyst is deoiled at 420°C (the oil content after deoiling is 5%), and then calcined at 700°C for 2 hours to obtain an oxidized calcined material; then the oxidized calcined material is ground into 60-100 meshes.

[0068] 2) Oxygen pressure alkali leaching Weigh 200 g of ground oxidized roasted material, add 67 g of soda ash, 23 g of flake caustic soda, 800 g of water, introduce an oxygen partial pressure of 0.60 MPa, heat to 190 ° C, stir at 300 rpm, and leaching at high temperature for 2 h; release the pressure when the temperature is below 60 ° C, take out the material and filter it to obtain 650 mL of alkaline leaching solution (filtrate) (pH = 12.5) and 162 g of alkaline leaching residue.

[0069] 3) Recovery of aluminum, vanadium and molybdenum in alkaline leaching solution: 3.1 Aluminum deposition 1 The alkaline leaching solution was heated to 85°C in a water bath, and carbon dioxide gas was introduced until the pH value of the system dropped to 10.5, resulting in pseudo-boehmite precipitation. The solution was cooled and aged for 3-4 hours, and filtered to obtain pseudo-boehmite with a wet weight of 51 g. After drying, 29 g of aluminum hydroxide was obtained. The filtrate after aluminum precipitation was processed in the next step.

[0070] 3.2 Impurity Removal The filtrate after aluminum precipitation still contains trace impurities such as P and Si. Add appropriate amounts of magnesium sulfate and ammonium sulfate to the filtrate, adjust the pH to 11, and stir the reaction at 70°C for 20 minutes to make the phosphate and silicate in the filtrate form magnesium ammonium phosphate, magnesium phosphate and magnesium silicate precipitation; the main chemical reactions involved are: 2Na3PO4 + 2MgSO4 + (NH4)2SO4 + 12H2O = 2MgNH4PO4·6H2O↓+ 3Na2SO4 2Na3PO4 + 3MgSO4 = Mg3(PO4)2↓+ 3Na2SO4 Na2SiO3 + MgSO4 = MgSiO3↓+ Na2SO4 The material after the above reaction is filtered to obtain a purified liquid (filtrate).

[0071] The pH value of the purified liquid was adjusted to 4.5, and passed through D314 anion exchange resin to adsorb molybdenum and vanadium. After saturated adsorption, it was eluted with 8% sodium hydroxide solution to obtain a molybdenum- and vanadium-rich analytical solution.

[0072] 3.3 Vanadium precipitation The molybdenum-vanadium-rich analytical solution was pumped into the reactor, the pH was adjusted to 8.5, and then an excess of 10% ammonium sulfate was added and reacted at 70°C for 1.5 hours to generate ammonium metavanadate precipitation; the chemical reaction involved is: 2NaVO3+(NH4)2SO4= 2NH4VO3↓+Na2SO4 The material after vanadium precipitation is filtered to obtain vanadium precipitation filtrate and ammonium metavanadate filter cake; the filter cake is washed to obtain 26.44 grams of ammonium metavanadate with a content of 98.5%, and the vanadium recovery rate reaches 92%.

[0073] 3.4 Molybdenum deposition The pH of the vanadium precipitation filtrate was adjusted to 3 with 30% sulfuric acid, and the reaction was carried out at 80°C for 40 minutes to obtain molybdic acid precipitate; the chemical reaction involved is: Na2MoO4+H2SO4 = H2MoO4↓+Na2SO4 The material after molybdenum precipitation was filtered to obtain molybdenum precipitation filtrate and molybdic acid filter cake; the molybdic acid filter cake was washed and dried to obtain 14.17 grams of molybdic acid, with a molybdenum content of 56.2% and a molybdenum recovery rate of 93%.

[0074] 3.5 Ion Exchange The molybdenum precipitation filtrate is enriched and concentrated and then returned to the resin adsorption process of step 3.2. The pH of the molybdenum precipitation filtrate is adjusted to 3.5-4.5 with sodium hydroxide solution, and then the above vanadium precipitation and molybdenum precipitation operations are repeated.

[0075] The main components of the liquid after ion exchange are ammonium sulfate and sodium sulfate, which enter the subsequent sulfate recovery process; the main component of the solution after analysis is molybdate, which returns to the molybdenum precipitation process.

[0076] 3.6 Sulfate Recovery In steps 3.2 and 3.5, the resin liquid after the anion exchange resin adsorbs molybdenum and vanadium is rich in sodium sulfate and ammonium sulfate. The resin liquid is concentrated, then cryogenically frozen and crystallized (freezing temperature is -5°C, time is 2h), and centrifuged to obtain sodium sulfate decahydrate and centrifuged mother liquor. Among them, sodium sulfate decahydrate is dried to obtain anhydrous sodium sulfate (sodium sulfate); the centrifuged mother liquor contains a large amount of ammonium sulfate, which can be reused in the vanadium precipitation process as an ammonium sulfate auxiliary material.

[0077] 4) Recovery of aluminum and nickel from alkaline leaching residue: 4.1 Sulfuric acid oxygen pressure acid leaching Weigh 180 g of alkaline leaching residue, add 1000 mL of 18% sulfuric acid, and introduce oxygen for pressurized leaching: the leaching temperature is 190 ° C, the oxygen is introduced at 0.6 MPa, the leaching time is 2 h, the stirring speed is 300 rpm, and the acid leaching residue precipitate and 800 mL of acid leaching solution are obtained by filtration; wherein, the concentrations of the main components in the acid leaching solution are: nickel 10.2 g / L, aluminum oxide 8.6 g / L.

[0078] The acid leaching residue is bauxite (sodium aluminum sulfate trihydrate), the main components of which are: sodium aluminum sulfate.

[0079] 4.2 Dealumination of acid leaching solution The acid leaching solution still contains 8.6 g / L of aluminum that needs to be removed. Sodium hydroxide solution is added to the acid leaching solution to adjust the pH to 5.5. The aluminum ions in the solution form aluminum hydroxide precipitation, which is incorporated into bauxite for subsequent treatment; the chemical reactions involved are: Al2(SO4)3 + 6NaOH = 2Al(OH)3↓+ 3Na2SO4 After the reaction is completed, suction filtration is performed to obtain aluminum hydroxide residue and filtrate after aluminum precipitation, and the filtrate after aluminum precipitation enters the nickel extraction process.

[0080] 4.3 Nickel deposition Sodium carbonate is added to the filtrate after aluminum precipitation to adjust the pH to 8.5 to form nickel carbonate precipitation. The main reactions that occur are: 10NiSO4+ 7Na2CO3+ 6NaOH+ H2O=7NiCO3·3Ni(OH)2·H2O +10Na2SO4 After the reaction is completed, suction filtration is performed to obtain a basic nickel carbonate product and a filtrate after nickel precipitation, and the filtrate after nickel precipitation enters a resin filtration process.

[0081] The basic nickel carbonate was washed with water and dried to obtain 11.07 grams, with a nickel content of 40% and a nickel recovery rate of 93%.

[0082] 4.4 Resin Filtration The filtrate after nickel precipitation still contains a small amount of nickel, which is adsorbed by nickel-type cation exchange resin D861. First, the pH of the filtrate after nickel precipitation is adjusted to 3.0-3.5, adsorbed by D861 resin, and analyzed by 4% sulfuric acid to obtain a nickel-containing analytical solution, which is reused in the oxygen pressure acid leaching process.

[0083] Comparative Example 1 Compared with Example 1, the difference lies in step 2), specifically:

[0084] Weigh 200g of ground roasted material, add 90g of soda ash and 700g of water, introduce an oxygen partial pressure of 0.55Mpa, heat to 180℃, stir at 250rpm, and leaching at high temperature for 1.5h; release the pressure when the temperature is below 60℃, take out the material and filter it to obtain 800mL of alkaline leaching solution (filtrate) (pH=11) and 150g of alkaline leaching residue.

[0085] The alkaline leaching residue was washed with water, and the analysis results of the leaching residue were: molybdenum 0.35%, vanadium pentoxide 0.78%, nickel 5.00%, and aluminum oxide 60.07%.

[0086] Comparative Example 2 Compared with Example 1, the difference lies in step 2), specifically: Weigh 200 g of ground roasted material, add 50 g of flake caustic soda, 700 g of water, introduce an oxygen partial pressure of 0.55 MPa, heat to 180 ° C, stir at 250 rpm, and leaching at high temperature for 1.5 h; release the pressure when the temperature is below 60 ° C, take out the material and filter it to obtain 800 mL of alkali leaching solution (filtrate) (pH = 11) and 150 g of alkali leaching residue.

[0087] The alkaline leaching residue was washed with water, and the analysis results of the leaching residue were: molybdenum 0.25%, vanadium pentoxide 0.78%, nickel 5.01%, and aluminum oxide 60.25%.

[0088] By comparing the alkali leaching effects of Example 1 and Comparative Examples 1-2, it can be seen that according to the different amounts of flake caustic soda and soda ash in Comparative Examples 1-2, both flake caustic soda and mixed caustic soda can achieve leaching effects. From the perspective of economy and effect, the use of mixed caustic soda in Example 1 can achieve a comprehensive effect, and the leaching effect for molybdenum and vanadium is better.

[0089] The above description is only a preferred specific implementation manner of the present invention, but the protection scope of the present invention is not limited thereto. Any technician familiar with the technical field can make equivalent replacements or changes according to the technical scheme and inventive concept of the present invention within the technical scope disclosed by the present invention, which should be covered by the protection scope of the present invention.

Claims

1. A method for recovering valuable metals from spent hydrogenation catalysts, characterized in that: The following steps are involved: 1) Oxidative roasting: The spent hydrogenation catalyst is subjected to oxidative roasting and crushed to obtain a roasting material; 2) Oxygen pressure alkali leaching: Place the roasted material in a pressure autoclave, add a mixed alkali consisting of sodium carbonate and sodium hydroxide, perform oxygen pressure alkali leaching, filter, and obtain alkali leaching solution and alkali leaching residue; 3) Recovery of aluminum, vanadium and molybdenum in alkaline leaching solution: 3.1) Aluminum precipitation 1: Add carbon dioxide to the alkaline leaching solution to adjust the pH of the system to 9.5-10.5, age, filter, and obtain pseudo-boehmite and filtrate; calcine the pseudo-boehmite to obtain activated alumina; 3.2) Impurity removal: adjust the pH of the filtrate after aluminum precipitation 1 to 10-11, add ammonium sulfate and magnesium sulfate to purify and remove impurities, stir the reaction, filter, and obtain a purified liquid after phosphorus, silicon and arsenic are removed; The pH of the purified liquid is adjusted to 3.5-4.5, the molybdenum and vanadium are adsorbed by anion exchange resin, and the liquid after the resin is collected; The resin adsorbing molybdenum and vanadium is analyzed to obtain an analysis solution rich in molybdenum and vanadium; 3.3) Vanadium precipitation: adjust the pH of the analytical solution to 7.5-8.5, add ammonium sulfate, stir to react, filter, and obtain ammonium metavanadate precipitate and vanadium precipitation filtrate; 3.4) Molybdenum precipitation: add sulfuric acid to the vanadium precipitation filtrate to adjust the pH value to 2.0-3.0, stir to react, filter, and obtain molybdic acid precipitate and molybdenum precipitation filtrate; 4) Recovery of aluminum and nickel from alkaline leaching residue: 4.1) Oxygen pressure acid leaching: Place the alkali leaching residue in a pressure autoclave for oxygen pressure acid leaching, add sulfuric acid to adjust the pH to 1.0-1.5, perform oxygen pressure acid leaching, filter, and obtain bauxite and acid leaching solution; 4.2) Aluminum precipitation 2: add alkali solution to the acid leaching solution to adjust the pH to 5, stir the reaction, filter, and obtain aluminum hydroxide and the filtrate after aluminum precipitation; 4.3) Nickel precipitation: Add sodium carbonate to the filtrate after aluminum precipitation, stir to react, filter, and obtain basic nickel carbonate precipitate.

2. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 1), when the oil content in the waste hydrogenation catalyst is ≥15%, it also includes performing thermal desorption pretreatment on the waste hydrogenation catalyst before oxidative roasting, specifically, heating the waste hydrogenation catalyst in a closed oxygen-proof environment to perform thermal desorption to remove oil; The oil content in the spent hydrogenation catalyst after thermal desorption treatment is controlled to be ≤8%.

3. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 1), the calcination temperature is 600-700°C and the calcination time is 2-5h; After the oxidation roasting, the roasted material is wet ball-milled and sieved.

4. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 2), the mass percentage of sodium carbonate and sodium hydroxide in the mixed alkali is 65-75:25-35; The conditions of oxygen pressure alkali leaching are: liquid-solid ratio 3-4:1, oxygen pressure 0.5-0.6MPa, temperature 170-190℃, leaching time 60-120min; the endpoint pH after oxygen pressure alkali leaching is controlled to be 11.5-12.

5.

5. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 3.1), the alkaline leaching solution is heated to 80-90° C., carbon dioxide is introduced into the solution to adjust the pH value of the solution to 9.5-10.5, and the solution is cooled and aged.

6. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 3.2), add ammonium sulfate and magnesium sulfate, and stir the mixture at 50-70°C for 20-40 minutes; Step 3.2) also includes recycling and reusing the post-resin liquid, which is to concentrate the post-resin liquid, freeze crystallize it, and centrifuge it to obtain sodium sulfate crystals and a centrifuge mother liquor rich in ammonium sulfate; wherein the centrifuge mother liquor is reused in the vanadium precipitation process of step 3.3).

7. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 3.3), the temperature of vanadium precipitation is 50-70°C.

8. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 3.4), the stirring reaction temperature is 70-80°C and the reaction time is 40-80min; Step 3.4) also includes enriching or concentrating the molybdenum precipitation filtrate and repeating the resin adsorption, vanadium precipitation and molybdenum precipitation operations in the above steps 3.2) to 3.4).

9. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 4.1), the conditions of oxygen pressure acid leaching are: liquid-solid ratio 4-6:1, oxygen pressure 0.4-0.6MPa, temperature 160-190°C, and leaching time 2-2.5h.

10. The method for recovering valuable metals from spent hydrogenation catalyst according to claim 1, characterized in that: In step 4.3), sodium carbonate is added to the filtrate after precipitation of aluminum 2 to control the pH at the end point of the reaction to be 8.0-8.5; Step 4.3) also includes adjusting the pH of the filtrate after nickel precipitation to 3.0-3.5, adsorbing nickel through a cation exchange resin, analyzing to obtain a nickel-containing analytical solution, and returning the solution to the oxygen pressure acid leaching process for treatment.

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