Method for resource utilization of sedimentary rare earth ore

Abstract

The present disclosure relates to the technical field of comprehensive utilization of sedimentary rare earth ore, and provides a resource utilization method of sedimentary rare earth ore, comprising: pre-treating and roasting the raw ore of the sedimentary rare earth ore to obtain pre-treated material; using sulfuric acid as a leaching agent to perform acid leaching on the pre-treated material to separate and obtain high-aluminum and iron-containing acid leaching solution and leaching residue; using ammonium bicarbonate as a neutralizing agent to perform multiple hydrolysis precipitations on the high-aluminum and iron-containing acid leaching solution to remove impurities step by step, to obtain mixed rare earth carbonate; and performing thermal decomposition on the mixed rare earth carbonate to obtain mixed rare earth oxide. The resource utilization method provided by the present disclosure at least achieves the effect of not only being able to improve the rare earth recovery rate, the associated element utilization rate and the impurity removal rate, but also being able to solve the problem of difficult treatment of mother liquor.

Description

Technical Field

[0001] This disclosure relates to the field of comprehensive utilization technology of sedimentary rare earth minerals, for example, to a method for resource utilization of sedimentary rare earth minerals. Background Technology

[0002] Sedimentary rare earth elements (REEs) are a new type of rare earth resource, with high-value key rare earth element oxides (such as praseodymium, neodymium, terbium, and dysprosium) accounting for over 20% of the total content. They possess advantages such as wide distribution, good continuity, and enormous resource potential. Compared to carbonate and alkaline rare earth elements, sedimentary REEs have certain advantages in terms of mining conditions and the proportion of key rare earth elements. Compared to ion-adsorption REEs and deep-sea rare earth-rich mud, sedimentary REEs also have advantages in terms of grade, scale, concentration, and environmental impact. Therefore, sedimentary REEs have significant development and utilization prospects.

[0003] Studies have shown that using traditional leaching processes based on ammonium salt and sulfuric acid systems, the direct leaching rate of rare earth elements at room temperature is extremely low (0.02%–24%). Prior to 2018, there had been no breakthrough in the efficient utilization technology of sedimentary rare earth elements. However, Chinese patent document CN201811407361.2, filed on 20181123 and published on 20190125, entitled "A Method for Selectively Leaching Sedimentary Rare Earth Ore," describes a method that involves crushing sedimentary rare earth ore into powder, roasting it at 550–850℃ for 0.5–5 hours, and then reacting it in an inorganic acid solution of 0.5–5 mol / L for 0.5–5 hours. This method achieves a rare earth leaching rate of over 90% and effectively inhibits the leaching of impurity elements. While this patent document promotes the efficient utilization of sedimentary rare earth from the perspective of selective leaching, it still suffers from problems such as low comprehensive utilization of associated elements and difficulties in mother liquor disposal. Furthermore, its rare earth recovery rate and impurity removal rate need further improvement.

[0004] In summary, there is an urgent need for a method to comprehensively utilize sedimentary rare earth elements, which can at least improve the rare earth recovery rate, the utilization rate of associated elements and the impurity removal rate, and solve the problem of difficult mother liquor disposal. Summary of the Invention

[0005] The purpose of this disclosure is to overcome the shortcomings of the prior art and provide a method for the resource utilization of sedimentary rare earth minerals, so as to at least achieve the effect of not only improving the rare earth recovery rate, the utilization rate of associated elements and the impurity removal rate, but also solving the problem of mother liquor disposal.

[0006] The purpose of this disclosure is achieved through the following technical solution:

[0007] On the one hand, a method for the resource utilization of sedimentary rare earth ore is provided. The method includes: pre-treating and roasting the sedimentary rare earth ore to obtain pre-treated material; using sulfuric acid as a leaching agent to perform acid leaching on the pre-treated material, separating a high-alumina iron-containing acid leaching solution and leaching residue; using ammonium bicarbonate as a neutralizing agent to perform multiple hydrolysis and precipitation processes on the high-alumina iron-containing acid leaching solution to remove impurities stepwise, obtaining mixed rare earth carbonates; and thermally decomposing the mixed rare earth carbonates to obtain mixed rare earth oxides.

[0008] It should be noted that the leaching residue can be directly stockpiled, backfilled, or used as a building material auxiliary material.

[0009] It should be understood that sedimentary rare earth deposits and ion-adsorption rare earth deposits have similar mineral compositions, both being dominated by clay minerals with kaolinite content exceeding 80%, and lacking selectable rare earth minerals of a fine particle size. Therefore, conventional physical beneficiation methods are insufficient for enriching rare earth elements. Furthermore, unlike ion-adsorption rare earth deposits where the direct leaching rate of rare earth elements is very high (typically 60%–90%), sedimentary rare earth deposits exhibit extremely low direct leaching rates (maximum only 24%). This is likely because the rare earth elements in these deposits are not primarily bound to clay minerals such as kaolinite through ion adsorption. In other words, although sedimentary and ion-adsorption rare earth deposits have similar mineral compositions, the binding mechanisms of rare earth elements differ. Therefore, methods suitable for ion-adsorption rare earth deposits are not applicable to sedimentary rare earth deposits. Consequently, extraction methods for rare earth elements from different mineral types do not necessarily provide mutual technological inspiration.

[0010] In some embodiments, the step of using ammonium bicarbonate as a neutralizing agent to perform multiple hydrolysis and precipitation of the high-alumina iron-containing acid leaching solution to remove impurities step by step and obtain mixed rare earth carbonates includes: adding the ammonium bicarbonate to the high-alumina iron-containing acid leaching solution, adjusting the pH value to 1.8-2.6, cooling and crystallizing, and separating the low-alumina rare earth solution and the first crystallization byproduct.

[0011] It should be noted that the first crystallization byproduct includes crude ammonium alum.

[0012] The crude ammonium alum product can be washed and purified to obtain ammonium alum byproducts.

[0013] In some embodiments, the temperature for cooling crystallization is -5 to 30°C, and the cooling crystallization time is 1 to 6 hours.

[0014] It should be emphasized that aluminum alum double salts have lower solubility than sulfates, and the temperature changes more significantly. The alum crystals formed during the aluminum precipitation process have good particle size, are easy to filter, and have high crystallinity. At the same time, the content of rare earth elements is low, resulting in high purity of the product. This is beneficial for the synergistic utilization of rare earth elements and aluminum elements in high-aluminum iron-containing acid leaching solutions.

[0015] 2NH4HCO3+H2SO4→2(NH4)2SO4+2CO2↑+2H2O(1)

[0016] Al2(SO4)3+(NH4)2SO4+12H2O→2NH4Al(SO4)2·12H2O(2)

[0017] According to chemical reaction equations (1) and (2), it can be seen that ammonium bicarbonate slow-release aluminum precipitation has two advantages over ammonium sulfate aluminum precipitation: on the one hand, the reaction of ammonium bicarbonate with sulfuric acid is a process of gradually generating ammonium sulfate. By adjusting the amount of ammonium bicarbonate and the feeding method, the growth of ammonium alum (NH4Al(SO4)2·12H2O) crystals can be more precisely controlled; on the other hand, the reaction of ammonium bicarbonate with sulfuric acid can reduce the acidity and does not reintroduce new sulfate ions. After cooling and crystallizing to precipitate ammonium alum, the salt concentration of the solution is reduced more efficiently, which is beneficial for subsequent impurity removal and preparation of rare earth carbonate products.

[0018] In some of the above embodiments, the acidity of the solution is reduced by reacting ammonium bicarbonate with residual sulfuric acid in the high-alumina iron-containing acid leaching solution to generate ammonium sulfate. The generated ammonium sulfate then reacts with aluminum ions provided by the leaching agent to generate aluminum sulfate, which further reacts to form ammonium alum. At the same time, by controlling the pH value and the temperature and time of cooling crystallization, not only can the crystal form of ammonium alum be effectively regulated, but also the rare earth loss rate can be guaranteed to be less than 5% and the aluminum removal rate greater than 85%, thereby further improving the rare earth recovery rate and the impurity removal rate, thus achieving efficient recycling of rare earth.

[0019] In some embodiments, the step of using ammonium bicarbonate as a neutralizing agent to perform multiple hydrolysis and precipitation of the high-alumina iron-containing acid leaching solution to remove impurities step by step and obtain mixed rare earth carbonates further includes: adding the ammonium bicarbonate to the low-alumina rare earth solution to adjust the pH value to 4.8-5.4, and separating the iron-alumina rare earth purification solution and the impurity residue; and adding the ammonium bicarbonate to the iron-alumina rare earth purification solution to adjust the pH value to 6.8-7.0, and separating the mixed rare earth carbonates and the precipitate mother liquor.

[0020] It should be noted that the slag can be directly stockpiled, backfilled, or used as building material auxiliary material.

[0021] In some of the above embodiments, by using ammonium bicarbonate to adjust the pH value of the low-alumina rare earth solution to 4.8-5.4, iron and aluminum impurities in the low-alumina rare earth solution can be further removed to obtain an iron-alumina rare earth purified solution; on this basis, by using ammonium bicarbonate to adjust the pH value of the iron-alumina rare earth purified solution to 6.8-7.0, precipitation can be achieved to obtain mixed rare earth carbonates.

[0022] It should be understood that the main component in the precipitate mother liquor is ammonium sulfate, which is formed by the reaction of ammonium ions and sulfate ions.

[0023] Therefore, the above embodiments, by using ammonium bicarbonate to adjust the pH value to 1.8–2.6, 4.8–5.4, and 6.8–7.0 in stages to perform multiple hydrolysis and precipitation, can improve the recovery rate of rare earth and the removal rate of impurities, thereby improving the purity of mixed rare earth oxides. The operation is simple and low-cost.

[0024] In some embodiments, the step of using ammonium bicarbonate as a neutralizing agent to perform multiple hydrolysis and precipitation of the high-alumina iron-containing acid leaching solution to remove impurities step by step and obtain mixed rare earth carbonates further includes: evaporating and crystallizing the precipitated mother liquor to obtain a second crystalline byproduct.

[0025] It should be noted that the second crystallization byproduct includes ammonium sulfate byproduct.

[0026] In some embodiments, the evaporation crystallization temperature is 50–100°C, and the evaporation crystallization time is 2–10 h.

[0027] In some embodiments, the pretreatment calcination temperature is 550–800°C, and the pretreatment calcination time is 1–4 hours.

[0028] In some embodiments, the concentration of the sulfuric acid is 5–8 mol / L.

[0029] In some embodiments, the acid leaching temperature is 70–99°C, and the acid leaching time is 0.5–4 hours.

[0030] In some embodiments, the thermal decomposition includes high-temperature calcination; wherein the high-temperature calcination temperature is 900–1100°C, and the high-temperature calcination time is 2–10 h.

[0031] It is worth noting that the resource utilization method provided in this disclosure can achieve comprehensive utilization of sedimentary rare earth minerals under efficient and low-cost conditions, which greatly improves the comprehensive development and utilization value of sedimentary rare earth minerals and helps the industrialization process of sedimentary rare earth mineral development and utilization.

[0032] First, the resource utilization method provided in this disclosure involves first performing blank roasting (i.e., pretreatment roasting), then using sulfuric acid as a leaching agent for acid leaching to obtain a high-alumina iron-containing acid leaching solution. Next, ammonium bicarbonate is used as a neutralizing agent for multiple hydrolysis and precipitation processes to produce mixed rare earth carbonates, which can be further used to obtain mixed rare earth oxides. In this process, multiple hydrolysis and precipitation processes are used to remove impurities step by step, which can improve the recovery rate of rare earths and the removal rate of impurities, thereby improving the purity of the mixed rare earth oxides. The method is simple to operate and has low cost.

[0033] Secondly, the resource utilization method provided in this disclosure can also produce ammonium alum by-products and ammonium sulfate by-products in stages, thereby improving the utilization rate of associated element aluminum while recovering and utilizing rare earth elements. In addition, the production of ammonium alum by-products and ammonium sulfate by-products can not only effectively reduce process costs, but also ensure that there is no ammonia nitrogen wastewater discharge throughout the entire process, solving the problem of difficult mother liquor disposal and meeting the requirements of green environmental protection.

[0034] The beneficial effects of this disclosure are:

[0035] 1. The resource utilization method of sedimentary rare earth ore provided in this disclosure can realize the comprehensive utilization of sedimentary rare earth ore under the conditions of high efficiency and low cost, which greatly improves the comprehensive development and utilization value of sedimentary rare earth ore and helps the industrialization process of sedimentary rare earth ore development and utilization.

[0036] 2. The present disclosure provides a method for the resource utilization of sedimentary rare earth minerals. By first using sulfuric acid as a leaching agent for acid leaching, and then using ammonium bicarbonate as a neutralizing agent for multiple hydrolysis and precipitation, a mixed rare earth oxide, ammonium alum by-product, and ammonium sulfate by-product can be produced. This method improves the utilization rate of associated element aluminum while recovering and utilizing rare earth elements, and solves the problem of difficult mother liquor disposal, thus meeting the requirements of green and environmental protection.

[0037] 3. The present disclosure provides a method for the resource utilization of sedimentary rare earth minerals, which improves the recovery rate of rare earth and the removal rate of impurities by multiple hydrolysis and precipitation to remove impurities in stages, thereby improving the purity of mixed rare earth oxides. The method is simple to operate and low in cost. Attached Figure Description

[0038] To more clearly illustrate the technical solutions in this disclosure, the accompanying drawings used in some embodiments of this disclosure will be briefly described below. Obviously, the drawings described below are merely drawings of some embodiments of this disclosure, and those skilled in the art can obtain other drawings based on these drawings. Furthermore, the drawings described below can be regarded as schematic diagrams and are not intended to limit the actual process of the methods involved in the embodiments of this disclosure.

[0039] Figure 1 This disclosure provides a process flow diagram for the resource utilization method of sedimentary rare earth minerals. Detailed Implementation

[0040] The technical solutions in some embodiments of this disclosure will be clearly and completely described below. Obviously, the described embodiments are only some embodiments of this disclosure, and not all embodiments. Based on the embodiments provided in this disclosure, all other embodiments obtained by those skilled in the art are within the scope of protection of this disclosure.

[0041] Hereinafter, the terms "first," "second," etc., are used for descriptive purposes only and should not be construed as indicating or implying relative importance or implicitly specifying the number of technical features indicated. Therefore, a feature defined with "first," "second," etc., may explicitly or implicitly include one or more of that feature.

[0042] Example 1

[0043] This embodiment provides a method for the resource utilization of sedimentary rare earth ore, which is used to process a sedimentary rare earth ore in Southwest China. The content of rare earth oxides (REO) in the sedimentary rare earth ore is 0.292%, and the quantitative results of its main minerals are shown in Table 1.

[0044] Table 1. Main mineral composition and content of the raw ore composite sample.

[0045] Mineral composition <![CDATA[Na2O]]> MgO <![CDATA[Al2O3]]> <![CDATA[SiO2]]> <![CDATA[P2O5]]> <![CDATA[SO2]]> Cl content(%) 0.077 0.078 35.795 41.069 0.081 0.026 0.018 Mineral composition <![CDATA[K2O]]> CaO <![CDATA[TiO2]]> <![CDATA[Fe2O3]]> NiO CuO TREO content(%) 0.143 0.024 5.145 1.673 0.023 0.023 0.292

[0046] like Figure 1 As shown, the resource utilization method includes:

[0047] S1. The raw ore is pretreated and roasted at 700℃ for 2 hours to obtain pretreated material;

[0048] S2. The pretreated material is acid leached with sulfuric acid at a concentration of 6 mol / L at a temperature of 95℃ for 2 hours. The solid and liquid are separated to obtain a high-alumina iron-containing acid leaching solution and leaching residue. The leaching residue can be directly stockpiled, backfilled, or used as an auxiliary material for building materials.

[0049] S3. Add ammonium bicarbonate to the high-alumina iron-containing acid leaching solution to adjust the pH to 2.2, and perform cooling crystallization at 5°C for 3 hours to remove aluminum impurities. Solid-liquid separation yields a low-alumina rare earth solution with low aluminum impurity content and a crude ammonium alum product. The crude ammonium alum product is then washed and purified to obtain an ammonium alum by-product. In this stage, the rare earth recovery rate is 96%, the aluminum removal rate is 92%, and the purity of the ammonium alum by-product is 99.2%.

[0050] S4. Add ammonium bicarbonate to the low-alumina rare earth solution to adjust the pH value to 5.0, further remove iron and aluminum impurities, and separate the solid and liquid to obtain an iron-aluminum-rare earth purification solution and a residue. The residue can be directly stockpiled, backfilled, or used as a building material auxiliary material. In this stage, the rare earth recovery rate is 90.8%, the aluminum removal rate is 99.6%, and the iron removal rate is 99.8%.

[0051] S5. Add ammonium bicarbonate to the iron, aluminum and rare earth purification solution to adjust the pH value to 7.0, precipitate rare earth, separate the solid and liquid to obtain mixed rare earth carbonate and precipitate mother liquor, then evaporate and crystallize the precipitate mother liquor at 70℃ for 3 hours to obtain ammonium sulfate byproduct with a total nitrogen content of 20.6%; in this stage, the rare earth recovery rate is 99.7%;

[0052] S6. The mixed rare earth carbonates were calcined at high temperature of 1000℃ for 4 hours to obtain mixed rare earth oxides with a REO content of 92.2%.

[0053] Example 2

[0054] This embodiment provides a method for the resource utilization of sedimentary rare earth ore, which is used to process a sedimentary rare earth ore in Southwest China. The content of rare earth oxides (REO) in the sedimentary rare earth ore is 0.236%, and the quantitative results of its main minerals are shown in Table 2.

[0055] Table 2. Main mineral composition and content of the raw ore composite sample

[0056] Mineral composition <![CDATA[Na2O]]> MgO <![CDATA[Al2O3]]> <![CDATA[SiO2]]> <![CDATA[P2O5]]> <![CDATA[SO2]]> Cl content(%) 0.223 0.377 35.338 42.019 0.079 0.0275 0.005 Mineral composition <![CDATA[K2O]]> CaO <![CDATA[TiO2]]> <![CDATA[Fe2O3]]> NiO CuO TREO content(%) 4.385 0.038 0.739 1.424 0.01 0.024 0.236

[0057] like Figure 1 As shown, the resource utilization method includes:

[0058] S1. The raw ore is pretreated and roasted at 550℃ for 4 hours to obtain pretreated material;

[0059] S2. The pretreated material is acid leached with sulfuric acid at a concentration of 5 mol / L at a temperature of 99℃ for 0.5 h. The solid and liquid are separated to obtain a high-alumina iron-containing acid leaching solution and leaching residue. The leaching residue can be directly stockpiled, backfilled, or used as an auxiliary material for building materials.

[0060] S3. Add ammonium bicarbonate to the high-alumina iron-containing acid leaching solution to adjust the pH to 1.8, and perform cooling crystallization at -5℃ for 6 hours to remove aluminum impurities. Solid-liquid separation yields a low-alumina rare earth solution with low aluminum impurity content and a crude ammonium alum product. The crude ammonium alum product is then washed and purified to obtain an ammonium alum by-product. In this stage, the rare earth recovery rate is 95%, the aluminum removal rate is 94%, and the purity of the ammonium alum by-product is 99.5%.

[0061] S4. Add ammonium bicarbonate to the low-alumina rare earth solution to adjust the pH value to 4.8, further remove iron and aluminum impurities, and separate the solid and liquid to obtain an iron-aluminum-rare earth purification solution and a residue. The residue can be directly stockpiled, backfilled, or used as a building material auxiliary material. In this stage, the rare earth recovery rate is 92.6%, the aluminum removal rate is 98.5%, and the iron removal rate is 99.7%.

[0062] S5. Add ammonium bicarbonate to the iron, aluminum and rare earth purification solution to adjust the pH value to 6.8, precipitate rare earth, separate the solid and liquid to obtain mixed rare earth carbonate and precipitate mother liquor, then evaporate and crystallize the precipitate mother liquor at 100℃ for 2 hours to obtain ammonium sulfate byproduct with a total nitrogen content of 20.7%; in this stage, the rare earth recovery rate is 98.5%;

[0063] S6. The mixed rare earth carbonates were calcined at high temperature (900℃) for 10 hours to obtain mixed rare earth oxides with a REO content of 90.6%.

[0064] Example 3

[0065] This embodiment provides a method for the resource utilization of sedimentary rare earth ore, which is used to process a sedimentary rare earth ore in Southwest China. The content of rare earth oxides (REO) in the sedimentary rare earth ore is 0.425%, and the quantitative results of its main minerals are shown in Table 3.

[0066] Table 3. Main mineral composition and content of the raw ore composite sample

[0067] Mineral composition <![CDATA[Na2O]]> MgO <![CDATA[Al2O3]]> <![CDATA[SiO2]]> <![CDATA[P2O5]]> <![CDATA[SO2]]> Cl content(%) 0.070 0.132 35.740 40.826 0.292 0.044 0.015 Mineral composition <![CDATA[K2O]]> CaO <![CDATA[TiO2]]> <![CDATA[Fe2O3]]> NiO CuO TREO content(%) 0.077 0.094 4.766 1.383 0.01 0.011 0.425

[0068] like Figure 1 As shown, the resource utilization method includes:

[0069] S1. The raw ore is pretreated and roasted at 800℃ for 1 hour to obtain pretreated material;

[0070] S2. The pretreated material is acid leached with sulfuric acid at a concentration of 8 mol / L at a temperature of 70℃ for 4 hours. The solid and liquid are separated to obtain a high-alumina iron-containing acid leaching solution and leaching residue. The leaching residue can be directly stockpiled, backfilled, or used as an auxiliary material for building materials.

[0071] S3. Add ammonium bicarbonate to the high-alumina iron-containing acid leaching solution to adjust the pH to 2.6, and perform cooling crystallization at 30°C for 1 hour to remove aluminum impurities. Solid-liquid separation yields a low-alumina rare earth solution with low aluminum impurity content and a crude ammonium alum product. The crude ammonium alum product is then washed and purified to obtain an ammonium alum by-product. In this stage, the rare earth recovery rate is 92%, the aluminum removal rate is 91%, and the purity of the ammonium alum by-product is 98.8%.

[0072] S4. Add ammonium bicarbonate to the low-alumina rare earth solution to adjust the pH value to 5.4, further remove iron and aluminum impurities, and separate the solid and liquid to obtain an iron-aluminum-rare earth purification solution and a residue. The residue can be directly stockpiled, backfilled, or used as a building material auxiliary material. In this stage, the rare earth recovery rate is 87.9%, the aluminum removal rate is 99.7%, and the iron removal rate is 99.9%.

[0073] S5. Add ammonium bicarbonate to the iron, aluminum and rare earth purification solution to adjust the pH value to 6.9, precipitate rare earth, separate the solid and liquid to obtain mixed rare earth carbonate and precipitate mother liquor, then evaporate and crystallize the precipitate mother liquor at 50℃ for 10h to obtain ammonium sulfate byproduct with a total nitrogen content of 20.9%; in this stage, the rare earth recovery rate is 98.9%;

[0074] S6. The mixed rare earth carbonates were calcined at high temperature of 1100℃ for 2 hours to obtain mixed rare earth oxides with a REO content of 92.8%.

[0075] Compare with Example 1

[0076] Example 1 and Comparative Example 1 were compared. The difference between Comparative Example 1 and Example 1 is as follows:

[0077] 1) Delete S3 to S4;

[0078] 2) Replace the iron, aluminum and rare earth purification solution in S5 with a high-alumina iron-containing acid leaching solution;

[0079] Other conditions, such as the selection of raw ore, the selection and dosage of remaining reagents, and the remaining process flow and condition parameters, are the same as in Example 1 (this comparative example, compared to Example 1, only involves one hydrolysis and precipitation, which is used to demonstrate that the resource utilization method disclosed herein is more effective).

[0080] The results showed that in S5, the rare earth recovery rate was 99.7%, and the total nitrogen content in the ammonium sulfate byproduct was 20.8%; in S6, the REO content in the mixed rare earth oxides was 0.2%.

[0081] As can be seen, since Comparative Example 1 did not perform stepwise impurity removal, but directly adjusted the pH value of the high-alumina iron-containing acid leaching solution to 7.0, aluminum, iron and rare earth co-precipitated, resulting in extremely high iron and aluminum impurity content and extremely low REO content in the mixed rare earth oxides.

[0082] Compare with Example 2

[0083] Example 1 and Comparative Example 2 were compared. The difference between Comparative Example 2 and Example 1 is as follows:

[0084] 1) Delete S3;

[0085] 2) Replace the low-alumina rare earth solution in S4 with a high-alumina iron-containing acid leaching solution;

[0086] Other conditions, such as the selection of raw ore, the selection and dosage of remaining reagents, and the remaining process flow and condition parameters, are the same as in Example 1 (this comparative example, compared to Example 1, only involves two hydrolysis and precipitation processes to demonstrate that the resource utilization method disclosed herein is more effective).

[0087] The results showed that in S4, the rare earth recovery rate was 30.6%, the aluminum removal rate was 99.4%, and the iron removal rate was 99.9%; in S5, the rare earth recovery rate was 99.8%, and the total nitrogen content in the ammonium sulfate byproduct was 20.4%; in S6, the REO content in the mixed rare earth oxides was 92.1%.

[0088] It can be seen that, since the pH value of the high-alumina iron-containing acid leaching solution was directly adjusted to 4.8 in Comparative Example 2, the recovery rate of rare earth in S4 was greatly reduced. This indicates that a large amount of rare earth was wasted because it was not recovered in S4, which led to a significant reduction in the total amount of mixed rare earth oxides.

[0089] Compare with Example 3

[0090] Example 1 and Comparative Example 3 were compared. The difference between Comparative Example 3 and Example 1 is as follows:

[0091] 1) Delete S4;

[0092] 2) Replace the iron, aluminum and rare earth purification solution in S5 with a low-aluminum rare earth solution.

[0093] Other conditions, such as the selection of raw ore, the selection and dosage of remaining reagents, and the remaining process flow and condition parameters, are the same as in Example 1 (this comparative example, compared to Example 1, only involves two hydrolysis and precipitation processes to demonstrate that the resource utilization method disclosed herein is more effective).

[0094] The results showed that in S3, the rare earth recovery rate was 96%, the aluminum removal rate was 92%, and the purity of the ammonium alum by-product was 99.1%; in S5, the rare earth recovery rate was 99.7%, and the total nitrogen content in the ammonium sulfate by-product was 20.6%; in S6, the REO content in the mixed rare earth oxides was 8.33%.

[0095] It can be seen that in Comparative Example 3, after adjusting the pH of the high-alumina iron-containing acid leaching solution to 2.2, the pH of the low-alumina rare earth solution was directly adjusted to 6.8. That is, Comparative Example 3 did not perform deep impurity removal of aluminum and iron, resulting in extremely low REO content in rare earth oxides.

[0096] Compare with Example 4

[0097] Example 1 and Comparative Example 4 were compared. The difference between Comparative Example 4 and Example 1 is as follows:

[0098] Replace ammonium bicarbonate in S3 to S5 with sodium bicarbonate;

[0099] Other conditions, such as the selection of raw ore, the selection and dosage of remaining reagents, and the remaining process flow and condition parameters, are the same as in Example 1 (compared to Example 1, this comparative example replaces ammonium bicarbonate with sodium bicarbonate to demonstrate that the resource utilization method of this disclosure is more effective).

[0100] The results showed that in S3, the rare earth recovery rate was 96%, the aluminum removal rate was 82%, and the purity of the sodium alum by-product was 98.7%; in S4, the rare earth recovery rate was 71.4%, the aluminum removal rate was 99.6%, and the iron removal rate was 99.9%; in S5, sodium sulfate was obtained by evaporation and crystallization; and in S6, the REO content in the mixed rare earth oxides was 89.8%.

[0101] As can be seen, in Comparative Example 4, replacing ammonium bicarbonate with sodium bicarbonate resulted in a decrease in the aluminum removal rate and a poorer impurity removal effect in S3 due to the high solubility of sodium alum byproducts. This led to a high residual amount of aluminum impurities in the resulting rare earth solution, which in turn increased the amount of rare earths carried over during the deep removal of aluminum and iron impurities in S4, thus significantly reducing the rare earth recovery rate. At the same time, the introduction of sodium ions during the precipitation of rare earths in S5 also reduced the purity of the final product, mixed rare earth oxides. Furthermore, the practicality of sodium alum byproducts and sodium sulfate was lower than that of ammonium alum byproducts and ammonium sulfate, respectively.

[0102] It should be noted that, compared with Example 1, in Comparative Example 4, the total amount of aluminum in the obtained rare earth solution increases due to the decrease in the aluminum removal rate in S3. On this basis, even though the aluminum removal rate in S4 is as high as 99.6% (i.e., equal to Example 1), the sodium content in the iron-aluminum rare earth purification solution increases due to the significant decrease in the rare earth recovery rate, which also leads to a decrease in the purity of the mixed rare earth oxides.

[0103] Therefore, the resource utilization method for sedimentary rare earth minerals provided in this disclosure not only improves the rare earth recovery rate, the utilization rate of associated elements and the impurity removal rate, but also solves the problem of difficult mother liquor disposal.

[0104] The above description is merely a preferred embodiment of this disclosure. It should be understood that this disclosure is not limited to the forms disclosed herein and should not be construed as excluding other embodiments. It can be used in various other combinations, modifications, and environments, and can be altered within the scope of the concept described herein through the above teachings or techniques or knowledge in related fields. Any modifications and variations made by those skilled in the art that do not depart from the spirit and scope of this disclosure should be within the protection scope of the appended claims.

Claims

1. A method for resource utilization of a deposit-type rare earth ore, characterized by, The method comprises the following steps: carrying out pretreatment roasting on raw ore of a sedimentary type rare earth ore to obtain pretreated material; carrying out acid leaching on the pretreated material by using sulfuric acid as a leaching agent to separate a high-aluminum and iron-containing acid leaching solution and a leaching residue; carrying out multiple hydrolysis precipitations on the high-aluminum and iron-containing acid leaching solution by using ammonium bicarbonate as a neutralizing agent to remove impurities step by step to obtain mixed carbonic rare earth; and carrying out thermal decomposition on the mixed carbonic rare earth to obtain mixed rare earth oxide. The multiple hydrolysis precipitations on the high-aluminum and iron-containing acid leaching solution by using ammonium bicarbonate as a neutralizing agent to remove impurities step by step to obtain mixed carbonic rare earth comprises the following steps: adding the ammonium bicarbonate to the high-aluminum and iron-containing acid leaching solution, adjusting the pH value to 1.8-2.6, carrying out cooling crystallization, and separating a low-aluminum rare earth solution and a first crystallization by-product; wherein the cooling crystallization is carried out at a temperature of-5-30 ℃ for 1-6 h; adding the ammonium bicarbonate to the low-aluminum rare earth solution, adjusting the pH value to 4.8-5.4, and separating an iron and aluminum-removed rare earth purification solution and an impurity-removing residue; and adding the ammonium bicarbonate to the iron and aluminum-removed rare earth purification solution, adjusting the pH value to 6.8-7.0, and separating the mixed carbonic rare earth and a precipitation mother liquor.

2. The method of claim 1, wherein, The multiple hydrolysis precipitations on the high-aluminum and iron-containing acid leaching solution by using ammonium bicarbonate as a neutralizing agent to remove impurities step by step to obtain mixed carbonic rare earth further comprises the following step: carrying out evaporation crystallization on the precipitation mother liquor to obtain a second crystallization by-product.

3. The method of claim 2, wherein, The evaporation crystallization is carried out at a temperature of 50-100 ℃ for 2-10 h.

4. The method of claim 1, wherein, The pretreatment roasting is carried out at a temperature of 550-800 ℃ for 1-4 h.

5. The method of claim 1, wherein, The concentration of the sulfuric acid is 5-8 mol / L.

6. The method of resource utilization of claim 1, wherein, The acid leaching is carried out at a temperature of 70-99 ℃ for 0.5-4 h.

7. The method of claim 1, wherein, The thermal decomposition comprises high-temperature roasting; wherein the high-temperature roasting is carried out at a temperature of 900-1100 ℃ for 2-10 h.

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