WO2020030084A1 - 一种联合法处理稀土精矿的冶炼分离工艺 - Google Patents

一种联合法处理稀土精矿的冶炼分离工艺 Download PDF

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WO2020030084A1
WO2020030084A1 PCT/CN2019/099932 CN2019099932W WO2020030084A1 WO 2020030084 A1 WO2020030084 A1 WO 2020030084A1 CN 2019099932 W CN2019099932 W CN 2019099932W WO 2020030084 A1 WO2020030084 A1 WO 2020030084A1
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rare earth
leaching
hydrochloric acid
roasting
smelting
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PCT/CN2019/099932
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English (en)
French (fr)
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黄小卫
冯宗玉
孙旭
徐旸
王猛
王良士
夏超
刘向生
赵龙胜
张永奇
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有研稀土新材料股份有限公司
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Priority to US17/267,481 priority Critical patent/US20210310100A1/en
Priority to CN201980052517.6A priority patent/CN112534072A/zh
Priority to AU2019316882A priority patent/AU2019316882B2/en
Publication of WO2020030084A1 publication Critical patent/WO2020030084A1/zh
Priority to ZA2021/01505A priority patent/ZA202101505B/en

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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
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • 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
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C22B1/06—Sulfating roasting
    • 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
    • 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 smelting and separation, and particularly relates to a smelting and separating process for treating rare earth concentrates by a combined method.
  • One or more mixed rare earth minerals are mixed rare earth minerals.
  • China's rare earth resources are mainly based on mineral light rare earth mineral resources, accounting for more than 90% of the total reserves.
  • Industrial rare earth minerals are mainly fluorocarbon cerium and monazite, whose light rare earth content is as high as 96% -98%. According to statistics, the light rare earth deposits with industrial application value are mainly Baotou Baiyun Ebo rare earth mine, Sichuan Panxi Mianning rare earth mine, and Shandong Weishan rare earth mine.
  • the fluorocarbon cerium ore generally adopts the oxidation roasting-hydrochloric acid leaching chemical treatment process.
  • the concentrate is decomposed by oxidation roasting to generate hydrochloric acid-soluble rare earth oxide, rare earth fluoride or rare earth oxyfluoride, and cerium is oxidized to tetravalent Ions, and during the leaching process of hydrochloric acid, trivalent rare earths are leached to obtain cerium-less rare earth chloride, cerium and some trivalent rare earths, fluorine, and thorium are left in the soluble slag, and the alkali is used to remove the fluorine to obtain the cerium-rich
  • the slag can be used to produce ferrosilicon alloys, or to produce cerium oxide with a purity of about 98% after reduction and leaching.
  • the rare earth cerium chloride is separated into a single rare earth by the P507 extractant.
  • the advantage of this process is that the investment is small and the production cost is low, but the disadvantage is that the process is discontinuous, and the cerium, praseodymium and fluorine do not dissolve in the slag during the hydrochloric acid leaching process. After the slag undergoes alkali conversion, the fluorine will The form of sodium fluoride enters the waste water, and thorium and fluorine are dispersed in the slag and the waste water, which is difficult to be recycled. As a result, the entire process not only pollutes the environment, but the recycling purity of the cerium product is only about 98%, which has a low use value.
  • the cerium oxide product recovered from the cerium-rich slag has low purity
  • the mixed concentrate processing process has a high concentration of fluorine and sulfur in flue gas emissions during sulfuric acid roasting, and a high treatment and recovery cost , Low concentration of rare earth leachate.
  • roasting and decomposing the rare earth concentrate in a certain roasting atmosphere to obtain a roasted ore
  • the rare earth concentrate according to the present invention includes, but is not limited to, fluorocarbon cerite or a mixed type rare earth ore of monazite or xenotime or one type of monazite or xenotime.
  • the roasting atmosphere in the roasting step includes water vapor or a weakly oxidizing atmosphere;
  • the weakly oxidizing atmosphere includes, but is not limited to, one or more of N 2 , CO, CO 2 , air, and an inert gas.
  • This kind of atmosphere can reduce the oxygen content by controlling the gas flow rate;
  • the purpose of the water vapor atmosphere is to defluorinate and obtain pure HF as the recovery product;
  • the purpose of the weakly oxidizing atmosphere is to reduce the oxidation rate of cerium and improve the rare earth Leaching yield.
  • the HF gas obtained after defluorination is adsorbed and recovered by using a rare earth oxide or a rare earth hydrated oxide to obtain a fluorinated rare earth product.
  • the principle is that the rare-earth oxide forms a polynuclear hydroxyl compound in water, causing ion exchange between OH - and fluoride ions on the rare earth oxide to achieve a double fluoride removal effect.
  • the HF produced in the defluorination roasting and decomposition process of the rare earth concentrate can be effectively recovered to obtain a fluorinated rare earth product, and the tail gas can be discharged in accordance with standards, which has significant environmental protection benefits.
  • the roasting temperature of the roasting step is 350-650 ° C, preferably 400-600 ° C, where the roasting temperature is increased within a certain range, which can increase the rare earth leaching rate; the time of the roasting step is 0.5-6h, which can extend the roasting time within a certain range, which can increase the rare earth leaching rate.
  • REFCO 3 REOF (CeOF) + CO 2 ⁇ .
  • REOF + H 2 O RE 2 O 3 + HF ⁇ .
  • the removed HF gas is recovered by an adsorbent such as a rare earth oxide.
  • the concentration of the hydrochloric acid is 3-10 mol / L, preferably 4-7 mol / L, and the ratio of the amount of hydrochloric acid to the roasted concentrate is 0.4-2.0 mol hydrochloric acid / 100 g of rare earth concentrate, preferably It is 0.7-1.5mol hydrochloric acid / 100g rare earth concentrate.
  • the hydrochloric acid leaching step is preferably two or more steps of hydrochloric acid counter-current leaching.
  • solid-liquid separation is used to obtain one rare earth leaching solution and one leaching slag.
  • Hydrochloric acid leaching is obtained through solid-liquid separation.
  • the rare earth leaching solution in this step is returned to be used as the bottom water for the hydrochloric acid leaching in the previous step.
  • the leaching slag in this step can be subjected to the next step of hydrochloric acid leaching.
  • the method for adding hydrochloric acid is to carry out continuous co-current leaching of 2-5 stages in the leaching process, and control the hydrochloric acid to be added in a concentration gradient during each stage of leaching, and the first stage is diluted.
  • Concentration of hydrochloric acid, higher levels of hydrochloric acid are added in the last few stages to maintain the acidity of the mixed solution at 0.01-0.6 mol / L, and preferably 0.05-0.3 mol / L. The lower the acidity, the more favorable the rare earth leaching.
  • the purpose is to ensure that the tetravalent Ce is not reduced after entering the solution, and increase the leaching rate of rare earth and fluorine; through stepwise leaching, a higher rare earth concentration can be obtained, and the rare earth concentration in the leaching solution reaches 150-250 g / L. At the same time, because the residual acid content of the leachate is effectively reduced, the neutralizer consumption in subsequent processes is also reduced.
  • hydrochloric acid leaching is performed at a lower temperature because F is mainly present in the solution as a [CeF x ] 4-x complex, and low temperature conditions are favorable for [CeF x ] 4-x
  • the steady state of the coordination compound can make more dissolution of rare earth and F, and the leaching rate of rare earth reaches 70% -95%.
  • the leaching temperature of the hydrochloric acid leaching step is controlled to be 10-75 ° C, preferably 20-65 ° C, and the total reaction time is controlled to be 0.5-10h, preferably 1-6h, mainly to improve the leaching rate of rare earth and F.
  • the dehydration step is a natural dehydration and / or drying method, and the moisture content of the dehydrated leaching slag after the treatment is preferably less than 10%, and the REO content of the dewatering leaching slag is 20%- 60%, mainly REPO 4 , can be mixed with other rare earth concentrates for sulfuric acid roasting.
  • the mass ratio (w / w) of the concentrated sulfuric acid to the leached slag after dehydration is 0.3-1.2: 1, and preferably 0.5: 1.
  • the scheme of the present invention is compared with the prior art. Technology, a large number of rare earths have been leached in the 1-2 step, and the sulfuric acid use amount is greatly reduced in the sulfuric acid roasting step.
  • the temperature of the sulfuric acid roasting step is 200-450 ° C, and preferably 200-220 ° C or 250-350 ° C, and the roasting time of the roasting step is 1-4h;
  • the temperature of the water leaching step is 20-50 ° C, and preferably 25-40 ° C, preferably the leaching time is 2-5h, and the leaching solution can be neutralized with a basic substance to a pH of 4-4.5, and the sulfuric acid obtained
  • the concentration of the rare earth solution is 25-45 g / L (REO).
  • the hydrochloric acid leaching slag is first washed with water, and the washing water: leaching slag ratio (w / w) is 0.5-10: 1, preferably 0.5-5: 1. 0-50%, preferably 0-30%.
  • the treated washing water contains a rare earth concentration of 5-50 g / L (REO) and a H + concentration of ⁇ 0.1 mol / L.
  • the purpose of the washing water is to wash the rare earth chloride entrained in the leaching slag into the solution, further increase the rare earth leaching rate, and remove the chloride ions that may cause corrosion to the subsequent sulfated roasting equipment.
  • a certain degree of dehydration treatment can reduce the
  • the sulfuric acid is used to strengthen the dilution of concentrated sulfuric acid in the roasting process, and the water washing liquid is used for the rare earth concentrate slurrying or the preparation of hydrochloric acid in step (2) to realize the closed-loop circulation of the water washing liquid.
  • the step (3) further includes a step of adding the obtained rare earth sulfate solution to iron powder for configuration, and the amount of the added iron powder is 2% -10% of the mass of the hydrochloric acid leaching slag.
  • the step (3) further includes a step of extracting and transforming the obtained rare earth sulfate solution to obtain a rare earth chloride solution, and extracting and separating to obtain a single rare earth compound.
  • the extraction transformation step is a transformation process by precipitation or extraction.
  • the step (2) further includes a step of subjecting the obtained rare earth leaching solution to aging treatment, solid-liquid separation to obtain a rare earth chloride solution and a fluorinated rare earth product; and combining the obtained rare earth chloride solution with the one described in step (3).
  • a rare earth chloride solution obtained by transformation of the rare earth sulfate solution is combined, and a single rare earth compound is obtained by extraction and separation.
  • the aging step is performed under standing or stirring conditions, and filtering is performed to obtain lanthanum cerium fluoride product; the temperature of the aging step is controlled to 60-90 ° C, further preferably 65-80 ° C, and the control is preferably performed.
  • the temperature of the aging step should be equal to or higher than the hydrochloric acid leaching temperature; the time of the aging step is 0.5-10h, preferably 1-4h.
  • the high-temperature aging step is used in the process of the present invention, which can effectively separate F from the leaching solution into the slag.
  • the F content in the leaching solution is ⁇ 8mg / L. Increasing the aging temperature and extending the aging time can further reduce the F content. To avoid the effect of F on subsequent extraction and separation.
  • the aging treatment can obtain a precipitate of rare earth fluoride, and it is preferable to control F ⁇ 8 mg / L in the leaching solution, and more preferably ⁇ 2 mg / L; the leaching residue in the solid-liquid separation obtained in step (2)
  • the ratio of the F residual amount to the F content in the rare earth concentrate is 5% or less, preferably 1% or less;
  • the rare earth concentration of the rare earth chloride solution obtained after the aging and filtering is 150-250 g / L (REO), and the rare earth is leached
  • the rate is 70% -95%, of which the leaching rate of Ce is 60% -95%.
  • Table 1 shows the comparison between this method and the traditional fluorocarbon cerite treatment method to obtain hydrochloric acid leaching solution.
  • the leaching solution concentration, total rare earth leaching rate, and Ce leaching rate obtained by this method are higher than those of the traditional fluorocarbon cerite method.
  • the H + concentration is relatively low, and the F in the leachate is basically free, which has obvious technical advantages.
  • the smelting separator for the combined method for processing rare earth concentrates further includes spraying the fluorine-containing tail gas generated in the step (1) with water or an alkaline liquid, or the rare earth oxides and rare earth hydrated oxides. Defluorination of one or two adsorbents to recover rare earth fluoride products; and / or, a step of desulfurizing and recovering the sulfur-containing tail gas generated in step (3) to obtain a sulfuric acid product;
  • the sulfur-containing tail gas generated in the sulfuric acid roasting process is subjected to desulfurization and recovery treatment, not only the tail gas emission reaches the standard, but also high-purity sulfuric acid products can be recovered and the concentration can reach more than 80%, which effectively solves the problem that the tail gas components in the traditional process contain F and equipment corrosion Severe attrition, difficult separation of F and S-containing substances, difficulty in meeting standards, and high operating costs.
  • the smelting separation process for treating rare earth concentrates by the combined method of the present invention adopts the method of atmosphere roasting-hydrochloric acid leaching-sulfuric acid roasting to treat rare earth concentrates containing fluorocarbon ceria, and controls the use of low-concentration hydrochloric acid in the hydrochloric acid leaching process.
  • a stepwise acid leaching method resulted in a higher rare earth chloride solution (150-250g / L REO).
  • the use of the [CeF x ] 4-x coordination compound to make more Ce enter the solution So that the leaching rate of Ce can reach 60% -95%, and the total rare-earth leaching rate can reach 70% -95%; further, the process of the present invention utilizes the relationship between the Cl - reducibility and the solubility product and temperature of fluorinated rare earths, and aging through high temperature. It further reduced the F - content of the leachate.
  • both the concentration of the leaching solution and the leaching rate of the rare earth are greatly improved, eliminating the step that the traditional process requires further evaporation and concentration to obtain a high concentration rare earth chloride solution, and the F in the solution
  • the content is very low, which prevents F from entering the extraction system to generate three phases, and can directly enter the P507-HCl system to separate and purify a single rare earth element.
  • the smelting and separation process for treating rare earth concentrates by the combined method of the present invention after the rare earth concentrates are roasted in atmosphere-hydrochloric acid leaching-aging at high temperature, only 5% -30% of the rare earths remain in the hydrochloric acid leaching slag, compared with
  • the traditional process for processing mixed rare-earth concentrates has greatly reduced the consumption of concentrated sulfuric acid, and also greatly reduced the consumption of water in the water leaching process. 70% -95% of the rare earths are directly sent to the chlorination system to extract and separate the rare earths. It also drastically reduces the acid-base consumption of sulfuric acid leaching solution extraction and transformation into a rare earth chloride solution.
  • the direction of fluorine is effectively controlled.
  • FIG. 1 is a flow chart of a smelting separation process for treating rare earth concentrates by a combined method according to the present invention.
  • the rare earth concentrate processed by the process described in this embodiment is a mixed rare earth ore of fluorocarbon cerium ore and monazite. According to the process flow chart shown in FIG. 1, the smelting separation of the rare earth concentrate is processed by the combined method described in this embodiment. The process includes the following steps:
  • the HF that escapes during the roasting process is treated with water spray.
  • the obtained roasted ore is added with hydrochloric acid at 25 ° C. for continuous continuous leaching in 4 stages.
  • the initial concentration of hydrochloric acid is 6 mol / L, and the ratio of the amount of hydrochloric acid to the roasted concentrate is 1.0 mol hydrochloric acid / 100 g of rare earth concentrate.
  • the rare-earth leaching solution and the leaching slag were collected, and the obtained rare-earth leaching solution had a rare-earth content of 238 g / L, a rare-earth leaching rate of 77%, and a Ce leaching rate of 70%;
  • the hydrochloric acid leaching step is preferably two or more steps of hydrochloric acid countercurrent leaching.
  • solid-liquid separation is used to obtain one step of rare earth leaching solution and one step of leaching slag.
  • the rare-earth leaching solution of this step and the leaching slag of this step are obtained by solid-liquid separation.
  • the rare-earth leaching solution of this step is returned to be used as the bottom water of the hydrochloric acid leaching in the previous step.
  • the method of adding hydrochloric acid for the leaching process is to perform continuous parallel leaching in 4 stages during the leaching process, and to control the concentration of hydrochloric acid, and add 1.5 mol / L, 2 mol / L, and 6 mol / L in the first to fourth stages, respectively.
  • L, 8mol / L hydrochloric acid, the acidity of the mixed solution decreases between 0.1-0.05mol / L gradient;
  • the rare earth leaching solution is aged at 65 ° C for 4 hours, and the solid-liquid separation obtains a rare earth chloride solution and a rare earth fluoride precipitate.
  • the F content in the rare earth chloride solution is 1.9 mg / L. Chemical rare earth products.
  • the roasted product was collected and added with water at 25 ° C. for 4 h, and after neutralization and impurity removal, a 32 g / L rare-earth sulfate solution was prepared; the total rare-earth yield was 97%.
  • the obtained rare-earth sulfate solution is subjected to extraction transformation to obtain a rare-earth chloride solution, which is combined with the rare-earth chloride solution in step (2) and subjected to extraction and separation to obtain a single rare-earth compound product.
  • the sulfur-containing waste gas generated during the sulfuric acid roasting process is recovered by spraying and absorbing the sulfuric acid product.
  • the rare earth concentrate processed by the process described in this embodiment is a mixed rare earth ore of fluorocarbon cerium ore and monazite.
  • the smelting and separation process of the rare earth concentrate processed by the combined method described in this embodiment includes the following steps:
  • the mixed rare earth ore of fluorocarbon cerium ore and monazite is roasted at 500 ° C for 4 hours under a weakly oxidizing atmosphere in the air (by adjusting the opening of the intake valve to control the oxygen content to 12%).
  • the opening degree of the valve is 50%, and the roasted ore is obtained;
  • the HF escaped during the roasting process is recovered by water spraying.
  • the obtained roasted ore is added with hydrochloric acid at 25 ° C. for continuous continuous leaching in 4 stages.
  • the initial concentration of hydrochloric acid is 6 mol / L, and the ratio of the amount of hydrochloric acid to the roasted concentrate is 1.0 mol hydrochloric acid / 100 g of rare earth concentrate.
  • the rare earth leaching solution and the leaching slag were collected, and the rare earth content of the obtained rare earth leaching solution was 250 g / L, the rare earth leaching rate was 80%, and the Ce leaching rate was 75%;
  • the hydrochloric acid leaching step is preferably two or more steps of hydrochloric acid countercurrent leaching.
  • solid-liquid separation is used to obtain one step of rare earth leaching solution and one step of leaching slag.
  • the rare-earth leaching solution of this step and the leaching slag of this step are obtained by solid-liquid separation.
  • the rare-earth leaching solution of this step is returned to be used as the bottom water of the hydrochloric acid leaching in the previous step.
  • the method of adding hydrochloric acid for the leaching process is to perform continuous parallel leaching in 4 stages during the leaching process, and to control the concentration of hydrochloric acid, and add 1.5 mol / L, 2 mol / L, and 6 mol / L in the first to fourth stages, respectively.
  • L, 8mol / L hydrochloric acid, the acidity of the mixed solution decreases between 0.1-0.05mol / L gradient;
  • the rare-earth leaching solution is aged at 80 ° C for 4 hours.
  • the solid-liquid separation results in a rare-earth chloride solution and a rare-earth fluoride precipitate.
  • the F content in the rare-earth chloride solution is 1.2 mg / L. Chemical rare earth products.
  • the roasted product was collected and added with water at 25 ° C. for 4 h, and after neutralization and impurity removal, a 32 g / L rare-earth sulfate solution was prepared; the total rare-earth yield was 97%.
  • the obtained rare-earth sulfate solution is subjected to extraction transformation to obtain a rare-earth chloride solution, which is combined with the rare-earth chloride solution in step (2) and subjected to extraction and separation to obtain a single rare-earth compound product.
  • the sulfur-containing waste gas generated during the sulfuric acid roasting process is recovered by spraying and absorbing the sulfuric acid product.
  • the rare earth concentrate processed by the process described in this embodiment is a mixed rare earth ore of fluorocarbon cerium ore and monazite.
  • the smelting and separation process of the rare earth concentrate processed by the combined method described in this embodiment includes the following steps:
  • the HF escaped during the roasting process is recovered by using a rare earth oxide adsorbent to obtain a fluorinated rare earth product.
  • the obtained roasted ore is added with hydrochloric acid at 25 ° C for continuous continuous leaching in 4 stages.
  • the initial concentration of hydrochloric acid is 6 mol / L, and the ratio of the amount of hydrochloric acid to the roasted concentrate is 1.0 mol hydrochloric acid / 100 g of rare earth concentrate.
  • the rare-earth leaching solution and the leaching slag were collected, and the obtained rare-earth leaching solution had a rare-earth content of 235 g / L, a rare-earth leaching rate of 75%, and a Ce leaching rate of 69%;
  • the hydrochloric acid leaching step is preferably two or more steps of hydrochloric acid countercurrent leaching.
  • solid-liquid separation is used to obtain one step of rare earth leaching solution and one step of leaching slag.
  • the rare-earth leaching solution of this step and the leaching slag of this step are obtained by solid-liquid separation, wherein the rare-earth leaching solution of this step is returned to be used as the bottom water of the previous step of hydrochloric acid leaching, and the leaching slag of this step can be subjected to the next step of hydrochloric acid leaching.
  • the method of adding hydrochloric acid for the leaching process is to perform continuous parallel leaching in 4 stages during the leaching process, and to control the concentration of hydrochloric acid, and add 1.5 mol / L, 2 mol / L, and 6 mol / L in the first to fourth stages, respectively.
  • L, 8mol / L hydrochloric acid, the acidity of the mixed solution decreases between 0.1-0.05mol / L gradient;
  • the rare earth leaching solution is aged at 80 ° C for 4 hours.
  • the solid-liquid separation results in a rare earth chloride solution and a rare earth fluoride precipitate.
  • the F content in the rare earth chloride solution is 1.5 mg / L. Chemical rare earth products.
  • the roasted products were collected and added with water at 25 ° C. for 4 h. After neutralization and impurity removal, a 32 g / L rare-earth sulfate solution was prepared, and the total yield of the rare-earth was 95%.
  • the obtained rare-earth sulfate solution is subjected to extraction transformation to obtain a rare-earth chloride solution, which is combined with the rare-earth chloride solution in step (2) and subjected to extraction and separation to obtain a single rare-earth compound product.
  • the sulfur-containing waste gas generated during the sulfuric acid roasting process is recovered by spraying and absorbing the sulfuric acid product.
  • Example 4-23 The steps of Example 4-23 are as in Example 1-3. The conditions of each step are shown in Table 2-4 below. The final total yield of rare earth is shown in Table 4:

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Abstract

一种联合法处理稀土精矿的冶炼分离工艺，采用气氛焙烧-盐酸浸出-硫酸焙烧的方法对包含氟碳铈矿的稀土精矿进行处理，并控制在盐酸浸出过程中采用低浓度盐酸步进式酸浸的方式，获得了高浓度稀土溶液(150-250g/L REO)，使得Ce的浸出率达到60％以上，并通过陈化降低了浸出液的F -含量，浸出渣经过硫酸焙烧-水浸进一步回收稀土，稀土总收率达95％以上。整个工艺的工业适应性更为广泛，可综合处理多种复杂稀土矿物，实现矿物型稀土精矿的绿色环保、高效清洁生产。

Description

一种联合法处理稀土精矿的冶炼分离工艺
本申请基于申请号为CN201810912079.3、申请日为2018年8月10日的中国专利申请提出，并要求该中国专利申请的优先权，该中国专利申请的全部内容在此引入本申请作为参考。
技术领域
本发明属于稀土冶炼分离技术领域，具体涉及一种联合法处理稀土精矿的冶炼分离工艺，适用于氟碳铈矿、氟碳铈矿与独居石、磷钇矿、磷灰石等稀土矿物中的一种或多种的混合型稀土矿物。
背景技术
我国稀土资源主要是以矿物型的轻稀土矿资源为主，约占总储量的90％以上，而工业稀土矿物主要是氟碳铈矿和独居石，其轻稀土含量高达96％-98％，据统计，具有工业应用价值的轻稀土矿床主要为包头白云鄂博稀土矿、四川攀西冕宁稀土矿和山东微山稀土矿。
目前，90％左右的包头混合型稀土矿采用有研总院开发的第三代硫酸法专利技术冶炼，即以浓硫酸进行强化焙烧分解、水浸、中和除杂、碳铵沉淀-盐酸溶解或P507和P204萃取转型与分离，该工艺具有简单可控、易于连续大规模生产、稀土回收率高等优势，并且对精矿品位的要求不高、运行成本较低。但在硫酸强化焙烧过程中会产生含硫和氟的复杂尾气，回收处理难度大，设备投资大，从而整体提高了运行成本。
而氟碳铈矿则普遍采用氧化焙烧-盐酸浸出化学法处理工艺，其精矿经过氧化焙烧分解生成可溶于盐酸的氧化稀土、氟化稀土或氟氧化稀土，而铈 则被氧化为四价离子，并且在盐酸浸出过程中，三价稀土被浸出得到少铈氯化稀土，铈和部分三价稀土、氟、钍则留在优溶渣中，再经过碱分解除氟，得到的富铈渣可用于生产硅铁合金，或经还原浸出生产纯度为98％左右的氧化铈，少铈氯化稀土则经过P507萃取剂分离为单一稀土。该工艺的优点是投资小、生产成本较低，但其缺陷即是工艺不连续，且盐酸浸出过程中铈、钍、氟不溶解则留在渣中，而渣经过碱转化后，氟会以氟化钠的形式进入废水，钍、氟则分散在渣和废水中难以被回收利用，导致整个工艺不仅对环境造成污染，而且铈产品的回收纯度仅为98％左右，利用价值低。
近年来，随着国内环保法规的逐步健全，各地对稀土行业污染物排放标准日趋严格。环保部于2011年1月24日颁布世界首部《稀土工业污染物排放标准》(GB26451-2011)，对现有和新建稀土工业企业生产设施水污染物和大气污染物排放限值、监测和监控均做出了明确要求。2011年5月10日，《国务院关于促进稀土行业持续健康发展的若干意见(国发〔2011〕12号)》。2015年1月1日起实施的《新环保法》明确规定对重点行业实行重点污染物排放总量控制制度。2016年10月，工信部发布的《稀土行业发展规划(2016-2020年)》对稀土行业“十三五”期间的生产指标和绿色发展指标均做出了明确要求。上述国家政策对稀土生产污染防治具有重大的战略意义，稀土行业也迫切需求绿色环保的冶炼分离新技术。
发明内容
为了解决现有氟碳铈矿处理工艺盐酸浸出液Ce含量低，富铈渣回收得到的氧化铈产品纯度低，混合型精矿处理工艺硫酸焙烧过程含氟、硫烟气排放高、处理回收成本高、稀土浸出液浓度低等问题。本发明所述的一种联合法处理稀土精矿的冶炼分离工艺，包括如下步骤：
(1)将稀土精矿在一定焙烧气氛下进行焙烧分解，得到焙烧矿；
(2)将所得焙烧矿加入盐酸浸出稀土，经固液分离，分别收集稀土浸出液和浸出渣；
(3)将所得浸出渣进行脱水处理后，加入浓硫酸进行焙烧，收集焙烧产物经水浸、中和除杂后，得到硫酸稀土溶液。
优选的，本发明所述稀土精矿包括但不限于氟碳铈矿或者氟碳铈矿与独居石或磷钇矿的一种或两种的混合型稀土矿。
所述步骤(1)中，所述焙烧步骤的焙烧气氛包括水蒸气或弱氧化气氛；所述弱氧化气氛包括但不限于N 2、CO、CO 2、空气、惰性气体中的一种或几种的气氛，其中可通过控制气体的通入量来降低氧气含量；水蒸气气氛的目的是为了脱氟，得到纯净的HF作为回收产物；弱氧化气氛的目的是降低铈的氧化率，提高稀土浸出收率。
所述步骤(1)中，脱氟后得到的HF气体，采用稀土氧化物，或稀土水合氧化物进行吸附回收得到氟化稀土产品。其原理是由于稀土氧化物在水中形成多核羟基化合物，使其上的OH -和氟离子发生离子交换而达到双重除氟效果。通过吸附回收，稀土精矿的脱氟焙烧分解过程产生的HF得到有效回收，得到氟化稀土产品，尾气实现达标排放，具有显著的环保效益。
所述步骤(1)中，所述焙烧步骤的焙烧温度为350-650℃，优选为400-600℃，其中一定范围内焙烧温度升高，可提高稀土浸出率；所述焙烧步骤的时间为0.5-6h，其中一定范围内延长焙烧时间，可提高稀土浸出率。
本发明工艺中，所述稀土精矿焙烧过程主要为精矿中的RECO 3F分解的过程，其反应式为：REFCO 3＝REOF(CeOF)+CO 2↑。在水蒸气存在的条件下，发生脱氟过程，其反应式为：REOF+H 2O＝RE 2O 3+HF↑。脱出的HF气体，通过稀土氧化物等吸附剂回收氟，其反应式为：6HF↑+RE 2O 3＝2REF3+3H 2O。
所述步骤(2)中，所述盐酸的浓度为3-10mol/L，优选为4-7mol/L，盐酸用量与所述焙烧精矿的比例0.4-2.0mol盐酸/100g稀土精矿，优选为0.7-1.5mol盐酸/100g稀土精矿。
所述步骤(2)中，所述盐酸浸出步骤优选为两步或多步盐酸逆流浸出， 第一步盐酸浸出后经固液分离得到一步稀土浸出液和一步浸出渣，一步浸出渣再进行下一步盐酸浸出，经固液分离得到本步稀土浸出液和本步浸出渣，其中本步稀土浸出液返回用作上一步盐酸浸出的底水，本步浸出渣可再进行下一步盐酸浸出。
所述步骤(2)中，所述盐酸的加入方式为在浸出过程中进行2-5级的连续并流浸出，且控制盐酸在每级浸出过程中呈浓度梯度加入，第一级加入较稀浓度的盐酸，后几级加入较高浓度的盐酸，以维持混合液的酸度为0.01-0.6mol/L，并优选为0.05-0.3mol/L，其中酸度越低越有利于稀土浸出。其目的是为了保证四价Ce进入溶液后不被还原，增加稀土和氟的浸出率；通过步进式浸出，可以获得较高的稀土浓度，浸出液中的稀土浓度达到150-250g/L。同时因为有效降低了浸出液的余酸含量，也减少后续流程的中和剂消耗。
在盐酸浸出焙烧矿的过程，采用较低的温度条件进行盐酸浸出，是因为F主要以[CeF x] 4-x配位化合物形式存在于溶液中，低温条件有利于[CeF x] 4-x配位化合物的稳态，可使稀土和F的更多的溶出，稀土浸出率达到70％-95％。
控制所述盐酸浸出步骤的浸出温度为10-75℃，优选为20-65℃，并控制总反应时间为0.5-10h，优选为1-6h，主要是为了提高稀土和F的浸出率。
所述步骤(3)中，所述脱水步骤为采用自然晾干和/或烘干的方式脱水处理，优选处理后的脱水浸出渣的含水率<10％，脱水浸出渣的REO含量20％-60％，主要为REPO 4，可与其他稀土精矿混合进行硫酸焙烧处理。
所述步骤(3)中，所述浓硫酸与脱水后的所述浸出渣的质量比(w/w)为0.3-1.2：1，并优选为0.5：1，本发明方案相比于现有技术，大量稀土在在1-2步骤中已经浸出，在硫酸焙烧步骤中，硫酸的使用量大幅减少。
所述步骤(3)中，所述硫酸焙烧步骤的温度为200-450℃，并优选200-220℃或250-350℃，所述焙烧步骤的焙烧时间为1-4h；
所述水浸步骤的温度为20-50℃，并优选为25-40℃，优选浸出时间2-5h，并优选的可将浸出液用碱性物质中和至pH为4-4.5，获得的硫酸稀土溶液浓度为25-45g/L(REO)。
所述步骤(3)中，盐酸浸出渣先用水洗涤，控制洗水：浸出渣比(w/w)为0.5-10：1，优选为0.5-5：1，烘干后浸渣含水率在0-50％，优选为0-30％。处理后的洗水中含有稀土浓度为5-50g/L(REO)，H +浓度<0.1mol/L。洗水的目的是为了将浸渣中夹带的氯化稀土洗进溶液，进一步提高稀土浸出率，同时去除可能会对后续硫酸化焙烧设备造成腐蚀的氯离子，一定程度的脱水处理则可以减少对硫酸强化焙烧过程浓硫酸的稀释，将水洗液回用于步骤(2)中的稀土精矿调浆或配制盐酸，实现水洗液的闭路循环。
所述步骤(3)中，还包括将所得硫酸稀土溶液加入铁粉进行配置的步骤，配加铁粉的量为盐酸浸出渣质量的2％-10％。
所述步骤(3)中，还包括将所得硫酸稀土溶液进行萃取转型的步骤，得到氯化稀土溶液，经萃取分离，得到单一稀土化合物。
所述萃取转型步骤为通过沉淀或萃取进行转型处理。
所述步骤(2)中，还包括将所得稀土浸出液进行陈化处理的步骤，固液分离得到氯化稀土溶液和氟化稀土产品；并将所得氯化稀土溶液与步骤(3)中所述硫酸稀土溶液经过转型得到的氯化稀土溶液进行合并，经萃取分离得到单一稀土化合物。
所述陈化步骤为在静置或搅拌条件下进行，过滤得到氟化镧铈产品；控制所述陈化步骤的温度为60-90℃，进一步优选为65-80℃，并优选控制所述陈化步骤的温度应该等于或高于所述盐酸浸出温度；所述陈化步骤的时间为0.5-10h，优选为1-4h。
由于浸出液中的F主要以[CeF x] 4-x配位化合物形式存在于溶液中，Cl 2/Cl –的电极电势随着温度升高降低，显著低于Ce 4+/Ce 3+的电极电位，使形成的[CeF x] 4–x配合物被Cl –还原，并释放出F –；生成的F –立即与RE 3+结合 (主要为F离子周边的Ce)，形成氟化稀土沉淀；此外，氟化稀土的溶度积与温度呈反相关的关系，温度越高，其溶度积越低，如CeF 3，25℃，Ksp＝8.0×10 -16；100℃，Ksp＝9.3×10 -18，进一步促进氟化稀土沉淀。以此，本发明工艺中使用高温陈化步骤，可有效将F从浸出液中分离进入渣中，浸出液中的F含量<8mg/L，升高陈化温度、延长陈化时间可进一步降低F含量，避免F对后续萃取分离的影响。
本步骤中，所述陈化处理可获得氟化稀土沉淀，并优选控制浸出溶液中的F<8mg/L，进一步优选为<2mg/L；步骤(2)固液分离获得的浸出渣中的F残留量相对于稀土精矿中F含量的比例为5％以下，优选为1％以下；所述陈化过滤后得到的氯化稀土溶液稀土浓度为150-250g/L(REO)，稀土浸出率70％-95％，其中Ce的浸出率60％-95％。如下表1给出了本方法与传统氟碳铈矿处理方法获得盐酸浸出液对比，可见，本方法获得的浸出液浓度、稀土总浸出率、Ce浸出率均高于传统氟碳铈矿方法，浸出液的H +浓度较低，浸出液中基本不含F，具有明显的技术优势。
表1本方法与传统氟碳铈矿处理方法获得盐酸浸出液对比
Figure PCTCN2019099932-appb-000001
所述的联合法处理稀土精矿的冶炼分离工，还包括将所述步骤(1)中产生的含氟尾气通过水或碱性液体喷淋处理，或稀土氧化物、稀土水合氧化物中的一种或两种吸附剂进行脱氟回收稀土氟化物产品；和/或，将所述步骤(3)中产生的含硫尾气进行脱硫回收处理的步骤，得到硫酸产品；
所述硫酸焙烧过程产生的含硫尾气经过脱硫回收处理后，不仅尾气排放达标，且可回收获得高纯度硫酸产品，浓度可达到80％以上，有效解决了传统工艺中尾气成分含F、设备腐蚀损耗严重、F与含S物质分离困难、难以达标处理、运行成本高昂等问题。
本发明所述联合法处理稀土精矿的冶炼分离工艺，采用气氛焙烧-盐酸浸出-硫酸焙烧的方法对包含氟碳铈矿的稀土精矿进行处理，并控制在盐酸浸出过程中采用低浓度盐酸步进式酸浸的方式，获得了较高的氯化稀土溶液(150-250g/L REO)；同时，利用[CeF x] 4-x配位化合物的存在特点，使更多的Ce进入溶液，使得Ce的浸出率达到60％-95％，总稀土浸出率达到70％-95％；再者，本发明工艺利用Cl -还原性与氟化稀土溶度积与温度的关系，通过高温陈化进一步降低了浸出液的F -含量。相较于传统处理氟碳铈矿的氧化焙烧-盐酸浸出工艺，浸出液浓度和稀土浸出率均大幅提高，省却了传统工艺需进一步蒸发浓缩才能获得高浓度氯化稀土溶液的步骤，且溶液中F含量很低，避免了F进入萃取体系产生三相，可直接进入P507-HCl体系分离提纯单一稀土元素。
本发明所述联合法处理稀土精矿的冶炼分离工艺，所述稀土精矿经气氛焙烧-盐酸浸出-高温陈化之后，仅5％-30％的稀土残留在盐酸浸渣中，相较于传统的处理混合型稀土精矿的工艺，大幅降低了浓硫酸的消耗，同时也大量减少了水浸过程用水的消耗，因有70％-95％的稀土均直接进入氯化体系萃取分离稀土，还大幅减少了硫酸浸出液萃取转型为氯化稀土溶液的酸碱消耗。
本发明所述联合法处理稀土精矿的冶炼分离工艺中，氟的走向得到有效控制，一是在气氛焙烧时，使大部分氟转化为氟化氢气体进入尾气，并经过吸附回收处理制备含氟产品；二是在盐酸浸出过程中，将少量氟浸出进入氯化稀土溶液中，经过陈化得到氟化稀土产品，使氯化稀土溶液中的F -浓度<8mg/L，避免氟进入浸出渣中，解决了硫酸焙烧工序含氟和含硫混合尾气难以治理的难题。因盐酸浸渣中剩余的稀土量大幅减少，硫酸焙烧过程中的浓硫酸用量也随之减少，硫酸焙烧过程中产生的SO 2烟气也减少排放60％以上，大幅降低了废气和废水处理成本，更加接近清洁生产的目标，生态友好，整体具有显著的经济和环保优势；而F的有效回收处理，也解决传统工艺含F废水量大、处理难以达标的难题。整个工艺的工业适应性更为广 泛，可综合处理多种复杂矿物，实现了包头混合型稀土精矿、氟碳铈精矿等矿物型稀土精矿绿色环保、高效清洁生产，经济和社会效益显著。
附图说明
为了使本发明的内容更容易被清楚的理解，下面根据本发明的具体实施例并结合附图，对本发明作进一步详细的说明，其中，
图1为本发明所述联合法处理稀土精矿的冶炼分离工艺的流程图。
具体实施方式
实施例1
本实施例所述工艺处理的稀土精矿为氟碳铈矿与独居石的混合型稀土矿，按照如图1所示的工艺流程图，本实施例所述联合法处理稀土精矿的冶炼分离工艺，包括如下步骤：
(1)将氟碳铈矿与独居石的混合型稀土精矿在空气气氛(氧气含量21％)下，于500℃进行焙烧处理4h，得到焙烧矿；
本步骤中，焙烧过程中逸出的HF，采用水喷淋进行处理。
(2)将所得焙烧矿加入盐酸于25℃进行4级连续并流浸出，盐酸初始浓度为6mol/L，盐酸用量与所述焙烧精矿的比例为1.0mol盐酸/100g稀土精矿。经固液分离，分别收集稀土浸出液和浸出渣，得到的稀土浸出液稀土含量238g/L，稀土浸出率77％，Ce浸出率70％；
本步骤中，所述盐酸浸出步骤优选为两步或多步盐酸逆流浸出，第一步盐酸浸出后经固液分离得到一步稀土浸出液和一步浸出渣，一步浸出渣再进行下一步盐酸浸出，经固液分离得到本步稀土浸出液和本步浸出渣，其中本步稀土浸出液返回用作上一步盐酸浸出的底水，本步浸出渣可再进行下一步盐酸浸出。
本步骤中，所述浸出处理的盐酸加入方式为在浸出过程中进行4级的连续并流浸出，且控制盐酸浓度，在第1-4级分别加入1.5mol/L、2mol/L、 6mol/L、8mol/L的盐酸，混合液的酸度在0.1-0.05mol/L之间梯度降低；
本步骤中，稀土浸出液在65℃陈化4h，固液分离得到氯化稀土溶液和氟化稀土沉淀，氯化稀土溶液中的F含量为1.9mg/L，氟化稀土沉淀烘干后获得氟化稀土产品。
(3)将所得浸出渣洗涤、烘干脱水处理至含水率9％后，加入浓硫酸于300℃进行焙烧3h，控制所述浓硫酸与浸出渣的质量比(w/w)为0.3：1；
收集焙烧产物加水于25℃进行水浸4h，随后经中和除杂后，制得32g/L的硫酸稀土溶液；稀土总收率97％。将所得硫酸稀土溶液进行萃取转型，得到氯化稀土溶液，与步骤(2)中的氯化稀土溶液合并，进行萃取分离，得到单一稀土化合物产品。
本步骤中，硫酸焙烧过程产生的含硫废气，采用洗水喷淋吸收的方式回收得到硫酸产品。
实施例2
本实施例所述工艺处理的稀土精矿为氟碳铈矿与独居石的混合型稀土矿，本实施例所述联合法处理稀土精矿的冶炼分离工艺，包括如下步骤：
(1)将氟碳铈矿与独居石的混合型稀土矿在空气弱氧化气氛(通过调节进气阀开度等方式控制氧气含量为12％)下，于500℃进行焙烧处理4h，进气阀的开度为50％，得到焙烧矿；
本步骤中，焙烧过程中逸出的HF，采用水喷淋进行回收处理。
(2)将所得焙烧矿加入盐酸于25℃进行4级连续并流浸出，盐酸初始浓度为6mol/L，盐酸用量与所述焙烧精矿的比例为1.0mol盐酸/100g稀土精矿。经固液分离，分别收集稀土浸出液和浸出渣，得到的稀土浸出液稀土含量250g/L，稀土浸出率80％，Ce浸出率75％；
本步骤中，所述盐酸浸出步骤优选为两步或多步盐酸逆流浸出，第一步盐酸浸出后经固液分离得到一步稀土浸出液和一步浸出渣，一步浸出渣再进行 下一步盐酸浸出，经固液分离得到本步稀土浸出液和本步浸出渣，其中本步稀土浸出液返回用作上一步盐酸浸出的底水，本步浸出渣可再进行下一步盐酸浸出。
本步骤中，所述浸出处理的盐酸加入方式为在浸出过程中进行4级的连续并流浸出，且控制盐酸浓度，在第1-4级分别加入1.5mol/L、2mol/L、6mol/L、8mol/L的盐酸，混合液的酸度在0.1-0.05mol/L之间梯度降低；
本步骤中，稀土浸出液在80℃陈化4h，固液分离得到氯化稀土溶液和氟化稀土沉淀，氯化稀土溶液中的F含量为1.2mg/L，氟化稀土沉淀烘干后获得氟化稀土产品。
(3)将所得浸出渣洗涤、烘干脱水处理至含水率9％后，加入浓硫酸于300℃进行焙烧3h，控制所述浓硫酸与浸出渣的质量比(w/w)为0.3：1；
收集焙烧产物加水于25℃进行水浸4h，随后经中和除杂后，制得32g/L的硫酸稀土溶液；稀土总收率97％。将所得硫酸稀土溶液进行萃取转型，得到氯化稀土溶液，与步骤(2)中的氯化稀土溶液合并，进行萃取分离，得到单一稀土化合物产品。
本步骤中，硫酸焙烧过程产生的含硫废气，采用洗水喷淋吸收的方式回收得到硫酸产品。
实施例3
本实施例所述工艺处理的稀土精矿为氟碳铈矿与独居石的混合型稀土矿，本实施例所述联合法处理稀土精矿的冶炼分离工艺，包括如下步骤：
(1)将氟碳铈矿与独居石的混合型稀土矿在水蒸气气氛下，于650℃进行焙烧处理4h，进气阀的开度为100％，得到焙烧矿；
本步骤中，焙烧过程中逸出的HF，采用稀土氧化物吸附剂进行回收，得到氟化稀土产品。
(2)将所得焙烧矿加入盐酸于25℃进行4级连续并流浸出，盐酸初始 浓度为6mol/L，盐酸用量与所述焙烧精矿的比例为1.0mol盐酸/100g稀土精矿。经固液分离，分别收集稀土浸出液和浸出渣，得到的稀土浸出液稀土含量235g/L，稀土浸出率75％，Ce浸出率为69％；
本步骤中，所述盐酸浸出步骤优选为两步或多步盐酸逆流浸出，第一步盐酸浸出后经固液分离得到一步稀土浸出液和一步浸出渣，一步浸出渣再进行下一步盐酸浸出，经固液分离得到本步稀土浸出液和本步浸出渣，其中本步稀土浸出液返回用作上一步盐酸浸出的底水，本步浸出渣可再进行下一步盐酸浸出。
本步骤中，所述浸出处理的盐酸加入方式为在浸出过程中进行4级的连续并流浸出，且控制盐酸浓度，在第1-4级分别加入1.5mol/L、2mol/L、6mol/L、8mol/L的盐酸，混合液的酸度在0.1-0.05mol/L之间梯度降低；
本步骤中，稀土浸出液在80℃陈化4h，固液分离得到氯化稀土溶液和氟化稀土沉淀，氯化稀土溶液中的F含量为1.5mg/L，氟化稀土沉淀烘干后获得氟化稀土产品。
(3)将所得浸出渣洗涤、烘干脱水处理至含水率9％后，加入浓硫酸于300℃进行焙烧3h，控制所述浓硫酸与浸出渣的质量比(w/w)为0.3：1；
收集焙烧产物加水于25℃进行水浸4h，随后经中和除杂后，制得32g/L的硫酸稀土溶液，稀土总收率95％。将所得硫酸稀土溶液进行萃取转型，得到氯化稀土溶液，与步骤(2)中的氯化稀土溶液合并，进行萃取分离，得到单一稀土化合物产品。
本步骤中，硫酸焙烧过程产生的含硫废气，采用洗水喷淋吸收的方式回收得到硫酸产品。
实施例4-23的步骤如实施例1-3，每步骤的条件如下表2-4所示，最终的稀土总收率如表4所示：
表2
Figure PCTCN2019099932-appb-000002
表3
Figure PCTCN2019099932-appb-000003
Figure PCTCN2019099932-appb-000004
表4
Figure PCTCN2019099932-appb-000005
Figure PCTCN2019099932-appb-000006
由此可见，通过本发明可综合处理多种复杂稀土矿物，整个工艺的工业适用性广泛。通过气氛焙烧使大部分氟转化为氟化氢气体进入尾气，并经过吸附回收处理制备含氟产品；通过多级连续盐酸浸出并调控酸度，获得了高浓度氯化稀土溶液(150-250g/L REO)，同时保证稀土浸出率达到70％以上；通过陈化降低了浸出液的F -含量，获得氟化稀土产品以避免氟进入浸出渣中，解决了硫酸焙烧工序含氟和含硫混合尾气难以治理的难题；浸出渣经过硫酸焙烧-水浸进一步回收稀土，稀土总收率达95％以上。实现了矿物型稀土精矿的绿色环保、高效清洁生产。
显然，上述实施例仅仅是为清楚地说明所作的举例，而并非对实施方式的限定。对于所属领域的普通技术人员来说，在上述说明的基础上还可以做出其它不同形式的变化或变动。这里无需也无法对所有的实施方式予以穷举。而由此所引伸出的显而易见的变化或变动仍处于本发明创造的保护范围之中。

Claims (11)

  1. 一种联合法处理稀土精矿的冶炼分离工艺，其特征在于，包括如下步骤：
    (1)将稀土精矿在一定焙烧气氛下进行焙烧分解，得到焙烧矿；
    (2)将所得焙烧矿加入盐酸浸出稀土，经固液分离，分别收集稀土浸出液和浸出渣；
    (3)将所得浸出渣进行脱水处理后，加入浓硫酸进行焙烧，收集焙烧产物经水浸、中和除杂后，得到硫酸稀土溶液。
  2. 根据权利要求1所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(1)中，所述焙烧步骤的焙烧气氛包括水蒸气、空气、CO、CO 2中的一种或多种。
  3. 根据权利要求1-2任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(1)中，所述焙烧步骤的焙烧温度为350-650℃。
  4. 根据权利要求1-3任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(2)中，所述盐酸浸出步骤优选为两步或多步盐酸逆流浸出，第一步盐酸浸出后经固液分离得到一步稀土浸出液和一步浸出渣，一步浸出渣再进行下一步盐酸浸出，经固液分离得到本步稀土浸出液和本步浸出渣，其中本步稀土浸出液返回用作上一步盐酸浸出的底水，本步浸出渣可再进行下一步盐酸浸出。
  5. 根据权利要求1-3任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(2)中，所述盐酸的加入方式为在浸出过程中进行2-5级的连续并流浸出，且控制盐酸在每级浸出过程中呈浓度梯度加入，以维持浸出过程混合液的酸度为0.01-0.6mol/L。
  6. 根据权利要求1-5任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(2)中，控制所述盐酸浸出步骤的浸出温度为 10-75℃。
  7. 根据权利要求1-6任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(3)中，所述浓硫酸与脱水后的所述浸出渣的质量比(w/w)为0.3-1.2：1。
  8. 根据权利要求1-7任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(3)中，所述硫酸焙烧步骤的温度为200-450℃，所述水浸步骤的温度为20-50℃。
  9. 根据权利要求1-8任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(3)中，还包括将所得硫酸稀土溶液进行萃取分离，或萃取转型或沉淀转型的步骤，得到氯化稀土溶液，经萃取分离，得到单一稀土化合物。
  10. 根据权利要求1-9所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，所述步骤(2)中，将所得稀土浸出液在60-90℃下进行陈化1-5小时，固液分离得到氯化稀土溶液和氟化稀土粉体产品；并将所得氯化稀土溶液与步骤(3)中所述硫酸稀土溶液转型得到的氯化稀土溶液进行合并，经萃取分离得到单一稀土化合物。
  11. 根据权利要求1-10任一项所述的联合法处理稀土精矿的冶炼分离工艺，其特征在于，将所述步骤(1)焙烧过程中产生的含氟尾气通过水或碱性液体喷淋处理，或稀土氧化物、稀土水合氧化物中的一种或两种吸附剂进行脱氟回收稀土氟化物产品；将所述步骤(3)硫酸焙烧过程中产生的含硫尾气进行脱硫回收处理，得到硫酸产品。
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