A method for recovering valuable metals from spent petroleum hydroprocessing catalysts

Abstract

This invention discloses a method for recovering valuable metals from waste petroleum hydrogenation catalysts, belonging to the field of hazardous waste resource utilization technology. The method includes: Step (A): mixing the waste hydrogenation catalyst with a collector, a slagging agent, and a reducing agent, and performing reduction smelting at 1100–1300°C to obtain smelting slag and an alloy; Step (B): crushing and grinding the alloy obtained in Step (A), and then performing pressurized alkaline leaching in the presence of an oxidizing gas to obtain an alkaline leaching solution and alkaline leaching residue; Step (C): performing acid leaching on the alkaline leaching residue obtained in Step (B) in the presence of a reducing agent to obtain an acid leaching solution; Step (D): neutralizing and hydrolyzing the acid leaching solution obtained in Step (C) to obtain a hydrolysis tail liquid and precipitates containing bismuth oxychloride and antimony oxychloride. This invention achieves reduced smelting temperature, easier alloy processing, efficient recovery of valuable metals, and zero pollutant emissions through a combined process of pyrometallurgical reduction smelting, hydrolytic stepwise leaching, and collector recycling.

Description

Technical Field

[0001] This invention relates to the field of hazardous waste resource utilization technology, specifically a method for recovering valuable metals from waste petroleum hydrogenation catalysts. Background Technology

[0002] Catalysts are core materials in the petroleum refining and chemical industries, with the global petroleum processing industry consuming millions of tons of catalysts annually. As catalysts deactivate due to poisoning, carbon buildup, sintering of active components, or loss, a large amount of waste petroleum hydrogenation catalysts, classified as hazardous waste, is generated. These waste catalysts contain organic matter and heavy metals, which can cause serious environmental pollution if not properly disposed of. Furthermore, the grades of valuable metals such as molybdenum (Mo), tungsten (W), vanadium (V), nickel (Ni), and cobalt (Co) are far higher than those in primary ore, making them highly valuable secondary resources of non-ferrous metals. Therefore, the efficient recovery of valuable metals from waste petroleum hydrogenation catalysts can achieve resource regeneration and reduce environmental burden, representing a key direction for practicing green and sustainable development.

[0003] Currently, the technologies for recovering valuable metals from waste petroleum hydrogenation catalysts, both domestically and internationally, mainly fall into two categories: hydrometallurgical processes and pyrometallurgical processes.

[0004] Wet processes: Typical technologies include pressure acid leaching, pressure alkaline leaching, oxidative roasting-alkaline leaching, sodium roasting-water leaching, and reducing ammonia leaching. These processes generally suffer from low metal recovery rates due to the encapsulation of valuable metals by the alumina support in the spent catalyst. Furthermore, wet processes consume large amounts of chemical reagents, generate large volumes of wastewater, and incur high subsequent wastewater treatment costs.

[0005] Pyrometallurgical processes, exemplified by iron capture smelting, concentrate valuable metals in ferroalloys through high-temperature smelting, followed by hydrometallurgical separation. For instance, the method disclosed in Chinese patent CN106282570A requires smelting at temperatures of 1550~1800℃, demanding sophisticated equipment and consuming significant energy. Furthermore, molybdenum and vanadium easily disperse into the leaching solution during subsequent acid leaching, resulting in low recovery rates. While Chinese patent CN113234930A lowers the smelting temperature to 1300~1500℃, it still requires secondary melting, water quenching, powdering, and oxidative roasting of the alloy. This process is lengthy, energy-intensive, and during pressurized oxidative acid leaching, some iron enters the leaching solution, interfering with the separation and purification of nickel and cobalt.

[0006] In summary, existing technologies suffer from drawbacks such as high smelting temperatures, high energy consumption, long process flow, low recovery rate of valuable metals, and large amounts of reagent consumption and wastewater generation. There is an urgent need to develop a low-energy-consumption, short-process, high-recovery-rate, and environmentally friendly recycling method.

[0007] Therefore, those skilled in the art have provided a method for recovering valuable metals from waste petroleum hydrogenation catalysts to solve the problems mentioned in the background art. Summary of the Invention

[0008] The purpose of this invention is to provide a method for recovering valuable metals from waste petroleum hydrogenation catalysts. By using a combined process of "pyrometallurgical reduction smelting - wet stepwise leaching - collector recycling", the smelting temperature is reduced, the alloy is easy to handle, the valuable metals are efficiently recovered, and the pollutants are zero-emission, thereby solving the problems mentioned in the background art.

[0009] To achieve the above objectives, the present invention provides the following technical solution:

[0010] A method for recovering valuable metals from waste petroleum hydrogenation catalysts includes the following steps:

[0011] Step (A): The waste hydrogenation catalyst is mixed with a collector, a slag-forming agent and a reducing agent, and then reduced and smelted at 1100-1300℃ to obtain smelting slag and alloy;

[0012] Step (B): After crushing and grinding the alloy obtained in step (A) into powder, it is subjected to pressure alkaline leaching in the presence of oxidizing gas to obtain alkaline leaching solution and alkaline leaching residue;

[0013] Step (C): The alkaline leaching residue obtained in step (B) is subjected to acid leaching in the presence of a reducing agent to obtain an acid leaching solution;

[0014] Step (D): The acid leaching solution obtained in step (C) is neutralized and hydrolyzed to obtain hydrolysis tail liquid and precipitates containing bismuth oxychloride and antimony oxychloride; the hydrolysis tail liquid is extracted, separated and evaporated to crystallize to obtain cobalt products and nickel products; the precipitates are returned to step (A) as a collector for recycling.

[0015] As a further aspect of the present invention: in step (A), the trapping agent includes an antimony component and a bismuth component; the antimony component is selected from one or more of antimony powder, antimony trioxide, antimony pentoxide, antimony trichloride, antimony pentachloride, antimony oxychloride, stibnite, and antimony concentrate; the bismuth component is selected from one or more of bismuth powder, bismuth trioxide, bismuth trichloride, bismuth oxychloride, bismuthite, and bismuthus ore.

[0016] As a further aspect of the present invention: in step (A), the mass ratio of antimony component to bismuth component in the trapping agent is 1:(0.5-2).

[0017] As a further aspect of the present invention: in step (A), the slag-forming agent is selected from one or more of calcium oxide, sodium carbonate, borax, silicon dioxide, calcium fluoride, sodium fluoride, and sodium silicate; the reducing agent is selected from one or more of coke, graphite powder, coal powder, flour, and sawdust.

[0018] As a further embodiment of the present invention: in step (A), the mass ratio of waste hydrogenation catalyst, scavenging agent, slag-forming agent and reducing agent is 100:(20-60):(50-200):(5-20); the reduction smelting time is 2-6 hours.

[0019] As a further aspect of the present invention: in step (B), the oxidizing gas is oxygen or ozone; the reagent used in the pressurized alkaline leaching is an alkaline reagent, which is selected from one or more of sodium carbonate, sodium hydroxide, potassium carbonate, potassium hydroxide, sodium bicarbonate, and potassium bicarbonate.

[0020] As a further aspect of the present invention: in step (B), the amount of alkaline reagent used is 1 to 4 times the sum of the molar amounts of molybdenum, vanadium, tungsten and antimony in the alloy.

[0021] As a further embodiment of the present invention: in step (B), the conditions for pressurized alkaline leaching are: leaching temperature 140-210℃, leaching time 1-5 hours, oxidizing gas partial pressure 0.5-3MPa, and liquid-solid mass ratio (5-20):1.

[0022] As a further embodiment of the present invention: in step (C), the reducing agent is selected from one or more of sulfur dioxide, sodium sulfite, sodium metabisulfite, oxalic acid, and potassium iodide; the acid reagent used in the acid leaching is hydrochloric acid; and the amount of the reducing agent is 1 to 1.5 times the molar amount of antimony in the alkaline leaching residue.

[0023] As a further embodiment of the present invention: the acid leaching conditions in step (C) are: leaching temperature 50-80°C, leaching time 1-3 hours, and liquid-solid mass ratio (4-10):1.

[0024] The neutralizing agent used in step (D) for the neutralization hydrolysis is sodium hydroxide, and the pH value at the end of the hydrolysis is 1.5 to 2.5.

[0025] Compared with the prior art, the beneficial effects of the present invention are:

[0026] 1. Low melting temperature and low energy consumption: By synergistically capturing antimony and bismuth to form a low-melting-point alloy phase, and combining it with a low-melting-point slag, the melting temperature is reduced to a level far below that of existing pyrometallurgical processes, significantly reducing energy consumption and extending the service life of the melting equipment.

[0027] 2. Short process flow and simple operation: Antimony and bismuth alloys are highly brittle and can be directly crushed and ground into powder without secondary melting, eliminating the steps of "secondary melting-water quenching-calcination" in the existing process and shortening the process flow.

[0028] 3. High recovery rate of valuable metals: Through optimization of slag type ratio and step-by-step leaching process, the capture rate (smelting stage) and total recovery rate (final product) of valuable metals are improved, which is significantly better than existing hydrometallurgical and pyrometallurgical processes.

[0029] 4. Green and environmentally friendly, and recyclable: Smelting slag is a harmless solid waste that can be used as building material; the bismuth oxychloride and antimony oxychloride produced by the hydrolysis of acid leaching solution can be recycled as a catching agent, with no secondary waste generated, realizing a closed loop of "resource-product-recycled resources" and reducing the amount of waste emissions. Attached Figure Description

[0030] Figure 1 This is a flowchart of a method for recovering valuable metals from waste petroleum hydrogenation catalysts. Detailed Implementation

[0031] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0032] As mentioned in the background section of this application, research has found that existing technologies suffer from technical defects such as high smelting temperature, high energy consumption, long process flow, low recovery rate of valuable metals, and large amounts of reagent consumption and wastewater generation. There is an urgent need to develop a low-energy-consumption, short-process, high-recovery-rate, and environmentally friendly recycling method.

[0033] To address the aforementioned deficiencies, this application discloses a method for recovering valuable metals from waste petroleum hydrogenation catalysts. Through a combined process of "pyrometallurgical reduction smelting - wet stepwise leaching - collector recycling", the method achieves reduced smelting temperature, easier alloy processing, efficient recovery of valuable metals, and zero emissions of pollutants.

[0034] The following will describe in detail, with reference to the accompanying drawings, how the solution of this application solves the above-mentioned technical problems.

[0035] Please see Figure 1 In this embodiment of the invention, a method for recovering valuable metals from waste petroleum hydrogenation catalysts includes the following steps:

[0036] Step (A): Reduction smelting;

[0037] Waste hydrogenation catalyst is mixed uniformly with a scavenging agent, a slagging agent, and a reducing agent in a specific ratio, and then subjected to reduction smelting at 1100–1300℃. After the reaction is completed, the mixture is cooled and separated to obtain smelting slag and an alloy; wherein:

[0038] The collector comprises an antimony component and a bismuth component; the antimony component is selected from one or more of antimony powder, antimony trioxide, antimony pentoxide, antimony trichloride, antimony pentachloride, antimony oxychloride, stibnite, and antimony concentrate; the bismuth component is selected from one or more of bismuth powder, bismuth trioxide, bismuth trichloride, bismuth oxychloride, bismuthite, and bismuthus ore; the mass ratio of the antimony component to the bismuth component is 1:(0.5-2);

[0039] Slag-forming agent: selected from one or more of calcium oxide, sodium carbonate, borax, silicon dioxide, calcium fluoride, sodium fluoride, and sodium silicate;

[0040] Reducing agent: selected from one or more of coke, graphite powder, coal powder, flour, and sawdust;

[0041] Material mass ratio: waste hydrogenation catalyst: collector: slag-forming agent: reducing agent = 100:(20~60):(50~200):(5~20);

[0042] Melting time: 2 to 6 hours.

[0043] In this step, antimony and bismuth, due to their low melting point characteristics, form low-melting-point (900-1100℃) alloy phases with W, Mo, V, Ni, and Co in the waste catalyst. Impurities such as alumina and silicon form low-melting-point (1000-1200℃) smelting slag with the slag-forming agent, achieving efficient separation of valuable metals and impurities. The smelting slag is a harmless solid waste that can be used as a raw material for building materials.

[0044] Step (B): Pressure oxidation and alkali leaching;

[0045] The alloy obtained in step (A) is crushed and ground into powder (preferably with a particle size of 100-200 mesh). Under pressure, it undergoes alkaline leaching in the presence of an oxidizing gas. After the reaction, solid and liquid are separated to obtain an alkaline leaching solution and an alkaline leaching residue.

[0046] Oxidizing gases: oxygen or ozone;

[0047] Alkaline reagent: selected from one or more of sodium carbonate, sodium hydroxide, potassium carbonate, potassium hydroxide, sodium bicarbonate, and potassium bicarbonate; the amount used is 1 to 4 times the sum of the molar amounts of molybdenum, vanadium, tungsten, and antimony in the alloy;

[0048] Leaching conditions: leaching temperature 140~210℃, leaching time 1~5 hours, oxidizing gas partial pressure 0.5~3MPa, liquid-solid mass ratio (5~20):1.

[0049] In this step, W, Mo, and V in the alloy powder react with alkaline reagents under pressure oxidation to form soluble salts that enter the alkaline leaching solution. Subsequently, vanadium, tungsten, and molybdenum products can be obtained through ammonium salt precipitation-extraction separation. Antimony, bismuth, nickel, and cobalt remain in the alkaline leaching residue in the form of sodium antimonate pyroantimonate, bismuth oxide, nickel oxide, and cobalt oxide.

[0050] Step (C): Reduction pickling;

[0051] The alkaline leaching residue obtained in step (B) is added to hydrochloric acid and acid leaching is carried out in the presence of a reducing agent. After the reaction is completed, solid and liquid are separated to obtain the acid leaching solution; wherein:

[0052] Reducing agent: selected from one or more of sulfur dioxide, sodium sulfite, sodium metabisulfite, oxalic acid, and potassium iodide; the amount used is 1 to 1.5 times the molar amount of antimony in the alkaline leaching residue;

[0053] Acid leaching conditions: leaching temperature 50-80℃, leaching time 1-3 hours, liquid-solid mass ratio (4-10):1.

[0054] In this step, the reducing agent reduces the high-valence antimony, bismuth, nickel, and cobalt in the alkaline leaching residue into soluble ions, allowing them to fully dissolve and enter the acid leaching solution, thus achieving deep extraction of valuable metals.

[0055] Step (D): Neutralization, hydrolysis, and product separation;

[0056] The acid leaching solution obtained in step (C) was adjusted to pH 1.5–2.5 using a neutralizing reagent and subjected to neutralization and hydrolysis at room temperature. After the reaction was completed, solid-liquid separation was performed to obtain the hydrolysis tail liquid and a precipitate containing bismuth oxychloride and antimony oxychloride; wherein:

[0057] Neutralizing agent: sodium hydroxide;

[0058] Subsequent processing: The precipitate is returned to step (A) as a collector for recycling; the hydrolysis tail liquid is extracted and separated (using P204 or P507 extractant), and then evaporated and crystallized to obtain cobalt and nickel products respectively.

[0059] To further illustrate the present invention, the following describes in detail, with reference to embodiments, a method for recovering valuable components from petrochemical waste.

[0060] Example 1:

[0061] Step (A1): Reduction smelting;

[0062] 1000g of spent hydrogenation catalyst (main components: Mo 5.2%, W 3.8%, V 2.1%, Ni 1.5%, Co 0.8%, Al2O3 65%) was mixed evenly with 130g of antimony oxychloride (antimony component), 70g of bismuth oxychloride (bismuth component, antimony component to bismuth component mass ratio 1:0.54), slagging agent (150g calcium oxide, 100g sodium carbonate, 200g borax, 50g calcium fluoride, 100g sodium silicate), and 200g flour (reducing agent). The mixture was placed in a vertical melting furnace and melted at 1120℃ for 2 hours. After cooling, the slag and alloy were separated. The collection rates of molybdenum, vanadium, tungsten, nickel, cobalt, antimony, and bismuth were 98.85%, 98.23%, 98.07%, 99.01%, 98.72%, 99.23%, and 99.38%, respectively.

[0063] Step (B1): Pressure oxidation and alkali leaching;

[0064] The alloy obtained in step (A1) was crushed and ground into 150-mesh powder, added to a high-pressure reactor, and ozone (partial pressure 3 MPa) was introduced. Sodium carbonate (the amount used was 4 times the sum of the molar amounts of Mo, V, W, and Sb in the alloy) and deionized water (liquid-solid ratio 20:1) were added, and the mixture was stirred and reacted at 200°C for 1 hour. The solid and liquid were separated to obtain an alkaline leaching solution and an alkaline leaching residue. The alkaline leaching solution was separated by ammonium chloride precipitation and P350 extraction to obtain molybdenum product (purity 99.5%), vanadium product (purity 99.3%), and tungsten product (purity 99.6%), with total recovery rates of 96.65%, 97.02%, and 96.11%, respectively.

[0065] Step (C1): Reduction pickling;

[0066] Add the alkaline leaching residue obtained in step (B1) to hydrochloric acid (mass fraction 30%), add sodium sulfite (the amount is 1.5 times the molar amount of Sb in the alkaline leaching residue), control the liquid-solid ratio to 4:1, stir and leach at 50°C for 1 hour, and separate the solid and liquid to obtain acid leaching solution.

[0067] Step (D1): Neutralization, hydrolysis, and product separation;

[0068] The acid leaching solution obtained in step (C1) was slowly added to sodium hydroxide solution at room temperature to adjust the pH value to 2.5. After stirring for 30 minutes, solid-liquid separation was performed to obtain precipitates containing bismuth oxychloride and antimony oxychloride (to be recycled back to step A1) and hydrolysis tail liquid. The hydrolysis tail liquid was extracted and separated by P204, evaporated and crystallized to obtain nickel product (purity 99.2%) and cobalt product (purity 99.4%), with total recovery rates of 96.79% and 96.24%, respectively. The recycling recovery rates of antimony and bismuth were 98.01% and 98.34%, respectively.

[0069] Example 2:

[0070] Step (A2): Reduction smelting;

[0071] 1000g of waste hydrogenation catalyst was mixed evenly with 70g of antimony concentrate (antimony component), 130g of bismuth trichloride (bismuth component, antimony component to bismuth component mass ratio 1:1.86), slagging agent (600g sodium carbonate, 400g borax, 200g calcium fluoride, 800g sodium silicate), and 50g of coal powder (reducing agent). The mixture was then smelted at 1300℃ for 6 hours, and the smelting slag and alloy were separated. The collection rates of molybdenum, vanadium, tungsten, nickel, cobalt, antimony, and bismuth were 99.05%, 98.93%, 98.52%, 98.56%, 98.42%, 99.09%, and 99.21%, respectively.

[0072] Step (B2): Pressure oxidation and alkali leaching;

[0073] The alloy was crushed and ground to 100 mesh, added to a high-pressure reactor, and oxygen (partial pressure 0.5 MPa) was introduced. Potassium hydroxide (the amount used was 1 times the sum of the molar amounts of Mo, V, W, and Sb in the alloy) and deionized water (liquid-solid ratio 5:1) were added, and the reaction was carried out at 150°C for 5 hours. The alkaline leaching solution and alkaline leaching residue were separated. The alkaline leaching solution was separated by ammonium nitrate precipitation and N235 extraction to obtain molybdenum product (purity 99.4%), vanadium product (purity 99.2%), and tungsten product (purity 99.5%), with total recovery rates of 97.21%, 97.34%, and 96.52%, respectively.

[0074] Step (C2): Reduction acid leaching;

[0075] Hydrochloric acid (30% by mass) and oxalic acid (1 times the molar amount of Sb in the alkali leaching residue) were added to the alkali leaching residue. The liquid-solid ratio was controlled at 10:1. The residue was leached at 80℃ for 3 hours, and the acid leaching solution was obtained by separation.

[0076] Step (D2): Neutralization, hydrolysis, and product separation;

[0077] The pH of the acid leaching solution was adjusted to 1.5 with sodium hydroxide, and after hydrolysis at room temperature, the precipitate (returned to the recycling system) and the hydrolysis tail liquid were separated. The tail liquid was extracted with P507 and evaporated and crystallized to obtain nickel products (purity 99.1%) and cobalt products (purity 99.3%), with total recovery rates of 96.54% and 96.38%, respectively. The recycling recovery rates of antimony and bismuth were 98.22% and 98.25%, respectively.

[0078] Example 3:

[0079] Step (A3): Reduction smelting;

[0080] 1000g of waste hydrogenation catalyst was mixed evenly with 100g of antimony powder (antimony component), 100g of bismuth powder (bismuth component, mass ratio 1:1), slagging agent (400g calcium oxide, 200g sodium carbonate, 100g borax, 100g calcium fluoride, 300g silicon dioxide, 100g sodium fluoride), and 150g coke (reducing agent). The mixture was then smelted at 1200℃ for 4 hours, and the smelting slag and alloy were separated. The collection rates of molybdenum, vanadium, tungsten, nickel, cobalt, antimony, and bismuth were 98.57%, 98.84%, 98.70%, 98.99%, 98.65%, 99.28%, and 99.13%, respectively.

[0081] Step (B3): Pressure oxidation and alkali leaching;

[0082] The alloy was crushed and ground to 200 mesh, then added to a high-pressure reactor. Ozone (partial pressure 2 MPa) was introduced, along with sodium hydroxide (2.5 times the sum of the molar amounts of Mo, V, W, and Sb in the alloy) and deionized water (liquid-solid ratio 13:1). The reaction was carried out at 175°C for 3 hours, and the alkaline leaching solution and alkaline leaching residue were separated. The alkaline leaching solution was then separated by ammonium sulfate precipitation and TBP extraction to obtain molybdenum (purity 99.3%), vanadium (purity 99.4%), and tungsten (purity 99.6%) products, with total recoveries of 96.74%, 96.46%, and 96.29%, respectively.

[0083] Step (C3): Reduction acid leaching;

[0084] Hydrochloric acid (30% by mass) and potassium iodide (1.3 times the molar amount of Sb in the alkali leaching residue) were added to the alkali leaching residue. The liquid-solid ratio was controlled at 7:1, and the residue was leached at 65°C for 2 hours to obtain the acid leaching solution.

[0085] Step (D3): Neutralization, hydrolysis, and product separation;

[0086] The pH of the acid leaching solution was adjusted to 2.0 with sodium hydroxide. After hydrolysis at room temperature, the precipitate (returned to the recycling system) and the hydrolysis tail liquid were separated. The tail liquid was extracted with P204 and evaporated for crystallization to obtain nickel products (purity 99.2%) and cobalt products (purity 99.5%), with total recovery rates of 96.28% and 96.07%, respectively. The recycling recovery rates of antimony and bismuth were 98.33% and 98.15%, respectively.

[0087] The above embodiments demonstrate that the method of the present invention can stably achieve efficient recovery of valuable metals from waste petroleum hydrogenation catalysts, and the process is green and environmentally friendly with low energy consumption, showing promise for industrial application.

[0088] This invention utilizes the synergistic capture of antimony and bismuth to form a low-melting-point alloy phase, coupled with a low-melting-point slag mold, to lower the smelting temperature significantly below that of existing pyrometallurgical processes, thereby greatly reducing energy consumption and extending the service life of smelting equipment. Furthermore, the antimony-bismuth alloy is highly brittle, allowing for direct crushing and grinding into powder without secondary melting, eliminating the "secondary melting-water quenching-calcination" steps in existing processes and shortening the process flow. In addition, through optimized slag mold ratios and a stepwise leaching process, the capture rate of valuable metals (in the smelting stage) and the total recovery rate (in the final product) are significantly improved, outperforming existing hydrometallurgical and pyrometallurgical processes. Simultaneously, the smelting slag is a harmless solid waste that can be used as building material; the bismuth oxychloride and antimony oxychloride produced by the hydrolysis of the acid leaching solution can be recycled as capture agents, generating no secondary waste and achieving a closed loop of "resource-product-recycled resource," thus reducing the discharge of waste gas, wastewater, and solid waste.

[0089] Obviously, those skilled in the art can make various modifications and variations to this invention without departing from its spirit and scope. Therefore, if these modifications and variations fall within the scope of the claims of this invention and their equivalents, this invention also intends to include these modifications and variations.

[0090] The above description is merely a preferred embodiment of the present invention, but the scope of protection of the present invention is not limited thereto. Any equivalent substitutions or modifications made by those skilled in the art within the scope of the technology disclosed in the present invention, based on the technical solution and inventive concept of the present invention, should be covered within the scope of protection of the present invention.

Claims

1. A method for recovering valuable metals from waste petroleum hydrogenation catalysts, characterized in that, Includes the following steps: Step (A): The waste hydrogenation catalyst is mixed with a collector, a slag-forming agent and a reducing agent, and then reduced and smelted at 1100-1300℃ to obtain smelting slag and alloy; Step (B): After crushing and grinding the alloy obtained in step (A) into powder, it is subjected to pressure alkaline leaching in the presence of oxidizing gas to obtain alkaline leaching solution and alkaline leaching residue; Step (C): The alkaline leaching residue obtained in step (B) is subjected to acid leaching in the presence of a reducing agent to obtain an acid leaching solution; Step (D): The acid leaching solution obtained in step (C) is neutralized and hydrolyzed to obtain hydrolysis tail liquid and precipitates containing bismuth oxychloride and antimony oxychloride; the hydrolysis tail liquid is extracted, separated and evaporated to crystallize to obtain cobalt products and nickel products; the precipitates are returned to step (A) as a collector for recycling.

2. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 1, characterized in that, In step (A), the collector includes an antimony component and a bismuth component; the antimony component is selected from one or more of antimony powder, antimony trioxide, antimony pentoxide, antimony trichloride, antimony pentachloride, antimony oxychloride, stibnite, and antimony concentrate; the bismuth component is selected from one or more of bismuth powder, bismuth trioxide, bismuth trichloride, bismuth oxychloride, bismuthite, and bismuthus ore.

3. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 2, characterized in that, In step (A), the mass ratio of antimony component to bismuth component in the trapping agent is 1:(0.5-2).

4. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 1, characterized in that, In step (A), the slag-forming agent is selected from one or more of calcium oxide, sodium carbonate, borax, silicon dioxide, calcium fluoride, sodium fluoride, and sodium silicate; the reducing agent is selected from one or more of coke, graphite powder, coal powder, flour, and sawdust.

5. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 1, characterized in that, In step (A), the mass ratio of waste hydrogenation catalyst, scavenging agent, slag-forming agent, and reducing agent is 100:(20-60):(50-200):(5-20); the reduction smelting time is 2-6 hours.

6. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 1, characterized in that, In step (B), the oxidizing gas is oxygen or ozone; the reagent used in the pressurized alkaline leaching is an alkaline reagent, which is selected from one or more of sodium carbonate, sodium hydroxide, potassium carbonate, potassium hydroxide, sodium bicarbonate, and potassium bicarbonate.

7. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 6, characterized in that, In step (B), the amount of alkaline reagent used is 1 to 4 times the sum of the molar amounts of molybdenum, vanadium, tungsten, and antimony in the alloy.

8. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 1, characterized in that, In step (B), the conditions for pressurized alkaline leaching are: leaching temperature 140-210℃, leaching time 1-5 hours, oxidizing gas partial pressure 0.5-3MPa, and liquid-solid mass ratio (5-20):

1.

9. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 1, characterized in that, In step (C), the reducing agent is selected from one or more of sulfur dioxide, sodium sulfite, sodium metabisulfite, oxalic acid, and potassium iodide; the acid reagent used in the acid leaching is hydrochloric acid; and the amount of the reducing agent is 1 to 1.5 times the molar amount of antimony in the alkaline leaching residue.

10. The method for recovering valuable metals from waste petroleum hydrogenation catalysts according to claim 1, characterized in that, The acid leaching conditions in step (C) are: leaching temperature 50-80℃, leaching time 1-3 hours, and liquid-solid mass ratio (4-10):

1. The neutralizing agent used in step (D) for the neutralization hydrolysis is sodium hydroxide, and the pH value at the end of the hydrolysis is 1.5 to 2.5.

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