Method for extracting and concentrating rare earths from phosphogypsum

Abstract

The present invention relates to a method for extracting rare earths by treating phosphogypsum. The process comprises basic leaching of the phosphogypsum followed by sonication-catalyzed acid leaching of the solid residue obtained at the end of the basic leaching.

Description

[0001] METHOD FOR EXTRACTION AND CONCENTRATION OF RARE EARTHS FROM PHOSPHOGYPSE

[0002] TECHNICAL FIELD OF THE INVENTION

[0003] The present invention relates to a process for extracting rare earth elements by processing phosphogypsum. The process comprises a basic leaching of the phosphogypsum followed by an acid leaching, catalyzed by sonication, of the solid residue obtained after the basic leaching.

[0004] STATE OF THE ART

[0005] Rare earth elements (REEs) generally refer to 17 metallic elements, i.e., the lanthanide series (lanthanum, cerium, praseodymium, neodymium, promethium, samarium, europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium, ytterbium, and lutetium), as well as scandium and yttrium. REEs and their alloys possess remarkable magnetic properties and are used in emerging technologies, particularly green digital technologies. Their sustainable supply is therefore a major challenge. Currently, developing a sustainable and cost-effective process that adheres to the principles of green chemistry by utilizing secondary resources or industrial by-products is of paramount importance.

[0006] Therefore, in order to preserve natural resources, it is essential to seek other potentially exploitable alternatives, especially in countries where primary resources are limited or even nonexistent. Various industries produce by-products containing rare earth elements in varying concentrations. The main advantage of these resources is the reduction of overall costs, the preservation of natural raw materials, and the valorization of waste.

[0007] In this context, phosphogypsum (PG) appears as an essential alternative to meet the demands of the rapidly growing market, while preserving natural resources and reducing environmental impact.

[0008] Phosphorus phosphate (PG) is a by-product of phosphoric acid production. It results from the reaction of a strong acid, such as hydrochloric acid or sulfuric acid, on phosphate ores, typically apatite, according to the following reaction: The quantities of PG produced annually are significant; i.e., the production of one tonne of phosphoric acid is accompanied by the production of five tonnes of PG. Globally, approximately 7 billion tonnes of PG have been produced to date, with a current production rate of 150 to 200 million tonnes per year (document [3]).

[0009] PG is generally found in the form of a fine, moist, and aggregated powder. It is mainly composed of calcium sulfate, sulfur oxide derivatives, and to a lesser extent, derivatives based on silicon, sodium, TRs, aluminum, iron, phosphorus pentoxide, fluorine, etc.

[0010] It is known that PG contains a relatively low content of TRs, i.e., the concentration of TRs elements is generally less than 0.1% by weight in PG from the processing of sedimentary phosphate ores, and can reach 0.3% by weight in PG from the processing of igneous phosphate ore (document [2]).

[0011] Numerous methods for extracting TRs from PG have been developed. These methods consist of concentrating the TRs in filtrates obtained from the acid leaching of PG.

[0012] Acid leaching (using nitric, hydrochloric or sulfuric acid) is the most widespread method for recovering rare earth elements contained in PG. The TRs present in phosphogypsum can be recovered with a yield of 52% by leaching the PG at a temperature of 40°C with a 10 to 15% H2SO4 solution for a liquid:solid ratio of 2 (document [5]).

[0013] It is important to note that it is difficult to recover all the TRs present in PG without destroying the crystal lattice. Thus, direct recovery of TRs is generally quite complex due to their simultaneous precipitation with calcium sulfate. The efficiency of acid leaching can be improved by various methods: mechanical activation of PG via ball milling, the use of Purolite C-160 type resins (trade name), in the presence of 20% sulfuric acid, or in a solid-solid-liquid leaching system, to achieve a maximum recovery of 72.5% of the TRs (document [6]).

[0014] RU2473708 C1 (document [7]) describes a process for leaching TRs with a yield exceeding 70%. The operating procedure consists of treating phosphogypsum with sulfuric acid in the presence of a sorbent containing a sulfonic acid-type functional group. The acid:PG and PG:sorbent ratio varies from 4 to 7. Furthermore, the contact time varies from 5 to 7 hours.

[0015] The use, on a laboratory scale, of a pretreatment of phosphogypsum with microwave at 1200W for 15 minutes before leaching with 1.5 M HCl for a period of 60 minutes at 85°C improves the leaching efficiency to a value of 92.5% (document [8]).

[0016] WO 2020 / 067856 A1 describes the acid leaching of phosphogypsum, previously washed with a potassium sulfate solution, using 2N nitric acid for 6 hours, in the presence of a potassium chlorate oxidizing agent and a zinc powder reducing agent, yields a filtrate with a rare earth concentration (REE) of 179 ppm. However, the PG:acid weight ratio is 4:6, and the amounts of potassium chlorate and zinc powder are 10 g / kg and 5 g / kg, respectively, relative to the leaching mixture. The addition of an oxidizing agent aims to promote the transfer of rare earths into the leachate, while the reducing agent minimizes the solubility of impurities. A further evaporation step increases the rare earth concentration in the filtrate to 260 ppm.

[0017] Generally, direct leaching of phosphogypsum results in an acidic filtrate with a low concentration of rare earth elements, which is a major limitation of these processes. Hence the need to develop new methods that improve rare earth element extraction.

[0018] EP 0419318 A1 [9] describes a process for concentrating rare earth elements in a residue obtained from the salt leaching of phosphogypsum. This process consists of leaching the phosphogypsum with a 25 g / L sodium chloride salt water solution, preceded by basic treatment of the resulting insoluble residues with a sodium bicarbonate solution, before leaching the residues with concentrated nitric acid. The dissolution is carried out under conditions in which the salt water-to-phosphogypsum ratio is 200. However, this process consumes water and sodium chloride and results in the release into the environment of large quantities of salt water containing, for example, chlorides, sulfates, fluorides, etc.

[0019] WO 2014 / 148945 A1

[0010] describes how the addition of potassium or sodium during the attack of phosphate rock with concentrated sulfuric acid promotes the syncrystallization of rare earth sulfates NaTR(SO4)2 or KTR(SO4)2 with the precipitated gypsum phase. The process is based on the addition of sodium sulfate or potassium sulfate with a mass ratio of K2O to Ln2O3 ranging from 0.25 to 5. More than 98% of the rare earths present in the apatite precipitate with the phosphogypsum. However, the transformation process used is of the hemihydrate type, which is increasingly being abandoned, and the concentration factor of rare earths in the phosphogypsum does not exceed 1.5, which constitutes a limitation of the process. The present invention therefore aims to overcome the aforementioned drawbacks and to provide a simple and sustainable process for isolating rare earths with a high concentration factor.

[0020] Listes des documents de l’art antérieur cités

[0021] Document [1] : S. Wu et al., « Recovery of rare earth elements from phosphate rock by hydrometallurgical processes - A critical review », Chemical Engineering Journal, vol. 335, p. 774-800, mars 2018, doi: 10.1016 / j.cej.2017.10.143.

[0022] Document [2] : W02020067856A1

[0023] Document [3] : I. Hammas-Nasri, S. Elgharbi, M. Ferhi, K. Horchani-Naifer, et M.

[0024] Férid, « Investigation of phosphogypsum valorization by the integration of the Merseburg method », New J. Chem., vol. 44, no 19, p. 8010-8017, 2020, doi: 10.1039 / D0NJ00387E. Document [4] : M. Walawalkar, C. K. Nichol, et G. Azimi, « Process investigation of the acid leaching of rare earth elements from phosphogypsum using HCl, HNO3, and H2SO4», Hydrometallurgy, vol. 166, p. 195-204, déc. 2016, doi: 10.1016 / j.hydromet.2016.06.008.

[0025] Document [5] : A. Jarosihski, J. Kowalczyk, et Cz. Mazanek, « Development of the Polish wasteless technology of apatite phosphogypsum utilization with recovery of rare earths », Journal of Alloys and Compounds, vol. 200, no 1-2, p. 147-150, oct. 1993, doi: 10.1016 / 0925-8388(93)90485-6.

[0026] Document [6] : V. N. Rychkov et al., « Recovery of rare earth elements from phosphogypsum », Journal of Cleaner Production, vol. 196, p. 674-681 , sept. 2018, doi: 10.1016 / j.jclepro.2018.06.114.

[0027] Document [7] : RU2473708C1

[0028] Document [8] : A. Lambert, J. Anawati, M. Walawalkar, J. Tam, et G. Azimi, «

[0029] Innovative Application of Microwave Treatment for Recovering of Rare Earth Elements from Phosphogypsum », ACS Sustainable Chemistry & Engineering, vol. 6, no 12, p. 16471-16481 , déc. 2018, doi: 10.1021 / acssuschemeng.8b03588.

[0030] Document [9] : EP0419318A1

[0031] Document

[0010] : WO 2014 / 148945 A1.

[0032] SUMMARY OF THE INVENTION

[0033] The present invention relates to a process for extracting rare earths by processing phosphogypsum, the process comprising the following steps: (a) basic leaching of phosphogypsum by a basic aqueous solution comprising a chelating agent, leading to the formation of a liquid filtrate (Fa) and a solid residue (Ra);

[0034] (b) separation of the liquid filtrate (Fa) and the solid residue (Ra);

[0035] (c) acid leaching of the solid residue (Ra) obtained in step (b) by an acidic aqueous solution leading to the formation of a liquid filtrate comprising the rare earths (Fc) and a solid residue (Rc), the acid leaching being catalyzed by sonication;

[0036] (d) separation of the liquid filtrate (Fd) and the solid residue (Rd).

[0037] Other aspects of the invention are as described below.

[0038] DEFINITIONS

[0039] The term "rare earths", abbreviated as TRs, refers to the rare earth elements called lanthanides with atomic numbers from 57 to 71 inclusive, yttrium with atomic number 39 and scandium with atomic number 21.

[0040] The term "phosphogypsum" refers to a by-product of the wet process of phosphoric acid production, that is, obtained by reacting fluorophosphate calcium ores with a strong acid such as sulfuric acid. This type of phosphogypsum is also called LREC phosphogypsum, from the initials of the English term "Low Rare Earth Case".

[0041] The term "leaching" in the context of the present invention refers to the extraction of a soluble compound from a solid residue by means of washing operations. When the washing is carried out using a basic aqueous solution, it is referred to as "basic leaching," while it is referred to as "acid leaching" when the washing is carried out using an acidic aqueous solution.

[0042] Unless otherwise stated, the ratios indicated in this application are mass ratios and the percentages are mass percentages.

[0043] DETAILED DESCRIPTION OF THE INVENTION

[0044] The aim of the present invention is to remedy the aforementioned drawbacks of the prior art and to propose an innovative process for extracting rare earths by leaching phosphogypsum.

[0045] The present invention relates to a process for extracting rare earths by processing phosphogypsum, the process comprising the following steps: (a) basic leaching of phosphogypsum by a basic aqueous solution comprising a chelating agent, leading to the formation of a liquid filtrate (Fa) and a solid residue (Ra);

[0046] (b) separation of the liquid filtrate (Fa) and the solid residue (Ra);

[0047] (c) acid leaching of the solid residue (Ra) obtained in step (b) by an acidic aqueous solution leading to the formation of a liquid filtrate comprising rare earths (Fc) and a solid residue (Rc), the acid leaching being catalyzed by sonication;

[0048] (d) separation of the liquid filtrate (Fd) and the solid residue (Rd).

[0049] The phosphogypsum useful in the context of the present invention typically results from the action of a strong acid, for example sulfuric acid, on a phosphate rock. Generally, phosphogypsum comprises 26% to 40% CaO, 30% to 60% SO3, less than 3% SiO2, less than 2% P2O5, less than 2% F, and 0.01% to 0.1% rare earth elements (% expressed by weight).

[0050] Typically, phosphogypsum has a rare earth content ranging from 314 to 344 ppm or even from 351 to 448 ppm, or even from 314 to 458 ppm.

[0051] Phosphogypsum may be subjected to one or more pretreatment steps before being sent to the basic leaching step (a).

[0052] The preprocessing step(s) can be chosen from the group consisting of the following steps:

[0053] - A calcination step of phosphogypsum at a temperature ranging from 700 to 850°C;

[0054] - A step of washing the phosphogypsum with water;

[0055] - A drying stage of phosphogypsum at a temperature ranging from 40 to 90°C, for example at 80°C;

[0056] - A grinding and sieving step to obtain phosphogypsum in the form of particles with a size less than or equal to 250 pm, typically ranging from 10 to 250 pm.

[0057] The process according to the invention includes a step (a) of basic leaching of phosphogypsum (optionally pretreated) with a basic aqueous solution comprising a chelating agent. The basic leaching step makes it possible to remove the major metallic species from the phosphogypsum, such as, for example, Ca2+ ions. 2+ , Mg 2+ Ba 2+ , in the presence of the basic solution, which causes their dissociation and solubilization.

[0058] The addition of a chelating agent promotes the complexation of Ca2+ ions 2+ , Mg 2+ Ba2+ , etc.

[0059] The chelating agent is preferably an octadentate pentaanionic ligand of the diethylenetriaminepentaacetic acid (DTPA) type, a tetradentate trianionic ligand of the nitrilotriacetic acid (NTA) type, a tetraanionic hexadentate ligand such as ethylenediaminetetraacetic acid (EDTA) or one of its salts, for example tetrasodium EDTA.

[0060] The chelating agent is advantageously present in the basic aqueous solution at a concentration ranging from 0.1 to 0.8M, preferably from 0.2 to 0.4M.

[0061] The stoichiometry between the chelating agent and the added phosphogypsum generally varies from 1 to 2, preferably from 1.1 to 1.4, even more preferably from 1.2.

[0062] Basic aqueous solutions generally include one or more base(s), for example, sodium hydroxide or sodium carbonate.

[0063] The basic solution can have a molar concentration in one or more base(s) ranging from 0.1 to 0.8 M, typically from 0.2 to 0.5 M or even from 0.3 to 0.4 M.

[0064] Typically, the basic leaching step (a) is carried out at a pH greater than or equal to 10, or greater than or equal to 11, 12, and less than or equal to 14. Such pH levels lead to the predominance of the basic form of EDTA (Y 4 -) in solution, which allows for strong complexation with Ca ions 2+ , Mg 2+ Ba 2+ .

[0065] Generally, the basic leaching step (a) is carried out under agitation, for example mechanical.

[0066] The basic leaching step (a) is typically carried out over a period of 5 to 15 minutes, or 15 to 30 minutes.

[0067] The basic leaching step (a) is typically carried out at a temperature ranging from 20 to 80°C, or from 20 to 70°C, or from 55 to 65°C. Generally, step (a) is carried out at atmospheric pressure. The solubilization rate of phosphogypsum in the basic aqueous solution can be at least 90% by weight relative to the initial mass, or even from 90 to 99% by weight. Generally, the solubilization rate is at least 90% by weight; typically it ranges from 91 to 99% or from 94 to 98% of the initial mass of phosphogypsum. In particular, with a basic solution comprising a sodium salt of EDTA, a strong dissociation of phosphogypsum is obtained: the solubilization rate is between 91 and 99%, preferably between 94 and 98% of the initial mass of phosphogypsum.

[0068] The solubilization rate therefore refers to the quantity by weight of phosphogypsum solubilized by the basic solution relative to the initial mass of phosphogypsum.

[0069] The basic leaching step (a) leads to the formation of a liquid filtrate (Fa) and a solid residue (Ra).

[0070] Typically, the liquid filtrate (Fa) is rich in mineral elements, mostly calcium sulfates.

[0071] Typically, the residue (Ra) is rich in TRs and may also contain other metallic elements such as Ca, Al, Sr, Mg.

[0072] Separation step (b)

[0073] The process includes a step (b) of separating the liquid filtrate (Fa) and the solid residue (Ra).

[0074] Step (b) of separation can be carried out by a solid / liquid separation allowing the solid residue (Ra) containing the rare earths to be separated from the liquid filtrate (Fa) formed from the basic leaching solution and the dissolved phosphogypsum elements.

[0075] Commonly, solid / liquid separation can be achieved by filtration, centrifugation, and / or decantation.

[0076] The separation of residues and filtrate is typically achieved by centrifugation at 3000 rpm for a period of 15 minutes.

[0077] The residue (Ra) obtained at the end of the liquid / solid separation typically has a concentration of rare earths higher than the initial concentration in the phosphogypsum, typically the concentration of TRs can range from 1559 to 1666 ppm or from 1650 to 1950 ppm, or even from 1665 to 1939 ppm, starting from a phosphogypsum having an initial concentration of TRs ranging from 351 to 448. The concentration factor thus generally varies from 4.3 to 4.5, or even from 4.3 to 4.7.

[0078] The filtrate (Fa) can commonly be redirected to a sodium sulfate production process as described for example in patent WO 2018 / 021900 A2.

[0079] Typically, the separated residue (Ra) is then subjected to the following processing steps:

[0080] - A washing step with water to remove traces of basic solution in the residue (Ra);

[0081] - A centrifugation step typically at a speed ranging from 2500 to 4000 rpm, for example 3000 rpm;

[0082] - A drying stage at a temperature ranging from 40 to 90°C, for example at 80°C; and

[0083] - A grinding and sieving step to obtain a residue (Ra) in the form of particles having a size less than or equal to 250 pm, typically ranging from 15 to 250 pm.

[0084] The residue (Ra) obtained at the end of the treatment steps (washing, centrifugation, drying, grinding) has a higher concentration of TRs than the concentration of the residue obtained at the end of the solid / liquid separation step, typically ranging from 3400 to 4300 ppm or even from 3427 to 4293 ppm, starting from a phosphogypsum with an initial concentration of TRs ranging from 314 to 458. The concentration factor thus generally varies from 9 to 11.

[0085] Acid leaching step (c)

[0086] The process according to the invention includes a step (c) of acid leaching of the solid residue (Ra) obtained at the end of step (b) (with or without additional treatments) with an acidic aqueous solution containing one or more strong acid(s) selected from the group consisting of sulfuric acid, nitric acid, and hydrochloric acid, leading to the formation of a liquid filtrate (Fc) and a solid residue (Rc).

[0087] Acid leaching allows the extraction of elements soluble in a strong acid from the solid residue (Ra). The acid leaching step thus leads to the dissolution of TRs, calcium hydroxide, and other metallic elements (e.g., Ca, Al, Sr, Mg...) present in the solid residue (Ra).

[0088] Step (c) of acid leaching is catalyzed by sonication, typically in an ultrasonic bath, using a probe sonicator or a sonotrode sonicator. Sonication typically enhances the solubilization of the TRs by the acid solution. For example, step (c) is catalyzed by an ultrasonic bath at a frequency of 45 kHz, or by a probe sonicator at various frequencies.

[0089] The acid leaching step (c) can be catalyzed by sonication for a period of 6 to 36 hours, for example for 24 hours.

[0090] Step (c) of acid leaching is typically carried out at a temperature ranging from 25 to 80°C, or from 40 to 70°C or from 55 to 65°C. Generally, step (c) is carried out at atmospheric pressure.

[0091] The strong acid solution(s) used in step (c) may have a mass concentration of acid ranging from 5 to 65 wt%. The strong acid solution(s) and the residue (Ra) may be present in an acid: residue (Ra) (volume of acid in L) / mass of residue Ra in kg ratio ranging from 15:1 to 1:1 and typically from 8:1 to 2:1.

[0092] In particular, the acid solution used in step (c) can be an aqueous solution of nitric acid with a concentration ranging from 10 to 65 wt%, and typically from 50 to 65 wt%. The citric acid solution and the residue (Ra) can be present in a citric acid:residue (Ra) ratio (volume of acid in L / mass of residue Ra in kg) ranging from 15:2 to 1:1 or from 4:1 to 2:1. For example, the nitric acid:residue (Ra) ratio (volume of acid in L / mass of residue Ra in kg) can be 3:1 with nitric acid at 65 wt%.

[0093] In particular, the acid solution used in step (c) can be an aqueous hydrochloric acid solution with a concentration ranging from 10 to 37 wt%, and commonly from 30 to 37 wt%. The hydrochloric acid solution and the residue (Ra) can be present in a hydrochloric acid:residue (Ra) ratio (volume of acid in L / mass of residue Ra in kg) ranging from 15:1 to 1:1 or from 4:1 to 2:1. For example, the hydrochloric acid:residue (Ra) ratio (volume of acid in L / mass of residue Ra in kg) can be 3:1 with 37 wt% hydrochloric acid.

[0094] In particular, the acid solution used in step (c) can be an aqueous solution of sulfuric acid with a concentration ranging from 5 to 30 wt%, and commonly from 10 to 15 wt%. The sulfuric acid solution and the residue (Ra) can be present in a sulfuric acid:residue (Ra) ratio (volume of acid in L / mass of residue Ra in kg) ranging from 8:1 to 3:1 or from 8:1 to 5:1. For example, the sulfuric acid:residue (Ra) ratio (volume of acid in L / mass of residue Ra in kg) can be 5:1 with 10 wt% sulfuric acid.

[0095] An increase in the acid-to-residue ratio (Ra) (volume of acid in L / mass of residue Ra in kg) above 3:1 slightly increases the extraction rate of TR elements in the leaching solution, but leads to a decrease in the concentration of TR elements in the leachate due to the increased volume of acid used, thus hindering their subsequent processing. A decrease in the acid-to-residue ratio (Ra) below 3:1 results in a decrease in the extraction rate of TR elements in the leaching solution, and furthermore, makes it difficult to separate the leachate, which is now enriched in TR elements, from the depleted residue.

[0096] Decreasing the concentration of nitric and hydrochloric acid leads to a decrease in the extraction rate. However, decreasing the concentration of sulfuric acid increases the extraction rate, partly due to the recrystallization effect of rare earth elements dissolved in concentrated sulfuric acid solutions.

[0097] Separation step (d)

[0098] The process according to the invention includes a separation step (d) leading to the obtaining of a liquid filtrate containing the acid(s) and rare earths (Fd) and a solid residue (Rd).

[0099] Step (d) of separation can be carried out by centrifugation and / or by filtration, for example using a crucible.

[0100] At the end of step (d) an acidic filtrate (Fd) is obtained containing strong acid and TRs. Optionally, the filtrate (Fd) may contain chloride ions and / or nitrate ions and / or metallic impurities based for example on iron and / or aluminum.

[0101] At the end of step (d), an acidic residue (Rd) is obtained containing less than 10% by weight of TRs compared to the amount of TRs initially contained in the phosphogypsum. Typically, the residue (Rd) contains less than 40% by weight of mineral impurities compared to the amount of mineral impurities contained in the solid residue (Ra) from step (b).

[0102] Advantageously, but optionally, a second acid leaching step can be carried out using the filtrate from the previous step, containing the strong acid enriched in TRs, to treat the residues resulting from the basic leaching step of the phosphogypsum. In this step, the nitric or hydrochloric acid solution from the first acid leaching is thus recycled.

[0103] In a preferred embodiment, part or all of the liquid filtrate (Fd) is sent to step (c) for use as a strong acid solution. Thus, the liquid filtrate (Fd) can constitute part or all of the strong acid solution. Typically, the filtrate (Fd) and the residue (Ra) have a filtrate (Fd):residue (Ra) ratio (filtrate volume in L / acid mass in kg) ranging from 1:1 to 8:1, or from 2:1 to 7:1, for example, 6:1. Typically, this embodiment can be used in a continuous or batch process. In a batch process, the process is repeated at least "n" times so as to use the liquid filtrate (Fd) obtained in implementation "n-1" in step (c) of implementation "n".

[0104] Advantageously, this embodiment increases the extraction rate of TRs from the residue (Ra) while reducing the volume of acid solution used. Thus, this embodiment increases the extraction efficiency of the process while simplifying and reducing the size of the equipment required to implement the process according to the invention.

[0105] Thus the process according to the invention allows obtaining an acid filtrate (Fd) with a concentration in TRs higher than the concentration of phosphogypsum, typically ranging from 668 to 1027 ppm or even from 817 to 1197 ppm or even ranging from 636 to 1197 ppm.

[0106] List of figures

[0107] Figure 1: Illustrative diagram of an embodiment of the extraction process according to the invention

[0108] Figure 2: X-ray diffractogram of phosphogypsum (a) and residue (Ra)

[0109] Figure 3: SEM image of phosphogypsum (A, B) and residue R3 (C, D)

[0110] EXAMPLES The following non-restrictive examples illustrate examples of implementation of the invention.

[0111] Materials and methods

[0112] The compositions of phosphogypsum and residue are determined by XRD.

[0113] The phosphogypsum used in the examples below comes from the phosphoric workshop, phosphate processing unit Jorf Lasfar / Morocco.

[0114] In the examples below, TRs denotes the sum of rare earth elements in ppm. OTRs denotes the sum of rare earth oxides in ppm. rpm means rotations per minute.

[0115] Example 1: Enrichment of phosphogypsum and preparation of a rare-earth-rich nitric acid filtrate

[0116] In this example, the extraction of TRs contained in phosphogypsum is carried out according to a procedure consisting of dissolving 230 g of phosphogypsum in 4 liters of a 0.4 M EDTA.Na4 solution (pH equal to 12.6), under mechanical stirring for 15 min at atmospheric pressure and ambient temperature (approximately 25°C).

[0117] The solid-liquid mixture (Ra / Fa) obtained after basic leaching was subjected to a separation step by centrifugation for 15 min at a rotation speed of 3000 rpm. Inductively coupled plasma spectrometry (ICP) results showed that the rare earth elements were concentrated. The residue (Ra) had an enrichment level ranging from 76.9% to 78.9% relative to phosphogypsum.

[0118] The results of the analysis of a phosphogypsum and their residues obtained are detailed in Table 1.

[0119]

[0120] The resulting residue (Ra) is washed with distilled water to remove traces of chelating agent and sodium sulfate. The washed residue (Ra) is then dried at a temperature ranging from 60°C to 80°C, and subsequently ground and sieved to obtain residue (Ra) in the form of particles with a size less than or equal to 250 µm.

[0121] The residue obtained (Ra) is in the form of a dark brown powder.

[0122] The residue obtained (Ra) is 9 to 11 times more concentrated in TRs than the starting phosphogypsum.

[0123] The mineralogical composition of phosphogypsum and washed residue (Ra) (Table 2) shows that the increase in the concentration of TRs in the residues is accompanied by an increase in the concentration of silicates, fluorides and phosphate pentoxide, and a decrease in the concentration of sulfates.

[0124] Table 2: Mineralogical composition of different phosphogypsum samples and their residues obtained (Ra)

[0125] * Elements that do not undergo enrichment

[0126] The composition of phosphogypsum and residue is determined by X-ray diffractometry (XRD).

[0127] The results in Figure 2 show that the washed residue (Ra) is a mixture of 2 phases (SiO2 and CaF2) corresponding to a different matrix from the initial phosphogypsum matrix which consists mainly of calcium sulfate dihydrate (CaSO4.2H2O) and silicon dioxide (SiO2).

[0128] Scanning electron microscopy (SEM) analysis (Figure 3) of phosphogypsum shows that the phosphogypsum crystals exhibited acicular, tabular, and small crystalline cluster morphology. However, analysis of the residue confirms the solubilization of phosphogypsum due to the disappearance of the phosphogypsum crystals.

[0129] Acid leaching of the residue (Ra) with a nitric acid solution is carried out by contacting 4 g of washed residue (Ra) with 12 ml of 65% nitric acid (by weight) in an acid-to-residue ratio of 3:1 (acid volume in L / residue mass in kg Ra). The mixture is stirred and catalyzed by an ultrasonic bath at a frequency of 45 kHz, at various intervals, preferably over 24 hours at a temperature of 55°C.

[0130] The operating acoustic power of the ultrasonic bath varies from 1 to 9. A value of 9 corresponds to the maximum power (100%), and a value of 1 corresponds to 40% of the maximum power. This work was carried out with a power value of 3. The resulting filtrate (Fa) is characterized by a dark orange color.

[0131] After the separation step, the filtrate (Fd) and the residue (Rd) were analyzed by ICP to determine the leaching efficiency and the rare earth element (REE) concentration. The results show that the filtrate (Fd) is rich in REEs, with a concentration exceeding 800 ppm. The leaching efficiency was 94%. Table 3 shows the distribution of rare earth elements in the filtrate. nitric (Fd)

[0132] The filtrate (Fd) is used as a strong acid solution for acid leaching of a new batch of residue (Ra) using a cascade procedure to decrease the amount of acid while increasing the concentration of TRs in the filtrate (Fd). The protocol involves treating a batch of residue (Ra) with filtrate (Fd) at a filtrate-to-residue ratio (filtrate volume in L / residue mass in kg) of 6:1 in an ultrasonic bath set to power level 3, temperature of 55°C, and a variable duration, preferably 24 hours. The TR analysis results are shown in Table 4.

[0133] Table 4: Composition in TRs elements and OTR oxide (ppm) of the residue (Ra), filtrate (Fd) and cascade filtrate (Fd) as well as the overall leaching yield with nitric acid

[0134] The ICP analysis results (Table 4) show that the use of the filtrate (Fd) as a strong acid solution in a so-called cascade mode advantageously increases the TRs content while maintaining an overall leaching yield of 95.4% in the presence of nitric acid.

[0135] Advantageously, the use of the filtrate (Fd) as a strong acid solution in a so-called cascade mode advantageously allows increasing the content of TRs in the filtrate and decreasing the amount of strong acid solution used and to be treated.

[0136] Example 2: Enrichment of phosphogypsum and preparation of a rare earth rich hydrochloric acid filtrate In this example, the Ra residue is obtained as in example 1.

[0137] Acid leaching is carried out with a hydrochloric acid solution as the leaching agent.

[0138] Acid leaching of the residue (Ra) is carried out with 37% hydrochloric acid by weight, ensuring that the volume of hydrochloric acid is three times the mass of the residue, corresponding to an acid-to-residue ratio (volume of acid in L / mass of residue in kg) of 3:1. The mixture is stirred in an ultrasonic bath at power level 3 for a variable time, preferably 24 hours, at a temperature of 55°C. A sonotrode sonicator can also be used to reduce reaction times. The resulting filtrate (Fd) is clear yellow, with a rare earth concentration of approximately 1000 ppm. The leaching yield is 96.2% (±2%) in the presence of hydrochloric acid. hydrochloric (Fd) The filtrate (Fd) is used as a strong acid solution for acid leaching of a new batch of residue (Ra) according to a cascade operating procedure to decrease the amount of acid while increasing the concentration of TRs in the filtrate (Fd).

[0139] The protocol implemented consists of treating a batch of residue (Ra) with a filtrate (Fd) at a filtrate-to-residue ratio (acid volume in L / residue mass in kg) of 6:1 in an ultrasonic bath set at power level 3, a temperature of 55°C, and a variable duration, preferably 24 hours. The results of the TR analyses are shown in Table 6. A sonotrode sonicator can also be used to reduce reaction times. The ICP analysis results (Table 6) show that using the filtrate (Fd) as a strong acid solution in a cascade mode advantageously increases the TR content while maintaining an overall leaching efficiency of 95.0% in the presence of hydrochloric acid.

[0140] (F3) and filtrate (F4) as well as the overall yield of leaching with hydrochloric acid

[0141] Example 3: Enrichment of phosphogypsum and preparation of a rare-earth-rich sulfuric acid filtrate

[0142] In this example, the residual Ra is obtained as in example 1.

[0143] Acid leaching is carried out using a sulfuric acid solution as the leaching agent.

[0144] Acid leaching of the residue (Ra) is carried out with 10% sulfuric acid by weight at an acid-to-residue ratio (volume of acid in L / mass of residue in kg) of 5:1. The mixture is activated by an ultrasonic bath, power level 3, for a variable time, preferably 24 hours, at a temperature of 55°C. A sonotrode sonicator can also be used to reduce reaction times.

[0145] The resulting Fd filtrate is clear yellow in color, with a rare earth concentration greater than 600 ppm (Table 7). The extraction yield by the process according to the invention is 79.0% (±3%) in the presence of sulfuric acid.

[0146] Table 7: Composition in TRs elements and OTR oxide (ppm) of the residue (Ra) and sulfuric filtrate (Fd)

Claims

DEMANDS 1. A process for extracting rare earth elements by processing phosphogypsum, the process comprising the following steps: (a) basic leaching of phosphogypsum by a basic aqueous solution comprising a chelating agent, leading to the formation of a liquid filtrate (Fa) and a solid residue (Ra); (b) separation of the liquid filtrate (Fa) and the solid residue (Ra); (c) acid leaching of the solid residue (Ra) obtained in step (b) by an acidic aqueous solution leading to the formation of a liquid filtrate comprising rare earths (Fc) and a solid residue (Rc), the acid leaching being catalyzed by sonication; (d) separation of the liquid filtrate (Fd) and the solid residue (Rd).

2. Extraction process according to claim 1, wherein the stoichiometry between the chelating agent and phosphogypsum varies from 1 to 2, preferably from 1.1 to 1.

4.

3. Extraction process according to claim 1 or 2, wherein the chelating agent is ethylenediaminetetraacetic acid (EDTA) or one of its salts.

4. Extraction process according to any one of claims 1 to 3, wherein step (b) is carried out by solid / liquid separation, preferably by filtration, centrifugation, and / or decantation.

5. Extraction method according to any one of claims 1 to 4, wherein step (c) is catalyzed by sonication in an ultrasonic bath or by means of a probe sonicator.

6. Extraction method of any one of claims 1 to 5, wherein step (c) is catalyzed by sonication for a period of 6 to 36 hours, preferably for 24 hours.

7. Extraction process according to any one of claims 1 to 6, wherein the strong acid(s) solution has a mass concentration of acid from 5 to 65 wt%.

8. An extraction process according to any one of claims 1 to 7, wherein the strong acid solution(s) and the residue (Ra) are present in an acid:residue (Ra) ratio (volume of acid / mass of residue Ra) ranging from 15:1 to 1:

1.

9. An extraction process according to any one of claims 1 to 8, wherein step (d) is carried out by centrifugation and / or filtration, preferably using a crucible.

10. Extraction process according to any one of claims 1 to 9, wherein part or all of the liquid filtrate (Fd) is sent to step (c) for use as a strong acid solution.

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