WO2022033041A1 - 一种离子吸附型稀土的提取方法 - Google Patents

一种离子吸附型稀土的提取方法 Download PDF

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WO2022033041A1
WO2022033041A1 PCT/CN2021/083462 CN2021083462W WO2022033041A1 WO 2022033041 A1 WO2022033041 A1 WO 2022033041A1 CN 2021083462 W CN2021083462 W CN 2021083462W WO 2022033041 A1 WO2022033041 A1 WO 2022033041A1
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rare earth
leaching
solution
aluminum sulfate
container
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PCT/CN2021/083462
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English (en)
French (fr)
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李永绣
李鸿阳
王康
周华娇
周雪珍
刘艳珠
李东平
李静
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南昌大学
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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/10—Hydrochloric acid, other halogenated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/40—Mixtures
    • C22B3/402—Mixtures of acyclic or carbocyclic compounds of different types
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B59/00—Obtaining rare earth metals
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • the invention belongs to the technical field of rare earth hydrometallurgy and environmental protection.
  • the invention relates to a high-efficiency green extraction method for successively leaching ion adsorption type rare earths with calcium chloride and aluminum sulfate as main leaching agents.
  • Ion-adsorbed rare earth is a unique rare earth deposit first discovered and named in Longnan, Jiangxi province. This rare earth can be exchange leached by various types of electrolyte solutions. Taking advantage of this property, a variety of ion adsorption-type rare earth extraction methods and processes have been developed. Among them, the first-generation sodium chloride leaching process with sodium chloride as the main leaching agent was the earliest application. The advantage is that the price of raw materials is low and the source is convenient. However, the leaching ability of sodium chloride is poor, and the required concentration is high. It needs more than 5wt% of sodium chloride to have a good leaching effect, and the actual concentration is about 7wt%.
  • Jiangxi University and other units have developed the second-generation ammonium sulfate leaching process using ammonium sulfate as the leaching agent.
  • the method requires a low concentration of leaching agent, 1wt% to 3wt%, the leaching efficiency is obviously improved, and the leaching efficiency of 1.5wt% ammonium sulfate can reach more than 90%.
  • the method has the advantages of short process flow, low consumption, small waste water volume, significantly reduced cost, high and stable product purity, low residual electrolyte concentration in tailings, and is conducive to plant growth. Therefore, it has been the main leaching technology for mining since it was popularized and applied in the early 1980s.
  • ammonia nitrogen pollution from ammonium sulfate leaching is a common concern.
  • the preparation of leaching agent and the collection and management of leachate can be done well in one cycle per day.
  • the pollution to the water body is not obvious, but it is beneficial to the restoration of the tailings vegetation.
  • the concentration of ammonium sulfate used is high, the seepage is difficult to control, and the leakage loss of leaching agent solution and leaching solution leads to serious excess of ammonia nitrogen and rare earth ions in the water body of the mining area.
  • the second is the "source technology” that takes into account the elimination of ammonia nitrogen pollution, the improvement of the yield of rare earth, and the stabilization of tailings.
  • the Chinese patent (201310594438.2) "a method for improving the leaching rate of ionic rare earth and the stability of tailings" proposes The staged leaching method, while exerting the advantages of ammonium sulfate leaching process, proposes a low-acidity leaching section for recovering colloidal adsorption phase rare earths in mines and a limewater tailing section for reducing ammonia nitrogen residue in tailings and pollutant overflow.
  • the present invention proposes a new method for leaching ion adsorption rare earth in stages by using calcium chloride as the leaching agent and then using aluminum sulfate as the leaching agent in the leaching process to solve the above problems.
  • the invention provides an extraction method of ion adsorption type rare earth, and the extraction method comprises the following steps:
  • the concentration of the calcium chloride solution is between 0.02 and 0.25 mol/L, and the pH value is between 4 and 8.
  • the liquid-solid ratio of the calcium chloride solution to the ion adsorption type rare earth ore is 0.6-1.0.
  • the calcium chloride solution is allowed to contain certain amounts of other monovalent and divalent non-ammonium cationic electrolytes and additives.
  • the concentration of the aluminum sulfate solution is between 0.02 and 0.20 mol/L, and the pH value is between 2 and 4.
  • the liquid-solid ratio of the aluminum sulfate solution to the ion adsorption type rare earth ore is 0.1-0.5.
  • the aluminum sulfate solution is allowed to contain a certain amount of other non-ammonium cationic electrolytes and additives.
  • the liquid-solid ratio of the lime water to the ion adsorption type rare earth tailings is 0.2-0.5.
  • step S3 the amount of lime aqueous solution in the lime water leaching is judged on the basis that the pH value of the effluent is greater than 5, and after the pH value reaches 5, top immersion with water is used to gradually increase the pH value to 5. After 6 stop water injection.
  • the calcium chloride solution includes calcium chloride and an electrolyte solution with calcium chloride as the main component recovered from the calcium-containing waste;
  • the aluminum sulfate solution includes aluminum sulfate and sulfuric acid recovered from the aluminum-containing waste
  • it also includes the following steps: - using lime precipitation to obtain rare earth hydroxide in the rare earth in the second container, using hydrochloric acid to dissolve the rare earth chloride solution with high concentration, low aluminum and low sulfate radicals, adding the insoluble slag into the third container, so that the The rare earth and aluminum hydroxide are dissolved, and then N1923-kerosene-iso-octanol mixed organic phase is used to extract and separate rare earth and aluminum, and the raffinate contains aluminum sulfate, which is recycled for the second stage of leaching; The rare earth is back-extracted with chloride, the back-extracted liquid is incorporated into the second container, and the rare earth hydroxide is prepared by precipitation with lime.
  • the obtained rare earth hydroxide can be used to saponify the p507 organic phase, and the acid generated by the saponification is recycled for dissolving the rare earth hydroxide, reducing the amount of acid required for optimally dissolving the rare earth hydroxide, without discharging saponification wastewater.
  • the total leaching efficiency is 99.76%, and the leaching efficiency is significantly improved;
  • the absolute value of the zeta potential of the first stage leaching is 13.2mv, which is close to the original ore;
  • the absolute value of the zeta potential after adding aluminum sulfate in the second stage is 2mv, which proves the stability of the tailings
  • aluminum sulfate is used in the second step, mainly from the recovery and recycling of leached ion adsorption aluminum, and the usage amount is small, thereby avoiding the subsequent extraction caused by the use of aluminum sulfate Increased separation workload and severe loss of organic phase.
  • Figure 1 The leaching efficiency of each leaching agent with a cation concentration of 0.128N.
  • Figure 2 The effect of calcium chloride cation concentration on the leaching efficiency (add 40ml of top water after calcium chloride leaching).
  • Figure 3 Influence of the amount of aluminum sulfate added on the leaching efficiency (calcium chloride 100-90-80-70-60-50ml + top water 10ml + aluminum sulfate 0-10-20-30-40-50ml + top water 40ml).
  • Figure 4 The proportion of rare earth ions leached in two stages (experimental conditions are the same as in Figure 3).
  • Figure 5 The effect of the amount of aluminum sulfate added on the zeta potential (experimental conditions are the same as in Figure 3).
  • Figure 6 Variation of leaching efficiency with the ratio of aluminum sulfate leaching without topping water between calcium chloride-aluminum sulfate leaching.
  • Figure 7 Consumption of lime water in the process of improving the pH of tailings leaching with aluminum sulfate (the usage amount of aluminum sulfate is 10-20-30ml).
  • Figure 8 Influence of the addition amount of top make-up water on the leaching efficiency after the second stage leaching (total mass of ore 100g calcium chloride 80ml+ top make-up water 10ml+ aluminum sulfate 20ml+ top make-up water 10/20/30/40ml).
  • Figure 9 The influence of the type of leaching agent on the leaching efficiency in the second stage, showing that aluminum sulfate has a higher leaching efficiency (the liquid-solid ratio of the leaching agent in the first stage (0.8:1) + 10ml top water, the leaching agent in the second stage Mineral agent (0.2:1) + 40ml top hydration).
  • Figure 10 Comparison of zeta potential values of tailings clay minerals after leaching rare earth with different leaching agents, it shows that the higher the cation valence, the lower the absolute value of zeta potential (the first stage zeta potential on the left is the second stage zeta potential on the right) .
  • Figure 11 XRD pattern of rare earth hydroxide obtained by calcium oxide precipitation.
  • Figure 12 SEM image of rare earth hydroxide obtained by calcium oxide precipitation.
  • Figure 13 The graph of the concentration of the organic phase of P507 extracted by rare earth saponification with the concentration of rare earth, the phase and the number of stages (the relationship between the concentration of rare earth in the loaded organic phase and the concentration and phase of the rare earth in the feed liquid).
  • Figure 14 The change of pH value of raffinate aqueous phase with rare earth concentration, phase and series after rare earth saponification extraction of P507 organic phase (the relationship between pH value of raffinate and rare earth concentration and phase of feed liquid).
  • Figure 15 The relationship between the concentration of rare earth ions in the feed liquid and the distribution ratio.
  • Figure 17 Ammonium chloride stripping of rare earth-supported N1923 organic phase (compared to the effect of ammonium chloride concentration on stripping rate).
  • Figure 18 Rare earth ion concentration of solution after stripping rare earth supported N1923 organic phase with ammonium chloride with different phases and concentrations.
  • Figure 19 Flow chart of ionic rare earth leaching, precipitation, extraction and separation, and recycling.
  • the aluminum sulfate solution with a concentration of 0.02mol/L and a pH value of 2 as the leaching agent, the aluminum sulfate solution and the ion-adsorption rare earth raw ore are leached for the secondary leaching of the ion adsorption rare earth raw ore at a liquid-solid ratio of 0.4:1.
  • the leachate with pH value greater than 4 is collected in the second container, and the leachate with pH value less than 4 is collected in the third container.
  • the ion adsorption type rare earth tailings are leached with lime water, and the liquid-solid ratio of the lime water to the ion adsorption type rare earth tailings is 0.2:1, so that the rare earth ions in the ion adsorption type rare earth tailings and the aluminum sulfate trapped between the rare earth particles outflow.
  • the ion adsorption type rare earth tailings are leached with lime water until the pH value of the leachate is greater than 5. Continue leaching with water until pH is greater than 6. Wherein, the leachate with pH value greater than 4 is collected in the first container, and the leachate with pH value less than 4 is collected in the third container.
  • the ion adsorption rare earth raw ore is washed with water to remove most of the calcium and rare earth ions in the gap of the ore layer to prevent aluminum sulfate from entering the ore layer to form sulfuric acid Calcium precipitation hinders the subsequent leaching process.
  • Said container refers to containers such as liquid storage tanks in mines, liquid storage tanks in workshops, and beakers in laboratories.
  • the rare earth content of the solution in the first container is very low, which can be directly used to prepare calcium chloride leaching agent and recycled for the leaching of the first stage, and can also be used for irrigation water for ecological restoration of tailings.
  • the leaching solution of calcium chloride as leaching agent calcium oxide is added to precipitate rare earth and aluminum, and then filtered to obtain rare earth hydroxide; then, rare earth is preferentially dissolved by hydrochloric acid, and the pH value of the solution is controlled to be 4.5.
  • High-concentration rare earth chloride solution the excellent slag is dissolved in an acidic solution leached by sulfuric acid and aluminum sulfate, the rare earth is extracted with a mixed organic phase of N1923-kerosene-iso-octanol, and aluminum sulfate is left in the water phase, and is recycled for the second stage.
  • the rare earth in the organic phase is back-extracted with hydrochloric acid and NH 4 Cl to obtain a rare earth chloride solution, and the total amount of rare earth ions is determined.
  • the leaching solution with aluminum sulfate as leaching agent N1923-kerosene-iso-octanol organic phase is used to extract rare earth directly, aluminum sulfate is left in the water phase, and recycled for the second stage leaching; rare earth in the organic phase is extracted with hydrochloric acid and NH 4 Cl stripping is complete, rare earth ions are obtained, and the total amount of rare earth ions is determined.
  • an aluminum sulfate solution with a concentration of 0.20 mol/L and a pH value of 4 was used as the leaching agent, and the aluminum sulfate solution and the ion-adsorbed rare earth ore were subjected to secondary leaching at a liquid-solid ratio of 0.5:1.
  • the leachate with pH value greater than 4 is collected in the second container, and the leachate with pH value less than 4 is collected in the third container.
  • the ion adsorption type rare earth tailings are leached with lime water, and the liquid-solid ratio of lime water and ion adsorption type rare earth tailings is 0.5:1, so that the rare earth ions in the ion adsorption type rare earth tailings and the aluminum sulfate trapped between the rare earth particles outflow.
  • the ion adsorption type rare earth tailings are leached with lime water until the pH value of the leachate is greater than 5. Continue leaching with water until pH is greater than 6. Wherein, the leachate with pH value greater than 4 is collected in the second container, and the leachate with pH value less than 4 is collected in the third container.
  • the ion adsorption rare earth raw ore is washed with water to remove most of the calcium and rare earth ions in the gap of the ore layer to prevent aluminum sulfate from entering the ore layer to form sulfuric acid Calcium precipitation hinders the subsequent leaching process.
  • the leaching solution of calcium chloride as the leaching agent calcium oxide is added to precipitate rare earth and aluminum, and then filtered to obtain rare earth hydroxide; then, the rare earth is preferentially dissolved by hydrochloric acid, and the pH value of the solution is controlled to be 5. After filtration, low aluminum and low sulfate radicals are obtained. High-concentration rare earth chloride solution; the excellent slag is dissolved in an acidic solution leached by sulfuric acid and aluminum sulfate, the rare earth is extracted with a mixed organic phase of N1923-kerosene-iso-octanol, and aluminum sulfate is left in the water phase, and is recycled for the second stage.
  • the rare earth in the organic phase is back-extracted with hydrochloric acid and NH 4 Cl to obtain a rare earth chloride solution, and the total amount of rare earth ions is determined.
  • the leaching solution with aluminum sulfate as leaching agent N1923-kerosene-iso-octanol organic phase is used to extract rare earth directly, aluminum sulfate is left in the water phase, and recycled for the second stage leaching; rare earth in the organic phase is extracted with hydrochloric acid and NH 4 Cl stripping is complete, rare earth ions are obtained, and the total amount of rare earth ions is determined.
  • the electrolyte solution with calcium chloride as the main component recovered from the calcium-containing waste with a concentration of 0.20mol/L and a pH value of 6 is used as the leaching agent.
  • the ion-adsorbed rare earth is leached at a ratio of 0.7:1, wherein the leaching solution containing no rare earth is initially collected in the first container, and then the leaching solution containing rare earth is collected in the second container.
  • the aluminum sulfate solution with a concentration of 0.10 mol/L and a pH value of 3 as the leaching agent, the aluminum sulfate solution and the ion adsorption rare earth raw ore are leached for the second time according to the liquid-solid ratio of 0.3:1 to the ion adsorption rare earth raw ore.
  • the leachate with pH value greater than 4 is collected in the second container, and the leachate with pH value less than 4 is collected in the third container.
  • the ion adsorption type rare earth tailings are leached with lime water, and the liquid-solid ratio of lime water to ion adsorption type rare earth tailings is 0.3:1, so that the rare earth ions in the ion adsorption type rare earth tailings and the aluminum sulfate retained between the rare earth particles outflow.
  • the ion adsorption type rare earth tailings are leached with lime water until the pH value of the leachate is greater than 5. Continue leaching with water until pH is greater than 6. Wherein, the leachate with pH value greater than 4 is collected in the second container, and the leachate with pH value less than 4 is collected in the third container.
  • the ion adsorption rare earth raw ore is washed with water to remove most of the calcium and rare earth ions in the gap of the ore layer to prevent aluminum sulfate from entering the ore layer to form sulfuric acid Calcium precipitation hinders the subsequent leaching process.
  • the leaching solution of calcium chloride as the leaching agent calcium oxide is added to precipitate rare earth and aluminum, and then filtered to obtain rare earth hydroxide; then, the rare earth is preferentially dissolved by hydrochloric acid, and the pH value of the solution is controlled to be 4.7.
  • High-concentration rare earth chloride solution the excellent slag is dissolved in an acidic solution leached by sulfuric acid and aluminum sulfate, the rare earth is extracted with a mixed organic phase of N1923-kerosene-iso-octanol, and aluminum sulfate is left in the water phase, and is recycled for the second stage.
  • the rare earth in the organic phase is back-extracted with hydrochloric acid and NH 4 Cl to obtain a rare earth chloride solution, and the total amount of rare earth ions is determined.
  • the leaching solution with aluminum sulfate as leaching agent N1923-kerosene-iso-octanol organic phase is used to extract rare earth directly, aluminum sulfate is left in the water phase, and recycled for the second stage leaching; rare earth in the organic phase is extracted with hydrochloric acid and NH 4 Cl stripping is complete, rare earth ions are obtained, and the total amount of rare earth ions is determined.
  • ion adsorption type rare earth raw ore Take ion adsorption type rare earth raw ore, sieve the raw ore with 20 mesh, take the fine particle part under 20 mesh, and use it after homogenization. The seepage flow is more uniform; the electrolyte solution with calcium chloride as the main component recovered from the calcium-containing waste with a concentration of 0.02mol/L and a pH value of 5 is used as the leaching agent.
  • the ion adsorption type rare earth is leached at a ratio of 0.9:1, wherein the leaching solution containing no rare earth is collected in the first container at first, and then the leaching solution containing rare earth is collected in the second container.
  • the electrolyte solution with aluminum sulfate as the main component in the aluminum-containing recycling waste with a concentration of 0.15 mol/L and a pH value of 3 was used as the leaching agent.
  • the ion adsorption type rare earth ore is subjected to secondary leaching, wherein the leaching solution with pH value greater than 4 is collected in the second container, and the leaching solution with pH value less than 4 is collected in the third container.
  • the ion adsorption type rare earth tailings are leached with lime water, and the liquid-solid ratio of lime water to ion adsorption type rare earth tailings is 0.4:1, so that the rare earth ions in the ion adsorption type rare earth ions and the aluminum sulfate trapped between the rare earth particles flow out.
  • the ion adsorption type rare earth tailings are leached with lime water until the pH value of the leachate is greater than 5. Continue leaching with water until pH is greater than 6. Wherein, the leachate with pH value greater than 4 is collected in the second container, and the leachate with pH value less than 4 is collected in the third container.
  • the ion adsorption rare earth is washed with water to remove most of the calcium and rare earth ions in the interstices of the ore layer to prevent aluminum sulfate from entering the ore layer to form calcium sulfate Precipitation hinders the subsequent leaching process.
  • the leaching solution with calcium chloride as the leaching agent calcium oxide is added to precipitate rare earth and aluminum, and then filtered to obtain rare earth hydroxide; then, the rare earth is preferentially dissolved by hydrochloric acid, and the pH value of the solution is controlled to be 4.8, and after filtration, a low-aluminum and low-sulfate solution is obtained.
  • High-concentration rare earth chloride solution the excellent slag is dissolved in an acidic solution leached by sulfuric acid and aluminum sulfate, the rare earth is extracted with a mixed organic phase of N1923-kerosene-iso-octanol, and aluminum sulfate is left in the water phase, and is recycled for the second stage.
  • Leaching that is, the leaching of aluminum sulfate as a leaching agent; the rare earth in the organic phase is back-extracted with hydrochloric acid and NH 4 Cl to obtain a rare earth chloride solution, and the total amount of rare earth ions is determined.
  • N1923-kerosene-iso-octanol organic phase is used to extract rare earth directly, aluminum sulfate is left in the water phase, and recycled for the second stage leaching; rare earth in the organic phase is extracted with hydrochloric acid and NH 4 Cl stripping is complete, rare earth ions are obtained, and the total amount of rare earth ions is determined.
  • a comparative method was adopted for the calculation of the leaching efficiency of rare earth ions obtained in the above examples.
  • the leaching efficiency is higher than that of calcium chloride leaching alone, and the amount of calcium chloride used is lower;
  • the leaching efficiency is the same as that of aluminum sulfate leaching alone, but the amount of aluminum sulfate is far less than the amount used alone, so that while ensuring high leaching efficiency, it can also reduce the amount of aluminum sulfate used, thereby reducing the subsequent extraction workload. .
  • the leaching efficiency of each leaching agent at 0.128N was measured. As shown in Figure 1, the results show that the leaching rate of aluminum sulfate is the highest, close to 98%, followed by ammonium sulfate, about 95%; the leaching rate of magnesium sulfate is less than 80%, and the leaching rate of calcium chloride is only about 70%.
  • the leaching efficiency can be gradually improved, from 70% to 88%; when the concentration is higher than 0.25mol/L, the leaching efficiency increases slowly, And blindly increasing the concentration of cations will cause excessive leaching agent to pollute the environment and increase the cost.
  • the optimal leaching agent concentration is 0.178mol/L.
  • the leaching efficiency is less than 90%, which proves that it is difficult to achieve a high leaching rate when the rare earth is leached by calcium chloride alone. This is also the main reason why calcium chloride leaching is not used in the past.
  • the experimental method is to add 0.256N calcium chloride leaching solution with a liquid-solid ratio of 1-0.5:1 in the first stage, then add a small amount of top water at a liquid-solid ratio of 0.1:1, and then add a liquid-solid ratio of 0 to 0.5 in the second stage. :1 Add 0.128N aluminum sulfate, keep the liquid-solid ratio of each leaching at 1, then add top replenishment water at a liquid-solid ratio of 0.4:1 to collect the leaching solution, measure the rare earth content, and calculate the leaching efficiency. The results are shown in Figure 3.
  • the results show that the leaching efficiency increases continuously with the addition of aluminum sulfate; the leaching efficiency increases from about 85% to nearly 100%.
  • the leaching efficiency can be increased to about 94% only by replacing 10% of calcium chloride solution, and the leaching efficiency can be increased to more than 96% when replacing 20%.
  • the leaching efficiency of rare earth in the first stage decreases, which is not large at the beginning, and then decreases significantly.
  • the ratio of the two leaching agents is equal, the total amount of rare earth ions leached in the aluminum sulfate leaching section is greater than that in the calcium chloride leaching section, and the total leaching efficiency of the two stages is significantly improved.
  • Example 6 A method similar to that in Example 5 was used, except that the top make-up water added between the two leaching stages was removed and supplemented after the aluminum sulfate leaching stage. The results are shown in Figure 6. It is found that the overall leaching efficiency is very close to the experimental data under the corresponding conditions in Example 5 under the condition of keeping the same amount of top make-up water, which proves that the top make-up water between the two leaching sections can be omitted.
  • Example 5 in the process of the second stage leaching, add aluminum sulfate solution at a liquid-solid ratio of 0.1-0.3:1, then add top supplementary water at a liquid-solid ratio of 0.4:1, and finally use saturated clarified lime water to drench Wash tailings to improve tailings pH, collect tailings effluent and measure its pH.
  • Figure 7 it can be seen that with the increase of the amount of aluminum sulfate added, the acidity of the soil increases, and the amount of lime water consumed to improve soil pH also increases, and considering that the amount of aluminum sulfate added is proportional to the workload of subsequent extraction, Reducing the addition of aluminum sulfate can reduce the amount of lime required for subsequent tailings treatment.
  • the residual rare earth is less than 0.01%.
  • the lime water for subsequent tail protection can also be used as a top make-up water shower.
  • the residual rare earth ions are washed, that is, the first 0.4% of the eluent after leaching will contain rare earth ions, which has recovery value.
  • the concentration of each leaching agent is 0.128N, and the volume ratio of the two leaching agents is 0.8:0.2, respectively combining calcium chloride + aluminum sulfate, magnesium sulfate + aluminum sulfate, calcium chloride + calcium chloride and ammonium chloride + Aluminum sulfate to measure the leaching efficiency, as shown in Figure 9, which proves that adding aluminum sulfate in the second stage will improve the leaching efficiency to varying degrees on the original basis, that is, staged leaching will not affect the leaching efficiency of the respective leaching agents.
  • the rare earth leached by calcium chloride is recovered by precipitation with calcium hydroxide, and the method for precipitating rare earth with calcium hydroxide (milk of lime) is to mix the clarified calcium chloride leaching solution with calcium hydroxide (milk of lime) under stirring.
  • the solution is simultaneously added to the suspension containing rare earth hydroxide crystal precipitate in a certain proportion, so that the formed precipitate can quickly form rare earth hydroxide crystal.
  • the supernatant is transferred to lime water Pool and calcium chloride leaching reagent configuration pool to achieve recycling. A part of the rare earth hydroxide remains in the sedimentation tank as a seed crystal for subsequent precipitation and continues to be used.
  • the rare earth is directly extracted with the acid extraction organic phase (P507-kerosene organic phase), and part of the rare earth enters the organic phase to reach the loaded rare earth concentration required for the extraction and separation of the rare earth, and the exchanged hydrogen ions enter the water phase, which is recycled for dissolving rare earth hydroxide,
  • the water phase is circulated in the system without discharging saponification waste water.
  • saponification experiments were carried out using different concentrations of rare earth feed liquid and cascade extraction stages and comparisons.
  • the rare earth ion concentration of the obtained aqueous and organic phases and the analysis results of typical impurity elements such as aluminum and thorium are shown in Table 2-4, which proves that when the concentration of the rare earth feed liquid is between 0.4 and 1.2, the organic phase can achieve the subsequent rare earth extraction.
  • the load required for separation is between 0.18-0.26mol/L.
  • Figure 14 shows the pH value of the raffinate aqueous phase corresponding to the conditions, and this low acidity raffinate aqueous phase can be used to dissolve rare earth hydroxides.
  • Table 4 shows the distribution ratio D of each element and the separation coefficient ⁇ of adjacent elements. It can be seen that the separation factor is beneficial to light and heavy rare earths. Grouping of elements.
  • Table 2 Concentration and distribution of each element during P507 extraction equilibrium under different feed concentrations.
  • Table 4 The distribution ratio D of each element and the separation coefficient ⁇ of adjacent elements at the time of extraction equilibrium.
  • the rare earth ions leached from aluminum sulfate in the second stage were extracted with an N1923 organic phase at a suitable ratio.
  • 100g of 20 mesh rare earth ore was used for leaching at a liquid-solid ratio of 1:1, and Al 2 (SO 4 ) 3 with a cation concentration of 0.128N was used for leaching.
  • the rare earth stripping rate increases with the increase of the concentration of ammonium chloride, and increases with the increase of the volume of ammonium chloride, which is because more and more chloride ions will be eluted by the extracted rare earth ion associates. .
  • the insoluble part of the solid is dissolved in sulfuric acid to obtain an aluminum sulfate solution containing rare earth, and then N1923 is used to extract and separate rare earth and Al 3+ .
  • Another method is to mix these solids with aluminum sulfate leaching solution, and control the pH of the solution to be around 3, so that both rare earth hydroxide and aluminum can be dissolved.
  • the supernatant is mainly composed of aluminum sulfate and a small amount of rare earth, and the insoluble matter is clay minerals and a small amount of calcium sulfate.
  • N1923 was used to extract rare earths from the supernatant, and the raffinate water phase was deoiled by air flotation, and the concentration and pH of aluminum sulfate were adjusted, and then continued to be used for leaching of ion-adsorbed rare earths.
  • the rare earth in the organic phase is back-extracted with NH 4 Cl solution, and the N1923 organic phase after the back-extraction of the rare earth continues to be recycled.
  • the back-extracted rare earth ions are precipitated with lime or ammonium bicarbonate to obtain the corresponding rare earth hydroxide or carbonate, and the rare earth oxide is obtained after calcination.
  • the method of Example 4 is used to prepare a high-concentration rare earth feed liquid and a saponified organic phase for subsequent extraction and separation. The whole process is shown in Figure 19.

Abstract

本发明提供一种离子吸附型稀土的提取方法，所述提取方法包括以下步骤：S1：以氯化钙溶液为浸取剂对离子吸附型稀土原矿进行浸出，稀土以离子形式进入浸出液；其中，最初不含稀土的浸出液收集于第一容器，后续含稀土的浸出液收集于第二容器；S2：以硫酸铝溶液为浸取剂对所述离子吸附型稀土原矿进行第二次浸出，稀土原矿中剩余的稀土以离子形式进入浸出液；其中，pH值大于4的浸出液仍收集于所述第二容器；pH值小于4的浸出液，收集于第三容器；S3：用石灰水对所述浸出后的离子吸附型稀土尾矿进行浸淋；然后，用水浸淋，直至浸出液的pH值大于6；其中，pH值小于4的浸出液，收集于所述第三容器，pH值大于4的浸出液仍收集于所述第一容器。

Description

一种离子吸附型稀土的提取方法 技术领域
本发明属于稀土湿法冶金和环境保护技术领域。涉及一种以氯化钙和硫酸铝为主要浸取剂先后浸取离子吸附型稀土的高效绿色提取方法。
背景技术
离子吸附型稀土是最早在江西龙南发现并命名的一类独特的稀土矿床，其稀土含量低且主要是以离子态被矿床中的各种黏土矿物和胶体粒子所吸附。这种稀土可以被各种类型的电解质溶液交换浸出。利用这一性质，开发了多种离子吸附型稀土提取方法和工艺。其中最早应用的是以氯化钠作为主体浸取剂的第一代氯化钠浸矿工艺。其优点是原料价格低、来源方便。但是，氯化钠的浸取能力差，要求的浓度高，需要5wt％以上的氯化钠才有较好的浸取效果，实际使用的浓度为7wt％左右。而且用草酸从这种高浓度氯化钠浸出液中沉淀稀土时，会形成草酸稀土与钠的复盐沉淀，煅烧后的产品中含有大量的氧化钠，需要经过水洗和再次煅烧才能得到稀土总量在92％以上的合格矿产品。工艺流程长，消耗大、废水量大，收率低，尾矿中残留的氯化钠浓度高，导致土地盐碱化，不利于尾矿植被恢复。
为此，江西大学等单位开发了第二代以硫酸铵作为浸取剂的硫酸铵浸矿工艺。该方法所需浸矿剂的浓度低，1wt％～3wt％即可，浸取效率明显提高，1.5wt％硫酸铵的浸矿效率就可达到90％以上。相较于氯化钠浸取工艺，该方法流程短、消耗小、废水量小、成本显著降低，产品纯度高且稳定，尾矿中电解质残留浓度低，且有利于植物生长。因而从上世纪八十年代初推广应用以来一直是矿山开采的主体浸取技术。
然而，硫酸铵浸矿的氨氮污染是人们普遍关注的问题。早期在用池浸方式开采时，每天一个周期，浸取剂的配制和浸出液的收集和管理都可以做得很好。除了尾矿中残留的铵氮外，对水体的污染不明显，反而对尾矿植被恢复有益。但是，当采用原地浸矿工艺时，使用的硫酸铵浓度高，渗流难以控制，浸矿剂溶液和浸出液的外泄损失导致矿区水体中氨氮和稀土离子超标严重。与此同时，硫酸铵原地浸矿过程的水土流失和山体崩塌事故也十分突出。南昌大学的研究证明导致水土流失和滑坡塌方的原因是由于浸矿过程和尾矿中残留的铵离子在黏土表面的吸附能力弱，容易从紧密层迁移到扩散层而产生较大的负zeta电位值，使尾矿在雨水的不断冲刷下出现滑坡，不仅导致了稀土资源的流失，而且残留的氨氮随雨水冲刷进入附近水体导致氨氮污染。
为解决硫酸铵浸取的氨氮污染问题，提出了许多解决方案。主要有两条基本途径。
一是简单的使用非铵浸取试剂替代硫酸铵作为浸取剂的“源头方案”。中国专利201010128302.9“一种从离子型稀土原矿回收稀土的方法”是以硫酸镁、氯化镁、氯化钙中的至少一种代替大部分甚至全部的硫酸铵、氯化铵或氯化钠作为浸取剂，用于浸取离子吸附型稀土；中国专利201310199034.3“一种离子吸附型稀土提取方法”是采用硫酸镁、或硫酸镁和/或硫酸铁、或硫酸镁和/或硫酸铝为主成分的水溶液作为浸取剂，浸取离子吸附型稀土；中国专利201310424572.8“离子吸附型稀土矿非按盐浸取稀土的工艺”则采用任意确定钙盐、镁盐、钠盐的配比，并按确定的配比配制形成复合盐作为浸取剂浸取离子吸附型稀土；这些方法对于替代铵盐浸取剂来消除氨氮污染是可取的，但浸取率不高，或因使用方法不当导致了浸取效率不高的问题。与此同时，钙、镁、铝、铁等高价离子在后续稀土沉淀分离过程中的共沉淀使产品纯度不高，收率降低，若全部采用萃取方法来分离，其成本高，有机相损失，废水中的磷和COD超标问题突出。
二是把消除氨氮污染和提高稀土收率、稳定尾矿等问题同时考虑的“源头 技术”，中国专利(201310594438.2)“一种提高离子型稀土浸取率和尾矿稳定性的方法”提出了分阶段的浸取方法，在发挥硫酸铵浸取工艺优势的同时，提出了回收矿中胶态吸附相稀土的低酸度浸取段和降低尾矿氨氮残留和污染物溢出的石灰水护尾段，同时也可以降低由于矿层胶态化趋势所导致的滑坡塌方风险；中国专利201610821052.4“一种以硫酸铝为浸取剂的离子型稀土高效绿色提取方法”提出了以硫酸铝作为新一代绿色浸取剂的离子吸附型稀土提取新工艺。首次证明硫酸铝的浸取效率是最高的，尾矿也更稳定。然而，如果单纯使用硫酸铝来浸取，浸出液中的稀土都需要用萃取法来分离，会导致矿山生产成本增加，有机相损失增大。历史上，有关从离子吸附型稀土浸出液中萃取回收富集稀土的研究报道有很多，包括酸性磷类萃取剂和羧酸类萃取剂；工业上也有过短暂的应用，其突出的问题是有机相损失导致的成本升高和水污染。因此，如何保持高的浸取效率又能改善尾矿稳定性且减少后续萃取的工作量，是目前离子吸附型稀土绿色开采亟待解决的关键技术问题。
发明内容
本发明提出在浸矿过程中先使用氯化钙作为浸取剂，再使用硫酸铝作为浸取剂来分段浸取离子吸附型稀土的新方法以解决上述问题。
本发明提供一种离子吸附型稀土的提取方法，所述提取方法包括以下步骤：
S1：以氯化钙溶液为浸取剂对离子吸附型稀土原矿进行浸出，稀土原矿中的稀土以离子形式进入浸出液；其中，最初不含稀土的浸出液收集于第一容器，后续含稀土的浸出液收集于第二容器；
S2：以硫酸铝溶液为浸取剂对所述离子吸附型稀土原矿进行第二次浸出，稀土原矿中剩余的稀土以离子形式进入浸出液；其中，pH值大于4的浸出液仍收集于所述第二容器；pH值小于4的浸出液，收集于第三容器；
S3：用石灰水对所述浸出后的离子吸附型稀土尾矿进行浸淋；然后，用水浸淋，直至浸出液的pH值大于6；其中，pH值小于4的浸出液，仍收集于所 述第三容器，pH值大于4的浸出液仍收集于所述第一容器。
优选地，在步骤S1中，所述氯化钙溶液的浓度在0.02～0.25mol/L之间，pH值在4～8之间。
优选地，在步骤S1中，所述氯化钙溶液与所述离子吸附型稀土原矿的液固比为0.6～1.0。
优选地，所述氯化钙溶液中允许含有一定量的其他一价和两价非铵阳离子电解质和添加剂。
优选地，在步骤S2中，所述硫酸铝溶液的浓度在0.02～0.20mol/L之间，pH值在2-4之间。
优选地，在步骤S2中，所述硫酸铝溶液与所述离子吸附型稀土原矿的液固比为0.1～0.5。
优选地，所述硫酸铝溶液中允许含有一定量的其他非铵阳离子电解质和添加剂。
优选地，步骤S3中，所述石灰水与所述离子吸附型稀土尾矿的液固比为0.2～0.5。
优选地，步骤S3中，所述的石灰水浸淋中的石灰水溶液用量是以流出液的pH值大于5为依据来判断，pH值达到5以后，用水顶浸，使pH值逐渐升高到6以后停止注水。
优选地，所述氯化钙溶液包括氯化钙以及从含钙废物中回收的以氯化钙为主要成分的电解质溶液；所述硫酸铝溶液包括硫酸铝以及从含铝废物中回收的以硫酸铝为主要成分的电解质溶液。
优选地，还包括以下步骤：-第二容器中的稀土采用石灰沉淀得到氢氧化稀土，用盐酸优溶得高浓度低铝低硫酸根的氯化稀土溶液，不溶渣加入第三容器，使其中的稀土和铝的氢氧化物溶解，再用N1923-煤油-异辛醇混合有机相萃取分离稀土与铝，萃余液中含硫酸铝，循环用于第二阶段的浸矿；有机相中的稀土用氯化物反萃，反萃液并入第二容器中，用石灰沉淀发制取氢氧化稀土。
优选地，所获得的氢氧化稀土可以用于皂化p507有机相，皂化产生的酸循环用于溶解氢氧化稀土，降低优溶氢氧化稀土所需的酸量，不排放皂化废水。
。
用本发明提供的提取方法来浸出离子吸附型稀土既可以保证高的稀土浸出率，又能减少后续萃取的工作量，节约成本，而且尾矿稳定，不影响植物生长。实验数据表明：在整个浸出的过程中，总浸出效率最高能达到99.76％。以使用100g的20目离子吸附性型稀土矿按照液固比0.8:1加入0.178mol/L的氯化钙、然后再按照液固比0.2:1加入0.043mol/L的硫酸铝浸出剂为例，总浸出效率为99.76％，浸出效率明显提升；第一段浸取zeta电位绝对值是13.2mv，与原矿的接近；第二段加入硫酸铝之后zeta电位绝对值为2mv，证明尾矿稳定性得到明显改善；且在本发明中，在第二步中使用硫酸铝，主要来源于浸出的离子吸附型铝的回收和循环使用，且使用量小，从而避免了由于硫酸铝的使用导致后续萃取分离工作量的加大，以及有机相的严重损失。
附图说明
图1：各浸取剂阳离子浓度为0.128N时的浸出效率。
图2：氯化钙阳离子浓度对浸出效率的影响(氯化钙浸取后顶补水加40ml)。
图3：硫酸铝加入量对浸出效率的影响(氯化钙100-90-80-70-60-50ml+顶补水10ml+硫酸铝0-10-20-30-40-50ml+顶补水40ml)。
图4：两阶段浸出稀土离子的占比(实验条件同图3)。
图5：硫酸铝加入量对zeta电位的影响(实验条件同图3)。
图6：氯化钙-硫酸铝浸取之间不加顶补水时浸出效率随硫酸铝浸取比例的变化。
图7：改善硫酸铝浸取尾矿pH过程中石灰水的消耗量(硫酸铝的使用量10-20-30ml)。
图8：第二阶段浸取后顶补水的加入量对浸出效率的影响(矿的总质量100g氯化钙80ml+顶补水10ml+硫酸铝20ml+顶补水10/20/30/40ml)。
图9：第二阶段浸矿剂种类对浸出效率的影响，显示硫酸铝又更高的浸取效率(第一阶段浸矿剂液固比(0.8:1)+10ml顶补水，第二阶段浸矿剂(0.2:1)+40ml顶补水)。
图10：不同浸取剂浸取稀土后尾矿黏土矿物的zeta电位值比较，显示阳离子价态越高，zeta电位绝对值越低(左边为第一阶段zeta电位右边为第二阶段zeta电位)。
图11：氧化钙沉淀所得氢氧化稀土的XRD图谱。
图12：氧化钙沉淀所得氢氧化稀土SEM图。
图13：稀土皂化萃取P507有机相的浓度随稀土浓度、相比和级数的变化图(负载有机相稀土浓度随料液稀土浓度和相比的变化关系)。
图14：稀土皂化萃取P507有机相后萃余水相的pH值随稀土浓度、相比和级数的变化图(萃余液pH值随料液稀土浓度和相比的变化关系)。
图15：料液稀土离子浓度与分配比的关系。
图16：连续萃取各指标的稳定性(a：O/A＝1:1单级萃取，萃余液溶解氢氧化稀土后稀土溶液的浓度；b：O/A＝1:1单级萃取有机相稀土浓度；c：O/A＝2:1两级萃取，萃余液溶解氢氧化稀土后稀土溶液的浓度；d：O/A＝2:1两级萃取有机相稀土浓度)。
图17：氯化铵反萃稀土负载N1923有机相(相比及氯化铵浓度对反萃率的影响)。
图18：不同相比、浓度氯化铵反萃稀土负载N1923有机相后溶液稀土离子浓度。
图19：离子型稀土浸取、沉淀、萃取分离、循环利用流程图。
具体实施方式
本发明将通过以下实施例作进一步说明。
实施例1
取离子吸附型稀土原矿，原矿使用20筛网筛分，取20目下细粒部分，均 匀化后使用，称取100g离子吸附型稀土原矿，装在砂芯玻璃柱中，上下方垫有滤纸使渗流更加均匀；采用浓度为0.12mol/L，pH值为4的氯化钙溶液作为浸取剂，氯化钙溶液与离子吸附型稀土原矿按液固比0.6:1对离子吸附型稀土原矿进行浸出，其中，最初不含稀土的浸出液收集于第一容器，后续含稀土的浸出液收集于第二容器。然后，采用浓度为0.02mol/L，pH值为2的硫酸铝溶液作为浸取剂，硫酸铝溶液与离子吸附型稀土原矿按液固比0.4:1对离子吸附型稀土原矿进行二次浸出，其中，pH值大于4的浸出液收集于第二容器，pH值小于4的浸出液收集于第三容器。采用石灰水浸淋离子吸附型稀土尾矿，石灰水与离子吸附型稀土尾矿的液固比为0.2:1，使离子吸附型稀土尾矿中的稀土离子和滞留在稀土颗粒间的硫酸铝流出。石灰水浸淋离子吸附型稀土尾矿直至浸出液的pH值大于5为止。继续用水浸淋，直至pH值大于6。其中，pH值大于4的浸出液收集于第一容器，pH值小于4的浸出液收集于第三容器。优选地，使用氯化钙浸取后，且在使用硫酸铝溶液浸取前，对离子吸附型稀土原矿进行水洗，以去除矿层空隙中的大部分钙和稀土离子以避免硫酸铝进入矿层形成硫酸钙沉淀而阻碍后续浸取过程。在石灰水浸淋和硫酸铝浸取之间，也需要用水洗去矿层空隙中的大部分硫酸铝，使其被粘土矿物充分吸附，避免石灰水进入矿层后形成硫酸钙和氢氧化铝而阻碍后续浸取过程。所述的容器是指矿山的储液池、车间的储液罐和实验室的烧杯等容器。第一容器的溶液稀土含量很低，可直接用于配制氯化钙浸矿剂而循环用于第一阶段的浸矿，也可用于尾矿生态修复用的浇灌用水。
对氯化钙作为浸取剂的浸出液，加入氧化钙来沉淀稀土和铝、过滤，得到氢氧化稀土；随后经过盐酸优先溶解稀土，控制溶液pH值为4.5，过滤后得到低铝低硫酸根的高浓度氯化稀土溶液；优溶渣再经硫酸和硫酸铝浸出的酸性溶液溶解、用N1923-煤油-异辛醇混合有机相萃取稀土、硫酸铝留在水相，循环用于第二阶段的浸取；有机相中的稀土用盐酸和NH 4Cl反萃，获取氯化稀土溶液，测定稀土离子总量。对硫酸铝作为浸取剂的浸出液，直接使用N1923-煤油-异 辛醇有机相萃取稀土，硫酸铝留在水相，循环用于第二阶段的浸取；有机相中的稀土用盐酸和NH 4Cl反萃完全，获取稀土离子，测定稀土离子总量。
实施例2
取离子吸附型稀土原矿，原矿使用20筛网筛分，取20目下细粒部分，均匀化后使用，称取100g离子吸附型稀土原矿，装在砂芯玻璃柱中，上下方垫有滤纸使渗流更加均匀；采用浓度为0.25mol/L，pH值为8的氯化钙溶液作为浸取剂，氯化钙溶液与离子吸附型稀土原矿按液固比1:1对离子吸附型稀土原矿进行浸出，其中，初始不含稀土的浸出液收集于第一容器，后续含稀土的浸出液收集于第二容器。然后，采用浓度为0.20mol/L，pH值为4的硫酸铝溶液作为浸取剂，硫酸铝溶液与离子吸附型稀土原矿按液固比0.5:1对离子吸附型稀土进行二次浸出，其中，pH值大于4的浸出液收集于第二容器，pH值小于4的浸出液收集于第三容器。采用石灰水浸淋离子吸附型稀土尾矿，石灰水与离子吸附型稀土尾矿的液固比为0.5:1，使离子吸附型稀土尾矿中的稀土离子和滞留在稀土颗粒间的硫酸铝流出。石灰水浸淋离子吸附型稀土尾矿直至浸出液的pH值大于5为止。继续用水浸淋，直至pH值大于6。其中，pH值大于4的浸出液收集于第二容器，pH值小于4的浸出液收集于第三容器。优选地，使用氯化钙浸取后，且在使用硫酸铝溶液浸取前，对离子吸附型稀土原矿进行水洗，以去除矿层空隙中的大部分钙和稀土离子以避免硫酸铝进入矿层形成硫酸钙沉淀而阻碍后续浸取过程。在石灰水浸淋和硫酸铝浸取之间，也需要用水洗去矿层空隙中的大部分硫酸铝，使其被粘土矿物充分吸附，避免石灰水进入矿层后形成硫酸钙和氢氧化铝而阻碍后续浸取过程。
对氯化钙作为浸取剂的浸出液，加入氧化钙来沉淀稀土和铝、过滤，得到氢氧化稀土；随后经过盐酸优先溶解稀土，控制溶液pH值为5，过滤后得到低铝低硫酸根的高浓度氯化稀土溶液；优溶渣再经硫酸和硫酸铝浸出的酸性溶液溶解、用N1923-煤油-异辛醇混合有机相萃取稀土、硫酸铝留在水相，循环用于第二阶段的浸取；有机相中的稀土用盐酸和NH 4Cl反萃，获取氯化稀土溶液， 测定稀土离子总量。对硫酸铝作为浸取剂的浸出液，直接使用N1923-煤油-异辛醇有机相萃取稀土，硫酸铝留在水相，循环用于第二阶段的浸取；有机相中的稀土用盐酸和NH 4Cl反萃完全，获取稀土离子，测定稀土离子总量。
实施例3
取离子吸附型稀土原矿，原矿使用20筛网筛分，取20目下细粒部分，均匀化后使用，称取100g离子吸附型稀土原矿，装在砂芯玻璃柱中，上下方垫有滤纸使渗流更加均匀；采用浓度为0.20mol/L，pH值为6的从含钙废物中回收的以氯化钙为主要成分的电解质溶液作为浸取剂，电解质溶液与离子吸附型稀土原矿按液固比0.7:1对离子吸附型稀土进行浸出，其中，开始不含稀土的浸出液收集于第一容器，而后含稀土的浸出液收集于第二容器。然后，采用浓度为0.10mol/L，pH值为3的硫酸铝溶液作为浸取剂，硫酸铝溶液与离子吸附型稀土原矿按液固比0.3:1对离子吸附型稀土原矿进行二次浸出，其中，pH值大于4的浸出液收集于第二容器，pH值小于4的浸出液收集于第三容器。采用石灰水浸淋离子吸附型稀土尾矿，石灰水与离子吸附型稀土尾矿的液固比为0.3:1，使离子吸附型稀土尾矿中的稀土离子和滞留在稀土颗粒间的硫酸铝流出。石灰水浸淋离子吸附型稀土尾矿直至浸出液的pH值大于5为止。继续用水浸淋，直至pH值大于6。其中，pH值大于4的浸出液收集于第二容器，pH值小于4的浸出液收集于第三容器。优选地，使用氯化钙浸取后，且在使用硫酸铝溶液浸取前，对离子吸附型稀土原矿进行水洗，以去除矿层空隙中的大部分钙和稀土离子以避免硫酸铝进入矿层形成硫酸钙沉淀而阻碍后续浸取过程。在石灰水浸淋和硫酸铝浸取之间，也需要用水洗去矿层空隙中的大部分硫酸铝，使其被粘土矿物充分吸附，避免石灰水进入矿层后形成硫酸钙和氢氧化铝而阻碍后续浸取过程。
对氯化钙作为浸取剂的浸出液，加入氧化钙来沉淀稀土和铝、过滤，得到氢氧化稀土；随后经过盐酸优先溶解稀土，控制溶液pH值为4.7，过滤后得到低铝低硫酸根的高浓度氯化稀土溶液；优溶渣再经硫酸和硫酸铝浸出的酸性溶 液溶解、用N1923-煤油-异辛醇混合有机相萃取稀土、硫酸铝留在水相，循环用于第二阶段的浸取；有机相中的稀土用盐酸和NH 4Cl反萃，获取氯化稀土溶液，测定稀土离子总量。对硫酸铝作为浸取剂的浸出液，直接使用N1923-煤油-异辛醇有机相萃取稀土，硫酸铝留在水相，循环用于第二阶段的浸取；有机相中的稀土用盐酸和NH 4Cl反萃完全，获取稀土离子，测定稀土离子总量。
实施例4
取离子吸附型稀土原矿，原矿使用20筛网筛分，取20目下细粒部分，均匀化后使用，称取100g离子吸附型稀土原矿，装在砂芯玻璃柱中，上下方垫有滤纸使渗流更加均匀；采用浓度为0.02mol/L，pH值为5的从含钙废物中回收的以氯化钙为主要成分的电解质溶液作为浸取剂，电解质溶液与离子吸附型稀土原矿按液固比0.9:1对离子吸附型稀土进行浸出，其中，起初不含稀土的浸出液收集于第一容器，而后含稀土的浸出液收集于第二容器。然后，采用浓度为0.15mol/L，pH值为3的含铝回收废物中以硫酸铝为主要成分的电解质溶液作为浸取剂，电解质溶液与离子吸附型稀土原矿按液固比0.1:1对离子吸附型稀土原矿进行二次浸出，其中，pH值大于4的浸出液收集于第二容器，pH值小于4的浸出液收集于第三容器。采用石灰水浸淋离子吸附型稀土尾矿，石灰水与离子吸附型稀土尾矿的液固比为0.4:1，使离子吸附型稀土中的稀土离子和滞留在稀土颗粒间的硫酸铝流出。石灰水浸淋离子吸附型稀土尾矿直至浸出液的pH值大于5为止。继续用水浸淋，直至pH值大于6。其中，pH值大于4的浸出液收集于第二容器，pH值小于4的浸出液收集于第三容器。优选地，使用氯化钙浸取后，且在使用硫酸铝溶液浸取前，对离子吸附型稀土进行水洗，以去除矿层空隙中的大部分钙和稀土离子以避免硫酸铝进入矿层形成硫酸钙沉淀而阻碍后续浸取过程。在石灰水浸淋和硫酸铝浸取之间，也需要用水洗去矿层空隙中的大部分硫酸铝，使其被粘土矿物充分吸附，避免石灰水进入矿层后形成硫酸钙和氢氧化铝而阻碍后续浸取过程。
对氯化钙作为浸取剂的浸出液，加入氧化钙来沉淀稀土和铝、过滤，得到 氢氧化稀土；随后经过盐酸优先溶解稀土，控制溶液pH值为4.8，过滤后得到低铝低硫酸根的高浓度氯化稀土溶液；优溶渣再经硫酸和硫酸铝浸出的酸性溶液溶解、用N1923-煤油-异辛醇混合有机相萃取稀土、硫酸铝留在水相，循环用于第二阶段的浸取，即硫酸铝作为浸取剂的浸取；有机相中的稀土用盐酸和NH 4Cl反萃，获取氯化稀土溶液，测定稀土离子总量。对硫酸铝作为浸取剂的浸出液，直接使用N1923-煤油-异辛醇有机相萃取稀土，硫酸铝留在水相，循环用于第二阶段的浸取；有机相中的稀土用盐酸和NH 4Cl反萃完全，获取稀土离子，测定稀土离子总量。对上述实施例中获得的稀土离子的浸出效率的计算采用对比的方法。
参见如下表一，采用0.064mol/L的氯化钙作为浸取剂按照液固比为1：1的比例对100g离子吸附型稀土原矿进行浸取，获取浸出液，并通过氧化钙沉淀、过滤，随后使用硫酸溶解、质子化N1923有机相振荡萃取15min、NH4Cl反萃振荡15min，最后测定稀土离子总量，作为对比例1；采用0.043mol/L硫酸铝作为浸取剂按液固比1：1对100g离子吸附型稀土原矿进行浸取，获取浸出液，并通过质子化N1923有机相萃取分离，NH 4Cl反萃测定稀土离子总量，作为对比例2。
表一
Figure PCTCN2021083462-appb-000001
如表一所示，采用本发明的两种浸取剂先后对离子吸附型稀土原矿进行浸取，相比单独氯化钙浸取的浸出效率更高，且使用氯化钙的用量更低；与单独使用硫酸铝浸取的浸出效率相同，但硫酸铝的用量远少于单独使用时的用量，从而在保证高浸出效率的同时，还能减少硫酸铝的使用量，从而降低后续萃取工作量。
分别使用0.128N的Al 2(SO 4) 3、MgSO 4、CaCl 2、(NH4) 2SO 4浸取、淋洗、萃取、反萃，测定的各浸取剂在0.128N时的浸矿效率如图1，结果证明：硫酸铝的浸取率最高，接近98％，硫酸铵次之，约95％；硫酸镁的浸取率不足80％，而氯化钙只有70％左右。
如图2所示，浸出效率随着氯化钙阳离子浓度的提升，浸出效率是可以得到逐步提升的，从70％上升到88％；在浓度高于0.25mol/L后，浸出效率增加缓慢，且一味增加阳离子浓度会引起过量的浸取剂污染环境、增加成本，综合考虑，最佳浸取剂浓度为0.178mol/L。浸取效率不足90％，证明单独依靠氯化钙来浸取稀土时，很难达到高的浸取率。这也是以往不用氯化钙浸取的主要原因。
实施例5
尽管单独用氯化钙不能达到很高的浸取率，但还是能把大部分稀土浸取出来。其优点是：氯化钙来源广泛，价格低，而且在稀土分离企业用石灰中和蒸氨法处理氯化铵废水时会产生大量的氯化钙溶液。如果用氯化钙来浸取稀土，可以消除用硫酸铵浸取稀土时带来的氨氮污染。为了提高浸取效率，我们需要联合其他方法来完成。图1的结果证明硫酸铝具有最好的浸取效率，因此，用少量硫酸铝溶液进行第二阶段的浸取，研究了硫酸铝加入量对浸出效率的影响。实验方法是第一阶段以液固比为1～0.5:1加入0.256N的氯化钙浸出液，随后以液固比0.1:1加入少量顶补水，然后第二阶段以液固比为0～0.5:1加入0.128N的硫酸铝，保持每次浸取的液固比为1，随后以液固比0.4:1加入顶补水收集浸出液，测定稀土含量，计算浸出效率。结果如图3。结果证明：随着硫酸铝加入量的增加浸出效率不断增加；浸取效率从85％左右提高到接近100％。只需取代10％的氯化钙溶液就可以使浸取效率提高到94％左右，取代20％时可以使浸取效率提高到96％以上。
如图4所示，随着氯化钙浸取剂用量的减少，第一阶段的稀土浸出效率降低幅度，在开始时并不大，随后就明显降低。而在两种浸矿剂比例相等的情况 下，硫酸铝浸取段浸出的稀土离子总量大于氯化钙浸取段浸出的稀土离子，两阶段的总浸出效率明显提升。
实施例6
使用与实施例5中类似的方法，只是去除两浸取阶段之间加入的顶补水，将其补充在硫酸铝浸取阶段之后，结果如图6。发现在保持加入顶补水量相同的情况下整体浸出效率与实施例5中对应条件下的实验数据非常接近，证明两浸取段之间的顶补水可以不加。
实施例7
按照实施例5的方法在第二阶段浸取的过程中分别以液固比加入0.1-0.3:1加入硫酸铝溶液，然后以液固比0.4:1加入顶补水，最后使用饱和澄清石灰水淋洗尾矿改善尾矿pH，收集尾矿流出液并测定其pH。如图7所示，可以看到随着硫酸铝加入量的增多，土壤酸度加大，消耗改善土壤pH的石灰水量也会增多，而且考虑到硫酸铝加入量与后续萃取的工作量成正比，减少硫酸铝的加入量可以减少后续尾矿处理所需的石灰量。
实施例8
第二阶段硫酸铝浸取后顶补水加入量对浸出效率的影响。按照上述结果，在第一阶段加入0.256N的氯化钙(0.8:1)收取浸出液后以液固比0.1加入顶补水，继续接收顶补液，第二阶段加入0.128N的硫酸铝(0.2:1)收取浸出液后再以液固比0.1-0.4顶补水并接收，测定顶补水量对浸出效率的影响。结果如图8，证明顶补水的增加会带出更多的稀土离子，加入0.4的顶补水后稀土残留低于0.01％，然而在实际操作中后续护尾的石灰水也可以当作顶补水淋洗残留的稀土离子，即浸取后前0.4淋洗液都会含有稀土离子，具有回收价值。
实施例9
按照上述结果各浸矿剂浓度均取0.128N，两种浸矿剂体积比为0.8:0.2分别组合氯化钙+硫酸铝、硫酸镁+硫酸铝、氯化钙+氯化钙和氯化铵+硫酸铝测定浸出效率，如图9所示，证明第二阶段加入硫酸铝会使浸出效率在原有基础上 有不同程度的提升，即分段浸取不会影响各自浸矿剂的浸出效率。
目前矿山广泛采用的硫酸铵、氯化钠等一价阳离子浸取试剂导致了很多地方的滑坡塌方和水土流失，导致水土流失和滑坡塌方的原因是这些一价阳离子与黏土矿物的吸附能力不足，导致黏土矿物粒子的zeta电位负值升高，一般在-20mv以上。这种带电粒子之间相互排斥力强，在水流冲刷下，很容易随水流失，使矿层的粘附力降低，导致滑坡塌方。采用硫酸铝来浸取稀土不仅可以提高浸取率，而且由于铝离子的价态高，对矿中黏土矿物的吸附能力强，会使黏土矿物的zeta电位绝对值减小，趋于0mv，这对于减少水土流失，降低尾矿滑坡塌方风险是非常有利的。如图10为采用不同浸矿剂的两阶段浸取时的zeta电位，可以看到第二阶段zeta电位绝对值会在加入硫酸铝后明显的降低，矿物的zeta电位绝对值明显减小。氯化钙浸取时，尾矿黏土矿物的zeta电位在-14mv左右，增加了硫酸铝浸取后，下降到-2mv左右(如图5所示)。这证明用硫酸铝浸取后的尾矿比未加入硫酸铝的相比会更为稳定。证明硫酸铝对于提升浸出效率、保护尾矿稳定性有显著的作用。
氯化钙浸出的稀土是用氢氧化钙来沉淀回收的，所述氢氧化钙(石灰乳)沉淀稀土的方法是在搅拌作用下，将澄清的氯化钙浸出液与氢氧化钙(石灰乳)溶液按一定的比例同步加入到含有氢氧化稀土结晶沉淀的悬浮液中，使形成的沉淀能快速形成氢氧化稀土结晶，沉淀完成后经进一步的陈化和澄清，上清液转入石灰水配置池和氯化钙浸取试剂配置池，实现循环使用。一部分氢氧化稀土留在沉淀池中作为后续沉淀的晶种继续使用。大部分的结晶氢氧化稀土用于过滤洗涤，产出氢氧化稀土中间产品，产品的X-射线衍射图和电镜照片如图11和12；证明产品是以结晶氢氧化稀土为主，颗粒大小在4-8微米之间。取5g洁净的氢氧化稀土使用盐酸全溶后进行化学分析结果证明其稀土氧化物含量占89.31wt％，铝含量为5.81wt％，另外不溶渣占据4.88wt％。证明铝和稀土一起沉淀下来，产品铝含量高不能直接用于萃取分离。用氢氧化钠浸洗后，铝含量可降低至2.26wt％，但成本增加。
由于氢氧化稀土沉淀中含有较高的铝，而用氢氧化钠又不能很好地把铝去除，且花费大。为此，我们将盐酸优溶与酸性萃取剂有机相的皂化相结合。设计了一套新的能够将有机相皂化与料液制备结合起来的新方法。采用串级逆流萃取的方式，用盐酸优溶从氯化钙浸出液中用氢氧化钙沉淀稀土所得的氢氧化稀土富集物，控制溶解过程中的pH＞4，制备出铝含量低的高浓度氯化稀土料液。用酸性萃取有机相(P507-煤油有机相)直接萃取稀土，部分稀土进入有机相，达到稀土萃取分离所需的负载稀土浓度，交换出来的氢离子进入水相，循环用于溶解氢氧化稀土，使水相在体系中循环，不用排放皂化废水。如图13，采用不同浓度的稀土料液和串级萃取级数、相比分别进行皂化实验。所得水相和有机相的稀土离子浓度和铝、钍等典型杂质元素的分析结果见表2-4，证明当稀土料液浓度在0.4-1.2之间时，均能使有机相达到后续稀土萃取分离所需的负载量要求，0.18-0.26mol/L之间。图14为对应条件萃余水相的pH值，这种低酸度的萃余水相可以用来溶解氢氧化稀土。
连续多轮逆流萃取有机相稀土离子负载浓度和出口水相酸度的稳定性。起始氯化稀土料液浓度C＝1.021372mol/L、pH＝4.6，每轮以皂化萃取过后的萃余水相加入氢氧化稀土经过搅拌溶解、澄清、过滤后得到的溶液再次接触有机相作为起始，依次进行第二轮、第三轮逆流皂化萃取操作如图16，萃余液溶解氢氧化稀土后稀土溶液的浓度与有机相稀土浓度均有所提升，证明该工艺可以连续进行萃取。在较高的水相稀土离子浓度条件下，有机相稀土离子负载量能很容易达到实际生产要求。
如表二表三所示为单级稀土皂化过程中各相各元素浓度、配分，表四为各元素的分配比D以及及相邻元素分离系数β，可以看出：分离因数有益于轻重稀土元素的分组。
表二：不同料液浓度下P507萃取平衡时各元素浓度及配分。
Figure PCTCN2021083462-appb-000002
表三：不同料液浓度下P507萃取平衡时各元素浓度(mol/L)及配分(续)。
Figure PCTCN2021083462-appb-000003
表四：萃取平衡时各元素分配比D及相邻元素分离系数β。
Figure PCTCN2021083462-appb-000004
第二阶段硫酸铝浸出的稀土离子使用N1923有机相以合适的相比进行萃取。使用100g的20目稀土原矿以液固比1:1，使用阳离子浓度为0.128N的Al 2(SO 4) 3进行浸出，得到的浸出液以相比O/A＝1:10进行萃取，所得萃取有机相 随后使用NH 4Cl进行反萃。结果发现：采用合适的相比，浸出液溶液中稀土离子的萃余率可以很容易低于0.1％，而稀土负载的N1923有机相需要使用1～5mol/L的氯化铵以O/A＝1～3:1进行反萃，结果见图17-18。稀土反萃率随氯化铵浓度的增大而增大，随着氯化铵体积的增大而增多，这是由于越来越多的氯离子将被萃取的稀土离子缔合物洗脱下来。氯化铵为4mol/L和相比O/A＝3:1时，氯离子能以最优效率反萃稀土离子；在O/A＝1:1，氯化铵浓度为5mol/L时，反萃率有96.79％。然而在增大相比的过程中，由于已经反萃下来的稀土离子浓度高，还有大量的硫酸根离子，会形成硫酸稀土沉淀而导致溶液中稀土离子浓度急速下降。
难溶部分固体，使用硫酸溶解得到含稀土的硫酸铝溶液，再使用N1923萃取分离稀土与Al 3+。另外一种方法是将这些固体与硫酸铝浸出液混合，控制溶液pH在3左右，使氢氧化稀土和铝都能溶解。上清液中主要是硫酸铝和少量稀土，不溶物是黏土矿物和少量硫酸钙。用N1923从上清液中萃取稀土，萃余水相利用气浮法除油后，调整硫酸铝浓度和pH值后，继续用于离子吸附型稀土的浸取。有机相中的稀土使用NH 4Cl溶液反萃，反萃稀土后的N1923有机相继续循环利用。反萃下来的稀土离子用石灰或者碳铵沉淀，得到相应的稀土氢氧化物或碳酸盐，煅烧后得到稀土氧化物。最后按实施例4的方法用于制备高浓度稀土料液和皂化有机相，用于后续萃取分离，整个过程如图19。
本领域的普通技术人员可以理解，上述各实施方式是实现本发明的具体实施方式，而在实际应用中，可以在形式上和细节上对其作各种改变，而不偏离本发明的精神和范围。

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  1. 一种离子吸附型稀土的提取方法，其特征在于，所述提取方法包括以下步骤：
    S1：以氯化钙溶液为浸取剂对离子吸附型稀土原矿进行浸出，稀土以离子形式进入浸出液；其中，最初不含稀土的浸出液收集于第一容器，后续含稀土的浸出液收集于第二容器；
    S2：以硫酸铝溶液为浸取剂对所述离子吸附型稀土原矿进行第二次浸出，剩余稀土以离子形式进入浸出液；其中，pH值大于4的浸出液仍收集于所述第二容器；pH值小于4的浸出液，收集于第三容器；
    S3：用石灰水对所述浸出后的离子吸附型稀土尾矿进行浸淋；然后，用水浸淋，直至浸出液的pH值大于6；其中，pH值大于4的浸出液仍收集于所述第一容器；pH值小于4的浸出液，收集于所述第三容器。
  2. 根据权利要求1所述的一种离子吸附型稀土的提取方法，其特征在于，在步骤S1中，所述氯化钙溶液的浓度在0.02～0.25mol/L之间，pH值在4～8之间。
  3. 根据权利要求1或2所述的一种离子吸附型稀土的提取方法，其特征在于，在步骤S1中，所述氯化钙溶液与所述离子吸附型稀土原矿的液固比为0.6～1.0。
  4. 根据权利要求1或2所述的一种离子吸附型稀土的提取方法，其特征在于，所述氯化钙溶液中允许含有一定量的其他一价和两价非铵阳离子电解质和添加剂。
  5. 根据权利要求1所述的一种离子吸附型稀土的提取方法，其特征在于，在步骤S2中，所述硫酸铝溶液的浓度在0.02～0.20mol/L之间，pH值在2-4之间。
  6. 根据权利要求1或5所述的一种离子吸附型稀土的提取方法，其特征在于，在步骤S2中，所述硫酸铝溶液与所述离子吸附型稀土原矿的液固比为0.1～0.5。
  7. 根据权利要求1或5所述的一种离子吸附型稀土的提取方法，其特征是所述硫酸铝溶液中允许含有一定量的其他非铵阳离子电解质和添加剂。
  8. 根据权利要求1所述的一种离子吸附型稀土的提取方法，其特征在于，步骤S3 中，所述石灰水与所述离子吸附型稀土尾矿的液固比为0.2～0.5。
  9. 根据权利要求1所述的一种离子吸附型稀土的提取方法，其特征在于，步骤S3中，所述的石灰水浸淋中的石灰水溶液用量是以流出液的pH值大于5为依据来判断，pH值达到5以后，用水顶浸，使pH值逐渐升高到6以后停止注水。
  10. 根据权利要求1所述的一种离子吸附型稀土的提取方法，其特征在于，所述氯化钙溶液包括氯化钙以及从含钙废物中回收的以氯化钙为主要成分的电解质溶液；所述硫酸铝溶液包括硫酸铝以及从含铝废物中回收的以硫酸铝为主要成分的电解质溶液。
  11. 根据权利要求1所述的一种离子吸附型稀土的提取方法，其特征在于，还包括以下步骤：第二容器中的稀土采用石灰沉淀得到氢氧化稀土，用盐酸优溶得高浓度低铝低硫酸根的氯化稀土溶液，不溶渣加入第三容器，使其中的稀土和铝的氢氧化物溶解，再用N1923-煤油-异辛醇混合有机相萃取分离稀土与铝，萃余液中含硫酸铝，循环用于硫酸铝的浸取；有机相中的稀土用氯化物反萃，反萃液并入第二容器中，用石灰沉淀法制取氢氧化稀土。
  12. 根据权利要求11所述的一种离子吸附型稀土的提取方法，其特征在于，所获得的氢氧化稀土可以用于皂化p507有机相，皂化产生的酸循环用于溶解氢氧化稀土，降低优溶氢氧化稀土所需的酸量，不排放皂化废水。
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CN117535535A (zh) * 2024-01-10 2024-02-09 矿冶科技集团有限公司 一种离子型稀土矿山原地复盐浸矿淋洗的无废开采方法
CN117535535B (zh) * 2024-01-10 2024-04-30 矿冶科技集团有限公司 一种离子型稀土矿山原地复盐浸矿淋洗的无废开采方法

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