Method for zirconium-hafnium deep separation and gradient recovery from hafnium-containing waste residue

Abstract

This invention discloses a method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste residue, belonging to the field of rare metal metallurgy and solid waste resource utilization technology. The method adopts a process route of "staged leaching pre-separation—deep purification—multi-component gradient recovery": firstly, the hafnium-containing waste residue is leached under normal pressure with dilute sulfuric acid to achieve preliminary gradient separation of zirconium and hafnium; the zirconium-rich leachate is extracted to remove impurities, precipitated, and calcined to obtain high-purity zirconium oxide; the hafnium-rich leaching residue is then leached under high acid pressure, and a P507-TOPO synergistic extraction system is used for deep separation of zirconium and hafnium to prepare high-purity hafnium oxide, while simultaneously using the siliceous tailings to prepare silica. This invention can significantly reduce acid consumption and processing costs, achieving a zirconium-hafnium separation coefficient ≥80, hafnium oxide purity ≥99.9%, and total zirconium and hafnium recovery rates of over 95% and 92%, respectively. It realizes efficient and clean utilization of hafnium-containing waste residue and full-component gradient recovery, and is suitable for the industrial treatment of solid waste from zirconium and hafnium smelting.

Description

Technical Field

[0001] This invention belongs to the field of rare metal metallurgy and solid waste resource utilization technology, specifically a method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag. Background Technology

[0002] Zirconium and hafnium are rare, high-melting-point metals with extremely similar properties, widely used in high-end fields such as nuclear industry, aerospace, and electronic ceramics. Zirconium and hafnium often coexist in nature, and large amounts of hafnium-containing waste are generated during zirconium ore processing and zirconium product manufacturing. This waste still contains certain amounts of valuable zirconium and hafnium metals, as well as impurities such as silicon, iron, and titanium. Direct stockpiling and disposal not only results in a serious waste of precious metal resources but also poses an environmental risk of heavy metal leakage.

[0003] Current hafnium-containing waste residue recovery processes mostly employ a full-volume acid leaching-solvent extraction route, which presents the following prominent problems: First, full-volume leaching requires high-concentration strong acids and high-temperature, high-pressure conditions, resulting in high acid and alkali consumption, high processing costs, and the simultaneous dissolution of a large number of impurities, increasing the load on subsequent separation processes; Second, the zirconium-hafnium separation coefficient of traditional single-extractant is relatively low, requiring dozens of extraction stages to achieve deep separation requirements, resulting in long processes, high energy consumption, and the hafnium product purity failing to meet the demands of high-end applications; Third, existing processes mainly focus on the recovery of zirconium and hafnium main metals, while associated components such as silicon and titanium in the waste residue are not utilized as resources, resulting in low comprehensive resource utilization and a still large amount of tailings emissions. Therefore, developing a hafnium-containing waste residue treatment process with high separation depth, excellent recovery rate, multi-component gradient recovery, and clean, low-energy consumption has significant industrial value and environmental significance. Summary of the Invention

[0004] The purpose of this invention is to provide a method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste residue, so as to solve the problems mentioned in the background art.

[0005] The objective of this invention can be achieved through the following technical solutions:

[0006] A method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag includes the following steps:

[0007] Step 1: Pre-treatment of hafnium-containing waste residue by graded leaching. After crushing and ball milling to a certain particle size, hafnium-containing waste residue is subjected to primary atmospheric pressure leaching with dilute sulfuric acid. After solid-liquid separation, zirconium-rich leachate and hafnium-rich leaching residue are obtained, realizing the initial gradient separation of zirconium and hafnium.

[0008] Step 2: Deep purification of zirconium-rich solution and recovery of zirconium products. The zirconium-rich leachate is subjected to extraction and impurity removal treatment to remove impurity metals such as iron, titanium, and aluminum. Then, it is subjected to zirconium-loaded organic back-extraction, oxalic acid precipitation, and high-temperature calcination to obtain high-purity zirconium oxide products.

[0009] Step 3: Enhanced leaching and deep separation and recovery of hafnium-rich slag. The hafnium-rich leaching slag is subjected to pressure leaching with high-concentration sulfuric acid to fully dissolve hafnium and residual zirconium, resulting in a hafnium-rich leachate. A synergistic extraction system is used to perform deep separation of zirconium and hafnium in the hafnium-rich leachate. After back-extraction, precipitation, and calcination, high-purity hafnium oxide is obtained. Simultaneously, silicon resources are recovered from the leaching tailings to achieve multi-component gradient recovery.

[0010] Furthermore, the specific steps of the graded leaching in step one are as follows:

[0011] After crushing the hafnium-containing waste residue, ball mill it to 150-200 mesh. Add a dilute sulfuric acid solution with a concentration of 1.0-1.5 mol / L at a liquid-to-solid ratio of 4-6:1 mL / g. Stir and leach for 1.5-2.5 h at 60-70℃ and 200-250 r / min. After the reaction is complete, vacuum filter and collect the filtrate, which is the zirconium-rich leachate, and the filter residue, which is the hafnium-rich leach residue.

[0012] Furthermore, the zirconium dioxide content in the hafnium-containing waste residue is 40%-55%, the hafnium dioxide content is 8%-15%, and the remainder is silicon, iron, titanium, and aluminum oxide impurities; the leaching rate of zirconium during the dilute sulfuric acid leaching process is 65%-75%, and the leaching rate of hafnium is 15%-25%.

[0013] Furthermore, the specific steps for the deep purification of the zirconium-rich solution in step two are as follows:

[0014] Di(2-ethylhexyl)phosphoric acid (P204) was used as the extractant, and sulfonated kerosene was used as the diluent. The volume fraction of the extractant was 20%-30%. The pH of the zirconium-rich leachate was adjusted to 0.8-1.2, and 3-5 stages of countercurrent extraction were performed to remove impurities at a ratio O / A = 1:2-3. The zirconium-loaded organic phase was back-extracted in 2.0-3.0 mol / L sulfuric acid solution at a ratio O / A = 3-4:1 in 2-3 stages to obtain a pure zirconium solution. Oxalic acid solution was added to the pure zirconium solution to adjust the pH to 1.5-2.0, and the reaction was stirred at 80-90℃ for 1-2 h. After filtration, the filter cake was calcined at 900-1000℃ for 2-3 h to obtain high-purity zirconium oxide.

[0015] Furthermore, the specific steps of the pressure leaching in step three are as follows:

[0016] Add hafnium-rich leaching residue to a sulfuric acid solution with a concentration of 4.0-6.0 mol / L at a liquid-to-solid ratio of 3-5:1 mL / g. After stirring evenly, transfer the solution to a high-pressure reactor and react at 150-170℃ and 0.4-0.6 MPa for 2-4 hours. After the reaction is completed, cool to room temperature and filter to obtain hafnium-rich leaching solution and siliceous tailings.

[0017] Furthermore, the leaching rate of hafnium during the pressurized leaching process is ≥92%, and the total leaching rate of zirconium is ≥95%; the silica tailings are washed with water, dissolved in alkali, and precipitated to prepare silica by-products, thereby realizing the recovery of silicon resources.

[0018] Furthermore, the specific steps of the synergistic extraction and depth separation in step three are as follows:

[0019] The synergistic extraction system is a mixture of 2-ethylhexyl di(2-ethylhexyl)phosphonic acid mono-2-ethylhexyl ester (P507) and trioctylphosphine oxide (TOPO), with sulfonated kerosene as the diluent. The volume fraction of P507 is 25%-35%, and the volume fraction of TOPO is 5%-10%. The acidity of the hafnium-rich leachate is adjusted to 1.0-1.5 mol / L H+, and 6-8 stages of countercurrent extraction are performed at a ratio O / A = 1:1-2. Zirconium preferentially enters the organic phase, while hafnium remains in the aqueous phase. The zirconium-loaded organic phase is back-extracted with 3.0-4.0 mol / L sulfuric acid and returned to the zirconium-rich solution purification step in step two. The hafnium-containing extract residue is adjusted to pH 2.0-2.5 with ammonia, and oxalic acid is added to precipitate hafnium. The filter cake is calcined at 950-1100℃ for 2-3 hours to obtain high-purity hafnium oxide.

[0020] Furthermore, the zirconium-hafnium separation coefficient of the synergistic extraction system is ≥80, which is much higher than that of the single extractant system; the purity of the obtained hafnium oxide product is ≥99.9%, and the purity of the zirconium oxide product is ≥99.5%.

[0021] Furthermore, the method also includes an extractant recycling step: the blank organic phase after back-extraction is saponified with sodium carbonate solution and then returned to the extraction process for recycling, with an extractant loss rate of ≤3% / batch.

[0022] The beneficial effects of this invention are:

[0023] This invention addresses the industry pain points of complex hafnium-containing waste residue composition, high difficulty in separating zirconium and hafnium, and low comprehensive resource utilization rate. It constructs a complete process system of "graded leaching pre-separation - fractional deep purification - multi-component gradient recovery", achieving multiple technological breakthroughs in separation efficiency, product purity, resource utilization rate and clean production, and has significant innovation and application value. In terms of separation logic, this invention pioneers a gradient progressive separation strategy, breaking through the traditional extensive mode of full leaching-full liquid separation: The first step utilizes the difference in leaching behavior of zirconium and hafnium in dilute acid to achieve preliminary separation at atmospheric pressure, allowing about 70% of zirconium to enter the liquid phase first, while most of the hafnium remains in the slag phase, significantly reducing the processing volume of subsequent deep separation; the second step performs simple impurity removal and purification on the zirconium-rich liquid to produce qualified zirconium products; the third step only performs enhanced leaching and deep extraction separation on the hafnium-rich slag. Through the tiered configuration of "easy first, difficult later, and quality-separated processing", the high-cost pressure leaching and deep extraction processes only process about 30% of the material volume, and the overall acid consumption and energy consumption are reduced by more than 35% compared with the traditional process, significantly reducing production and operating costs.

[0024] This invention develops a synergistic extraction system that utilizes the coordination synergistic effect of two extractants to increase the zirconium-hafnium separation coefficient to over 80, more than twice that of the single P507 system. Only 6-8 stages of countercurrent extraction are required to achieve nuclear-grade purity. This significantly reduces the number of extraction stages and dramatically improves separation efficiency, completely solving the technical bottlenecks of traditional processes such as numerous separation stages, large footprint, and large fluctuations in product purity. It can stably produce high-purity hafnium oxide products with a purity of over 99.9%. In terms of resource utilization, this invention achieves full-gradient recovery of the main metal and associated components: the total recovery rate of zirconium and hafnium main metals reaches over 95% and 92% respectively, far exceeding the industry average. Simultaneously, the silicon component in the leaching tailings is converted into silica byproducts, achieving a tailings reduction rate of over 80%. This truly realizes the reduction and high-value utilization of solid waste, avoiding the environmental risks of large-scale tailings stockpiling. In terms of clean production, this invention achieves closed-loop recycling and regeneration of the extractant, with an extractant loss rate of less than 3% per batch. Both the back-extraction acid and washing water can be reused in stages, and the entire process generates no high-concentration waste liquid discharge, aligning with the development direction of green metallurgy. In summary, this invention features a reasonable process route, strong technological innovation, and outstanding economic and environmental benefits. It provides a novel technological path for the efficient resource utilization of hafnium-containing waste residue and possesses extremely high prospects for industrialization and promotion. Detailed Implementation

[0025] The technical solutions of the present invention will be clearly and completely described below with reference to the embodiments of the present invention. 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 of ordinary skill in the art without creative effort are within the scope of protection of the present invention.

[0026] Example 1: A method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste residue, comprising the following steps:

[0027] S1: After crushing the hafnium-containing waste residue, ball mill it to 150 mesh, add a 1.0 mol / L dilute sulfuric acid solution at a liquid-to-solid ratio of 4:1 mL / g, stir and leach for 1.5 h at 60℃ and 200 r / min, and after the reaction is complete, vacuum filter and collect the zirconium-rich leachate and hafnium-rich leach residue separately.

[0028] S2: Using 20% ​​P204-sulfonated kerosene extractant, the pH of the zirconium-rich leachate was adjusted to 0.8, and three-stage countercurrent extraction was performed to remove impurities at a ratio O / A = 1:2. The zirconium-loaded organic phase was subjected to two-stage countercurrent back-extraction with 2.0 mol / L sulfuric acid solution at a ratio O / A = 3:1 to obtain a pure zirconium solution. Oxalic acid solution was added to the pure zirconium solution to adjust the pH to 1.5, and the reaction was stirred at 80℃ for 1 h. After filtration, the filter cake was calcined at 900℃ for 2 h to obtain a high-purity zirconium oxide product.

[0029] S3: The hafnium-rich leaching residue was added to a 4.0 mol / L sulfuric acid solution at a liquid-to-solid ratio of 3:1 mL / g. After stirring evenly, the mixture was transferred to a high-pressure reactor and reacted at 150℃ and 0.4 MPa for 2 hours. After cooling to room temperature, the mixture was filtered to obtain hafnium-rich leaching solution and silica tailings. A synergistic extraction system of 25% P507 and 5% TOPO was used to adjust the acidity of the hafnium-rich leaching solution to 1.0 mol / L H+. Six-stage countercurrent extraction was performed at a ratio of O / A = 1:1. The zirconium-loaded organic phase was back-extracted with 3.0 mol / L sulfuric acid and returned to the S2 process. The hafnium-containing leaching residue was adjusted to pH 2.0 with ammonia water, and oxalic acid was added to precipitate hafnium. The filter cake was calcined at 950℃ for 2 hours to obtain high-purity hafnium oxide product. The silica tailings were processed to prepare silica.

[0030] Example 2: A method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste residue, comprising the following steps:

[0031] S1: After crushing the hafnium-containing waste residue, ball mill it to 180 mesh, add a 1.2 mol / L dilute sulfuric acid solution at a liquid-to-solid ratio of 5:1 mL / g, stir and leach for 2.0 h at 65℃ and 220 r / min, and after the reaction is complete, vacuum filter and collect the zirconium-rich leachate and hafnium-rich leach residue separately.

[0032] S2: Using 25% P204-sulfonated kerosene extractant, the pH of the zirconium-rich leachate was adjusted to 1.0, and a four-stage countercurrent extraction was performed to remove impurities at a ratio of O / A = 1:2.5. The zirconium-loaded organic phase was subjected to a two-stage countercurrent back-extraction with 2.5 mol / L sulfuric acid solution at a ratio of O / A = 3.5:1 to obtain a pure zirconium solution. Oxalic acid solution was added to the pure zirconium solution to adjust the pH to 1.8, and the reaction was stirred at 85℃ for 1.5 h. After filtration, the filter cake was calcined at 950℃ for 2.5 h to obtain a high-purity zirconium oxide product.

[0033] S3: The hafnium-rich leaching residue was added to a 5.0 mol / L sulfuric acid solution at a liquid-to-solid ratio of 4:1 mL / g. After stirring evenly, the mixture was transferred to a high-pressure reactor and reacted at 160℃ and 0.5 MPa for 3 hours. After cooling to room temperature, the mixture was filtered to obtain hafnium-rich leaching solution and silica tailings. A synergistic extraction system of 30% P507 and 7.5% TOPO was used to adjust the acidity of the hafnium-rich leaching solution to 1.2 mol / L H+. Seven-stage countercurrent extraction was performed at a ratio of O / A = 1:1.5. The zirconium-loaded organic phase was back-extracted with 3.5 mol / L sulfuric acid and returned to the S2 process. The hafnium-containing leaching residue was adjusted to pH 2.2 with ammonia water, and oxalic acid was added to precipitate hafnium. The filter cake was calcined at 1000℃ for 2.5 hours to obtain high-purity hafnium oxide product. The silica tailings were processed to prepare silica.

[0034] Example 3: A method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste residue, comprising the following steps:

[0035] S1: After crushing the hafnium-containing waste residue, ball mill it to 200 mesh, add a 1.5 mol / L dilute sulfuric acid solution at a liquid-to-solid ratio of 6:1 mL / g, stir and leach for 2.5 h at 70℃ and 250 r / min, and after the reaction is complete, vacuum filter and collect the zirconium-rich leachate and hafnium-rich leach residue separately.

[0036] S2: Using 30% P204-sulfonated kerosene extractant, the pH of the zirconium-rich leachate was adjusted to 1.2, and five-stage countercurrent extraction was performed to remove impurities at a ratio O / A = 1:3. The zirconium-loaded organic phase was subjected to three-stage countercurrent back-extraction with 3.0 mol / L sulfuric acid solution at a ratio O / A = 4:1 to obtain a pure zirconium solution. Oxalic acid solution was added to the pure zirconium solution to adjust the pH to 2.0, and the reaction was stirred at 90℃ for 2 hours. After filtration, the filter cake was calcined at 1000℃ for 3 hours to obtain a high-purity zirconium oxide product.

[0037] S3: The hafnium-rich leaching residue was added to a 6.0 mol / L sulfuric acid solution at a liquid-to-solid ratio of 5:1 mL / g. After stirring evenly, the mixture was transferred to a high-pressure reactor and reacted at 170℃ and 0.6 MPa for 4 hours. After cooling to room temperature, the mixture was filtered to obtain hafnium-rich leaching solution and silica tailings. A synergistic extraction system of 35% P507 and 10% TOPO was used to adjust the acidity of the hafnium-rich leaching solution to 1.5 mol / L H+. Eight-stage countercurrent extraction was performed at a ratio of O / A = 1:2. The zirconium-loaded organic phase was back-extracted with 4.0 mol / L sulfuric acid and returned to the S2 process. The hafnium-containing leaching residue was adjusted to pH 2.5 with ammonia water, and oxalic acid was added to precipitate hafnium. The filter cake was calcined at 1100℃ for 3 hours to obtain high-purity hafnium oxide product. The silica tailings were processed to prepare silica.

[0038] Comparative Example 1: Based on Example 3, the staged leaching process was eliminated, and the hafnium-containing waste residue was directly leached under full pressure using 6.0 mol / L sulfuric acid. The subsequent purification and separation processes were the same as in Example 3.

[0039] Comparative Example 2: Based on Example 3, the synergistic extraction system was replaced with a single 35% volume fraction of P507 extractant, while the remaining process parameters and steps remained the same as in Example 3.

[0040] Comparative Example 3: Based on Example 3, no resource recovery of siliceous tailings was carried out, and the tailings were directly discharged. All other process parameters and steps remained the same as in Example 3.

[0041] Performance tests were conducted on Examples 1-3 and Comparative Examples 1-3, and the results are shown in Table 1:

[0042] Table 1

[0043]

[0044] As shown in Table 1, Examples 1-3 all achieved excellent zirconium and hafnium recovery effects and product purity, while significantly reducing acid consumption and tailings emissions. Example 3 had the optimal parameters, achieving a total zirconium recovery rate of 97.5%, a total hafnium recovery rate of 94.8%, and a hafnium oxide purity of 99.96%. Comparative Example 1 used full-volume pressure leaching, achieving a recovery rate close to Example 3, but with a significantly increased acid consumption of over 45%, resulting in a substantial increase in processing costs. Comparative Example 2 used a single extractant, leading to a significant decrease in zirconium and hafnium separation and a substantial reduction in product purity, failing to meet the requirements of high-end applications. Comparative Example 3 did not utilize tailings resources, resulting in tailings generation more than five times that of Example 3, leading to low resource utilization and a large environmental impact. In summary, this invention, through the synergistic effect of staged leaching, synergistic extraction, and multi-component recovery, significantly reduces production costs and environmental impact while ensuring high recovery rates and high product purity, demonstrating significant technical advantages.

[0045] Although embodiments of the invention have been shown and described, it will be understood by those skilled in the art that various changes, modifications, substitutions and variations can be made to these embodiments without departing from the principles and spirit of the invention.

Claims

1. A method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag, characterized in that, Includes the following steps: Step 1: Pre-treatment of hafnium-containing waste residue by graded leaching. After crushing and ball milling to a certain particle size, hafnium-containing waste residue is subjected to primary atmospheric pressure leaching with dilute sulfuric acid. After solid-liquid separation, zirconium-rich leachate and hafnium-rich leaching residue are obtained, realizing the initial gradient separation of zirconium and hafnium. Step 2: Deep purification of zirconium-rich solution and recovery of zirconium products. The zirconium-rich leachate is subjected to extraction and impurity removal treatment to remove impurity metals such as iron, titanium, and aluminum. Then, it is subjected to zirconium-loaded organic back-extraction, oxalic acid precipitation, and high-temperature calcination to obtain high-purity zirconium oxide products. Step 3: Enhanced leaching and deep separation and recovery of hafnium-rich slag. The hafnium-rich leaching slag is subjected to pressure leaching with high-concentration sulfuric acid to fully dissolve hafnium and residual zirconium, resulting in a hafnium-rich leachate. A synergistic extraction system is used to perform deep separation of zirconium and hafnium in the hafnium-rich leachate. After back-extraction, precipitation, and calcination, high-purity hafnium oxide is obtained. Simultaneously, silicon resources are recovered from the leaching tailings to achieve multi-component gradient recovery.

2. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 1, characterized in that, The specific steps of the graded leaching in step one are as follows: After crushing the hafnium-containing waste residue, ball mill it to 150-200 mesh. Add a dilute sulfuric acid solution with a concentration of 1.0-1.5 mol / L at a liquid-to-solid ratio of 4-6:1 mL / g. Stir and leach for 1.5-2.5 h at 60-70℃ and 200-250 r / min. After the reaction is complete, vacuum filter and collect the filtrate, which is the zirconium-rich leachate, and the filter residue, which is the hafnium-rich leach residue.

3. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 2, characterized in that, The hafnium-containing waste residue contains 40%-55% zirconium dioxide and 8%-15% hafnium dioxide by mass, with the remainder being impurities such as silicon, iron, titanium, and aluminum oxides; the leaching rate of zirconium during the dilute sulfuric acid leaching process is 65%-75%, and the leaching rate of hafnium is 15%-25%.

4. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 1, characterized in that, The specific steps for the deep purification of the zirconium-rich solution in step two are as follows: Di(2-ethylhexyl)phosphoric acid (P204) was used as the extractant, and sulfonated kerosene was used as the diluent. The volume fraction of the extractant was 20%-30%. The pH of the zirconium-rich leachate was adjusted to 0.8-1.2, and 3-5 stages of countercurrent extraction were performed to remove impurities at a ratio O / A = 1:2-3. The zirconium-loaded organic phase was back-extracted in 2.0-3.0 mol / L sulfuric acid solution at a ratio O / A = 3-4:1 in 2-3 stages to obtain a pure zirconium solution. Oxalic acid solution was added to the pure zirconium solution to adjust the pH to 1.5-2.0, and the reaction was stirred at 80-90℃ for 1-2 h. After filtration, the filter cake was calcined at 900-1000℃ for 2-3 h to obtain high-purity zirconium oxide.

5. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 1, characterized in that, The specific steps of the pressure leaching in step three are as follows: Add hafnium-rich leaching residue to a sulfuric acid solution with a concentration of 4.0-6.0 mol / L at a liquid-to-solid ratio of 3-5:1 mL / g. After stirring evenly, transfer the solution to a high-pressure reactor and react at 150-170℃ and 0.4-0.6 MPa for 2-4 hours. After the reaction is completed, cool to room temperature and filter to obtain hafnium-rich leaching solution and siliceous tailings.

6. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 5, characterized in that, The leaching rate of hafnium during the pressurized leaching process is ≥92%, and the total leaching rate of zirconium is ≥95%. The silica tailings are washed with water, dissolved in alkali, and precipitated to prepare silica by-products, thereby realizing the recovery of silicon resources.

7. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 1, characterized in that, The specific steps of the synergistic extraction and deep separation in step three are as follows: The synergistic extraction system is a mixture of 2-ethylhexyl di(2-ethylhexyl)phosphonic acid (P507) and trioctylphosphine oxide (TOPO), with sulfonated kerosene as the diluent. The volume fraction of P507 is 25%-35%, and the volume fraction of TOPO is 5%-10%. The acidity of the hafnium-rich leachate is adjusted to 1.0-1.5 mol / L H+, and 6-8 stages of countercurrent extraction are performed at a ratio O / A = 1:1-2. Zirconium preferentially enters the organic phase, while hafnium remains in the aqueous phase. The zirconium-supported organic phase is back-extracted with 3.0-4.0 mol / L sulfuric acid and returned to the zirconium-rich solution purification process in step two. The hafnium-containing extract residue is adjusted to pH 2.0-2.5 with ammonia water, oxalic acid is added to precipitate hafnium, and the filter cake is calcined at 950-1100℃ for 2-3 hours to obtain high-purity hafnium oxide.

8. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 7, characterized in that, The zirconium-hafnium separation coefficient of the synergistic extraction system is ≥80, which is much higher than that of the single extractant system; the purity of the obtained hafnium oxide product is ≥99.9%, and the purity of the zirconium oxide product is ≥99.5%.

9. The method for deep separation and gradient recovery of zirconium and hafnium in hafnium-containing waste slag according to claim 1, characterized in that, The method also includes an extractant recycling step: the blank organic phase after back-extraction is saponified with sodium carbonate solution and then returned to the extraction process for recycling, with an extractant loss rate of ≤3% / batch.

10. A product containing high-purity zirconium oxide and high-purity hafnium oxide, characterized in that, It is prepared by the method of deep separation and gradient recovery of zirconium hafnium in hafnium-containing waste slag as described in any one of claims 1-9.

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