Method for efficiently and synergistically recovering iron, aluminum and magnesium from red mud and seawater desalination strong brine

Abstract

The invention discloses a method for efficiently and synergistically recovering iron, aluminum and magnesium from red mud and seawater desalination strong brine. According to the method, iron, aluminum and other valuable components are recycled from red mud through the steps of drying, acid leaching, extraction, back extraction and the like, finally, ferric chloride is recycled to serve as a coagulant, and the recycled aluminum and seawater desalination strong brine cooperate to synthesize magnesium-aluminum hydrotalcite. According to the method, main valuable metal iron and aluminum in the red mud can be synergistically recycled, the recycled ferric chloride serves as a coagulant, the performance of the recycled ferric chloride is equivalent to that of a commercial standard product, the recycled aluminum can be synergistically synthesized into magnesium-aluminum hydrotalcite with concentrated seawater, and the method has the potential of synergistically increasing the value of industrial byproducts and converting the industrial byproducts into market commodities, has remarkable economic and environmental benefits and is worthy of popularization and application. The method has an important application prospect in the field of industrial solid waste resource utilization.

Description

Technical Field

[0001] This invention belongs to the field of industrial solid waste resource utilization technology, specifically relating to a method for the efficient and synergistic recovery of iron, aluminum and magnesium from red mud and concentrated seawater desalination brine. Background Technology

[0002] Red mud is a highly alkaline solid waste generated during alumina production. In China, red mud reserves exceed 1 billion tons, with an annual output exceeding 100 million tons; however, its utilization rate remains below 10%. Red mud solid waste is primarily stored in landfills near alumina plants, placing a heavy burden on land resources and the surrounding environment. Damming and stockpiling not only occupies large amounts of land, but its high alkalinity and potential leaching risk of heavy metals also pose a long-term threat to the regional ecological environment. Red mud is rich in valuable metals such as iron (Fe), aluminum (Al), and titanium, as well as trace rare earth elements. The combined mass fraction of iron oxides (calculated as Fe2O3) and aluminum oxides (calculated as Al2O3) typically reaches 30-50%, making it a highly promising secondary metallurgical resource.

[0003] Currently, the main technologies for recovering valuable metals from red mud are pyrometallurgy and hydrometallurgy. Compared to pyrometallurgy, hydrometallurgical methods are typically carried out under milder conditions, offering the core advantages of precise separation capabilities and lower energy consumption. However, existing hydrometallurgical processes are mostly geared towards "single metal extraction": iron is preferentially extracted via acid leaching-extraction, and the resulting aluminum-containing leaching residue is either neutralized and discharged due to high subsequent impurity removal costs and low product added value, resulting in aluminum resource loss and additional consumption of alkali agents. Simultaneously, with the rapid expansion of the seawater desalination industry, approximately 1.5 cubic meters of concentrated brine are produced as a byproduct for every 1 cubic meter of freshwater produced. These concentrated brines are rich in chemical resources, such as magnesium ions (Mg²⁺). 2+ Concentrations of concentrated brine can reach levels of several grams per liter (g / L). Currently, concentrated brine is mostly disposed of by discharge into the sea. However, insufficient diffusion may lead to an increase in salinity in the discharge area, posing a potential threat to the local marine ecosystem. Therefore, promoting the resource utilization of concentrated brine (such as magnesium and bromine extraction) has become an important direction for sustainable industrial development. If the aluminum-rich leaching residue after red mud acid leaching for iron can be coupled with the magnesium source in seawater desalination concentrated brine, and co-precipitated under mild conditions to synthesize magnesium-aluminum hydrotalcite (Mg-Al LDH), not only can the simultaneous high-value utilization of aluminum and magnesium be achieved, but the traditional neutralization process can also be eliminated, reducing alkali and salt consumption in the red mud treatment process, forming a closed-loop cycle of "red mud-concentrated brine" dual solid waste synergy and iron, aluminum, and magnesium ternary co-recovery. Therefore, developing a short-process technology based on "acid leaching-extraction-back-extraction for iron" and "leaching residue-concentrated brine synergy for magnesium-aluminum hydrotalcite" has significant economic and environmental implications for improving the overall resource utilization rate of red mud, alleviating the supply pressure of natural bauxite and magnesium ore, and reducing the environmental burden of concentrated brine. Summary of the Invention

[0004] To address the shortcomings of existing technologies, the present invention aims to provide a method for the efficient synergistic recovery of iron, aluminum, and magnesium from red mud and concentrated seawater desalination brine, thereby overcoming the low resource recovery efficiency and high cost of existing technologies. This method, through the synergistic reaction of two waste materials, achieves simultaneous and efficient recovery and value-added transformation of multiple components such as iron, aluminum, and magnesium, converting industrial by-products into marketable products, thus yielding significant economic and environmental benefits.

[0005] The objective of this invention is achieved through the following technical solution:

[0006] A method for the efficient synergistic recovery of iron, aluminum, and magnesium from red mud and concentrated seawater desalination brine is disclosed. This method involves steps such as drying, acid leaching, extraction, and back-extraction to recover valuable components like iron and aluminum from red mud. Ferric chloride is ultimately recovered as a coagulant, and the recovered aluminum is synergistically reacted with the concentrated seawater desalination brine to synthesize magnesium aluminum hydrotalcite. The specific steps are as follows:

[0007] S1: After the red mud is dried to constant weight, it is mixed with hydrochloric acid and stirred. Then, it is sealed and heated in a constant temperature stirring water bath to obtain an acidic leachate.

[0008] S2: Using kerosene-diluted tributyl phosphate (TBP) as the extractant, the acidic leachate obtained in step S1 was extracted, and the upper layer loaded with Fe was collected separately. 3+ The organic phase and the lower layer contain Al 3+ The residual extract;

[0009] S3: The organic phase obtained in step S2 is back-extracted with hydrochloric acid, and the concentrated solution containing FeCl3 in the lower layer is collected. This concentrated FeCl3 solution can be used as a liquid flocculant, or the solution can be further evaporated and concentrated to obtain FeCl3 solid coagulant.

[0010] S4: Concentrated seawater (containing Mg) – a byproduct of seawater desalination. 2+ (concentrated brine), and the Al-containing solution obtained in step S2 3+ The raffinate was mixed with the mixture, and the mixture was co-precipitated with an alkaline aqueous solution of NaOH-Na2CO3 by titration and stirring.

[0011] S5: The slurry obtained in step S4 is sealed and continuously stirred and aged at a constant temperature. Then, the solid precipitate is collected by centrifugation. The precipitate is then ultrasonically dispersed in deionized water and anhydrous ethanol. Finally, the precipitate is dried to constant weight, cooled to room temperature, and ground into powder to obtain magnesium aluminum hydrotalcite.

[0012] Preferably, in step S1, the drying temperature of the red mud is 50°C-70°C, the concentration of hydrochloric acid used for acid leaching is 4-10 mol / L, and the liquid-to-solid ratio is 10-50 mL / g.

[0013] Preferably, the conditions for constant temperature stirring water bath heating in step S1 are: stirring speed of 500-1000 rpm, temperature of 60°C-100°C, and heating time of 1-2 h.

[0014] Preferably, the kerosene in step S2 is sulfonated kerosene, the volume fraction of TBP in the extractant is 30%-90%, the ratio of extractant to water is 0.75-2:1, the extraction time is 3-5 min, the number of extraction stages is 1-2, and the extraction temperature is 25°C-35°C.

[0015] Preferably, in step S3, the concentration of hydrochloric acid used for back-extraction is 0.5-1 mol / L, the back-extraction temperature is 25°C-60°C, and the number of back-extraction stages is 1-4.

[0016] Preferably, the Al in the raffinate in step S4 3+ The concentration ranges from 14.8 to 54.1 mmol / L. The molar ratio of magnesium to aluminum in the mixture of concentrated brine and raffinate is 2:1 to 4:1. The OH- concentration in the alkaline aqueous solution is... - The molar concentration of CO3 is 1-2 times the total molar concentration of aluminum and magnesium in the mixture. 2- The molar concentration of aluminum is 1-2 times that of aluminum in the mixture.

[0017] Preferably, in step S4, the titration rate of the mixture of concentrated brine and raffinate and the alkaline aqueous solution is 1-2 mL / min, and the pH during the titration process is precisely maintained at 10.0 ± 0.1.

[0018] Preferably, in step S5, the stirring temperature of the slurry is 50°C, the stirring time is 16-24 h, and the crystallization temperature for drying the precipitate to constant weight is 60°C-80°C.

[0019] The present invention has the following beneficial effects:

[0020] This invention can co-recover the main valuable metals iron and aluminum from red mud. The recovered ferric chloride can be used as a coagulant with performance comparable to commercial benchmark products. The recovered aluminum can be co-synthesized with seawater desalination byproducts—concentrated seawater desalination brine—to form magnesium aluminum hydrotalcite. This invention has the potential to co-value-added industrial byproducts and transform them into marketable commodities, and has significant economic and environmental benefits. It also has important application prospects in the field of industrial solid waste resource utilization. Attached Figure Description

[0021] Figure 1 : Process flow diagram of the present invention.

[0022] Figure 2 Leaching efficiency of metal elements in red mud under different types of acid (a), liquid-to-solid ratio (b), temperature (c), and hydrochloric acid concentration (d).

[0023] Figure 3 Iron extraction efficiency and partition coefficient under different TBP volume fractions (a), O / A ratios (b), extraction times (c), extraction stages (d), and temperatures (e).

[0024] Figure 4 Iron back-extraction efficiency under different back-extraction O / A and hydrochloric acid concentrations (a); iron back-extraction efficiency under different back-extraction stages and temperatures (b).

[0025] Figure 5 Phosphorus removal (a) and turbidity removal (b) efficiencies of synthetic FeCl3 coagulant and commercial FeCl3 coagulant at different dosages.

[0026] Figure 6 XRD analysis of hydrotalcite samples synthesized under different magnesium-aluminum ratios (a), aluminum concentrations (b), and magnesium sources (c). Detailed Implementation

[0027] The following embodiments are further illustrations of the present invention, but not limitations thereof.

[0028] Example 1

[0029] Red mud samples were obtained from an alumina plant in Shandong Province, China. The red mud samples were dried to constant weight in a 60°C oven and then ground to a particle size of less than 500 µm. One g of dried red mud was mixed with 20 mL of 6 mol / L hydrochloric acid in a beaker. The mixture was stirred at 500 rpm and heated in an 80°C water bath for 60 minutes. After heating, the leachate was filtered through a 0.45 µm filter membrane, and the concentrations of iron and aluminum in the filtrate sample were determined by inductively coupled plasma optical emission spectrometry (ICP-OES).

[0030] Comparative Example 1

[0031] Referring to Example 1, the difference is that the acids used are replaced with 9.2 mol / L sulfuric acid, 7.5 mol / L phosphoric acid and 7.9 mol / L nitric acid, and the other steps are the same as in Example 1.

[0032] Comparative Example 2

[0033] Referring to Example 1, the difference is that the volume of hydrochloric acid is changed to 10 mL, 30 mL and 50 mL, and the other steps are the same as in Example 1.

[0034] Comparative Example 3

[0035] Referring to Example 1, the difference is that the temperature of the constant temperature water bath is changed to 40°C, 50°C, 60°C, 70°C and 100°C, and the other steps are the same as in Example 1.

[0036] Comparative Example 4

[0037] Referring to Example 1, the difference is that the hydrochloric acid concentration is changed to 4 mol / L, 8 mol / L and 10 mol / L, and the other steps are the same as in Example 1.

[0038] like Figure 2 As shown in Figure a, all acids except nitric acid can efficiently leach aluminum. Sulfuric acid has the highest leaching efficiency for aluminum, exceeding 90%, but its leaching efficiency for iron is only about 55%. Compared to others, hydrochloric acid achieves a significantly higher iron leaching efficiency of 88.78%, while also achieving a leaching efficiency of 84.96% for aluminum. Overall, HCl demonstrates significant advantages in co-leaching efficiency, process simplicity, and versatility, achieving efficient co-leaching of iron and aluminum while also exhibiting good leaching efficiency for calcium and magnesium. From the perspective of comprehensive resource recovery, hydrochloric acid promotes the simultaneous extraction of multiple metals from red mud, providing an ideal feedstock for subsequent separation and purification steps.

[0039] like Figure 2 As shown in b, increasing the liquid-to-solid ratio reduces the viscosity of the leaching system, which enhances the diffusion of hydrogen ions to the surface of red mud particles and the transfer of dissolved metal ions to the bulk solution. However, increasing the liquid-to-solid ratio from 20 mL / g to 30 mL / g only slightly improves the leaching efficiency (4.46% for Fe and 1.08% for Al), indicating that the reaction is close to equilibrium at a liquid-to-solid ratio of 20 mL / g. Furthermore, increasing the liquid-to-solid ratio leads to increased acid consumption and wastewater generation, thereby increasing process costs.

[0040] like Figure 2 As shown in Figure c, the leaching efficiency of aluminum is already high at relatively low temperatures, and fluctuates around 78% as the temperature further increases. The leaching efficiency of iron, however, increases with temperature, reaching 88.46% at 80°C. Further increasing the temperature to 100°C results in only a slight increase in iron leaching efficiency but a significant rise in energy costs.

[0041] like Figure 2 As shown in Figure d, when the hydrochloric acid concentration is 4 mol / L, the leaching efficiency of iron is around 70%. Increasing the hydrochloric acid concentration to 6 mol / L can increase the leaching efficiency of iron to 88.78%. Further increasing the hydrochloric acid concentration does not significantly improve the leaching of iron. In contrast, the leaching efficiency of aluminum reaches over 80% at all tested hydrochloric acid concentrations.

[0042] Example 2

[0043] Tributyl phosphate (TBP) was diluted with sulfonated kerosene to obtain an extract with a TBP volume fraction of 50%. The filtrate obtained in Example 1 was extracted with the extract at a volume ratio of 1:1. Single-stage extraction was performed at 25°C for 5 min, followed by standing in a separatory funnel for 30 min to achieve complete phase separation. The organic and aqueous phases were collected separately. The iron concentration in the raffinate was determined by ICP-OES, and the iron concentration in the organic phase sample was calculated based on mass balance.

[0044] Comparative Example 5

[0045] Referring to Example 2, the difference is that the volume fraction of TBP in the extract is changed to 10%, 30%, 70% and 90%, while the other steps are the same as in Example 2.

[0046] Comparative Example 6

[0047] Referring to Example 2, the difference is that the volume ratio of extract to filtrate is changed to 0.25:1, 0.5:1, 0.75:1 and 2:1, and the other steps are the same as in Example 2.

[0048] Comparative Example 7

[0049] Referring to Example 2, the difference is that the extraction time is changed to 1 min, 2 min, 3 min and 4 min, and the other steps are the same as in Example 2.

[0050] Comparative Example 8

[0051] Referring to Example 2, the difference is that the number of extraction stages is changed to 2, 3 and 4, while the other steps are the same as in Example 2.

[0052] Comparative Example 9

[0053] Refer to Example 2, except that the extraction temperature is changed to 40℃, 50℃, and 60℃, while the other steps are the same as in Example 2.

[0054] like Figure 3 As shown in figure a, the iron extraction efficiency increases significantly with increasing TBP volume fraction. At a TBP volume fraction of 10%, the iron extraction efficiency is less than 30%. However, at a volume fraction of 50%, the efficiency rises significantly to 97.13%, indicating that iron is almost completely extracted from the aqueous phase. When the TBP volume fraction exceeds 50%, the viscosity and density of the organic phase increase, leading to a decrease in interfacial tension and a slower phase separation rate.

[0055] like Figure 3As shown in b, when the ratio of extract to filtrate (organic phase to aqueous phase, O / A ratio) increases from 0.25 to 2, the iron extraction efficiency gradually increases from 32.96% to nearly 100%. When the O / A ratio is ≤ 0.75, the iron extraction efficiency is still below 90%, and the formed complex has limited solubility in the organic phase, leading to the formation of a third phase and emulsification, making phase separation difficult. When the O / A ratio is 1, the iron extraction efficiency reaches 97.26%, and the partition coefficient increases sharply to 35.46. Further increasing the O / A ratio only marginally improves the iron extraction efficiency.

[0056] like Figure 3 As shown in c, TBP exhibits rapid kinetics in iron extraction, achieving an extraction efficiency of 96.86% after only 5 minutes of mixing, indicating that the system has essentially reached chemical equilibrium. Figure 3 As shown in d, increasing the number of extraction stages only slightly increases the extraction efficiency of iron, but significantly increases the partition coefficient. Figure 3 As shown in Figure e, the extraction efficiency of iron is 96.76% at 25°C. Further increasing the temperature leads to a slight decrease in both the extraction efficiency and the partition coefficient.

[0057] Example 3

[0058] The organic phase prepared in Example 2 was back-extracted using 0.5 mol / L hydrochloric acid as the back-extraction agent, with a volume ratio of organic phase to back-extraction agent of 1.5:1. Single-stage back-extraction was performed at 25°C for 5 min. The sample was then allowed to stand in a separatory funnel for 30 min to achieve complete phase separation. The lower aqueous phase containing FeCl3 was collected. The iron concentration in the aqueous phase sample was determined by ICP-OES.

[0059] Comparative Example 10

[0060] Refer to Example 3, except that the hydrochloric acid concentration is changed to 0 (deionized water), 1, 2 and 3 mol / L, and the other steps are the same as in Example 3.

[0061] Comparative Example 11

[0062] Referring to Example 3, the difference is that the back-extraction temperature is changed to 40°C, 50°C and 60°C, and the other steps are the same as in Example 3.

[0063] like Figure 4 As shown in Figure a, at a hydrochloric acid concentration of 0.5 mol / L, the back-extraction efficiency of iron is 100%. However, as the hydrochloric acid concentration increases, the back-extraction efficiency of iron gradually decreases, and when the hydrochloric acid concentration reaches 2 mol / L, the back-extraction of iron is almost completely inhibited. Figure 4As shown in b, the effect of temperature on the iron back-extraction efficiency is negligible; as the temperature increases, the iron back-extraction efficiency only decreases slightly from 97.62% to 95.85%. Similarly, the number of back-extraction stages has a very small effect on the iron back-extraction efficiency; when the back-extraction technology is increased from one stage to four stages, the iron back-extraction efficiency increases slightly from 96.7% to 99.52%.

[0064] Example 4

[0065] The aqueous phase containing FeCl3 prepared in Example 3 was dried in an oven at 60°C until the liquid was completely evaporated, yielding black FeCl3 as a solid coagulant. The coagulation performance of the prepared FeCl3 solid coagulant was determined using phosphorus-containing and turbid synthetic wastewater.

[0066] The iron content in the synthesized FeCl3 ranged from 31.48% to 32.04%, close to the specifications (32.0%-33.0%) required by the Chinese national standard (GB / T 4482-2018) for water treatment agents—ferric chloride. Figure 5 As shown, within the range of coagulant dosage of 20 to 120 mg / L (Fe / P = 0.88-5.27), the phosphorus removal efficiency of the synthesized FeCl3 coagulant increased from 8.66% to 90.24%, and the turbidity removal efficiency increased from 2.80% to 91.40%, both of which are comparable to the effects of commercial FeCl3 coagulants.

[0067] Example 5

[0068] Seawater was collected from Ganjingzi District, Dalian, and concentrated to 50% of its original volume to obtain concentrated brine for desalination. Besides magnesium, the concentrated brine also contained impurities such as potassium, sodium, and calcium; its main chemical components are shown in Table 1. The concentrated brine was mixed with the aqueous phase prepared in Example 2. The aluminum-magnesium mixture had an aluminum concentration of 54.1 mmol / L and a magnesium to aluminum molar ratio of 3:1. A mixed alkaline aqueous solution of NaOH and Na₂CO₃ was prepared, wherein OH⁻... - The molar concentration of CO3 is 1.6 times the total molar concentration of aluminum and magnesium in the aluminum-magnesium mixture. 2- The molar concentration of the aluminum-magnesium mixture was twice that of aluminum. 100 mL of deionized water was placed in a titration beaker, and the initial pH was adjusted to 10.0 ± 0.1 by adding an alkaline aqueous solution. Then, while stirring at 500-600 rpm, the aluminum-magnesium mixture and the alkaline aqueous solution were simultaneously titrated into the beaker at a controlled rate of 2 mL / min. By fine-tuning the relative addition rate, the pH was precisely maintained at 10.0 ± 0.1 throughout the titration process.

[0069] The titrated slurry was aged under sealed, continuous stirring at 50°C for 24 h. The solid product was collected by centrifugation at 3000 rpm for 15 min, sonicated for 3 min, washed three times with deionized water, and then three times with anhydrous ethanol to ensure complete dispersion. The product was dried to constant weight in a 70°C oven, cooled to room temperature, and then ground into a fine powder using an agate mortar to obtain magnesium aluminum hydrotalcite. The mineral phase of the synthesized hydrotalcite was identified by X-ray diffraction (XRD).

[0070] Table 1. Main chemical components of concentrated brine

[0071] Comparative Example 12

[0072] Referring to Example 5, the difference is that the molar ratio of magnesium to aluminum in the aluminum-magnesium mixture is changed to 2:1 and 4:1, while the other steps are the same as in Example 5.

[0073] Comparative Example 13

[0074] Referring to Example 5, the difference is that the aluminum concentration in the aluminum-magnesium mixture is changed to 14.8 and 37.0 mmol / L, while the other steps are the same as in Example 5.

[0075] Comparative Example 14

[0076] Referring to Example 5, the difference is that the concentrated brine from seawater desalination is replaced with a commercial MgCl2 solution, while the other steps are the same as in Example 5.

[0077] like Figure 6 As shown in Figure a, the XRD patterns of samples with magnesium-aluminum ratios of 4 and 3 exhibit sharp, strong diffraction peaks, indicating a highly ordered layered structure. In contrast, the sample with a magnesium-aluminum ratio of 2 shows broadened and weakened peaks, indicating poor crystallinity and lower structural order. Figure 6 b shows that the intensity of all characteristic diffraction peaks increases with increasing aluminum concentration. For example... Figure 6 As shown in Figure c, the peak intensity and crystallinity of hydrotalcite synthesized from concentrated brine from seawater desalination are similar to those of hydrotalcite synthesized from commercial MgCl2 solution.

[0078] The above detailed description is a specific description of the embodiments of the present invention. These embodiments are not intended to limit the patent scope of the present invention. All equivalent implementations or modifications that do not depart from the present invention should be included in the patent scope of this case.

Claims

1. A method for the efficient synergistic recovery of iron, aluminum, and magnesium from red mud and concentrated seawater desalination brine, characterized in that, Includes the following steps: S1: After the red mud is dried to constant weight, it is mixed with hydrochloric acid and stirred. Then, it is sealed and heated in a constant temperature stirring water bath to obtain an acidic leachate. S2: Using tributyl phosphate diluted with kerosene as the extractant, the acidic leachate obtained in step S1 was extracted, and the upper layer loaded with Fe was collected separately. 3+ The organic phase and the lower layer contain Al 3+ The residual extract; S3: The organic phase obtained in step S2 is back-extracted with hydrochloric acid, and the concentrated solution containing FeCl3 in the lower layer is collected. This concentrated FeCl3 solution can be used as a liquid flocculant, or the solution can be further evaporated and concentrated to obtain FeCl3 solid coagulant. S4: Contains Mg 2+ The seawater desalination brine and the Al-containing brine obtained in step S2 3+ The raffinate was mixed with the mixture, and the mixture was co-precipitated with an alkaline aqueous solution of NaOH-Na2CO3 by titration and stirring. S5: The slurry obtained in step S4 is sealed and continuously stirred and aged at a constant temperature. Then, the solid precipitate is collected by centrifugation. The precipitate is then ultrasonically dispersed in deionized water and anhydrous ethanol. Finally, the precipitate is dried to constant weight, cooled to room temperature, and ground into powder to obtain magnesium aluminum hydrotalcite.

2. The method according to claim 1, characterized in that, In step S1, the drying temperature of the red mud is 50°C-70°C, the concentration of hydrochloric acid used for acid leaching is 4-10 mol / L, and the liquid-to-solid ratio is 10-50 mL / g.

3. The method according to claim 1, characterized in that, The conditions for constant temperature stirring water bath heating in step S1 are: stirring speed of 500-1000 rpm, temperature of 60°C-100°C, and heating time of 1-2 h.

4. The method according to claim 1, characterized in that, The kerosene mentioned in step S2 is sulfonated kerosene, the volume fraction of tributyl phosphate in the extractant is 30%-90%, the ratio of extractant to water is 0.75-2:1, the extraction time is 3-5 min, the number of extraction stages is 1-2, and the extraction temperature is 25°C-35°C.

5. The method according to claim 1, characterized in that, In step S3, the concentration of hydrochloric acid used for back-extraction is 0.5-1 mol / L, the back-extraction temperature is 25°C-60°C, and the number of back-extraction stages is 1-4.

6. The method according to claim 1, characterized in that, In step S4, Al in the raffinate 3+ The concentration ranges from 14.8 to 54.1 mmol / L. The molar ratio of magnesium to aluminum in the mixture of concentrated brine and raffinate is 2:1 to 4:

1. The OH- concentration in the alkaline aqueous solution is... - The molar concentration of CO3 is 1-2 times the total molar concentration of aluminum and magnesium in the mixture. 2- The molar concentration of aluminum is 1-2 times that of aluminum in the mixture.

7. The method according to claim 1, characterized in that, In step S4, the titration rate of the mixture of concentrated brine and raffinate, as well as the alkaline aqueous solution, is 1-2 mL / min, and the pH during the titration process is precisely maintained at 10.0 ± 0.

1.

8. The method according to claim 1, characterized in that, In step S5, the stirring temperature of the slurry is 50°C, the stirring time is 16-24 h, and the temperature for drying the sediment to constant weight is 60°C-80°C.

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