A method for recovering uranium and aluminum from ionic rare earth ore slag

Abstract

The present invention provides a method for recovering uranium and aluminum from ionic rare earth ore slag, relating to the field of hydrometallurgy. The method for recovering uranium and aluminum from ionic rare earth ore slag provided by the present invention comprises the following steps: mixing the ionic rare earth ore slag with concentrated sulfuric acid and performing a stirring leaching to obtain a leachate; mixing the leachate with an ammonium-containing reagent and performing an ammonium-aluminum reaction to obtain an ammonium aluminum sulfate precipitate and a precipitate mother liquor; extracting uranium from the precipitate mother liquor to obtain a uranium-loaded organic phase; and stripping the uranium-loaded organic phase to obtain a qualified uranium solution. The present invention can effectively solve the problem of the difficulty in separating valuable elements such as aluminum, uranium, and rare earths from the slag, reduce the cost of recovering valuable elements from the slag, improve the operating conditions of the separation process, and achieve volume and quantity reduction of radioactive slag, radioactivity exemption of the leached slag, and recovery of valuable elements.

Description

Technical Field

[0001] The invention relates to the technical field of hydrometallurgy, and in particular to a method for recovering uranium and aluminum from ionic rare earth ore slag. Background Art

[0002] Ionic rare earth ore is a unique mineral resource in my country. Rare earth exists in the weathering crust in the form of ions, mainly medium and heavy rare earths, and is concentrated in Hunan, Jiangxi, Guangdong and other places in my country. In the current ionic rare earth ore extraction process, the pH value is adjusted by adding ammonium bicarbonate or ammonia water to the leaching solution to make Al 3+ 、Fe 3+ The precipitate of impurity ions such as uranium and thorium, as well as some rare earth elements, from ionic rare earth ores, also accumulates in the slag during the purification process. Currently, manufacturers primarily store this radioactive slag in slag storage, which creates significant safety and environmental pressures while also wasting uranium resources.

[0003] There are relatively few studies on the extraction and recovery of radioactive elements from ionic rare earth ore residues, and the research mainly focuses on the extraction and recovery of rare earth and thorium. Sun Xiaoqi et al. from the Xiamen Rare Earth Materials Research Institute proposed a "step-by-step leaching method for radioactive waste residues from ionic rare earth mines (CN112458291A)" that extracts thorium and rare earth from the waste residues by roasting the leached residues and then performing three-stage countercurrent leaching with concentrated hydrochloric acid; and a "method for recovering thorium and rare earth from waste residue leachate (CN112458320A)" that extracts thorium from the leachate by adding CA-12, strips thorium with inorganic acid, and adjusts the acidity of the thorium solution by adding alkali. Thorium is hydrolyzed and precipitated to produce Th(OH)4; saponified CA-12 is then added to the thorium extraction residue to extract the lanthanides. This method involves roasting the radioactive waste residue and performing multi-stage leaching, resulting in complex processes, high processing costs, and the generation of roasting waste gas. The hydrochloric acid extraction system differs from the ionic rare earth production system and requires more sophisticated equipment. The efficiency of the staged extraction separation needs to be improved. The extraction and recovery of radioactive uranium is not involved.

[0004] Chinese patent CN104711424A proposes a method for recovering rare earths and aluminum from the slag removed from the leachate of a weathering crust-eluvial rare earth ore. The slag is leached with a 1-10% acid solution, and the leachate is adjusted to a pH of 5.0-5.5 by adding an alkaline solution to form an aluminum hydroxide precipitate. The rare earth-containing filtrate is then adjusted to a pH of 6.5-8.0 by adding an alkaline solution to obtain the rare earth precipitate. This method does not involve the separation and recovery of radioactive elements from the slag. Instead, it uses a neutralization precipitation method to separate aluminum and rare earths. During the formation of amorphous aluminum hydroxide, it co-precipitates with other metal elements. Aluminum hydroxide has strong adsorption and a high water content, which inevitably causes the adsorption and entrainment of uranium and rare earths. This makes solid-liquid separation difficult, affecting product quality.

[0005] Chinese patent CN114250367A proposes a method for extracting and recovering uranium, thorium, aluminum, and rare earths from ionic rare earth impurity removal residue. The impurity removal residue is leached with concentrated sulfuric acid; the leachate is subjected to uranium extraction using N235+TRPO, and the uranium-loaded organic phase is stripped to obtain a uranium-enriched solution; thorium is extracted from the uranium extraction residue, and the rare earth and aluminum-containing residue is subjected to an ammonium-aluminum reaction; cooling and crystallization are performed, and ammonium aluminum sulfate and a mother liquor containing rare earths are filtered and washed to obtain the ammonium aluminum sulfate product. This method directly extracts uranium from the leachate, but suffers from large extraction throughput, high specific gravity and viscosity of the leachate, poor operating conditions, and low extraction efficiency. Furthermore, TRPO, as a co-extractant, is relatively expensive. Using cooling crystallization to separate and recover aluminum requires large equipment investment and demanding operating conditions.

[0006] In summary, existing technologies have little research on the radioactive element uranium in impurity removal slag, or have not involved technically and economically feasible methods for recovering uranium from impurity removal slag. In particular, since ionic rare earth ore impurity removal slag has low uranium content, high aluminum and iron content, and contains impurities such as clay and humic acid, existing technologies are unable to achieve economical and efficient separation and recovery of uranium. Existing technologies use neutralization precipitation and cooling crystallization methods to separate aluminum from impurity removal slag. The neutralization precipitation method has problems such as impurity carryover and adsorption during the separation process, and the cooling crystallization method has problems such as high production investment and high process control precision requirements, which limits its production application. Summary of the Invention

[0007] The object of the present invention is to provide a method for recovering uranium and aluminum from ionic rare earth ore slag. The present invention can effectively solve the problem of difficult separation of valuable elements such as aluminum, uranium, and rare earth in the slag, reduce the cost of recovering valuable elements from the slag, improve the operating conditions of the separation process, and achieve volume and amount reduction of radioactive slag, exemption of radioactivity of leaching slag, and recovery of valuable elements.

[0008] In order to achieve the above-mentioned object of the invention, the present invention provides the following technical solutions:

[0009] The present invention provides a method for recovering uranium and aluminum from ionic rare earth ore impurity removal residue, comprising the following steps:

[0010] Mixing the ionic rare earth ore impurity-removed residue with concentrated sulfuric acid, and performing stirring leaching to obtain a leachate;

[0011] The leachate is mixed with an ammonium-containing reagent to carry out an ammonium-aluminum reaction to obtain an ammonium aluminum sulfate precipitate and a precipitate mother liquor;

[0012] Extracting uranium from the precipitated mother liquor to obtain an organic phase loaded with uranium;

[0013] The uranium-loaded organic phase is stripped to obtain a qualified uranium solution.

[0014] Preferably, the concentration of the concentrated sulfuric acid is 98 wt %; and the mass ratio of the ionic rare earth ore impurity removal residue to the concentrated sulfuric acid is 1:0.8-1.5.

[0015] Preferably, the stirring rate of the stirring leaching is 200 to 400 r / min; and the reaction time of the stirring leaching is 20 to 60 min.

[0016] Preferably, the pH value of the leachate is 0.8-2.5.

[0017] Preferably, the amount of the ammonium-containing reagent is calculated based on an ammonium to aluminum ratio of 2.0 to 5.0:1.

[0018] Preferably, the ammonium-aluminum reaction is carried out under stirring conditions; the stirring rate is 90 to 180 r / min; and the time of the ammonium-aluminum reaction is 60 to 240 min.

[0019] Preferably, the extractants used in the uranium extraction include N235, tributyl phosphate and kerosene.

[0020] Preferably, the O / A ratio in the extractant is 1:10-20.

[0021] Preferably, the reagent used in the back extraction is sodium carbonate solution; the concentration of the sodium carbonate solution is 100-300 g / L.

[0022] Preferably, the phase ratio O / A during the back extraction is 5 to 10:1.

[0023] The present invention provides a method for recovering uranium and aluminum from ionic rare earth ore impurity removal residue. The method adopts concentrated sulfuric acid for stirring leaching to achieve simultaneous leaching of uranium, rare earth and other elements from the impurity removal residue. In view of the high aluminum content and high viscosity of the leachate, an ammonium aluminum reaction is used to remove aluminum to generate ammonium aluminum sulfate for precipitation and separation. Uranium in the precipitated mother liquor is then recovered by extraction, thereby achieving the recovery of uranium and aluminum from the ionic rare earth ore impurity removal residue.

[0024] The present invention combines the reagents used in the production of ionic rare earth ores, realizes the resource disposal of impurity removal slag under the original system, improves the operating phenomena of the separation process, effectively improves the separation efficiency of valuable metals, reduces the cost of equipment and reagents, and realizes the comprehensive recovery of valuable elements such as uranium and aluminum in the impurity removal slag. DETAILED DESCRIPTION

[0025] The present invention provides a method for recovering uranium and aluminum from ionic rare earth ore impurity removal residue, comprising the following steps:

[0026] Mixing the ionic rare earth ore impurity-removed residue with concentrated sulfuric acid, and performing stirring leaching to obtain a leachate;

[0027] The leachate is mixed with an ammonium-containing reagent to carry out an ammonium-aluminum reaction to obtain an ammonium aluminum sulfate precipitate and a precipitate mother liquor;

[0028] Extracting uranium from the precipitated mother liquor to obtain an organic phase loaded with uranium;

[0029] The uranium-loaded organic phase is stripped to obtain a qualified uranium solution.

[0030] The present invention mixes ionic rare earth ore impurity removal residue with concentrated sulfuric acid and performs stirring leaching to obtain a leachate. In the present invention, the ionic rare earth ore impurity removal residue comprises, by mass fraction, 0.01-0.05% U, 1.00-5.00% REO, and 35-50% Al2O3. In the present invention, the ionic rare earth ore impurity removal residue preferably has a moisture content of 65-80%.

[0031] In the present invention, before mixing the ionic rare earth ore slag with concentrated sulfuric acid, the ionic rare earth ore slag preferably further comprises: grinding the ionic rare earth ore slag and then mixing it with water to break up and dissociate the solid particles of the slag, thereby increasing the surface area in contact with the solution; and stirring the mixture with water for preliminary dissolution, thereby increasing the reaction rate with sulfuric acid. In the present invention, the mass ratio of the ionic rare earth ore slag to water is preferably 1:2.8-5.6, more preferably 1:3.0-5.0. In the present invention, the mass of the ionic rare earth ore slag is measured as the mass of the dry slag.

[0032] In the present invention, the concentration of the concentrated sulfuric acid is preferably 98wt%; the mass ratio of the ionic rare earth ore impurity removal slag to the concentrated sulfuric acid is preferably 1:0.8-1.5. In the present invention, the mass of the ionic rare earth ore impurity removal slag is calculated as the mass of the dry slag.

[0033] In the present invention, the stirring rate of the stirring leaching is preferably 200-400 r / min, more preferably 250-350 r / min; the reaction time of the stirring leaching is preferably 20-60 min, more preferably 30-50 min. In the present invention, the reaction time of the stirring leaching refers to the reaction time from the addition of all sulfuric acid to the end of the stirring leaching.

[0034] In the stirring leaching process of the present invention, no additional heating is required, and the stirring leaching process utilizes the dilution heat of concentrated sulfuric acid and the reaction heat to maintain the leaching temperature at 60-80°C.

[0035] In the present invention, after the stirring leaching, solid-liquid separation is preferably performed to obtain leachate and leach residue.

[0036] In the present invention, the pH value of the leachate is preferably 0.8 to 2.5, more preferably 1.2 to 1.8.

[0037] In the present invention, after obtaining the leached residue, the leached residue is preferably subjected to countercurrent washing. The washing liquid for the countercurrent washing is preferably water or a 20g / L dilute sulfuric acid solution. The washing ratio is preferably 0.5-1.0:1, the number of washing stages is preferably 3-5, and the washing rates for uranium and rare earth elements are preferably both above 97%, with the uranium washing rate preferably being 98-99% and the rare earth washing rate preferably being 97-99%. The washing liquid from the countercurrent washing is preferably recycled as water for the agitated leaching process. The washing ratio refers to the volume of washing liquid used to wash a unit weight of solids in the slurry.

[0038] This method uses concentrated sulfuric acid and a low liquid-to-solid ratio for agitated leaching, enabling the simultaneous leaching of valuable elements from the impurity-removing residue. Heat from sulfuric acid dilution and reaction is utilized to provide heat. High acidity prevents silicic acid and humic acid from colloidal states, improving the solid-liquid separation performance of the leached slurry. Countercurrent washing of the leached residue, with the washing liquid returned to the leaching water, increases the recovery rate of valuable elements, facilitates their separation, and achieves radioactivity immunity for the leached residue.

[0039] After obtaining the leachate, the present invention mixes the leachate with an ammonium-containing reagent to perform an ammonium-aluminum reaction to obtain an ammonium aluminum sulfate precipitate and a precipitate mother liquor. In the present invention, the ammonium-containing reagent is preferably used in an ammonium-aluminum ratio of 2.0 to 5.0:1, and the ammonium-aluminum ratio is more preferably 3.0 to 4.0:1. In the present invention, the ammonium-aluminum ratio refers to the molar ratio of ammonium ions to aluminum ions.

[0040] In the present invention, the ammonium-containing reagent preferably comprises one or more of ammonium sulfate, ammonia water, ammonium bicarbonate and ammonium carbonate. In the present invention, the ammonium-containing reagent is preferably added to the leachate in the form of a solid or a saturated solution.

[0041] In the present invention, the ammonium-aluminum reaction is preferably carried out under stirring conditions; the stirring rate is preferably 90 to 180 r / min, more preferably 120 to 150 r / min; the ammonium-aluminum reaction time is preferably 60 to 240 min, more preferably 90 to 180 min. In the present invention, the temperature of the ammonium-aluminum reaction is preferably room temperature. In the present invention, the aluminum precipitation rate of the ammonium-aluminum reaction is preferably greater than 90%, more preferably 95 to 96%.

[0042] In the present invention, after the ammonium aluminum reaction, solid-liquid separation is preferably performed to obtain ammonium aluminum sulfate precipitate and precipitation mother liquor.

[0043] In the present invention, after obtaining the ammonium aluminum sulfate precipitate, the present invention preferably performs displacement washing on the ammonium aluminum sulfate precipitate. In the present invention, the detergent used in the displacement washing is preferably a saturated solution of an ammonium reagent; the washing conditions are preferably 3 to 5 stages of countercurrent washing, and the washing ratio of each stage is preferably 0.5 to 1.5:1, more preferably 1.0:1; the washing liquid of the displacement washing is preferably returned as an ammonium aluminum reaction reagent. In the present invention, the ammonium reagent in the saturated solution of the ammonium reagent used in the displacement washing is consistent with the ammonium reagent used in the ammonium aluminum reaction. In the displacement washing process of the present invention, the aluminum washing rate in the ammonium aluminum sulfate is preferably <3.0%, more preferably 1.0 to 1.3%; the uranium and rare earth washing rates are preferably independently greater than 97.5%, wherein the uranium washing rate is preferably 98.7 to 99%, and the rare earth washing rate is preferably 98.8 to 99%; the ammonium aluminum sulfate product obtained after displacement washing meets national standards.

[0044] The present invention not only removes aluminum from the leachate through the ammonium-aluminum reaction, but also adjusts the sulfate concentration in the solution, reducing the solution viscosity and creating conditions for uranium extraction. In a specific embodiment of the present invention, the sulfate concentration in the precipitation mother liquor is reduced from 150-200 g / L to 30-50 g / L.

[0045] The invention separates aluminum elements in a leachate by selective precipitation of ammonium aluminum sulfate, and countercurrently washes the ammonium aluminum sulfate precipitate with a saturated solution containing an ammonium reagent, thereby achieving selective separation of aluminum in the leachate, effectively reducing the entrainment of impurity elements, and reducing the dissolution loss of aluminum.

[0046] After obtaining the precipitation mother liquor, the present invention performs uranium extraction on the precipitation mother liquor to obtain a uranium-loaded organic phase. In the present invention, the extractant used for the uranium extraction preferably includes N235, tributyl phosphate (TBP), and kerosene. In the present invention, the extractant preferably includes, by volume, 5-10% N235, 10-20% TBP, and the remainder kerosene, more preferably 7.5% N235, 15% TBP, and the remainder kerosene. In the present invention, the O / A ratio in the extractant is preferably 1:10-20, more preferably 1:15.

[0047] In the present invention, the uranium extraction preferably utilizes multi-stage countercurrent extraction, more preferably 3-5 stages of countercurrent extraction. In the present invention, the uranium extraction rate is preferably greater than 98.0%, and the uranium concentration in the raffinate is preferably less than 1.0 mg / L, more preferably 0.65-0.78 mg / L. The raffinate from the uranium extraction is preferably returned to the rare earth production line for rare earth recovery.

[0048] The present invention uses N235+TBP+kerosene to extract and recover uranium from a precipitation mother liquor, has good operating phenomena (fast phase separation, no third phase or three-phase material is generated), low reagent cost, and the raffinate is returned to the rare earth production line, the valuable elements are recovered and utilized, and no wastewater is discharged in the recovery process.

[0049] After obtaining the uranium-loaded organic phase, the present invention performs back extraction on the uranium-loaded organic phase to obtain a qualified uranium solution. In the present invention, the reagent used for the back extraction is preferably a sodium carbonate solution; the concentration of the sodium carbonate solution is preferably 100-300 g / L, more preferably 150-200 g / L. In the present invention, the phase ratio O / A during the back extraction is preferably 5-10:1, more preferably 8:1. In the present invention, the back extraction preferably utilizes multi-stage countercurrent back extraction, more preferably 3-5 stages of countercurrent back extraction. In the present invention, the organic phase after the back extraction is preferably returned to be used as an extractant for uranium extraction. In the present invention, the uranium recovery rate is preferably 95-96%.

[0050] In the present invention, ionic rare earth ore impurity removal residue has the characteristics of a wide variety of valuable elements, radioactivity, a low uranium / aluminum grade ratio, and inclusions of clay and humic acid. The existing technology has problems such as insufficient research objects, high reagent costs, and low metal separation and recovery efficiency. The present invention proposes a method for recovering uranium and aluminum from ionic rare earth ore impurity removal residue based on factors such as the element occurrence state, the metal separation and recovery order, and process water circulation. The method has the following advantages: high metal recovery rate, selective precipitation through ammonium-aluminum reaction, multi-stage countercurrent washing and wash water return, and multi-stage countercurrent uranium extraction, achieving separation and recovery of uranium and aluminum, producing ammonium aluminum sulfate product and qualified uranium solution, with aluminum recovery rates exceeding 85% and uranium recovery rates greater than 94%. In order to maximize resources, the solution after separating uranium and aluminum is returned to the rare earth production line to realize the recovery of valuable elements such as rare earths; the reagent cost is low, and in order to combine the ionic rare earth ore production process and reduce equipment requirements and investment, the reagents used in this method are all reagents used in rare earth production, and the N235 and TBP extractants used are relatively low in cost. The extraction stock solution is the mother liquor precipitated after aluminum removal, and the processing capacity is greatly reduced relative to the leaching solution, and the aluminum concentration, acidity, and viscosity are all adjusted. The extraction operation phenomenon is significantly improved, and the loss of the extractant is reduced; the technology is highly applicable, and the precipitation method and multi-stage countercurrent extraction method used in the method are simple to control and operate, have low equipment requirements, and have strong applicability for technical production applications. In the present invention, washing water and raffinate water are both recycled, no wastewater is discharged, the waste residue discharge amount is small, the leaching residue can achieve radioactive exemption, and the impact on the environment is greatly reduced.

[0051] The following will be combined with the embodiments of the present invention to clearly and completely describe the technical solutions of the present invention. Obviously, the embodiments described are only some of the embodiments of the present invention, not all of them. Based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without making creative efforts are within the scope of protection of the present invention.

[0052] Example 1

[0053] The moisture content of the impurity-removed slag of a certain ionic rare earth ore is 67%, the U content in the impurity-removed slag is 0.02wt%, the REO content is 1.44wt%, and the Al2O3 content is 42.08wt%.

[0054] S1. 500 g of ionic rare earth ore impurity removal residue was ground and mixed with water, the weight ratio of water to impurity removal residue was 3.0:1, and concentrated sulfuric acid with a concentration of 98 wt% was slowly added, the mass ratio of impurity removal residue to concentrated sulfuric acid was 1:0.8, the stirring rate was 200 r / min, the reaction time was 30 min, and the pH value of the leachate was 1.2;

[0055] S2. After leaching is completed, solid-liquid separation is performed to obtain leachate and leach residue; the leach residue is countercurrently washed with water, with a washing ratio of 0.5:1, 3 washing stages, a uranium washing rate of 98%, and a rare earth washing rate of 97%, and the washing liquid is used as leaching water.

[0056] S3. Add saturated ammonium sulfate solution to the leachate, with an ammonium to aluminum ratio of 2.0:1, a stirring rate of 120 r / min, a reaction time of 90 min, and an aluminum precipitation rate of 95%.

[0057] S4. After the precipitation reaction, solid-liquid separation is performed to obtain ammonium aluminum sulfate precipitate and precipitation mother liquor; the ammonium aluminum sulfate precipitate is subjected to three-stage countercurrent washing with a saturated ammonium sulfate solution, with a washing ratio of 0.5:1, an aluminum washing rate of 1.2%, a uranium washing rate of 98.7%, and a rare earth washing rate of 98.8%, and the washing liquid is returned as an aluminum precipitation reagent.

[0058] S5. Extract uranium from the precipitated mother liquor using 5 vol% N235 + 10 vol% TBP + kerosene as the extractant, with an O / A ratio of 1:10, and three-stage countercurrent extraction. The uranium concentration in the raffinate is 0.78 mg / L, and the raffinate is returned to the rare earth production line.

[0059] S6. The loaded organic phase is stripped of uranium using a 100 g / L sodium carbonate solution with an O / A ratio of 5:1. After three stages of countercurrent stripping, a qualified uranium solution is obtained with a uranium recovery rate of 95%. After stripping, the organic phase is returned and used as an extractant for uranium extraction.

[0060] Example 2

[0061] The moisture content of the impurity-removed slag of a certain ionic rare earth ore is 73%, the U content in the impurity-removed slag is 0.03wt%, the REO content is 1.74wt%, and the Al2O3 content is 40.27wt%.

[0062] S1. 500 g of ionic rare earth ore impurity removal residue was ground and mixed with water, the weight ratio of water to impurity removal residue was 4.0:1, and concentrated sulfuric acid with a concentration of 98 wt% was slowly added, the mass ratio of impurity removal residue to concentrated sulfuric acid was 1:0.8, the stirring rate was 250 r / min, the reaction time was 40 min, and the pH value of the leachate was 1.7;

[0063] S2. After leaching is completed, solid-liquid separation is performed to obtain leachate and leach residue; the leach residue is countercurrently washed with water, with a washing ratio of 1.0:1, 3 washing stages, a uranium washing rate of 98.5%, and a rare earth washing rate of 97.8%, and the washing liquid is used as leaching water.

[0064] S3. Add saturated ammonium sulfate solution to the leachate, with an ammonium to aluminum ratio of 3.0:1, a stirring rate of 150 r / min, a reaction time of 120 min, and an aluminum precipitation rate of 96%.

[0065] S4. After the precipitation reaction, solid-liquid separation is performed to obtain ammonium aluminum sulfate precipitate and precipitation mother liquor; the ammonium aluminum sulfate precipitate is subjected to four-stage countercurrent washing with a saturated ammonium sulfate solution, with a washing ratio of 1.0:1, an aluminum washing rate of 1.3%, a uranium washing rate of 98.9%, and a rare earth washing rate of 99.0%, and the washing liquid is returned as an aluminum precipitation reagent.

[0066] S5. Extract uranium from the precipitated mother liquor using 7.5 vol% N235 + 15 vol% TBP + kerosene as the extractant, with an O / A ratio of 1:10, and 4-stage countercurrent extraction. The uranium concentration in the raffinate is 0.65 mg / L, and the raffinate is returned to the rare earth production line.

[0067] S6. The loaded organic phase is stripped of uranium using 150 g / L sodium carbonate solution with an O / A ratio of 8:1. After four stages of countercurrent stripping, a qualified uranium solution is obtained with a uranium recovery rate of 96%. After stripping, the organic phase is returned and used as an extractant for uranium extraction.

[0068] Example 3

[0069] The moisture content of the impurity-removed slag of a certain ionic rare earth ore is 77%, the U content in the impurity-removed slag is 0.04wt%, the REO content is 1.80wt%, and the Al2O3 content is 40.49wt%.

[0070] S1. 500 g of ionic rare earth ore impurity removal residue was ground and mixed with water, the weight ratio of water to impurity removal residue was 5.0:1, and concentrated sulfuric acid with a concentration of 98 wt% was slowly added, the mass ratio of impurity removal residue to concentrated sulfuric acid was 1:0.8, the stirring rate was 350 r / min, the reaction time was 50 min, and the pH value of the leachate was 1.8;

[0071] S2. After leaching is completed, solid-liquid separation is performed to obtain leachate and leach residue; the leach residue is countercurrent washed with 20 g / L dilute sulfuric acid solution, with a washing ratio of 1.0:1, 5 washing stages, a uranium washing rate of 99%, and a rare earth washing rate of 99%, and the washing liquid is used as leaching water.

[0072] S3. Add saturated ammonium sulfate solution to the leachate, with an ammonium to aluminum ratio of 4.0:1, a stirring rate of 150 r / min, a reaction time of 180 min, and an aluminum precipitation rate of 96%.

[0073] S4. After the precipitation reaction, solid-liquid separation is performed to obtain ammonium aluminum sulfate precipitate and precipitation mother liquor; the ammonium aluminum sulfate precipitate is subjected to four-stage countercurrent washing with a saturated ammonium sulfate solution, with a washing ratio of 1.0:1, an aluminum washing rate of 1.0%, a uranium washing rate of 99%, and a rare earth washing rate of 99%, and the washing liquid is returned as an aluminum precipitation reagent.

[0074] S5. Extract uranium from the precipitated mother liquor using 10 vol% N235 + 20 vol% TBP + kerosene as the extractant, with an O / A ratio of 1:15, and 5-stage countercurrent extraction. The uranium concentration in the raffinate is 0.76 mg / L, and the raffinate is returned to the rare earth production line.

[0075] S6. The loaded organic phase is stripped of uranium using 200 g / L sodium carbonate solution with an O / A ratio of 8:1. After four stages of countercurrent stripping, a qualified uranium solution is obtained with a uranium recovery rate of 95%. After stripping, the organic phase is returned and used as an extractant for uranium extraction.

[0076] The present invention provides a method for recovering uranium and aluminum from ionic rare earth ore slag. The method sequentially performs concentrated sulfuric acid agitation leaching of the slag, ammonium-aluminum reaction precipitation of the leachate, countercurrent washing with ammonium aluminum sulfate, and uranium extraction and separation. This method achieves comprehensive recovery of valuable elements such as uranium and aluminum from the ionic rare earth ore slag, with aluminum recovery rates exceeding 85% and uranium recovery rates exceeding 94%. Compared with existing technologies, the present invention has a high recovery rate for valuable metals, low reagent costs, strong technical applicability, and fully recycled process water. The leached slag can achieve radioactivity exemption, achieving harmless disposal and resource utilization of the slag.

[0077] The above is only a preferred embodiment of the present invention. It should be pointed out that for ordinary technicians in this technical field, several improvements and modifications can be made without departing from the principles of the present invention. These improvements and modifications should also be regarded as within the scope of protection of the present invention.

Claims

1. A method for recovering uranium and aluminum from ionic rare earth ore residue, comprising the following steps: Mixing the ionic rare earth ore impurity-removed residue with concentrated sulfuric acid, and performing stirring leaching to obtain a leachate; The stirring rate of the stirring leaching is 200-400 r / min; The leachate and the ammonium-containing reagent are mixed to carry out an ammonium-aluminum reaction, and solid-liquid separation is performed to obtain an ammonium aluminum sulfate precipitate and a precipitate mother liquor; the ammonium aluminum reaction is carried out under stirring conditions; the stirring rate is 90 to 180 r / min; the ammonium aluminum reaction time is 60 to 240 minutes; the temperature of the ammonium aluminum reaction is room temperature; after obtaining the ammonium aluminum sulfate precipitate, the ammonium aluminum sulfate precipitate is subjected to displacement washing; the detergent used in the displacement washing is a saturated solution of the ammonium-containing reagent; The washing conditions are 3 to 5 stages of countercurrent washing, with a washing ratio of 0.5 to 1.5:1 for each stage; Extracting uranium from the precipitated mother liquor to obtain an organic phase loaded with uranium; The uranium extraction agent used is N235, tributyl phosphate and kerosene; in terms of volume fraction, the extractant includes 5-10% N235, 10-20% TBP and the balance kerosene; the O / A ratio in the extractant is 1:10-20; The uranium-loaded organic phase is stripped to obtain a qualified uranium solution.

2. The method according to claim 1, characterized in that The concentration of the concentrated sulfuric acid is 98 wt %; the mass ratio of the ionic rare earth ore impurity removal residue to the concentrated sulfuric acid is 1:0.8-1.

5.

3. The method according to claim 1 or 2, characterized in that The reaction time of the stirring leaching is 20 to 60 minutes.

4. The method according to claim 1 or 2, characterized in that The pH value of the leaching solution is 0.8-2.

5.

5. The method according to claim 1, wherein The amount of the ammonium-containing reagent is calculated based on an ammonium to aluminum ratio of 2.0 to 5.0:

1.

6. The method according to claim 1, characterized in that The reagent used in the back extraction is sodium carbonate solution; the concentration of the sodium carbonate solution is 100-300 g / L.

7. The method according to claim 6, characterized in that The phase ratio O / A during the back extraction is 5 to 10:1.