Immersion mining method for potassium in polyhalite

Abstract

The invention belongs to the technical field of salt chemical industry, and particularly relates to a leaching mining method for potassium in polyhalite. A sodium chloride solution and pure water are used as leaching agents, efficient extraction of potassium resources is achieved through three-stage alternate leaching of induction, strengthening and stabilization, the sodium chloride solution is adopted at low pressure in the induction stage, the solubility of CaSO4 is reduced through the salting-out effect, deposition is inhibited, meanwhile, Na < + > is preferentially combined with Mg < 2 + > and SO4 < 2-> released by polyhalite to generate anhydrous natroedite, and the sodium chloride solution and the sodium chloride solution are separated out; the polyhalite is pushed to be dissolved, balanced and positively moved to realize rapid initial dissolution of potassium; in the strengthening stage, pure water is switched and pressure is increased in a gradient mode, permeation is enhanced, the leaching area is enlarged, the thermodynamic driving force of polyhalite dissolution is improved, and potassium dissolution is accelerated by means of the effects of low viscosity, high diffusivity and high pressure of the pure water; in the stable stage, high pressure is maintained, a low-concentration sodium chloride solution is adopted, calcium ion deposition and potassium ion reverse adsorption are inhibited, deep recovery of residual potassium is achieved, and therefore the leaching rate of potassium and the comprehensive benefits of the process are improved.

Description

TECHNICAL FIELD

[0001] The present application relates to the technical field of salt chemical industry, and particularly relates to a method for leaching and mining potassium in polyhalite. BACKGROUND

[0002] Polyhalite (K2SO4·MgSO4·2CaSO4·2H2O) is a common potassium, magnesium and calcium sulfate mineral, and the theoretical potassium content can reach 28.9% (calculated as K2SO4). The polyhalite mines in China are mainly distributed in Sichuan Basin, Jianghan Basin, Qaidam Basin, Tarim Basin (Ruo'bo, Kuqa and Shache sub-basins) and Dawenkou Basin, among which the polyhalite in Sichuan Basin is the most widely distributed and has the largest potential resource quantity. There are about more than 10 billion tons of polyhalite resources (equivalent to K2O) in Northeast Sichuan area. The huge resource quantity of polyhalite is in sharp contrast with the shortage of soluble potassium salt resources in China, and the efficient development and utilization of polyhalite resources has extremely important significance for guaranteeing the national potassium salt safety.

[0003] The polyhalite mines in Sichuan Basin are marine sediment deep-buried polyhalite mines (the ore body burial depth is mostly 1500-3500 meters), which are often layered in the gypsum and anhydrite rock or distributed in the granular, lump and banded salt matrix. The polyhalite has the characteristics of stable structure, high enthalpy and poor water solubility, and it is difficult to directly extract potassium sulfate. In view of the occurrence characteristics and the characteristics of polyhalite, at present, the leaching process is an effective way to exploit the deep-buried polyhalite mine. At present, the researches mainly focus on the selection of leaching agent, and the polyhalite leaching experiments are mostly carried out under normal pressure and stirring conditions, which cannot effectively solve the essential characteristics of the stable polyhalite lattice structure and poor water solubility, and cannot meet the comprehensive requirements of efficiency, cost and stability for large-scale production. SUMMARY

[0004] In view of the deficiencies of the prior art, the purpose of the present application is to provide a method for leaching and mining potassium in polyhalite, using sodium chloride solution and pure water as leaching agent, and realizing the efficient extraction of potassium resources through three-stage alternating leaching of induction, strengthening and stabilization: in the induction stage, sodium chloride solution is used under low pressure, the solubility of CaSO4 is reduced by salting-out effect to inhibit deposition, and at the same time, Na + Preferentially releasing Mg 2+ , SO4 2- Combined to generate anhydrous sodium magnesium alum, and promoting the forward movement of polyhalite dissolution balance to realize the rapid preliminary dissolution of potassium; in the strengthening stage, the pure water is switched and the pressure is increased, the low viscosity, high diffusivity and high pressure of pure water are used to enhance the penetration, expand the leaching area and improve the thermodynamic driving force of polyhalite dissolution, so as to accelerate the potassium dissolution; in the stabilization stage, the high pressure is maintained and the low-concentration sodium chloride solution is used to inhibit the deposition of calcium ions and the reverse adsorption of potassium ions, so as to realize the deep recovery of residual potassium, thereby improving the leaching rate of potassium and the comprehensive benefit of the process.

[0005] To achieve the above objectives, the present invention provides a method for leaching potassium from carnallite, comprising the following steps:

[0006] S1. The carnallite ore sample is crushed and screened to obtain a pretreated ore sample;

[0007] S2. Add the pretreated mineral sample to the high-pressure reactor, inject the first sodium chloride solution according to the liquid-solid ratio, raise the temperature to the first set temperature, adjust the pressure, perform static leaching, and obtain the induced stage leachate;

[0008] S3. Add pure water to maintain the liquid-solid ratio, maintain the temperature, increase the pressure to the first set pressure, perform static leaching, and obtain the leaching solution for the enhanced stage.

[0009] S4. Add the second sodium chloride solution to maintain the liquid-solid ratio, raise the temperature to the second set temperature, maintain the pressure, perform static leaching, cool down to room temperature, separate the solid and liquid, and obtain the stable stage leaching solution;

[0010] S5. Combine the induction, strengthening, and stabilization stage leachates to obtain a potassium-rich leachate.

[0011] In one feasible implementation, in S1, the mesh size of the crushing and screening is 200 mesh.

[0012] Pretreatment of mineral samples is based on the physical principle of mass transfer efficiency of solid particles: when the raw carnallite ore particles are large, the potassium element inside is encapsulated in a dense mineral lattice, making it difficult for the leaching medium to penetrate and contact. Through crushing and sieving, the particle size of the mineral sample is reduced, and the specific surface area is significantly increased. This not only increases the exposed surface area of ​​the minerals but also forms abundant micropores and fissures within the particles, providing ample channels for the subsequent penetration and diffusion of the leaching solution, thus breaking down the mass transfer barrier of the raw ore. Its role is to lay a foundation for efficient mass transfer in the entire leaching process, eliminate the limitation of the physical form of the mineral sample on leaching efficiency, ensure that the subsequent leaching agent can fully interact with the potassium-containing minerals in the mineral sample, avoid the phenomenon of external solubility but internal insolubility caused by excessively large particles, and improve the overall uniformity and efficiency of the leaching process.

[0013] In one feasible implementation, in step S2, the liquid-to-solid ratio is (7-10):1; the mass concentration of the first sodium chloride solution is 4wt%-8wt%; the first set temperature is 120-140℃, the heating rate is 8-15℃ / h; the adjusting pressure is 0.1-0.2MPa; and the static leaching time is 3-5h.

[0014] The induction stage utilizes the principle of chemical equilibrium shift driven by ion competition and precipitation: the high concentration of sodium ions provided by the sodium chloride solution synergistically interacts with the magnesium and sulfate ions dissolved from the carnallite, preferentially combining to form a stable anhydrous sodium magnesium alum precipitate. The formation of this precipitate continuously reduces the concentration of free sulfate ions in the system, thereby disrupting the dissolution equilibrium of the carnallite. The carnallite lattice continuously dissociates to replenish sulfate ions in the solution, while simultaneously releasing a large amount of potassium ions into the solution. On one hand, this achieves preliminary and efficient dissolution of potassium ions, constructing a virtuous cycle from dissolution to precipitation through a precipitation-induced mechanism, avoiding the problem of insufficient motive force in simple leaching; on the other hand, it lays the foundation for subsequent stages. Through preliminary dissolution under mild conditions, it protects the porous structure of the mineral sample from damage and provides a sufficient mass transfer pathway for subsequent enhanced leaching.

[0015] In one feasible implementation, in step S3, the first set pressure is 6-9 MPa, the pressure increase rate is 0.6-1 MPa / h, and the static leaching time is 6-8 h.

[0016] The enhancement stage relies on the physical mass transfer enhancement effect of pressure regulation: After the induction stage, some crystal lattices on the surface of the mineral sample have dissociated, but a large number of potassium ions encapsulated in micropores remain inside. At this point, relying solely on the concentration of the leaching agent is insufficient for efficient extraction. By supplementing with pure water to maintain a stable liquid-solid environment, and gradually increasing the pressure at a constant temperature to form a high-pressure system, the high pressure compresses the air gap space of the micropores inside the mineral sample, while reducing the diffusion resistance of potassium ions at the solid-liquid interface, allowing the solution to penetrate deeper into the mineral and fully contact the encapsulated potassium-containing crystal lattice. In addition, the high-pressure environment can maintain the structural stability of the mineral sample particles, preventing particle agglomeration and pore blockage. Its role is to connect the initial dissolution of the induction stage, achieve efficient extraction of deep residual potassium, significantly improve the cumulative potassium leaching rate, and balance leaching efficiency and energy consumption costs through a gentle pressure increase method, ensuring the economy and continuity of the process.

[0017] In one feasible implementation, in step S4, the mass concentration of the second sodium chloride solution is 1wt%-3wt%; the second set temperature is 135-145℃, and the heating rate is 8-15℃ / h; the static leaching time is 12-20h; the step of cooling to room temperature is: cooling to room temperature under the condition of circulating water flow rate of 50-80L / min; the solid-liquid separation method is vacuum filtration, and the vacuum degree of the vacuum filtration is 0.08-0.1MPa.

[0018] The stabilization phase utilizes a mass transfer optimization mechanism that combines ionic strength regulation and temperature assistance: a low-concentration sodium chloride solution provides suitable ionic strength, avoiding the inhibitory effect of high-concentration ions on potassium ion dissolution while maintaining active sites on the mineral sample surface and promoting the continuous dissociation of residual potassium-containing lattices; increasing temperature enhances the kinetic energy of solution molecules, accelerating the migration rate of potassium ions at the solid-liquid interface and further strengthening the dissolution efficiency of deep potassium. After leaching, a steady cooling process prevents solution supersaturation and secondary precipitation of potassium salts. Vacuum filtration efficiently separates the leachate from the slag, ensuring complete potassium ion entry into the liquid phase. This further achieves deep potassium leaching, minimizing potassium residue in the slag. Subsequent treatment ensures the stability and purity of the leachate, providing a qualified feedstock for subsequent potassium resource recovery, thus completing the closed loop of the entire leaching process.

[0019] In one feasible implementation, in step S5, the concentration of K+ in the potassium-rich leaching solution is 9.78-12.58 g / L, and the leaching rate of K+ in the potassium-rich leaching solution is 76.5%-89.8%.

[0020] Combining the leachates from each stage essentially involves consolidating dispersed potassium resources. By homogenizing the mixture, concentration differences between solutions at different stages are eliminated, resulting in a potassium-rich system with stable composition and concentrated potassium ion concentration. This maximizes the collection of potassium resources, avoiding resource losses or increased management costs associated with staged storage. Furthermore, it provides a stable raw material system for subsequent potassium salt separation and purification, reducing the adverse effects of solution concentration fluctuations on subsequent processes and improving the stability and efficiency of the entire potassium resource recovery process.

[0021] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0022] This method uses sodium chloride solution and pure water as leaching agents, and achieves efficient and low-cost extraction of potassium resources from carnallite through a staged alternating leaching process. In the induction stage, sodium chloride solution is used as the leaching agent, which, under low pressure, promotes full wetting and penetration of the leaching agent into the shallow pores of the carnallite, resulting in a high concentration of Na+ in the solution. + and Cl - The salting-out effect significantly reduces the solubility of CaSO4, while Na + Mg that can preferentially react with carnallite 2+ SO4 2- This process combines to form anhydrous sodium magnesium alum with a lower crystallization energy barrier. This process is achieved through the active consumption of SO4. 2- And fix part of Ca through ion exchange 2+ This inhibits the supersaturation deposition of CaSO4 at the source, preventing pore blockage; on the other hand, the continuous formation of anhydrous sodium magnesium alum constantly consumes Mg. 2+The process begins by shifting the dissolution equilibrium of carnallite in the positive direction, enabling rapid initial potassium leaching. In the intensification stage, pure water is used as the leaching agent, and a gradient pressurization is implemented. This gradient pressurization enhances the gradual penetration of water molecules into the micro-fractures of the ore. Combined with the low viscosity and high diffusivity of pure water, the effective leaching area is continuously expanded, increasing the mass transfer rate of potassium ions. Furthermore, by increasing the pressure gradient to high pressure, the thermodynamic driving force of the carnallite dissolution reaction is further enhanced, strengthening lattice disruption and ion dissociation processes, thereby accelerating potassium leaching and increasing the total leaching amount. In the stabilization stage, high pressure is maintained, and a low-concentration sodium chloride solution is used as the leaching agent. High pressure ensures sufficient contact between the leaching agent and the residual ore sample, maintaining the ionic strength of the system, inhibiting the deposition of residual calcium ions, and preventing the reverse adsorption of potassium ions. This provides a continuous and stable reaction environment for the deep recovery of residual potassium, ultimately achieving a dual optimization of potassium resource leaching rate and overall process efficiency. Attached Figure Description

[0023] Figure 1 This is a flowchart of a method for leaching potassium from carnallite according to the present invention.

[0024] Figure 2 This is an XRD pattern of potassium extraction during the leaching process of carnallite in this invention. Detailed Implementation

[0025] To make the technical problems, technical solutions, and beneficial effects to be solved by this application clearer, the application will be further described in detail below with reference to embodiments. However, this should not be construed as limiting the scope of this application to the following examples. All other embodiments obtained by those skilled in the art without creative effort without departing from the above-described methodological spirit of this application are within the scope of protection of this application.

[0026] The singular forms “for,” “or,” “a,” “any,” and “described” used in this application are intended to include the plural forms unless the context clearly indicates otherwise. Furthermore, the terms “first” and “second” are used for descriptive purposes only and should not be construed as indicating or implying relative importance.

[0027] Example 1

[0028] like Figure 1 As shown, a method for leaching potassium from carnallite includes the following steps:

[0029] S1. The Sichuan mixed halide ore sample was crushed and passed through a 200-mesh sieve to obtain a pretreated ore sample;

[0030] S2. Add the pretreated mineral sample to the high-pressure reactor, inject 0.8L of sodium chloride solution with a mass concentration of 6wt% at a liquid-to-solid ratio of 8:1, heat to 130℃ at a rate of 10℃ / h, adjust the pressure to 0.16MPa, and perform static leaching for 4h. After leaching, discharge the potassium-containing leaching solution into recovery tank A to obtain the leaching solution of the induction stage.

[0031] S3. Add pure water to the reactor to maintain a liquid-solid ratio of 8:1 and keep the temperature at 130℃. Increase the pressure to 8MPa at a rate of 0.8MPa / h and perform static leaching for 7h. After leaching, discharge the potassium-containing leaching solution into recovery tank B to obtain the enhanced stage leachate.

[0032] S4. Add a 2wt% sodium chloride solution to the reactor to maintain a liquid-solid ratio of 8:1. Increase the temperature to 140℃ at a rate of 10℃ / h and maintain the pressure at 8MPa. Perform static leaching for 18h. After leaching, allow the reactor to cool naturally to room temperature with a circulating water flow rate of 60L / min. Use vacuum filtration to separate the solid and liquid components of the leached system to obtain the leachate in the stable stage.

[0033] S5. Combine the leaching solutions from the induction, strengthening, and stabilization stages to obtain a potassium-rich leaching solution, wherein K + The concentration was 12.58 g / L, K + The leaching rate was 89.8%.

[0034] Example 2

[0035] like Figure 1 As shown, a method for leaching potassium from carnallite includes the following steps:

[0036] S1. The Sichuan mixed halide ore sample was crushed and passed through a 200-mesh sieve to obtain a pretreated ore sample;

[0037] S2. Add the pretreated mineral sample to the high-pressure reactor, inject 0.7L of sodium chloride solution with a mass concentration of 5wt% at a liquid-to-solid ratio of 7:1, heat to 120℃ at a rate of 8℃ / h, adjust the pressure to 0.1MPa, and perform static leaching for 3h. After leaching, discharge the potassium-containing leaching solution into recovery tank A to obtain the induced stage leachate.

[0038] S3. Add pure water to the reactor to maintain a liquid-solid ratio of 7:1, keep the temperature at 120℃, increase the pressure to 6MPa at a rate of 0.6MPa / h, and perform static leaching for 6h. After leaching, discharge the potassium-containing leaching solution into recovery tank B to obtain the enhanced stage leachate.

[0039] S4. Add a 1 wt% sodium chloride solution to the reactor to maintain a liquid-solid ratio of 7:1. Increase the temperature to 135°C at a rate of 8°C / h and maintain the pressure at 6 MPa. Perform static leaching for 12 hours. After leaching, allow the reactor to cool naturally to room temperature with a circulating water flow rate of 50 L / min. Use vacuum filtration to separate the solid and liquid components of the leached system and collect the potassium-rich brine produced during the stabilization phase.

[0040] S5. Combine the leaching solutions from the induction, strengthening, and stabilization stages to obtain a potassium-rich leaching solution, wherein K + The concentration was 12.24 g / L, K + The leaching rate was 76.5%.

[0041] Example 3

[0042] like Figure 1 As shown, a method for leaching potassium from carnallite includes the following steps:

[0043] S1. The Sichuan mixed halide ore sample was crushed and passed through a 200-mesh sieve to obtain a pretreated ore sample;

[0044] S2. Add the pretreated mineral sample to the high-pressure reactor, inject 1L of sodium chloride solution with a mass concentration of 8wt% at a liquid-to-solid ratio of 10:1, heat to 140℃ at a rate of 15℃ / h, adjust the pressure to 0.2MPa, and perform static leaching for 5h. After leaching, discharge the potassium-containing leaching solution into recovery tank A to obtain the induced stage leachate.

[0045] S3. Add pure water to the reactor to maintain a liquid-solid ratio of 10:1 and keep the temperature at 140℃. Increase the pressure to 9MPa at a rate of 1MPa / h and perform static leaching for 8 hours. After leaching, discharge the potassium-containing leaching solution into recovery tank B to obtain the enhanced stage leachate.

[0046] S4. Add a 3wt% sodium chloride solution to the reactor to maintain a liquid-solid ratio of 10:1. Increase the temperature to 140℃ at a rate of 15℃ / h and maintain the pressure at 9MPa. Perform static leaching for 20h. After leaching, allow the reactor to cool naturally to room temperature with a circulating water flow rate of 80L / min. Use vacuum filtration to separate the solid and liquid components of the leached system and collect the potassium-rich brine produced during the stabilization phase.

[0047] S5. Combine the leaching solutions from the induction, strengthening, and stabilization stages to obtain a potassium-rich leaching solution, wherein K + The concentration was 9.78 g / L, K + The leaching rate was 87.3%.

[0048] The key parameters of the staged leaching process were investigated to further optimize the leaching and mining process of potassium in carnallite. The results are shown in Tables 1, 2 and 3.

[0049] 1) By exploring the key parameters of the induction stage, the optimal sodium chloride concentration and pressure in the induction stage were determined to promote the formation of anhydrous sodium magnesium alum and improve the initial leaching rate.

[0050] Fixed conditions: mineral sample particle size <200 mesh, liquid-solid ratio 8:1, leaching agent dosage 0.8L, temperature 130℃, leaching time 4h.

[0051] Variable design: Sodium chloride concentration was set to three parameters: 4wt%, 6wt%, and 8wt%, and pressure was set to two parameters: 0.16MPa and 2MPa. A total of 6 crossover experiments were conducted.

[0052] Table 1. Optimization of sodium chloride concentration and pressure during the induction phase.

[0053]

[0054] In the induction stage, the high concentration of sodium ions in the sodium chloride solution combines with magnesium and sulfate ions dissolved from the carnallite to form anhydrous sodium magnesium alum precipitate under low pressure (0.16 MPa). This precipitate formation significantly reduces the sulfate ion concentration in the system, shifting the carnallite dissolution equilibrium to the positive side and accelerating potassium ion leaching from the ore sample. Table 1 shows that the precipitate formation is highest at a sodium chloride concentration of 6 wt%, with a potassium leaching rate of 63.12%. Too low a concentration results in insufficient sodium ions, low precipitate formation, and weak leaching driving force; too high a concentration increases the ionic strength of the system, inhibiting potassium ion diffusion and decreasing the leaching rate. When the pressure increases to 2 MPa, the system is unfavorable for the precipitation of anhydrous sodium magnesium alum, significantly reducing the leaching rate. Precipitation-induced, efficient initial leaching of potassium ions lays the foundation for a high leaching rate in the subsequent enhancement stage.

[0055] 2) Based on the above-obtained optimal conditions for the induction phase, K + The leaching rate was 63.12%. Based on the optimal conditions of the induction stage, the key parameters of the strengthening stage were explored to determine the optimal pressure increase rate and leaching time of the strengthening stage, and to enhance the diffusion and leaching of potassium ions from the interior of the ore sample under high pressure.

[0056] Fixed conditions: liquid-to-solid ratio 8:1, temperature 130℃, final pressure 8MPa, mineral sample particle size <200 mesh.

[0057] Variable design: The pressure increase rate was set to three parameters: 0.6MPa / h, 0.8MPa / h, and 1MPa / h; the leaching time was set to three parameters: 6h, 7h, and 8h. A total of 9 cross-experiments were conducted.

[0058] Table 2. Optimization of pressurization rate and leaching time during the intensification stage.

[0059]

[0060] According to Table 2, the optimal conditions for the enhanced stage investigation were a pressurization rate of 6 wt% and a leaching time of 8 h. The optimal K values ​​under these conditions were then obtained. + The initial leaching rate was 16.65%. The enhanced leaching stage leverages the promoting effect of high pressure on potassium ion diffusion. Building upon the partial dissolution of the ore sample surface and the initial formation of the pore structure during the induction stage, the 8 MPa high pressure compresses the internal pore space of the ore sample, reducing potassium ion diffusion resistance while maintaining the stability of the ore sample particle structure and preventing agglomeration and channel blockage. Controlling the pressurization rate is crucial; a rate of 0.8 MPa / h allows the ore sample pore structure to gradually adapt to pressure changes. Too fast a rate leads to particle agglomeration, while too slow a rate prolongs the process time, both reducing leaching efficiency. As the leaching time increases, potassium ions diffuse more fully from the ore sample into the solution, but further extension of the time results in minimal leaching gain. This stage serves as a bridge between the induction and leaching stages, enhancing the diffusion effect through high pressure to further improve the potassium leaching rate. Simultaneously, by optimizing the pressurization rate and leaching time, it balances leaching effectiveness with production efficiency, creating conditions for deep leaching in the stable stage.

[0061] 3) Based on the above-obtained optimal conditions for the induction phase, K + The leaching rate was 63.12%, and K was under optimal conditions in the intensification stage. + The leaching rate was 16.65%, based on the optimal conditions for the induction and enhancement stages and their K values. + Based on the sum of leaching rates (79.77%), the key parameters of the stabilization stage were explored to determine the optimal sodium chloride concentration, temperature and leaching time in the stabilization stage, so as to achieve deep leaching of potassium ions and improve the total leaching rate.

[0062] Fixed conditions: liquid-to-solid ratio 8:1, pressure 8MPa, mineral sample particle size <200 mesh.

[0063] Variable design: Sodium chloride concentration was set to three parameters: 1wt%, 2wt%, and 3wt%; temperature was set to three parameters: 130℃, 140℃, and 150℃; leaching time was set to three parameters: 16h, 18h, and 20h. The core combination was selected for the experiment, with a total of 6 groups.

[0064] Table 3. Optimization of sodium chloride concentration, temperature, and leaching time during the steady-state phase.

[0065]

[0066] According to Table 3, the optimal conditions for the stable phase investigation were a sodium chloride concentration of 2 wt%, a temperature of 150 °C, and a leaching time of 16 h. The optimal K values ​​under these conditions were obtained. + The stage leaching rate was 13.15%, and the total leaching rate was 90.29%. The stabilization stage involved comprehensive control of sodium chloride concentration, temperature, and time to achieve deep dissolution of potassium ions: a low concentration of 2 wt% sodium chloride maintained suitable ionic strength, preventing excessive concentration from inhibiting potassium ion diffusion while maintaining surface activity of the mineral sample through a small amount of sodium ions, promoting the dissolution of residual potassium ions; when the temperature reached 140℃, the viscosity of the leaching medium decreased, mass transfer efficiency improved, and the average molecular kinetic energy increased, accelerating the migration of potassium ions from the micropores inside the mineral sample into the solution, significantly increasing the leaching rate compared to 130℃; a leaching time of 16 hours allowed for the complete dissolution of the sparingly soluble potassium ions in the mineral sample, with subsequent extensions only slightly increasing the leaching rate. Building upon the enhancement stage, a mild and efficient combination of parameters was used to achieve deep potassium leaching while ensuring a suitable potassium concentration in the leaching solution for subsequent separation and purification, balancing leaching effect, energy consumption cost, and process stability, thus completing the closed loop of the entire staged leaching process.

[0067] The above results demonstrate and describe the basic principles and main features of this application, as well as its advantages.

[0068] The embodiments of the present invention have been described above with reference to the accompanying drawings. However, the present invention is not limited to the specific embodiments described above. The specific embodiments described above are merely illustrative and not restrictive. Those skilled in the art can make many other forms under the guidance of the present invention without departing from the spirit and scope of the claims. All of these forms are within the protection scope of the present invention.

Claims

1. A method for the leaching of potassium from carnallite, characterized in that, The method comprises the following steps: S1, crushing and screening the polyhalite ore sample to obtain a pretreated ore sample; S2, adding the pretreated ore sample into a high-pressure reaction kettle, adding a first sodium chloride solution according to a liquid-solid ratio, heating to a first set temperature, adjusting the pressure, static leaching, and obtaining an induced stage leaching solution; S3, supplementing pure water to maintain the liquid-solid ratio, maintaining the temperature, increasing the pressure to a first set pressure, static leaching, and obtaining a strengthened stage leaching solution; S4, supplementing a second sodium chloride solution to maintain the liquid-solid ratio, heating to a second set temperature, maintaining the pressure, static leaching, cooling to room temperature, and solid-liquid separation, and obtaining a stable stage leaching solution; S5, compounding the induced stage, strengthened stage and stable stage leaching solutions to obtain a potassium-rich leaching solution.

2. A process for the solution mining of potassium from carnallite according to claim 1, characterised in that, In the S1, the crushing and screening has a mesh size of 200 mesh.

3. A process for the solution mining of potassium from carnallite according to claim 1, characterised in that, In the S2, the liquid-solid ratio is (7-10):1, and the mass concentration of the first sodium chloride solution is 4wt%-8wt%.

4. A process for the solution mining of potassium from carnallite according to claim 1, characterized in that, In the S2, the first set temperature is 120-140℃, the heating rate is 8-15℃ / h, the adjusted pressure is 0.1-0.2MPa, and the static leaching time is 3-5h.

5. A process for the solution mining of potassium from carnallite according to claim 1, characterized in that, In the S3, the first set pressure is 6-9MPa, the pressure increasing rate is 0.6-1MPa / h, and the static leaching time is 6-8h.

6. A process for the solution mining of potassium from carnallite according to claim 1, characterized in that, In the S4, the mass concentration of the second sodium chloride solution is 1wt%-3wt%, the second set temperature is 135-145℃, and the heating rate is 8-15℃ / h.

7. A process for the solution mining of potassium from carnallite according to claim 1, characterized in that, In the S4, the static leaching time is 12-20h, and the cooling to room temperature is performed under the condition that the circulating water flow rate is 50-80L / min.

8. A process for the solution mining of potassium from carnallite according to claim 1, characterized in that, In the S4, the solid-liquid separation is performed by vacuum filtration, and the vacuum degree of the vacuum filtration is 0.08-0.1MPa.

9. A process for the solution mining of potassium from carnallite according to claim 1, characterized in that, In S5, the potassium-rich leaching solution contains K + The concentration of potassium in the potassium-rich leaching solution is 9.78-12.58 g / L. + The leaching rate is 76.5%-89.8%.

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