A method for recovering co-associated low-grade valuable elements in lepidolite flotation concentrate by using normal-pressure alkali leaching method

Abstract

The application discloses a method for recovering co-associated low-grade valuable elements in a lithium mica flotation concentrate by using a normal-pressure alkali leaching method, and belongs to the technical field of lithium, rubidium and cesium resource recovery and recycling. The normal-pressure alkali leaching method is used to recover lithium, rubidium and cesium and other valuable elements in a lithium mica flotation concentrate, and the green, efficient and high synergistic leaching rate of lithium, rubidium and cesium is achieved. In view of the problems of resource waste, high energy consumption and secondary environmental pollution caused by the prior art, a high-efficiency and synergistic recovery of lithium, rubidium and cesium in lithium mica ore is achieved by using the high reactivity provided by unconventional alkali. The process flow of the application has the advantages of short time consumption, simple and safe equipment operation, high synergistic recovery rate of lithium, rubidium and cesium, good impurity removal effect, reduced secondary pollution, medium recycling in the subsequent process, greatly reduced lithium slag quantity in the subsequent process, 80% reduction of lithium slag quantity generated per ton of lithium mica ore, and provision of a new theoretical reference for low-grade co-associated lithium, rubidium and cesium resource extraction and resource waste reduction in China.

Description

TECHNICAL FIELD

[0001] The application belongs to the technical field of lithium-rubidium-cesium resource recovery and recycling, and particularly relates to a method for recovering co-associated low-grade valuable elements in lepidolite flotation concentrate by using a normal-pressure alkali leaching method. BACKGROUND

[0002] Lepidolite contains various rare metal resources such as lithium, rubidium and cesium. Lithium, rubidium and cesium are important strategic metal resources and are widely used in military and other fields as "new energy" and "new materials". In addition, lepidolite has low mining cost and rich domestic reserves. In order to meet the demand of high-tech industry for lithium, rubidium and cesium and other rare metals, how to efficiently and economically extract lithium, rubidium and cesium and other rare metals from lepidolite has become a prominent problem. Reasonable development and utilization of co-associated low-grade lithium, rubidium and cesium and other valuable elements in lepidolite can reduce resource waste, realize rational utilization of resources and take the road of green and sustainable development.

[0003] Currently, extracting lithium, rubidium and cesium from lepidolite has high economic value, but the current technical method for extracting and recovering lithium, rubidium and cesium has the problems of difficulty in co-dissolution of lithium, rubidium and cesium, difficulty in purification, enrichment and separation of rubidium and cesium solution, and inability to recycle resources, resulting in resource waste and other key technical bottlenecks. In addition, the existing technology also has the problems of high energy consumption of process flow and generation of waste gas in the process, which harms the environment and conflicts with the current "double carbon" target. It is necessary to research more green and low-carbon treatment technologies to realize high-quality utilization of valuable elements such as lithium, rubidium and cesium in lepidolite ore.

[0004] Lithium, rubidium and cesium are mainly extracted and recovered from pegmatite and salt lake brine, but the content and concentration of rubidium and cesium in salt lake brine are low and have no independent development and utilization value with the current technology. Pegmatite represented by spodumene and lepidolite has simple ore composition and high lithium content with Li2O grade of 6% to 8%. However, the domestic spodumene mineral reserves are not high and mainly rely on imports, which cannot be used for large-scale industrial production. Although the lithium, rubidium and cesium content of lepidolite is low and the extraction process has the problem of resource waste caused by the inability to utilize valuable elements, lepidolite has low development cost and large reserves, which has high economic value. If lithium, rubidium and cesium and other valuable elements in lepidolite can be efficiently developed and utilized, not only the production cost can be greatly reduced, but also the supply of lithium, rubidium and cesium resources can be ensured to support the development of emerging industries and realize sustainable development of resources.

[0005] Lithium, rubidium and cesium in lepidolite mainly exist in the form of oxides, and the mineral contains 2%-4% F, which can form insoluble fluoride and affect the leaching rate of lithium, rubidium and cesium. At present, the acid method, alkali method, salt method and pressure cooking method are mainly used to recover valuable elements such as lithium, rubidium and cesium. For example, Yubo Liu et al. used the acid method to determine the experimental parameters through thermodynamic calculation simulation, so that the extraction rates of Li, Rb and Cs reached 90.5%, 91.2% and 89.4% respectively, realizing the selective extraction of lithium, rubidium and cesium. ZHANG used the sulfuric acid method, so that Li, Rb and Cs existed in the form of soluble sulfate in the leaching solution, and the cooperative leaching rate of Li, Rb and Cs reached more than 95%, but there were disadvantages such as difficulty in deep removal of impurities and serious corrosion of equipment. MULWANDA et al. used the mixed alkali method, using NaOH and Ca(OH)2 as mixed leaching agents, through pressure leaching, water washing, precipitation separation and other processes, the leaching rates of Li, Rb and Cs reached 94%, 96% and 90% respectively. Because the impurity cations in the solution can cause coprecipitation, affecting the purity of valuable elements, and the concentrated alkali waste liquid is difficult to recover. Guo Chunping et al. used the sulfate method, mixed potassium sulfate, calcium sulfate and barium sulfate as mixed sulfates, mixed with lepidolite at a ratio of 1:1, and then subjected to high-temperature roasting at 900°C and dilute acid leaching. The leaching rate of lithium was 92.2%, although the leaching solution could directly precipitate 0-grade lithium carbonate, but the barium in the mixed salt was similar in nature to rubidium and cesium and could only replace part of rubidium and cesium, so the leaching rates of rubidium and cesium were only 61.5% and 63.8% respectively. Yan et al. and Qiu Shiyuan et al. used lime milk pressure cooking method and salt pressure cooking method respectively, both of which used high temperature and high pressure conditions to pressure cook and dissolve valuable elements such as lithium, rubidium and cesium, which had harsh process conditions and was limited in industrialization. Jinliang Wang et al. used the mixed salt roasting method, mixed Na2SO4 and CaCl2, and used a thermodynamic model to simulate and calculate the optimal reaction conditions. After roasting and filtration separation, the extraction efficiencies of Li, Rb and Cs were 98.70%, 97.27% and 98.40% respectively. Yan et al. used the mixed salt roasting method, and after water leaching at room temperature, five-stage countercurrent washing and other processes, the leaching rates of Li, Rb and Cs reached 94.8%, 93.5% and 90.1% respectively. The salt method has high extraction efficiency, but the process produces toxic and harmful gases, which conflicts with the current “double carbon” goal, so it is necessary to develop more green and low-carbon technologies to realize the efficient and cooperative extraction and recovery of valuable elements in lepidolite.

[0006] Based on the above reasons, the present application is proposed. SUMMARY

[0007] The lithium mica flotation concentrate has complex composition, stable structure, and rubidium, cesium and lithium are difficult to be simultaneously and efficiently leached. The existing pressure boiling method, sulfuric acid / salt roasting method and other associated lithium, rubidium and cesium resources have a series of key technical bottleneck problems such as poor resource endowment, difficulty in collaborative leaching of lithium, rubidium and cesium, difficulty in purification, enrichment and separation of lithium, rubidium and cesium solution, and easy to cause environmental pollution. The present application provides a method for recovering associated low-grade valuable elements in lithium mica flotation concentrate by using atmospheric alkali leaching method, which solves or at least partially solves the above technical defects existing in the prior art: the present application uses the high chemical reactivity provided by the sub-molten salt unconventional alkali to realize the phase transformation of the lithium mica flotation concentrate under atmospheric pressure and recover the valuable elements such as lithium, rubidium and cesium in it, improve the utilization efficiency of associated rubidium and cesium resources, and realize the efficient and green leaching of lithium, rubidium and cesium resources.

[0008] In order to achieve the above-mentioned purposes of the present application, the technical solutions adopted by the present application are as follows:

[0009] A method for recovering associated low-grade valuable elements in lithium mica flotation concentrate by using atmospheric alkali leaching method, the method specifically comprises the following steps:

[0010] S1, roasting the lithium mica flotation concentrate;

[0011] S2, adding alkali medium to the roasted lithium mica flotation concentrate, and placing the obtained mixed reactants in an atmospheric pressure reaction container, heating to 120-300 DEG C constant temperature reaction for 1-10h;

[0012] S3, water washing the product obtained after S2 reaction, solid-liquid separation, to obtain leaching solution and leaching residue;

[0013] S4, adjusting the pH value of the leaching solution obtained in S3 to 7.5-10, filtering and removing impurities, heating the leaching solution after removing impurities to 50-100 DEG C, then adding a precipitating agent, and continuing constant temperature reaction until no crystals are precipitated; filtering to obtain valuable element crystal precipitate and filtrate;

[0014] S5, evaporating and concentrating the filtrate in S4, then repeating steps S1-S4 to extract valuable elements for the second time, and repeating the cycle.

[0015] Further, in the above technical solution, the valuable elements are at least one of lithium, rubidium, cesium and the like.

[0016] Further, in the above technical solution, the roasting temperature in step S1 is 800-1000 DEG C, and the roasting time is 1-6h.

[0017] Further, in the above technical solution, the alkali medium in step S2 is an alkali water mixed solution composed of alkaline substances and deionized water. The mass concentration of the alkali medium is 50%-80%.

[0018] Further, the above technical solution, the mass ratio of the lithium mica flotation concentrate and the alkali medium in step S2 is 1:1-5.

[0019] Further, the above technical solution, the alkali substance is one or more of sodium hydroxide, potassium hydroxide, aluminum hydroxide, sodium peroxide, and calcium hydroxide.

[0020] Further, the above technical solution, hydrochloric acid, sulfuric acid, oxalic acid or nitric acid is added to adjust the pH value in step S4.

[0021] Further, the above technical solution, the precipitant in step S4 is at least one of sodium carbonate solution or sodium bicarbonate solution.

[0022] Further, the above technical solution, the mass concentration of the sodium carbonate solution or the sodium bicarbonate solution is 30%-50%.

[0023] Compared with the prior art, the present application has the following beneficial effects:

[0024] (1) The present application uses atmospheric pressure alkali leaching method to recover valuable elements such as lithium, rubidium and cesium in lithium mica flotation concentrate, and the green, efficient and high leaching rate of lithium, rubidium and cesium is achieved. In view of the problems of resource waste, high energy consumption and environmental secondary pollution caused by the prior art, a high reaction activity provided by unconventional alkali medium is used to achieve efficient and synergistic recovery of lithium, rubidium and cesium in lithium mica flotation concentrate. The process flow of the present application has less time consumption, the equipment operation is simple and safe, the synergistic recovery rate of lithium, rubidium and cesium is high, the impurity removal effect is good, the secondary pollution is reduced, the medium can be recycled in the subsequent process, the amount of lithium slag in the subsequent process is sharply reduced, the amount of lithium slag generated per ton of lithium mica ore is reduced by 80%, and a new theoretical reference is provided for the extraction of low-grade co-associated lithium, rubidium and cesium resources in China and the reduction of resource waste.

[0025] (2) The unconventional alkali medium used in the present application (the unconventional medium is an alkali solution with a concentration different from the alkali solution, and the medium with an alkali concentration higher than 50% is a sub-molten salt medium, which is an intermediate medium between solution and molten salt, and has high reaction activity) has higher reaction activity and activity coefficient, and the use of high-concentration alkali medium with an alkali concentration greater than 50% reduces the wrapping on the surface of lithium fluoride, so that the reaction can continue. The large amount of active oxygen components in the unconventional alkali medium is the core of its high chemical reaction activity, and the active oxygen in the medium can cause homologous substitution of O on the surface of the mineral, causing lattice distortion and gradually destroying the mineral structure, thereby achieving leaching of useful elements in the mineral particles. Compared with the traditional pressure cooking method, acid / alkali / salt method, the high chemical reaction provided by the unconventional alkali medium makes the alkali medium more advantageous in thermodynamics and kinetics, in addition, the unconventional alkali medium can be recycled to reduce resource waste, and the process is green and efficient. BRIEF DESCRIPTION OF DRAWINGS

[0026] In order to more clearly illustrate the technical solutions in the embodiments of the present application or the prior art, the following will briefly introduce the drawings needed to be used in the embodiments or prior art description. Obviously, the drawings in the following description only constitute some embodiments of the present application, and for those skilled in the art, other drawings can also be obtained from these drawings without creative labor.

[0027] Figure 1 The process flow chart of the method for recovering the co-associated low-grade valuable elements in the lepidolite flotation concentrate by using the normal-pressure alkali leaching method. DETAILED DESCRIPTION

[0028] The present application proposes to recover lithium, rubidium, cesium and other valuable elements in lepidolite flotation concentrate by using a sub-molten salt method, so as to reduce energy consumption, improve the recovery rate of valuable elements, and realize green, low-carbon and efficient leaching of lithium, rubidium, cesium and other valuable elements in lepidolite flotation concentrate.

[0029] The technical solutions in the embodiments of the present application will be described clearly and completely in combination with the drawings in the embodiments of the present application. Obviously, the described embodiments only constitute some embodiments of the present application, rather than all the embodiments. Based on the embodiments in the present application, all other embodiments obtained by those skilled in the art without creative labor fall within the scope of protection of the present application.

[0030] It should be noted that the professional terms used in the present application are only for the purpose of describing the specific embodiments, and are not intended to limit the scope of protection of the present application. Unless otherwise specified, the various raw materials, reagents, instruments and equipment used in the following embodiments of the present application can be purchased from the market or prepared by the existing method.

[0031] Embodiment 1

[0032] The source of the lepidolite flotation concentrate in this embodiment is as follows: from Jiangxi Yichun tantalum-niobium mine, the content of lithium, rubidium and cesium oxides Li2O, Rb2O and Cs2O is 2.88%, 0.55% and 0.051% respectively. The lepidolite ore is subjected to sorting and pre-enrichment treatment to obtain lepidolite flotation concentrate; wherein:

[0033] The sorting and pre-enrichment treatment is specifically as follows: a pre-desliming-flotation combined process is adopted, the lepidolite ore is ground to 15-18um and then subjected to pre-desliming treatment by a hydrocyclone, then a flotation process flow is adopted, hydrogen fluoride (200g / t) is added according to the mineral quality to activate the mineral surface, the mineral surface is treated by washing and adding caustic soda (10Kg / t) after activation for 20min, and then oleic acid collector is added to make the lepidolite float, thereby obtaining the lepidolite flotation concentrate.

[0034] A method for recovering co-associated low-grade valuable elements in lepidolite flotation concentrate by using atmospheric alkali leaching method, comprising the following steps:

[0035] S1, put the lepidolite flotation concentrate into the muffle furnace for high temperature roasting at 900℃ for 2 hours, then take 10g of the roasted lepidolite concentrate, add sodium hydroxide sub-molten salt alkali medium made of 15g sodium hydroxide and 10ml deionized water, mix uniformly, and place in an atmospheric pressure reaction container, react at 190℃ for 5 hours to obtain a reaction product;

[0036] S2, wash the reaction product obtained in S1 with tap water to 600ml and dilute to 1000ml to obtain an alkaline leaching solution and a leaching residue;

[0037] S3, add 30wt% hydrochloric acid solution to the alkaline leaching solution obtained in S2 to adjust the pH to 8.5, filter out impurities, heat the impurity-removed leaching solution to 60℃, then add 2.5ml of 50% Na2CO3 solution (precipitating agent), continue to react at 60℃ for 1h until the crystals are completely precipitated; filter to obtain carbonate crystal precipitate and filtrate;

[0038] S4, evaporate and concentrate the filtrate, then perform S1, S2 and S2 steps in sequence to perform secondary extraction of lithium, rubidium and cesium valuable elements.

[0039] Example 2

[0040] The source of the lepidolite flotation concentrate in this example is as follows: a certain place lepidolite ore content, in which the content of lithium, rubidium and cesium oxides Li2O, Rb2O and Cs2O is 3.27%, 1.02% and 0.08% respectively.

[0041] S1, the lepidolite ore is subjected to pre-enrichment treatment by sorting to obtain lepidolite flotation concentrate; wherein:

[0042] The pre-enrichment treatment by sorting is specifically as follows: it is determined to use pre-desliming-flotation combined process, the lepidolite ore is ground to 15-18um, then pre-desliming treatment is performed by hydrocyclone, then flotation process is used, Ca(OCl)2(200g / t) is added according to the mineral quality to activate the mineral surface, after 15min of activation, caustic soda (10Kg / t) is added to treat the mineral surface, and then oleic acid collector is added to make lepidolite float to obtain lepidolite flotation concentrate.

[0043] A method for recovering co-associated low-grade valuable elements in lepidolite flotation concentrate by using atmospheric alkali leaching method, comprising the following steps:

[0044] The lithium mica flotation concentrate is placed in a muffle furnace for high-temperature roasting at 600 DEG C for 4 hours, then 15g of the roasted lithium mica concentrate is added into an alkali medium prepared by 10g of sodium hydroxide, 10g of potassium hydroxide and 10g of deionized water, mixed uniformly, and placed in an atmospheric pressure reaction container for reaction at 220 DEG C for 7 hours to obtain a reaction product;

[0045] S2, the reaction product obtained in S1 is washed with deionized water to 200ml and diluted to 500ml to obtain an alkali leaching solution and a leaching residue;

[0046] S3, a 40wt% nitric acid solution is added to the alkali leaching solution obtained in S2 to adjust the pH to 8, and after filtration, impurities are removed, and the impurity-removed leaching solution is heated to 50 DEG C, then 1.5ml of a 60wt% Na2CO3 solution (precipitating agent) is added, and the reaction is continued at 50 DEG C for 2 hours until the crystals are completely precipitated; filtration is performed to obtain carbonate crystal precipitate and filtrate;

[0047] S4, the filtrate is evaporated and concentrated, and then the steps S1, S2 and S3 are sequentially performed to perform secondary extraction of lithium, rubidium and cesium valuable elements.

[0048] Example 3

[0049] The lithium mica flotation concentrate of this example is as follows:

[0050] From a certain concentrator in Jiangxi, the content of lithium, rubidium and cesium oxides Li2O, Rb2O and Cs2O is 2.86%, 1.75% and 0.36% respectively

[0051] S1, the lithium mica ore is sequentially sorted and pre-enriched to obtain a lithium mica flotation concentrate and a tailing; wherein:

[0052] The pre-enrichment sorting treatment is as follows: a pre-desliming-flotation combined process is determined, the lithium mica ore is ground to 15-18um, then pre-desliming treatment is performed by a hydrocyclone, then a flotation process is adopted, sodium silicate (200g / t) is added according to the mineral quality to activate the mineral surface, the activation is performed for 25 minutes, then washing is performed, hydrochloric acid (10Kg / t) is added to treat the mineral surface to adjust the pH of the slurry to 3-9, then oleic acid collector is added to make the lithium mica float to obtain the lithium mica flotation concentrate.

[0053] The method for recovering the co-occurring low-grade valuable elements in the lithium mica flotation concentrate by atmospheric pressure alkali leaching method in this example includes the following steps:

[0054] The lithium mica flotation concentrate is placed in a muffle furnace for high-temperature roasting at 850 DEG C for 3 hours. 5g of the roasted lithium mica flotation concentrate is added to an alkali medium prepared from 20g of sodium hydroxide, 1g of aluminum hydroxide and 15g of deionized water, mixed uniformly, and placed in an atmospheric pressure reaction container for reaction at 180 DEG C for 5h to obtain a reaction product;

[0055] S2, the reaction product obtained in S1 is washed with deionized water to 600ml and diluted to 1000ml, to obtain an alkaline leaching solution and a leaching residue;

[0056] S3, a 30wt% nitric acid solution is added to the alkaline leaching solution obtained in S2 to adjust the pH to 8.5, and after filtration, impurities are removed, and the impurity-removed leaching solution is heated to 80 DEG C, and then 1ml of a 30wt% Na2CO3 solution (precipitating agent) is added, and the reaction is continued at 80 DEG C for 1h until the crystals are completely precipitated; filtration is performed to obtain carbonate crystal precipitate and filtrate;

[0057] S4, the filtrate is evaporated and concentrated, and then the steps S1, S2 and S3 are sequentially performed to perform secondary extraction of lithium, rubidium and cesium valuable elements.

[0058] The lithium, rubidium and cesium valuable element precipitate obtained in Examples 1-3 is analyzed, and the results are shown in Table 1.

[0059] Table 1 Product quality of lithium, rubidium and cesium precipitate

[0060]

[0061] As can be seen from Table 1, the lithium, rubidium and cesium precipitate obtained by appropriately adjusting the implementation scheme of the patent examples 1-3 within the scope of the patent has a product quality of more than 90%, and the lithium, rubidium and cesium valuable element recovery rate is high. The same implementation scheme as that adopted in Example 1 has the best effect, and the process flow of the present application is simple, the process is green and low-carbon, and the sub-molten salt medium is recycled, reducing resource waste, and finally realizing green and efficient extraction of lithium mica co-associated low-grade lithium, rubidium and cesium.

[0062] It should be noted that when numerical ranges are involved in the present application, both endpoints of each numerical range and any number between the two endpoints can be selected. Since the same steps and examples are used, in order to prevent repetition, the present application describes preferred examples of appropriate changes in the implementation scheme within the scope of the invention. Although preferred examples of the present application have been described, those skilled in the art can make further changes and modifications to these examples once they understand the basic inventive concept. Therefore, the appended claims are intended to include the preferred examples and all changes and modifications falling within the scope of the present application.

[0063] Obviously, many modifications and variations of the present application are possible in light of the above teachings. It is, therefore, to be understood that within the scope of the appended claims and their equivalents, the application can be practiced otherwise than as specifically described.

Claims

1. A method for recovering co-associated low-grade valuable elements from lepidolite flotation concentrate using an atmospheric pressure alkaline leaching process, characterized in that: The method specifically comprises the following steps: S1, take the lepidolite flotation concentrate and put it into a muffle furnace for high-temperature roasting at 900 ° C for 2 hours, then take 10g of the roasted lepidolite concentrate, add 15g of sodium hydroxide and 10ml of deionized water into a sodium hydroxide sub-molten salt alkaline medium, mix well, and place it in a normal pressure reaction vessel, and react at 190 ° C for 5 hours to obtain a reaction product; S2. Rinse the reaction product obtained in S1 with tap water until the volume reaches 600 ml, filter and adjust the volume to 1000 ml to obtain an alkaline leachate and leach residue; S3, adding a 30 wt% hydrochloric acid solution to the alkaline leachate obtained in S2 to adjust the pH to 8.5, filtering and removing impurities, heating the leachate after impurity removal to 60° C., adding 2.5 ml of a 50% Na2CO3 solution by mass, and continuing to react at a constant temperature of 60° C. for 1 hour until crystals are completely precipitated; filtering to obtain a carbonate crystal precipitate and a filtrate; S4, evaporating and concentrating the filtrate and then sequentially performing steps S1, S2, and S3 to perform secondary extraction of lithium, rubidium, and cesium valuable elements; The valuable elements are specifically lithium, rubidium and cesium; The lepidolite ore is sorted and pre-enriched to obtain lepidolite flotation concentrate; wherein: The separation and pre-enrichment treatment is specifically as follows: a pre-desliming-flotation combined process is adopted to grind the lepidolite ore to 15-18 μm and then perform pre-desliming treatment in a hydrocyclone; then a flotation process is adopted to add 200 g / t of hydrogen fluoride to activate the mineral surface according to the mineral mass; after activation for 20 minutes, the mineral surface is washed and 10 kg / t of caustic soda is added to treat the mineral surface; and finally, an oleic acid collector is added to float the lepidolite to obtain a lepidolite flotation concentrate.

Images