Method for comprehensively recovering valuable elements in clay type lithium ore

Abstract

The invention provides a method for comprehensively recovering valuable elements in clay type lithium ore, which comprises the following steps: crushing the clay type lithium ore to obtain an ore powder raw material; the mineral powder raw material is subjected to pressurized acid leaching to obtain first recovery liquid and silicon-containing slag; the first recovery liquid is subjected to oxygen pressure iron removal, and second recovery liquid and iron-containing slag are obtained; an alkaline reagent is added into the second recovery liquid, and third recovery liquid and aluminum-containing slag are obtained after a neutralization reaction; a precipitation reagent is added into the third recovery liquid, and fourth recovery liquid and nickel-cobalt-containing slag are obtained after precipitation reaction; and adding an oxidation reagent into the fourth recovery liquid, and carrying out oxidation reaction to obtain a lithium-containing solution and manganese-containing slag. According to the method, all the process steps in the recycling method are precisely designed, high-selectivity leaching is conducted on the clay type lithium ore in a multi-stage leaching and multi-stage separation combined mode, and meanwhile efficient comprehensive recycling of valuable elements in the clay type lithium ore is achieved.

Description

Technical Field

[0001] This invention relates to the field of hydrometallurgy, and more specifically, to a method for the comprehensive recovery of valuable elements from clay-type lithium ores. Background Technology

[0002] Lithium resources have received significant attention as an important resource for developing new energy sources. Lithium resources mainly exist in the form of minerals such as spodumene, lithium feldspar, and lithium phosphate. Currently, the most common method for extracting lithium from lithium-containing minerals is alkaline leaching. However, this method has a low lithium recovery rate, and the recovery rate is directly proportional to the solubility of the lithium mineral.

[0003] Among lithium-bearing minerals, clay-type lithium ore refers to clay minerals or rocks containing lithium. These ores are typically composed of clay minerals such as lepidolite, montmorillonite, and illite, and are commonly found in sedimentary basins, weathering crusts, lacustrine sediments, and submarine sediments. Lithium exists in these clay minerals in the form of adsorption or intracrystalline bonding, and its content is usually low, but the total amount can be considerable, especially in large-scale geological units. Traditional processing of clay ore often involves high-temperature roasting, which requires sophisticated equipment, consumes a lot of energy, and is environmentally unfriendly. The lithium extraction process from clay-type lithium ore usually involves chemical treatment methods such as acid or alkali leaching, ion exchange, or solvent extraction to dissolve or extract lithium. However, because clay-type lithium ore contains more iron, silicon, aluminum, nickel, cobalt, and manganese than lithium, directly using the above methods for lithium extraction would result in excessively high concentrations of these impurities, which would adversely affect the lithium extraction process and the quality of the lithium concentrate. Therefore, effectively removing impurity elements from lithium-containing clay minerals is key to improving lithium recovery rates.

[0004] CN115725858B discloses a method for extracting lithium from clay-type lithium ore. This method involves mixing clay-type lithium ore powder with an aqueous solution to form a slurry, allowing it to stand, collecting the upper slurry, drying it to produce ore material, then obtaining concentrate powder through electrostatic separation, and finally obtaining lithium salt through roasting, acid leaching, and sodium carbonate conversion. CN111893318A discloses a method for extracting lithium from lithium-containing clay. This method involves mixing clay lithium, calcium carbonate, sodium sulfate, and potassium sulfate, followed by roasting and water leaching to obtain a lithium-containing leachate. However, as reflected in these two existing technologies, current clay-type lithium ore processing technologies mainly focus on lithium extraction, rarely mentioning other elements in the ore. Therefore, the economical and efficient development of clay minerals to achieve comprehensive recovery of all elements in these minerals is of great significance.

[0005] Therefore, how to provide a comprehensive recovery method for valuable metal elements in clay-type lithium ore, achieving high recovery rates for lithium extraction, while simultaneously recovering other high-value metal elements besides lithium, is one of the important technical problems that need to be solved in this field. Summary of the Invention

[0006] The main objective of this invention is to provide a method for the comprehensive recovery of valuable elements in clay-type lithium ores, in order to solve the problems of low recovery rate of lithium in clay-type lithium ores and difficulty in effectively recovering other valuable metal elements therein in the prior art.

[0007] To achieve the above objectives, the present invention provides a method for the comprehensive recovery of valuable elements from clay-type lithium ore, comprising: step S1, crushing the clay-type lithium ore to obtain ore powder raw material; the clay-type lithium ore includes lithium, silicon, iron, aluminum, nickel, cobalt, and manganese; step S2, subjecting the ore powder raw material to pressure acid leaching to obtain a first recovery liquid and silicon-containing slag; step S3, subjecting the first recovery liquid to oxygen pressure iron removal to obtain a second recovery liquid and iron-containing slag; step S4, adding an alkaline reagent to the second recovery liquid, and after a neutralization reaction to obtain a third recovery liquid and aluminum-containing slag; step S5, adding a precipitating reagent to the third recovery liquid, and after a precipitation reaction to obtain a fourth recovery liquid and nickel-cobalt-containing slag; step S6, adding an oxidizing reagent to the fourth recovery liquid, and after an oxidation reaction to obtain a lithium-containing solution and manganese-containing slag.

[0008] Furthermore, by weight, the clay-type lithium ore includes 0.1 to 1.0 parts of Li, 20 to 35 parts of Si, 1 to 5 parts of Fe, 15 to 25 parts of Al, 0.001 to 0.5 parts of Ni, 0.001 to 0.2 parts of Co, 0.001 to 0.1 parts of Mn, 0.01 to 0.5 parts of Mg and 0.01 to 0.5 parts of Ca.

[0009] Furthermore, in step S1, the particle size of the mineral powder raw material is 200 mesh to 500 mesh.

[0010] Further, step S2 includes: adding mineral powder raw material to acid solution at an acid-to-mineral ratio of 200g / kg to 500g / kg and a liquid-to-solid ratio of (3 to 6): 1 to obtain a mixed slurry; the mixed slurry is subjected to pressurized acid leaching to obtain a first recovery liquid and silica-containing slag; preferably, in step S2, the acid-to-mineral ratio is 450g / kg to 500g / kg and the liquid-to-solid ratio is (3 to 4): 1; more preferably, the acid solution is selected from one or more of sulfuric acid solution, hydrochloric acid solution and nitric acid solution.

[0011] Furthermore, in step S2, the pressure acid leaching is carried out for 2.0h to 4.0h under conditions of pressure of 0.1MPa to 4.0MPa, reaction temperature of 90℃ to 250℃, and stirring speed of 500r / min to 800r / min; preferably, the pressure acid leaching is carried out for 3.0h to 4.0h under conditions of pressure of 3.4MPa to 3.5MPa, reaction temperature of 240℃ to 250℃, and stirring speed of 700r / min to 800r / min.

[0012] Further, in step S3, the oxygen pressure iron removal is carried out for 1.0h to 3.0h under the conditions of oxygen pressure of 1.0MPa to 3.0MPa, reaction temperature of 90℃ to 200℃, and stirring speed of 500r / min to 800r / min; preferably, the oxygen pressure iron removal is carried out for 2.5h to 3.0h under the conditions of oxygen pressure of 2.0±0.2MPa, reaction temperature of 160℃ to 180℃, and stirring speed of 700r / min to 800r / min.

[0013] Further, in step S4, the neutralization reaction is carried out at a pH of 4.6 to 4.8, and the reaction temperature is 25°C to 100°C, with a reaction time of 1.0 h to 2.0 h; preferably, the reaction temperature is 40 ± 5°C; more preferably, the neutralizing agent is selected from one or more of sodium hydroxide, sodium carbonate, calcium hydroxide, and magnesium hydroxide, and the neutralizing agent is added in the form of an aqueous solution with a mass concentration of 5% to 20%.

[0014] Further, in step S5, the reaction temperature of the precipitation reaction is 50℃~100℃, and the reaction time is 0.5h~2.0h; preferably, the reaction temperature of the precipitation reaction is 50℃~55℃; more preferably, the precipitating agent is selected from one or more of sodium hydrosulfide, sodium sulfite and sodium metabisulfite, and the amount of precipitating agent added is 0.01g / L~5g / L.

[0015] Further, in step S6, the oxidation reaction is carried out at a pH of 4.0 to 8.0, and the reaction temperature is 25°C to 80°C, with a reaction time of 2.0 h to 6.0 h; preferably, the oxidation reaction is carried out at a pH of 4.0 to 4.5, and the reaction temperature is 30 ± 5°C, with a reaction time of 3.0 h to 4.0 h; more preferably, the oxidizing agent is selected from one or more of oxygen, sulfur dioxide, and carbon dioxide, and even more preferably oxygen.

[0016] Furthermore, the recovery rates for lithium are 40%–95%, silicon 75%–90%, iron 3%–12%, aluminum 10%–25%, nickel and cobalt 30%–99.9%, and manganese 70%–95%.

[0017] By applying the technical solution of this invention, and through precise design of each process step in the recovery method, a combination of multi-stage leaching and multi-stage separation is used to achieve highly selective leaching of clay-type lithium ore, while simultaneously realizing the efficient and comprehensive recovery of valuable elements. This not only effectively separates and recovers key metals such as lithium, nickel, cobalt, and manganese, but also treats impurities such as silicon, iron, and aluminum, improving the purity of the final lithium leachate. Furthermore, the entire process avoids roasting, resulting in lower equipment requirements and lower energy consumption. Detailed Implementation

[0018] It should be noted that, unless otherwise specified, the embodiments and features described in this application can be combined with each other. The present invention will now be described in detail with reference to the embodiments.

[0019] As described in the background section, existing technologies suffer from low lithium recovery rates in clay-type lithium ores and difficulty in effectively recovering other valuable metal elements. To address these technical problems, this invention provides a method for the comprehensive recovery of valuable elements from clay-type lithium ores, comprising: Step S1, crushing the clay-type lithium ore to obtain ore powder; the clay-type lithium ore includes lithium, silicon, iron, aluminum, nickel, cobalt, and manganese; Step S2, subjecting the ore powder to pressure acid leaching to obtain a first recovery liquid and silicon-containing slag; Step S3, subjecting the first recovery liquid to oxygen pressure iron removal to obtain a second recovery liquid and iron-containing slag; Step S4, adding an alkaline reagent to the second recovery liquid, followed by a neutralization reaction to obtain a third recovery liquid and aluminum-containing slag; Step S5, adding a precipitating reagent to the third recovery liquid, followed by a precipitation reaction to obtain a fourth recovery liquid and nickel-cobalt-containing slag; Step S6, adding an oxidizing reagent to the fourth recovery liquid, followed by an oxidation reaction to obtain a lithium-containing solution and manganese-containing slag.

[0020] The recycling method provided by this invention, through precise design of each process step, employs a combination of multi-stage leaching and multi-stage separation to comprehensively recover clay-type lithium ore, achieving efficient recovery of valuable elements. Specifically: First, in step S1, a crushing process is performed, using mechanical force to break large pieces of clay-type lithium ore into smaller particles, allowing for more effective subsequent chemical processing. In step S2, pressurized acid leaching involves contacting the ore powder with the acid solution under high pressure to dissolve metallic elements such as lithium, iron, aluminum, nickel, cobalt, and manganese. In step S3, oxygen is introduced into the acid leaching solution, promoting the recovery of Fe under pressure. 2+ Oxidized to Fe 3+ Subsequently, the obtained Fe was removed 3+In step S4, an alkaline reagent is added to the iron-removed solution to adjust the pH to a certain range, thereby precipitating aluminum. Step S5 involves adding a precipitant to the resulting third recovery solution, during which nickel and cobalt precipitate as sulfides. Then, in step S6, an oxidant is added to the nickel-cobalt precipitated solution to remove Mn. 2+ The manganese is oxidized into a high-valence manganese compound that is insoluble in water and then removed. Ultimately, the recovery of the other valuable metal elements mentioned above is achieved while significantly improving the purity of the lithium leachate and the lithium recovery rate.

[0021] More importantly, the leaching recovery method provided by this invention integrates the aforementioned steps and corresponding conditions into a unified technical solution, forming a complete technical route of "pressure acid leaching → oxygen pressure iron removal → neutralization aluminum removal → precipitation nickel and cobalt removal → oxidation manganese removal → finally obtaining a lithium-containing solution." Beyond the design of the process flow, it also includes parameter optimization for each process step. Compared to conventional methods in the art that involve only a single process step, this method more effectively achieves the efficient recovery of various metal elements from clay-type lithium ore. In other words, the aforementioned steps in the recovery method provided by this invention, in addition to their individual advantages, are more importantly able to work together as a unified solution to achieve efficient recovery of various valuable elements from clay-type lithium ore.

[0022] In particular, changing the order of the above process steps will also affect the separation and recovery efficiency of each element. For example, if aluminum is neutralized and precipitated first, followed by oxygen pressure removal of iron, some iron will precipitate during the aluminum precipitation process, reducing the purity of the aluminum slag. In addition, the precipitation of aluminum ions may consume some oxygen, resulting in incomplete oxidation of iron ions and a decrease in iron recovery rate. If nickel and cobalt are precipitated first, followed by neutralization and aluminum precipitation, other metal ions will precipitate together, reducing the purity of the nickel and cobalt recovery and affecting the recovery rate of other elements. Furthermore, during this process, aluminum in the solution is prone to forming insoluble colloids, increasing the difficulty of separation and filtration. If manganese is oxidized and removed first, followed by nickel and cobalt precipitation, i.e., manganese is oxidized before nickel and cobalt precipitation, the byproducts generated by manganese oxidation, such as MnO2, will change the pH value of the solution, thereby affecting the precipitation selectivity of the precipitating reagent for nickel and cobalt, leading to a decrease in nickel and cobalt recovery rate. In particular, the nickel-cobalt leaching process must be placed after the iron removal and aluminum leaching processes. This involves first applying pressure acid leaching to selectively leach lithium while maximizing the dissolution of valuable elements such as iron, aluminum, nickel, cobalt, and manganese. Then, iron is recovered through oxygen pressure iron removal, and aluminum is recovered by adding a neutralizing agent. Throughout the process, the solution's pH is controlled, avoiding repeated and drastic fluctuations, significantly reducing reagent consumption and operating costs, while also mitigating inter-element interference, ultimately achieving comprehensive recovery of all elements.

[0023] In summary, the method provided by this invention can simultaneously recover multiple valuable elements such as lithium, silicon, iron, aluminum, nickel, cobalt, and manganese. Compared with traditional lithium extraction processes, it not only focuses on the extraction of lithium but also achieves efficient recovery of other valuable elements, significantly improving resource utilization efficiency.

[0024] For the clay-type lithium ore being processed, the preferred composition by weight includes 0.1-1.0 parts of Li, 20-35 parts of Si, 1-5 parts of Fe, 15-25 parts of Al, 0.001-0.5 parts of Ni, 0.001-0.2 parts of Co, 0.001-0.1 parts of Mn, 0.01-0.5 parts of Mg, and 0.01-0.5 parts of Ca, in order to better adapt to the above-mentioned recovery method and improve the efficiency and purity of the recovery of each metal element. More preferably, the clay-type lithium ore comprises, by weight, 0.10-0.20 parts Li, 30-35 parts Si, 4.0-4.5 parts Fe, 22-25 parts Al, 0.001-0.25 parts Ni, 0.001-0.1 parts Co, 0.001-0.05 parts Mn, 0.1-0.35 parts Mg, and 0.2-0.3 parts Ca. That is to say, the comprehensive recovery method provided by this invention is more suitable for clay-type lithium ores with low lithium and high iron and aluminum content, thereby further increasing the lithium recovery rate.

[0025] To increase the contact area between the mineral powder and the acid solution during pressurized acid leaching, thereby accelerating the reaction process and improving leaching efficiency, the particle size of the mineral powder raw material in step S1 is preferably 200-500 mesh. More preferably, the particle size of the mineral powder raw material is 200-350 mesh, which can achieve sufficient acid leaching and improve the recovery rate of each element while reducing energy consumption during crushing and grinding, and reducing the difficulty and cost increase in subsequent processing caused by excessively fine particle size.

[0026] In several typical implementations, step S2 includes: adding mineral powder raw materials to acid solution at an acid-to-ore ratio of 200 g / kg to 500 g / kg and a liquid-to-solid ratio of (3 to 6):1 to obtain a mixed slurry; the mixed slurry is then subjected to pressurized acid leaching to obtain a first recovery liquid and silica-containing slag. By optimizing the acid-to-ore ratio, liquid-to-solid ratio, and pressurized acid leaching conditions, lithium is leached with high selectivity while leaching of other elements to a greater extent, while also reducing reagent consumption and improving economic efficiency. More preferably, in step S2, the acid-to-ore ratio is 450 g / kg to 500 g / kg, and the liquid-to-solid ratio is (3 to 4):1, to further improve the recovery rate of elements such as lithium, nickel, and cobalt, promoting efficient resource utilization. Simultaneously, reducing the liquid-to-solid ratio decreases the amount of acid used, thereby reducing the burden on wastewater treatment in subsequent recovery processes.

[0027] The acid solution used for pressure acid leaching is selected from one or more of sulfuric acid solution, hydrochloric acid solution, and nitric acid solution, and is not limited to these, thereby improving the flexibility and adaptability of the method provided by the present invention. Furthermore, a sulfuric acid solution with a mass concentration of 10.0% to 12.5% ​​is preferred because this concentration of sulfuric acid solution can significantly improve the leaching efficiency of each metal element, while not introducing other impurities, making the subsequent processing of the residue product containing each element simpler and its resource utilization value higher.

[0028] Furthermore, in step S2, the pressure acid leaching is carried out for 2.0h to 4.0h under conditions of 0.1MPa to 4.0MPa pressure, 90℃ to 250℃ reaction temperature, and 500r / min to 800r / min stirring rate. By optimizing the pressure, reaction temperature, and stirring rate of the pressure acid leaching as described above, the pressure acid leaching process can be achieved more efficiently and economically, while reducing equipment requirements and environmental impact. More preferably, the pressure acid leaching is carried out for 3.0h to 4.0h under conditions of 3.4MPa to 3.5MPa pressure, 240℃ to 250℃ reaction temperature, and 700r / min to 800r / min stirring rate, thereby achieving a higher overall recovery rate by better adapting to the characteristics of the elements to be extracted under conditions of lower energy consumption.

[0029] For oxygen pressure iron removal, i.e., step S3, it is preferably carried out for 1.0h to 3.0h under the following conditions: oxygen pressure of 1.0MPa to 3.0MPa, reaction temperature of 90℃ to 200℃, and stirring rate of 500r / min to 800r / min. Under the above oxygen pressure environment and reaction temperature, the solubility of oxygen increases, which can more effectively oxidize iron ions, converting them into insoluble Fe(III) compounds, thereby improving the iron removal rate, reducing the need for iron ion treatment in subsequent processes, and improving the overall process efficiency. Furthermore, this process can precipitate iron in a strong acid medium, while Ni... 2+ Co 2+ Li + Plasma remains in the solution, enabling more effective element separation. The optimal stirring rate and reaction time ensure sufficient contact between iron ions and oxygen, promoting a uniform reaction. Even better, oxygen pressure iron removal, conducted at an oxygen pressure of 2.0 ± 0.2 MPa, a reaction temperature of 160℃~180℃, and a stirring rate of 700 r / min~800 r / min for 2.5 h~3.0 h, achieves more efficient and thorough iron ion oxidation, significantly improving iron removal and recovery rates while reducing reagent consumption and byproduct formation, and simultaneously achieving high recovery rates of subsequent valuable elements.

[0030] In step S4, the neutralization reaction is preferably carried out at a pH of 4.6–4.8, with a reaction temperature of 25°C–100°C and a reaction time of 1.0–2.0 h. Within this optimized and precise pH range, aluminum ions can precipitate as Al(OH)3, while other valuable metals such as lithium, nickel, and cobalt remain in the solution, thus achieving efficient separation and recovery of aluminum. The optimized reaction temperature and time are designed to promote the complete conversion of aluminum ions into precipitate while reducing the co-precipitation of other metal elements. Based on this, the inventors, through extensive experimentation, have further optimized the neutralization reaction temperature to 40±5°C to achieve higher precipitation efficiency of aluminum while maintaining the stability of other valuable metals in the solution, thereby improving the recovery rate and purity of aluminum and enhancing the recovery rate of other elements, including lithium.

[0031] In several typical implementation methods, the neutralizing agent is selected from one or more of sodium hydroxide, sodium carbonate, calcium hydroxide, and magnesium hydroxide, and is added in the form of an aqueous solution with a mass concentration of 5% to 20% to facilitate more precise control of the pH value of the recovered solution and avoid affecting the recovery rate and purity of aluminum due to excessive or insufficient amounts.

[0032] In step S5, to optimize the precipitation process of nickel sulfide and cobalt sulfide, the preferred reaction temperature is 50℃~100℃, and the reaction time is 0.5h~2.0h. Under these preferred temperature and time conditions, the precipitation efficiency and selectivity of sulfides can be effectively improved, their loss can be reduced, and the formation of precipitates can be promoted, which is beneficial to subsequent solid-liquid separation and resource recovery, thereby improving the nickel and cobalt recovery rate. A more preferred reaction temperature of 50℃~55℃ is to minimize the increase in energy consumption and potential loss of other valuable elements due to excessively high temperatures while precipitating nickel and cobalt, thereby simultaneously improving the recovery rate and purity of the nickel and cobalt precipitate, as well as subsequent lithium and manganese, promoting better overall resource utilization of clay-type lithium ore. Furthermore, the preferred precipitating agent is selected from one or more of sodium hydrosulfide, sodium sulfite, and sodium metabisulfite, and the amount of precipitating agent added is 0.01g / L~5g / L, thereby promoting complete precipitation of nickel and cobalt elements and improving their recovery rate. Further optimization of the precipitating agent dosage to 0.8 g / L~1.2 g / L aims to find a better balance between economy and nickel-cobalt recovery efficiency, more efficiently converting nickel-cobalt ions into sulfide precipitates, and achieving high recovery rate and high purity precipitation of nickel and cobalt. In practical applications, it is preferred that the dosage of the precipitant be 1.2~1.5 g / L, based on the total weight of nickel and cobalt elements as 1.

[0033] In step S6, the oxidation reaction is carried out at a pH of 4.0–8.0, a reaction temperature of 25°C–80°C, and a reaction time of 2.0 h–6.0 h. This preferred pH range considers the efficient oxidation and separation of manganese, as well as the protection of other valuable elements such as lithium. Based on extensive experimental verification, the inventors have further preferred the oxidation reaction to be carried out at a pH of 4.0–4.5, a reaction temperature of 30 ± 5°C, and a reaction time of 3.0 h–4.0 h. Under these preferred conditions, manganese can be oxidized to MnO2 or other forms of precipitate, thereby separating it from the lithium solution. This also reduces lithium loss due to over-oxidation, improves the selectivity and efficiency of the oxidation reaction, and more effectively achieves manganese recovery and lithium purification.

[0034] In several typical implementations, the oxidizing agent is selected from one or more of oxygen, sulfur dioxide, and carbon dioxide. To avoid introducing other impurity elements, oxygen is further preferred as the oxidizing agent, thereby achieving higher lithium and manganese recovery rates.

[0035] Furthermore, the recovery rates for lithium are 40%–95%, silicon 75%–90%, iron 3%–12%, aluminum 10%–25%, the combined recovery rates for nickel and cobalt 30%–99.9%, and manganese 70%–95%. In other words, the recovery method provided by this invention possesses high selectivity, high efficiency, and economic viability, enabling the effective recovery of multiple valuable elements from clay-type lithium ores with relatively high recovery rates, thus providing technical support for resource recycling and sustainable development.

[0036] The present application will be further described in detail below with reference to specific embodiments, which should not be construed as limiting the scope of protection claimed in the present application.

[0037] Unless otherwise defined, all technical terms used herein have the same meaning as commonly understood by those skilled in the art. The technical terms used herein are for the purpose of describing particular embodiments only and are not intended to limit the scope of the invention.

[0038] It should be noted in advance that the composition of the several clay-type lithium ores processed by this invention is shown in Table 1 below (some impurity elements are not shown).

[0039] Table 1

[0040] Example 1

[0041] A method for comprehensive recovery of valuable elements from clay-type lithium ores: (1) Lithium ore 1 in Table 1 was used as the clay-type lithium ore to be processed. It was crushed, screened and ground to obtain mineral powder with a particle size of 200 mesh as mineral powder raw material. (2) The mineral powder raw material is mixed with a sulfuric acid aqueous solution with a mass concentration of 7.5% at an acid-to-mineral ratio of 300 g / kg and a liquid-to-solid ratio of 4:1 to obtain a mixed slurry. The obtained mixed slurry is added to a high-pressure reactor and subjected to a pressurized acid leaching reaction for 3 h at a pressure of 3.3 MPa, a reaction temperature of 240 °C, and a stirring speed of 800 r / min. After the reaction is completed, the product is separated into solid and liquid after the high-pressure reactor is cooled to obtain the first recovery liquid and silicon-containing slag.

[0042] (3) The first recovered liquid obtained in step (2) is pumped into a pressure vessel, oxygen is introduced, the oxygen pressure is controlled at 2.0 MPa, the reaction temperature is 200℃, the stirring rate is 800 r / min, and oxygen pressure iron removal is carried out under these conditions for 2 hours. During this process, hydrolysis and iron removal occur. After the reaction is completed, the second recovered liquid and iron-containing slag are obtained by solid-liquid separation. A portion of the second recovered liquid obtained later can be separated and returned to step (2), and a certain amount of sulfuric acid solution is added to it to serve as a leachate. This leachate is then mixed with a new batch of raw materials and subjected to pressure acid leaching again.

[0043] (4) Add an alkaline reagent, namely a 10% sodium hydroxide aqueous solution, to the second recovered liquid obtained in step (3) to control its pH at 4.6~4.8. Under these conditions, carry out a neutralization reaction at 40℃ for 2 hours. After the reaction is completed, the third recovered liquid and aluminum-containing slag are obtained by solid-liquid separation.

[0044] (5) Add sodium hydrosulfide, a precipitating agent, to the third recovered liquid obtained in step (4) at a dosage of 0.03 g / L (i.e., the dosage of the precipitant is 1.2 based on the total weight of nickel and cobalt elements). React at a reaction temperature of 50°C for 0.5 h. After the reaction is completed, the fourth recovered liquid and nickel-cobalt slag are obtained by solid-liquid separation.

[0045] (6) Oxygen is introduced into the fourth recovery liquid obtained in step (5), and the reaction temperature is controlled at 30℃ and the pH value is 4.0 for oxidation reaction for 3 hours. After the reaction is completed, lithium-containing solution and manganese-containing slag are obtained by solid-liquid separation.

[0046] Example 2

[0047] A method for comprehensive recovery of valuable elements from clay-type lithium ores: (1) Lithium ore 2 in Table 1 was used as the clay-type lithium ore to be processed. It was crushed, screened and ground to obtain mineral powder with a particle size of 200 mesh as mineral powder raw material. (2) The mineral powder raw material is mixed with a sulfuric acid aqueous solution with a mass concentration of 10% at an acid-to-mineral ratio of 400 g / kg and a liquid-to-solid ratio of 4:1 to obtain a mixed slurry. The obtained mixed slurry is added to a high-pressure reactor and subjected to a pressurized acid leaching reaction for 3 h at a pressure of 3.5 MPa, a reaction temperature of 240 °C, and a stirring speed of 800 r / min. After the reaction is completed, the product is separated into solid and liquid after the high-pressure reactor is cooled to obtain the first recovery liquid and silicon-containing slag.

[0048] (3) The first recovered liquid obtained in step (2) is pumped into a pressure vessel, oxygen is introduced, the oxygen pressure is controlled at 2.0 MPa, the reaction temperature is 200℃, the stirring rate is 800 r / min, and oxygen pressure iron removal is carried out under these conditions for 2 hours. During this process, hydrolysis and iron removal occur. After the reaction is completed, the second recovered liquid and iron-containing slag are obtained by solid-liquid separation. A portion of the second recovered liquid obtained later can be separated and returned to step (2), and a certain amount of sulfuric acid solution is added to it to serve as a leachate. This leachate is then mixed with a new batch of raw materials and subjected to pressure acid leaching again.

[0049] (4) Add an alkaline reagent, namely a 10% sodium hydroxide aqueous solution, to the second recovered liquid obtained in step (3) to control its pH at 4.6~4.8. Under these conditions, carry out a neutralization reaction at 40℃ for 2 hours. After the reaction is completed, the third recovered liquid and aluminum-containing slag are obtained by solid-liquid separation.

[0050] (5) Add sodium hydrosulfide, a precipitating agent, to the third recovered liquid obtained in step (4) at a dosage of 0.87 g / L. React at a reaction temperature of 50 °C for 0.5 h. After the reaction is completed, the fourth recovered liquid and nickel-cobalt slag are obtained by solid-liquid separation.

[0051] (6) Oxygen is introduced into the fourth recovery liquid obtained in step (5), and the reaction temperature is controlled at 30℃ and the pH value is 4.0 for oxidation reaction for 3 hours. After the reaction is completed, lithium-containing solution and manganese-containing slag are obtained by solid-liquid separation.

[0052] Example 3

[0053] A method for comprehensive recovery of valuable elements from clay-type lithium ores: (1) Lithium ore 3 in Table 1 was used as the clay-type lithium ore to be processed. It was crushed, screened and ground to obtain mineral powder with a particle size of 350 mesh as mineral powder raw material. (2) The mineral powder raw material is mixed with a sulfuric acid aqueous solution with a mass concentration of 12.5% ​​at an acid-to-mineral ratio of 500 g / kg and a liquid-to-solid ratio of 4:1 to obtain a mixed slurry. The obtained mixed slurry is added to a high-pressure reactor and subjected to a pressurized acid leaching reaction for 4 hours at a pressure of 3.5 MPa, a reaction temperature of 240℃, and a stirring speed of 800 r / min. After the reaction is completed, the product is separated into solid and liquid after the high-pressure reactor is cooled to obtain the first recovery liquid and silicon-containing slag.

[0054] (3) The first recovered liquid obtained in step (2) is pumped into a pressure vessel, oxygen is introduced, the oxygen pressure is controlled at 2.0 MPa, the reaction temperature is 180℃, the stirring rate is 800 r / min, and oxygen pressure iron removal is carried out under these conditions for 3 hours. During this process, hydrolysis and iron removal occur. After the reaction is completed, the second recovered liquid and iron-containing slag are obtained by solid-liquid separation. A portion of the obtained second recovered liquid can be separated and returned to step (2), and a certain amount of sulfuric acid solution is added to it to serve as a leachate. This leachate is then mixed with a new batch of raw materials and subjected to pressure acid leaching again.

[0055] (4) Add an alkaline reagent, namely a 10% sodium hydroxide aqueous solution, to the second recovered liquid obtained in step (3) to control its pH at 4.6~4.8. Under these conditions, carry out a neutralization reaction at 40℃ for 2 hours. After the reaction is completed, the third recovered liquid and aluminum-containing slag are obtained by solid-liquid separation.

[0056] (5) Add sodium hydrosulfide, a precipitating agent, to the third recovered liquid obtained in step (4) at a dosage of 0.91 g / L. React at a reaction temperature of 50 °C for 0.5 h. After the reaction is completed, the fourth recovered liquid and nickel-cobalt slag are obtained by solid-liquid separation.

[0057] (6) Oxygen is introduced into the fourth recovery liquid obtained in step (5), and the reaction temperature is controlled at 30℃ and the pH value is 4.0 for oxidation reaction for 3 hours. After the reaction is completed, lithium-containing solution and manganese-containing slag are obtained by solid-liquid separation.

[0058] Example 4

[0059] A method for comprehensive recovery of valuable elements from clay-type lithium ores: (1) Lithium ore 4 in Table 1 was used as the clay-type lithium ore to be processed. It was crushed, screened and ground to obtain mineral powder with a particle size of 200 mesh as mineral powder raw material. (2) The mineral powder raw material is mixed with a sulfuric acid aqueous solution with a mass concentration of 12.5% ​​at an acid-to-mineral ratio of 500 g / kg and a liquid-to-solid ratio of 4:1 to obtain a mixed slurry. The obtained mixed slurry is added to a high-pressure reactor and subjected to a pressurized acid leaching reaction for 3 h at a pressure of 3.5 MPa, a reaction temperature of 240 °C, and a stirring speed of 800 r / min. After the reaction is completed, the product is separated into solid and liquid after the high-pressure reactor is cooled to obtain the first recovery liquid and silicon-containing slag.

[0060] (3) The first recovered liquid obtained in step (2) is pumped into a pressure vessel, oxygen is introduced, the oxygen pressure is controlled at 2.0 MPa, the reaction temperature is 180℃, the stirring rate is 800 r / min, and oxygen pressure iron removal is carried out under these conditions for 3 hours. During this process, hydrolysis and iron removal occur. After the reaction is completed, the second recovered liquid and iron-containing slag are obtained by solid-liquid separation. A portion of the second recovered liquid obtained later can be separated and returned to step (2), and a certain amount of sulfuric acid solution is added to it to serve as a leachate. This leachate is then mixed with a new batch of raw materials and subjected to pressure acid leaching again.

[0061] (4) Add an alkaline reagent, namely a 10% sodium hydroxide aqueous solution, to the second recovered liquid obtained in step (3) to control its pH at 4.6~4.8. Under these conditions, carry out a neutralization reaction at 40℃ for 2 hours. After the reaction is completed, the third recovered liquid and aluminum-containing slag are obtained by solid-liquid separation.

[0062] (5) Add sodium hydrosulfide, a precipitating agent, to the third recovered liquid obtained in step (4) at a dosage of 1.18 g / L. React at a reaction temperature of 50 °C for 0.5 h. After the reaction is completed, the fourth recovered liquid and nickel-cobalt slag are obtained by solid-liquid separation.

[0063] (6) Oxygen is introduced into the fourth recovery liquid obtained in step (5), and the reaction temperature is controlled at 30℃ and the pH value is 4.0 for oxidation reaction for 3 hours. After the reaction is completed, lithium-containing solution and manganese-containing slag are obtained by solid-liquid separation.

[0064] Example 5

[0065] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The only difference between this embodiment and embodiment 3 is that in step (1), lithium ore 5 from Table 1 is used as the clay-type lithium ore to be processed, instead of lithium ore 3.

[0066] Example 6

[0067] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The only difference between this embodiment and embodiment 3 is that in step (1), lithium ore 3 is crushed, screened, and ground to obtain mineral powder with a particle size of 100 mesh as mineral powder raw material.

[0068] Example 7

[0069] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The only difference between this embodiment and embodiment 3 is that in step (2), the acid-to-ore ratio is changed to 200 g / kg, the liquid-to-solid ratio is changed to 6:1, and the mass concentration of the sulfuric acid aqueous solution is changed to 5%.

[0070] Example 8

[0071] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The difference between this embodiment and embodiment 3 lies only in step (3), specifically: The first recovered liquid obtained in step (2) is pumped into a pressure vessel, oxygen is introduced, the oxygen pressure is controlled at 3.0 MPa, the reaction temperature is 90℃, the stirring rate is 500 r / min, and under these conditions, oxygen pressure iron removal is carried out for 1 hour.

[0072] Example 9

[0073] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The difference between this embodiment and embodiment 3 lies only in step (4), specifically: An alkaline reagent, namely a 5% sodium hydroxide aqueous solution, was added to the second recovered liquid obtained in step (3) to control its pH at 5.0 ± 0.1. A neutralization reaction was carried out under these conditions at 100°C for 1 hour.

[0074] Example 10

[0075] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The difference between this embodiment and embodiment 3 lies only in step (5), specifically: Sodium hydrosulfide, a precipitating agent, was added to the third recovered liquid obtained in step (4) at a dosage of 0.1 g / L. The reaction was carried out at a temperature of 100 °C for 2.0 h. After the reaction was completed, the fourth recovered liquid and nickel-cobalt slag were obtained by solid-liquid separation.

[0076] Example 11

[0077] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The difference between this embodiment and embodiment 3 lies only in step (6), specifically: Oxygen was introduced into the fourth recovery liquid obtained in step (5), and the oxidation reaction was carried out at a controlled temperature of 80℃ and a pH of 8.0 for 6 hours. After the reaction was completed, a lithium-containing solution and a manganese-containing slag were obtained by solid-liquid separation.

[0078] Comparative Example 1

[0079] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The only difference between this comparative example and Example 1 is that steps (5) and (6) were not performed; instead, the third recovered liquid obtained in step (4) was directly used as the final lithium-containing solution.

[0080] Comparative Example 2

[0081] A method for comprehensive recovery of valuable elements from clay-type lithium ores: The only difference between this comparative example and Example 1 is that step (3) was not performed, but the first recovered liquid obtained in step (2) was directly used as the second recovered liquid and subsequent steps (4) to (6) were performed.

[0082] Test methods

[0083] The recovery rates of each element obtained from the above embodiments and comparative examples are shown in Table 2.

[0084] Table 2

[0085] As can be seen from the above description, the embodiments of the present invention achieve efficient recovery of lithium from clay-type lithium ore, and simultaneously achieve efficient recovery of other valuable elements besides lithium. The resulting silicon-containing slag, iron-containing slag, aluminum-containing slag, nickel-cobalt-containing slag, and manganese-containing slag can all be utilized for resource recovery.

[0086] Specifically, the main difference between Examples 1, 2, and 4 is the acid-to-mineral ratio. Increasing the acid-to-mineral ratio increases the contact amount between sulfuric acid and the mineral, accelerating the dissolution of lithium from the crystal lattice and improving the lithium leaching rate. Furthermore, under the acid-to-mineral ratio conditions provided in this application, exceeding a specific value can suppress the leaching of elements such as iron and aluminum, reducing competitive consumption of sulfuric acid and ensuring sufficient sulfuric acid for lithium leaching, thereby significantly improving the lithium leaching selectivity.

[0087] Compared with Example 4, Example 3 reduced the particle size of the mineral powder. The reduced particle size can significantly increase the contact area between the mineral and the leachate, promote the efficiency of chemical reaction, and thus further improve the leaching rate of elements.

[0088] Compared with Example 3, Example 5 uses high-lithium clay ore as the leaching target, and the elemental leaching rate and recovery rate are similar to those of Example 3, indicating that the method provided in this application also has a good treatment effect on high-lithium clay ore. It is also evident that the above-mentioned comprehensive recovery method provided by this invention is more suitable for clay-type lithium ores with low lithium and high iron and aluminum content, thereby further increasing the lithium recovery rate.

[0089] Compared with Example 3, the mineral powder particle size in Example 6 is less than 100 mesh, which exceeds the preferred particle size range of the present invention, resulting in a low activation effect on the ore and a slight reduction in the leaching effect of each element.

[0090] Compared to Example 3, Example 7 uses an acid-to-ore ratio of 200 g / kg in high-pressure acid leaching. The limited ability of sulfuric acid to damage the ore structure makes it more difficult for lithium to dissolve from the crystal lattice, resulting in a lower lithium leaching rate. In other words, by optimizing the conditions of pressure acid leaching in step S2, a higher overall recovery rate can be achieved by better adapting to the characteristics of the elements to be extracted while consuming less energy.

[0091] Compared to Example 3, Example 8 uses an oxygen pressure iron removal reaction temperature of 90°C, which reduces the hydrolysis rate of iron ions and leads to a decrease in iron recovery rate. In other words, by optimizing the oxygen pressure iron removal conditions in step S3, more efficient and thorough iron ion oxidation can be achieved, thereby significantly improving the iron removal and recovery rate while reducing reagent consumption and byproduct generation, and simultaneously achieving a high recovery rate of subsequent valuable elements.

[0092] Compared to Example 3, Example 9, with a neutralization pH of 5.0, did not show a significant effect on aluminum precipitation. In fact, in step S4, aluminum begins to precipitate at around pH 3.5, and precipitation is almost complete at pH 4.8. Further increases in pH do not enhance aluminum precipitation but instead increase reagent consumption. A pH above 7 causes nickel and cobalt precipitation, resulting in poor element separation. Therefore, in this step, a pH < 5 is preferred, especially a pH of 4.6–4.8, which promotes better precipitation of aluminum ions as Al(OH)3, while other valuable metals such as lithium, nickel, and cobalt remain in the solution, thus achieving more efficient separation and recovery of aluminum.

[0093] Compared to Example 3, Example 10 used only 0.1 g / L of nickel-cobalt precipitating reagent. This insufficient reagent resulted in incomplete nickel-cobalt precipitation, leading to a lower recovery rate. In other words, by optimizing the parameters of the precipitation process in step S5, the recovery rate of nickel-cobalt can be further improved.

[0094] Compared to Example 3, Example 11 uses a manganese precipitation reaction temperature of 80°C and a pH of 8.0. While these conditions have little impact on the precipitation effect of manganese, they do affect the precipitate morphology, forming colloids or particles, thus affecting the solid-liquid separation effect. In other words, optimizing the oxidation reaction conditions in step S6 can more effectively achieve manganese recovery and lithium purification.

[0095] Compared with Example 1, Comparative Example 1 did not perform the nickel-cobalt and manganese precipitation steps, resulting in the nickel, cobalt and manganese elements remaining in the lithium-rich solution and being difficult to separate.

[0096] Compared with Example 1, Comparative Example 2 did not perform oxygen pressure iron removal, which resulted in iron precipitating along with aluminum in step (4), and effective separation could not be achieved.

[0097] It should be noted that the terms "first," "second," etc., used in the specification and claims of this application are used to distinguish similar objects and are not necessarily used to describe a specific order or sequence. It should be understood that such terms can be used interchangeably where appropriate so that the embodiments of this application described herein can be implemented, for example, in a sequence other than those described herein.

[0098] The above description is merely a preferred embodiment of the present invention and is not intended to limit the invention. Various modifications and variations can be made to the present invention by those skilled in the art. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the scope of protection of the present invention.

Claims

1. A method for comprehensively recovering valuable elements from clay-type lithium ores, characterized in that, include: Step S1: The clay-type lithium ore is crushed to obtain mineral powder raw material; the clay-type lithium ore mainly includes lithium, silicon, iron, aluminum, nickel, cobalt and manganese. Step S2: The mineral powder raw material is subjected to pressurized acid leaching to obtain a first recovery liquid and silicon-containing slag; Step S3: The first recovered liquid is subjected to oxygen pressure to remove iron, resulting in a second recovered liquid and iron-containing slag; Step S4: Add an alkaline reagent to the second recovery liquid, and after neutralization reaction, obtain the third recovery liquid and aluminum-containing slag; Step S5: Add a precipitating agent to the third recovered liquid, and after precipitation reaction, obtain the fourth recovered liquid and nickel-cobalt slag; Step S6: Add an oxidizing agent to the fourth recovery liquid, and after oxidation reaction, obtain a lithium-containing solution and a manganese-containing slag.

2. The method for comprehensive recovery of valuable elements in clay-type lithium ores according to claim 1, characterized in that, By weight, the clay-type lithium ore comprises 0.1 to 1.0 parts of Li, 20 to 35 parts of Si, 1 to 5 parts of Fe, 15 to 25 parts of Al, 0.001 to 0.5 parts of Ni, 0.001 to 0.2 parts of Co, 0.001 to 0.1 parts of Mn, 0.01 to 0.5 parts of Mg and 0.01 to 0.5 parts of Ca.

3. The method for comprehensive recovery of valuable elements in clay-type lithium ores according to claim 1 or 2, characterized in that, In step S1, the particle size of the mineral powder raw material is 200 mesh to 500 mesh.

4. The method for comprehensive recovery of valuable elements in clay-type lithium ores according to any one of claims 1 to 3, characterized in that, Step S2 includes: adding the mineral powder raw material to an acid solution at an acid-to-mineral ratio of 200g / kg to 500g / kg and a liquid-to-solid ratio of (3 to 6):1 to obtain a mixed slurry; subjecting the mixed slurry to pressurized acid leaching to obtain the first recovered liquid and the silica-containing slag. Preferably, in step S2, the acid-to-ore ratio is 450 g / kg to 500 g / kg, and the liquid-to-solid ratio is (3-4):1; More preferably, the acid solution is selected from one or more of sulfuric acid solution, hydrochloric acid solution and nitric acid solution.

5. The method for comprehensive recovery of valuable elements in clay-type lithium ores according to claim 4, characterized in that, In step S2, the pressurized acid leaching is carried out for 2.0h to 4.0h under the conditions of a pressure of 0.1MPa to 4.0MPa, a reaction temperature of 90℃ to 250℃, and a stirring rate of 500r / min to 800r / min. Preferably, the pressurized acid leaching is carried out for 3.0h to 4.0h under the conditions of a pressure of 3.4MPa to 3.5MPa, a reaction temperature of 240℃ to 250℃, and a stirring rate of 700r / min to 800r / min.

6. The method for comprehensive recovery of valuable elements from clay-type lithium ores according to any one of claims 1 to 5, characterized in that, In step S3, the oxygen pressure iron removal is carried out for 1.0h to 3.0h under the conditions of oxygen pressure of 1.0MPa to 3.0MPa, reaction temperature of 90℃ to 200℃, and stirring speed of 500r / min to 800r / min. Preferably, the oxygen pressure iron removal is carried out for 2.5h to 3.0h under the conditions of oxygen pressure of 2.0±0.2MPa, reaction temperature of 160℃~180℃, and stirring speed of 700r / min~800r / min.

7. The method for comprehensive recovery of valuable elements from clay-type lithium ores according to any one of claims 1 to 6, characterized in that, In step S4, the neutralization reaction is carried out at a pH of 4.6 to 4.8, and the reaction temperature is 25°C to 100°C, with a reaction time of 1.0 h to 2.0 h. Preferably, the neutralization reaction temperature is 40±5℃; More preferably, the neutralizing agent is selected from one or more of sodium hydroxide, sodium carbonate, calcium hydroxide, and magnesium hydroxide, and the neutralizing agent is added in the form of an aqueous solution with a mass concentration of 5% to 20%.

8. The method for comprehensive recovery of valuable elements from clay-type lithium ores according to any one of claims 1 to 7, characterized in that, In step S5, the precipitation reaction temperature is 50℃~100℃, and the reaction time is 0.5h~2.0h. Preferably, the reaction temperature of the precipitation reaction is 50℃~55℃; More preferably, the precipitating agent is selected from one or more of sodium hydrosulfide, sodium sulfite, and sodium metabisulfite, and the amount of the precipitating agent added is 0.01 g / L to 5 g / L.

9. The method for comprehensive recovery of valuable elements from clay-type lithium ores according to any one of claims 1 to 8, characterized in that, In step S6, the oxidation reaction is carried out at a pH of 4.0 to 8.0, and the reaction temperature is 25°C to 80°C, with a reaction time of 2.0 h to 6.0 h. Preferably, the oxidation reaction is carried out at a pH of 4.0 to 4.5, and the reaction temperature is 30 ± 5 °C, with a reaction time of 3.0 h to 4.0 h. More preferably, the oxidizing agent is selected from one or more of oxygen, sulfur dioxide, and carbon dioxide, and is even more preferably oxygen.

10. The method for comprehensive recovery of valuable elements in clay-type lithium ores according to any one of claims 1 to 9, characterized in that, The recovery rate of lithium is 40%~95%, the recovery rate of silicon is 75%~90%, the recovery rate of iron is 3%~12%, the recovery rate of aluminum is 10%~25%, the total recovery rate of nickel and cobalt is 30%~99.9%, and the recovery rate of manganese is 70%~95%.

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