Method for recovering rare earths and coproducts from phosphogypsum and phosphoric acid sludge

Abstract

The present disclosure relates to a method for treating phosphogypsum or phosphoric acid sludge in order to recover rare earths, the method comprising the following consecutive steps: (a) treating the phosphogypsum or the phosphoric acid sludge according to the following steps (a1) or (a2), respectively: (a1) washing the phosphogypsum with a polar solvent and collecting the obtained solid residue; or (a2) separating the phosphoric acid sludge into a liquid phase and a solid residue, and separately collecting the obtained solid residue and the obtained liquid phase; washing the solid residue obtained at the end of the treatment step with a polar solvent at room temperature and separately collecting the obtained solid residue and the obtained liquid phase; and washing the solid residue obtained at the end of the washing step with a hot polar solvent and collecting the obtained solid residue and the obtained liquid phase; (b) drying, grinding and sieving the solid residue obtained at the end of step (a); (c) alkaline leaching the solid residue obtained at the end of step (b) using a strong inorganic base, and collecting the obtained solid residue and the obtained liquid phase; (d) drying, grinding and sieving the solid residue obtained at the end of step (c); (e) acid leaching the solid residue obtained at the end of step (d) using nitric acid and collecting the obtained solid residue and the obtained liquid phase; (f) precipitating calcium nitrate crystals from the liquid phase obtained at the end of step (e) and collecting the solid residue and the obtained liquid phase; (g) adjusting the pH of the liquid phase obtained at the end of step (f) to a value of between 1 to 1.8 by adding a strong base and collecting the obtained solid residue and the obtained liquid phase; (h) drying the solid residue obtained at the end of step (g) to obtain a solid product containing rare earths.

Description

Process for recovering rare earths and co-products from phosphogypsum and phosphoric sludge TECHNICAL FIELD The invention relates to a process for treating phosphogypsum and phosphoric sludge to recover rare earth elements. More specifically, the present invention relates to a process enabling the comprehensive recovery of various high-value by-products present in phosphogypsum or phosphoric sludge, aiming for zero waste. STATE OF THE ART Rare earth elements (REEs) constitute a set of 17 metallic elements, i.e., lanthanides plus yttrium and scandium. Rare earths and their alloys are of particular interest because their physico-chemical properties make them materials of choice in various applications such as optics (lighting, radiography), mechanics due to their high hardness, or magnetism (manufacture of permanent magnets). Given the increasing demand for TRs, sustainable TR supply has become an important economic and strategic issue. Thus, to address the overexploitation of primary resources, it becomes urgent to substitute the recovery of TRs from primary ores with their recovery from secondary ores or industrial by-products, especially in countries where resources are dwindling or non-existent. Currently, the development of a clean, sustainable, economically viable, and green chemistry-compliant process for recovering trace elements from industrial by-products is of critical urgency. Examples of potentially exploitable industrial by-products include phosphate by-products such as phosphogypsum (PGs) and phosphoric sludge (BPs). Managing by-products from the phosphate industry is a major environmental and economic challenge for the coming years. Indeed, natural phosphate processing industries generate by-products containing high-value elements, such as triglycerides, phosphorus (P2O5), and fluorides, which could be exploited at a lower cost while preserving natural resources. In this context, phosphoric sludge (PS) and phosphogypsum (PG) represent an essential alternative to meet the intensive market demand, particularly in TRs. PG is the by-product resulting from the wet process of manufacturing phosphoric acid 29% P2O5 by attacking natural calcium phosphates, apatites, with concentrated sulfuric acid, according to reaction (1). Cai0F2(PO4)6 + 10H2SO4+ 10nH2O - 6H3PO4+ 10CaSO4.nH2O + 2HF (1) In general, PG is presented as a fine, moist and aggregated powder, consisting mainly of a dihydrate gypsum phase CaSO4.2H2O. BPs are the precipitated by-product resulting from the evaporation / concentration of phosphoric acid from 29% to 54% P2O5

[0001] . Phosphoric acid (PA) is a pasty slurry with two phases: an anhydrous gypsum solid phase (CaSO4) and a 54% phosphoric acid liquid phase (P2O5). Silicates, triglycerides, aluminum, iron, fluorine, etc., are minor phases present in PA at relatively high concentrations compared to phosphoric acid (PG). It is known that phosphate ores contain varying levels of valuable elements, i.e., TRs, fluorides, silicates, etc. In most cases, these elements are concentrated in the PG and BPs. In the case of the dihydrate process, 55 to 66% of the TRs co-precipitate with the PG during the attack of the phosphates by concentrated sulfuric acid [2]. The concentration of TRs in the PGs and BPs, resulting from the transformation of sedimentary phosphate ores, is less than 0.1 and 0.22%, respectively [3], [4]. Numerous processes have been developed to valorize the gypsum phase for subsequent use in the construction industry and, to a lesser extent, to recover the TRs [5]. Among the processes described in the literature, document CN1 11560521 A describes an acid leaching process of gypsum polypropylene (GP) using a solution containing 20 wt% sulfuric acid, in the presence of an additive: aminotrimethylenephosphonic acid at 0.002 wt%, hydroxyethylidene diphosphonic acid at 0.0015 wt%, or polyethylene glycol 400 at 0.006 wt%. The GP was treated with a liquid-to-solid ratio of 10⁻¹ at 100°C, at 50 revolutions per minute (rpm) for a reaction time of 12 h. After separation, the solid phase is washed with boiling water. This process allows the production of alpha-hemihydrate gypsum with a TRs leaching rate greater than 90% [6]. Application CN1 12408452A describes a process for extracting TRs combined with purifying PG. High TR leaching was achieved by treating the PG with a solution composed of: 4 wt% nitric acid, 25 wt% glycerol-butanol (glycerol-to-butanol mass ratio 3:1), 8 wt% potassium nitrate, and 2.5 × 10⁻² mol / kg of an additive consisting of maleic acid, sodium maleate, and Al(NO)₃ in a molar ratio of 2:8:1. The resulting mixture of the prepared solution and the PG (liquid-to-solid ratio 10⁻¹) is poured into a three-necked round-bottom flask and heated in an oil bath at 95 °C with a stirring speed of 50 rpm for 6 h. After filtration, the phase The solid is washed twice, first with boiling water and then with acetone. The rare earth leaching rate reaches 97.3% and the resulting gypsum is of high strength [7]. Application WO 2020 / 067856 A1 demonstrates that acid leaching of a PG pre-washed with a potassium sulfate solution, using 2N nitric acid for 6 hours, in the presence of a potassium chlorate oxidizing agent to promote the dissolution of TRs, and a zinc powder reducing agent to minimize the leaching of impurities, yields a TRs leachate with a concentration of 179 ppm. The PG:acid weight ratio is 4:6, and the amount of potassium chlorate and zinc is 10 g / kg and 5 g / kg, respectively, relative to the leaching mixture. Post-treatment of the leachate by evaporation promotes an increase in the TRs concentration to 260 ppm [4]. Besides acid leaching, other purification techniques can be implemented. Magnetic separation also allows for high PG purification. According to CN1 10918251 A, after impurity separation, the purified PG is centrifuged to remove residual water. The main elements removed are: phosphorus impurities (95-97%), fluorine impurities (82-92%), organic impurities (80-90%), heavy metal oxides, and radioactive metal impurities (70-87%). The TR recovery rate varies between 70 and 83% [8]. Other processes describe the purification of PGs using basic solutions. The use of a 25% ammonium sulfate solution at a temperature above 92°C for a duration ranging from 10 to 20 min and a solid-liquid ratio ranging from 1-5 to 1-20 is thus described [9]. Application RU2739409C1

[0010] describes a leaching-precipitation process for TRs. Initially, 100g of PG is leached for 6 h in 193.7 g of ammonium carbonate at a temperature of 40-55°C, according to the following reaction: CaSO4 + (NH4)2CO3— > CaCO3 + (NH4)2SO4 (2) The second step consists of dissolving 65 g of the calcium carbonate obtained in 58% nitric acid at a stoichiometry of 120, 130, or 150%. After filtration to remove insoluble compounds, the leaching solution is cooled to between -10 and -15 °C to crystallize the calcium nitrate tetrahydrate. After separation of the calcium nitrate, the leachate is neutralized with lime, ammonia, or milk of lime to a pH ranging from 0.7 to 1.3. The resulting precipitate contains 16% TRs oxide (RTs)

[0010] . It is also known from patent RU 2258036 C1

[0011] , that the treatment of PG with sodium carbonate of concentration 2.0-2.5mol / L at 60-80°C for 30-45 min, with a liquid-solid ratio of 2.0-2.5, promotes the precipitation of TR-rich calcium carbonate. The precipitated calcium carbonates are then transformed into calcium oxide (CaO) by calcination at 900-950°C. The calcium oxide then undergoes solubilization with ammonium chloride to produce calcium chlorides and an insoluble precipitate rich in TRs. The latter is leached with a mixture of 5-6% hydrochloric acid and ascorbic acid at 80-90°C for 30-60 minutes, the weight ratio of ascorbic acid to TR elements being (0.4-0.5):1. The resulting solid phase contains 30-32% strontium, which is recovered as a commercial product. The leachate, neutralized with ammonia to a pH of 9.0-9.5, precipitates the TRs. The precipitate undergoes post-treatment with a sodium sulfate solution at pH 0.3-0.5 at 80-90°C for 60-90 minutes to produce a mixed precipitate containing calcium sulfate and TRs. Generally, the conversion of PGs with sodium carbonates, ammonium carbonates, etc., leads to the generation of carbon dioxide in subsequent TR recovery stages, which is a major limitation of these processes. Hence the need to develop a greener, more environmentally friendly conversion technique. It is also known from EP 0419318 A1

[0012] that solubilizing PG with a saline solution containing 25 g / L of sodium chloride, followed by pretreatment of the insoluble residues rich in TRs with a sodium bicarbonate solution and then concentrated nitric acid, promotes the dissolution of TRs. However, this process consumes water; for example, treating one (1) kg of PG requires 200 liters of water, and results in the discharge of large quantities of saline water containing fluorides, sulfates, chlorides, etc., into the environment. Another process relies on leaching the BPs with nitric acid of concentration 7.2 mol / L, preheated to 72 °C, for 1 h at an acid-BP ratio of 4:1 under stirring at 400 rpm. This technique allows the leaching of 58.1% of the TRs

[0013] . However, prior art processes are mainly focused on the valorization of the gypsum phase and on the extraction of TRs and do not seek to individually isolate other valuable elements in order to then valorize them. Furthermore, none of the documents presented deal with the valuation of valuable elements (e.g. fluoride, silica, etc.) in BPs. From an industrial point of view, there remains a need for the provision of a process for the extraction and valorization of rare earths but also of other valuable elements such as phosphorus, fluorine, silicon, calcium or sulfur contained in PG and BPs, which is as environmentally friendly as possible. BRIEF DESCRIPTION OF THE INVENTION The present invention relates to a process for treating phosphogypsum or phosphoric sludge to recover rare earth elements, the process comprising the following successive steps: (a) treatment of phosphogypsum or phosphoric sludge respectively according to the following steps (a1) or (a2): (a1) washing the phosphogypsum with a polar solvent and collecting the resulting solid residue; or (a2) separation of phosphoric sludge into a liquid phase and a solid residue and separate collection of the solid residue obtained and the liquid phase obtained; washing of the solid residue obtained after treatment with a polar solvent at ambient temperature and separate collection of the solid residue obtained and the liquid phase obtained; and washing of the solid residue obtained after washing with a hot polar solvent and collection of the solid residue obtained and the liquid phase obtained; (b) drying, grinding and sieving of the solid residue obtained at the end of step (a); (c) basic leaching of the solid residue obtained at the end of step (b) with a strong mineral base, and collection of the solid residue obtained and of the liquid phase obtained; (d) drying, grinding and sieving of the solid residue obtained at the end of step (c); (e) acid leaching of the solid residue obtained at the end of step (d) with nitric acid and collection of the solid residue obtained and of the liquid phase obtained; (f) precipitation of calcium nitrate crystals from the liquid phase obtained at the end of step (e) and collection of the solid residue and the liquid phase obtained; (g) adjusting the pH of the liquid phase obtained at the end of step (f) to a value ranging from 1 to 1.8 by adding a strong base and collecting the solid residue obtained and the liquid phase obtained; (h) drying of the solid residue obtained at the end of step (g) to obtain a solid product containing rare earths. Other aspects of the invention are as described below and in the claims. DESCRIPTION OF FIGURES Figure 1 represents a synoptic diagram of an example embodiment of the invention. Figure 2 represents the evolution of the concentration of P2O5 in the filtrates (noted F1 to F6), i.e. the liquid phases, as a function of the number of cold washes during step (a) according to example 1. Figure 3 represents the X-ray diffractogram of the solid residue (noted B2) obtained after the cold washes and of the solid residue (noted B4) obtained after step (b) according to example 1. Figure 4 represents the X-ray diffractogram of the solid residue (denoted R2) obtained at the end of step (d) according to example 1. Figure 5 represents the X-ray diffractogram of the precipitated sodium fluorosilicate (noted C1) from the cooling of the liquid phase obtained at the end of step (b) according to example 1. Figure 6 represents the X-ray diffractogram of sodium fluorosilicate (noted C2) from the treatment of the liquid phase obtained at the end of step (a) cooled and separated from the precipitated crystals and treated with a NaCl solution (A) or a Na2CO3 solution (B). Figure 7 represents the X-ray diffractogram of sodium sulfate (noted C3) obtained by neutralizing the liquid phase obtained at the end of step (c) and then drying it and grinding the solid residue. Definitions The term "rare earths" 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. The term "ambient temperature" refers, for the purposes of this invention, to a temperature within the range of 20°C to 25°C. When a step of the process according to the invention is carried out "cold," it should be understood that it is carried out at ambient temperature as defined above. The two terms are interchangeable. The term "hot wash" refers to a washing operation at a temperature above ambient temperature as defined above and below boiling point (T eb in °C) of the polar solvent used for the washing operation. Advantageously, the temperature of the hot polar solvent is (T eb -40°C) at (T eb -10°C), preferably of (T eb -20°C) at (T eb -10°C). The term "phosphogypsum" refers to a by-product of the wet process of phosphoric acid production, that is, obtained by reacting fluorophosphated calcium ores with sulfuric acid. This type of phosphogypsum is also called LREC phosphogypsum, from the acronym "Low Rare Earth Case." More 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. The term "phosphoric sludge" refers to the by-product obtained during the concentration stage of wet-process phosphoric acid, that is, phosphoric acid obtained by reacting fluorophosphate calcium ores with sulfuric acid. The concentration stage is preferably carried out to obtain phosphoric acid with a mass fraction ranging from 50% to 54% P₂O₅. Phosphoric sludge is also referred to as HREC phosphogypsum, an acronym for "High Rare Earth Case." More generally, phosphoric sludge comprises 22% to 30% CaO, 25% to 35% SO₃, less than 2% SiO₂, 25% to 45% P₂O₅, less than 2% F, and 0.1% to 0.4% rare earth elements (percentages expressed by weight). The term "leaching" in the context of the present invention means the extraction of a soluble compound from a solid residue by means of washing operations. When the washing is carried out using a basic 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 solution. The term "liquid phase" in the context of the present invention means a solution isolated from a solid residue by any step of separation of the liquid phase and the solid residue. Unless otherwise stated, the ratios indicated in this application are mass ratios and the percentages are mass percentages. DETAILED DESCRIPTION OF THE INVENTION The inventors have thus developed a process for recovering rare earth elements from phosphogypsum or phosphoric sludge. In addition to rare earth elements, the process according to the invention allows for the recovery and valorization of phosphorus, fluorine, silicon, calcium, and sulfur (collectively referred to as co-products) contained in PG and BPs. The proposed process is simple to implement and allows for the valorization of co-products while being environmentally sound. The invention relates to a process for treating phosphogypsum or phosphoric sludge to recover rare earth elements. The process comprises the following successive steps: (a) treatment of phosphogypsum or phosphoric sludge respectively according to the following steps (a1) or (a2): (a1) washing the phosphogypsum with a polar solvent and collecting the resulting solid residue; or (a2) separation of phosphoric sludge into a liquid phase and a solid residue and separate collection of the solid residue obtained and the liquid phase obtained; washing of the solid residue obtained after separation with a polar solvent at ambient temperature and separate collection of the solid residue obtained and the liquid phase obtained; and washing of the solid residue obtained after washing with a hot polar solvent and collection of the solid residue obtained and the liquid phase obtained; (b) drying, grinding and sieving of the washed solid residue obtained at the end of step (a); (c) basic leaching of the ground solid residue obtained at the end of step (b) with a strong mineral base, and separate collection of the solid residue obtained and of the liquid phase obtained; (d) drying, grinding and sieving of the solid residue obtained at the end of step (c); (e) acid leaching of the solid residue obtained at the end of step (d) with nitric acid and separate collection of the solid residue obtained and of the liquid phase obtained; (f) precipitation of calcium nitrate crystals from the liquid phase obtained at the end of step (e) and separate collection of the solid residue and the liquid phase obtained; (g) adjusting the pH of the liquid phase obtained at the end of step (f) to a value ranging from 1 to 1.8 by adding a strong base and collecting the solid residue obtained and the liquid phase obtained; (h) drying of the solid residue obtained at the end of step (g) to obtain a solid product containing rare earths. The steps of the process of the invention can be as described in detail below. The process of the present invention advantageously allows for the isolation and thus the valorization other elements contained in phosphopypse and BPs when one or more of the optional steps described below are implemented (e.g. phosphorus pentoxide, sodium fluorosilicate, calcium nitrate, sodium or ammonium nitrate, sodium sulfate). The process can be implemented using phosphogypsum or BPs. Steps (a) and (b) allow the content of soluble impurities in phosphoric sludge or phosphogypsum to be reduced. When the process is implemented using phosphogypsum, step (a) is a phosphogypsum treatment step involving washing the phosphogypsum with a polar solvent (step (a1)). The phosphogypsum washing step removes impurities from the wet phosphoric acid production process, such as organic matter, traces of P2O5, and macroscopic impurities, thus preventing their enrichment in subsequent stages of the process. This step also concentrates the TRs (transient residues). The polar solvent is preferably water, methanol, ethanol, acetone, an ether or a mixture thereof. When the polar solvent is an ether, it can be a petroleum ether, diethyl ether, or diisopropyl ether. Advantageously, the polar solvent is water. The polar solvent can be used at room temperature or hot, preferably at room temperature which is more economically advantageous. The washing of phosphogypsum is carried out, as an example, under low agitation for 5 min, the solid and the polar solvent being mixed in a solid:liquid mass ratio of 1:1. The phosphogypsum washing operation can be carried out in cascade, that is to say that the solid obtained at the end of a washing operation is then isolated and undergoes a new washing operation with the polar solvent. The solid-to-liquid mass ratio can vary from 1:1 to 1:5 for these subsequent washing operations. The washing time is generally 1 to 15 minutes, preferably 5 to 10 minutes. The phosphogypsum washing operation can thus be carried out several times until the liquid phase separated from the solid residue no longer contains P2O5 or only traces of it. The measurement of the liquid phase conductivity is indicative of the presence of P2O5. Advantageously, the phosphogypsum washing operation is thus carried out until the The conductivity value of the liquid phase is stabilized at a value of 2.35 mS / cm. Advantageously, the washing operation is carried out twice. Washing allows the residual phosphoric acid to be recovered from the isolated liquid phases and removed. When the process is implemented using phosphoric sludge, step (a) is a phosphoric sludge treatment step (a2) comprising: - the separation of phosphoric sludge into a liquid phase and a solid residue, and separate collection of the resulting solid residue and liquid phase, - washing the solid residue obtained after separation with a polar solvent at room temperature and separate collection of the solid residue and the resulting liquid phase; and - washing the solid residue obtained after washing with a hot polar solvent and collecting the solid residue obtained and the liquid phase obtained. Several separation techniques can be used to separate phosphoric sludge into a liquid phase (phosphoric acid with a mass fraction ranging from 45% to 54% P2O5) and a solid residue. Separation can be achieved by sedimentation, vacuum filtration, or centrifugation, particularly centrifugation. The solid residue obtained includes calcium sulfate (CaSC₂), sodium fluorosilicate (NasSiFe), residual P₂O₅, TRs and other metallic elements (see Table 1 in the experimental section). The resulting residue is then washed at room temperature using a polar solvent. Washing allows the recovery of residual P2O5 and the removal of impurities in order to concentrate the TRs. The polar solvent used for washing the solid residue can be either a protic or aprotic polar solvent. The preferred polar solvent is water, methanol, acetone, ethers, or a mixture thereof. Generally, the polar solvent is water. The contact time during washing at room temperature, preferably carried out under agitation, generally varies from 1 to 15 min, preferably from 5 to 10 min. After washing with the polar solvent at room temperature, the resulting solid residue and liquid phase are collected separately. "Collected separately" means that the solid residue and liquid phase are separated and then individually collected. This separation can be achieved by vacuum filtration. Washing at room temperature can be a cascade wash, meaning that several successive washes can be performed, for example 2, 3, 4, 5, or 6 washes. These successive washes can be carried out with the same polar solvent. Advantageously, the washes are carried out with increasing liquid / solid mass ratios, typically ranging from 1:1 to 5:1 or 10:1. This optimizes the recovery of P2O5. The washing operation can be repeated until the liquid phase separated from the solid residue no longer contains P2O5 or only in trace amounts, i.e. at a content of less than 0.52%. The solid residue thus obtained after washing at room temperature with the polar solvent includes calcium sulfate, sodium fluorosilicate, TRs and other metallic elements. After washing at room temperature, the solid residue is washed with a hot polar solvent. Since the solubility of sodium fluorosilicate is increased in the hot polar solvent, the hot washing step reduces the sodium fluorosilicate (Na2SiFe) content in the solid residue. The polar solvent used may be the same as that used for room temperature washing or it may be different. Typically, the polar solvent is water. In some embodiments, both room temperature and hot washing are performed with water. Typically, hot washing is carried out at a temperature ranging from 60°C to 100°C, preferably from 80°C to 90°C when the polar solvent is water. Hot washing with polar solvent is typically carried out with a liquid / solid mass ratio ranging from 5:1 to 15:1, preferably from 10:1 to 15:1. The contact time during hot washing, preferably carried out under agitation, generally varies from 5 to 20 min, preferably from 10 to 15 min. The separation of the solid residue from the liquid phase can be achieved by vacuum filtration. The solid residue obtained after hot washing includes calcium sulfate, TRs, and other metallic elements (see Table 1 in the experimental section). It has a reduced sodium fluorosilicate content. Fluorine and silicon are recovered in the liquid phase, mainly in the form of sodium fluorosilicate. Step (b) is a drying, grinding and sieving step of the washed solid residue obtained at the end of step (a), whether the latter is carried out from phosphogypsum or phosphoric sludge. The washed solid residue is typically dried at a temperature ranging from 60 to 80°C. Grinding reduces the particle size of the residue to facilitate residue treatment in step (c) of the process. The ground solid residue is typically sieved through a 250 µm mesh, and the fraction smaller than 250 µm is collected. The ground and sieved residue typically consists of particles with a size less than or equal to 250 µm, preferably ranging from 100 to 250 µm. Optional step for processing the liquid phase obtained in step (a) When the process is implemented from phosphoric sludge, the process may include a step of crystallization / precipitation of fluorosilicate salts, mainly in the form of sodium fluorosilicate, contained in the liquid phase obtained at the end of step (a) and collection of the salts obtained. The crystallization / precipitation of fluorosilicate salts can be obtained by cooling to room temperature the liquid phase obtained at the end of step (a). Sodium fluorosilicate crystals precipitated after cooling to room temperature can be isolated from the liquid phase. One or more sodium salts, preferably sodium chloride (NaCl) or sodium carbonate (NasCOs), particularly sodium chloride, can be added to the liquid phase (filtrate) saturated with fluorine, silicates, and sodium to induce the precipitation of residual fluorosilicate salts, which can then be isolated. Sodium chloride (NaCl) is particularly preferred because it yields fluorosilicate salts with a purity greater than 96%, typically ranging from 96% to 99%. Steps (c) to (f) allow the TRs to be concentrated and leached. Step (c) is a basic leaching step of the solid residue (dried, ground and sieved) obtained at the end of step (b). The leaching is carried out using a strong mineral base, preferably a solution of sodium hydroxide (NaOH), potassium (KOH), lithium (LiOH) or rubidium (RbOH), particularly preferably a solution of sodium hydroxide. When the strong base is a sodium hydroxide solution, the sodium hydroxide is typically present in the solution at a concentration ranging from 5 to 7 mol / L, preferably a concentration of 5.4 to 6.2 mol / L. The mass ratio of sodium hydroxide solution to solid residue obtained at the end of step (b) is 1 to 3, preferably 1.5 to 2.5. Leaching is generally carried out at a temperature of 30 to 50 °C, preferably 32 to 37 °C in order to optimize the conversion of calcium sulfates. The leaching stage typically lasts from 1 to 10 minutes, preferably from 1 to 5 minutes. During step (c), the calcium sulfates present in the solid residue are converted into soluble divalent alkali sulfates of formula X2SO4 (I) with X= Na, K, Li or Rb and into insoluble calcium hydroxide. The solid residue and the liquid phase obtained after basic leaching are collected separately. The resulting solid residue includes calcium hydroxide, TRs and other metallic elements. Advantageously, the conversion rate is between 95 and 99.9%, preferably between 97.5 and 99.9% when the basic solution is a sodium hydroxide solution. The conversion rate T of sulfates during the basic leaching step (c) is defined as follows: 100 where mi is the initial mass and calcium sulfate content of the ground residue before leaching obtained at the end of step (b) and mf is the final mass and T f the calcium sulfate content of the residue obtained at the end of step (d). The basic leaching stage has the advantage of being rapid. Furthermore, the small amount of caustic soda used makes the process environmentally friendly. This is followed by a drying, grinding, and sieving stage of the solid residue obtained at the end c). The conditions of step (d) are as described in step (b). The crushed and sieved residue typically consists of particles with a size less than or equal to 400 µm, preferably ranging from 100 to 400 µm. Optional step for processing the liquid phase obtained in step (c) The process may further include a neutralization step of the liquid phase obtained at the end of step (c). Neutralization can be carried out by adding sulfuric acid to the liquid phase obtained at the end of step (c) to obtain a pH ranging from 5 to 8, preferably from 5 to 7. This step converts the residual base present in the liquid phase (filtrate) (NaOH, KOH, LiOH, or RbOH) into the corresponding alkali metal sulfate (X₂SO₄(I) where X = Na, K, Li, or Rb). This step yields crystals of the corresponding alkali metal sulfate. The sulfate crystals obtained are collected. Advantageously, the mineral base used in step (c) is soda and the recovered sodium sulfate has a mass purity of 98% to 99.9%, preferably 99.5% to 99.9%. Step (e): Step (e) is an acid leaching step of the solid residue obtained at the end of step (d) using nitric acid. Acid leaching allows the extraction of nitric acid-soluble elements from the solid residue. 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. Preferably, nitric acid is present as a solution with an acid concentration of 45 to 65% by weight, preferably 50 to 60% by weight. Nitric acid and calcium hydroxide are typically present in a stoichiometric ratio ranging from 0.5 to 2, advantageously equal to 1. When nitric acid and calcium hydroxide are present in a stoichiometric ratio of less than 1, the dissolution of calcium hydroxide is not maximal. When nitric acid and calcium hydroxide are present in a stoichiometric ratio of less than 1, the solubilization of rare earths is reduced. Nitric acid allows strong dissociation of the solid residue with a solubilization rate of 80 to 99%, preferably 85 to 95% of the initial mass of the solid residue. The leaching stage typically lasts from 1 to 10 minutes, preferably from 1 to 5 minutes. The solid residue and liquid phase obtained after acid leaching are collected separately, typically by centrifugation. The collected liquid phase is rich in rare earths, calcium nitrates and other metallic elements. Step (f) Step (f) is a precipitation step of calcium nitrate crystals from the liquid phase obtained at the end of step (e). Precipitation can be achieved by cooling, preferably at a temperature below 0°C. Cooling allows the crystallization of calcium salts, in particular calcium nitrate tetrahydrate (Ca(NO3)2.4H2O). Seeding with calcium salt crystals, for example with calcium nitrate tetrahydrate, can be carried out in order to accelerate the rate of crystallization. The liquid phase can be cooled at a temperature ranging from -20 to 0°C, preferably from -15 to -5°C, for a period typically ranging from 8 to 24 hours. advantageously from 16 hours to 24 hours, which allows the obtaining of calcium nitrate tetrahydrate crystals (solid residue). The resulting solid residue (e.g., calcium nitrate tetrahydrate) and liquid phase are collected separately. Separation can be achieved by vacuum filtration. The collected liquid phase is rich in TRs and other metallic elements. Optional step of washing the calcium nitrate crystals contained in the solid residue obtained at the end of step (f) The collected solid residue (precipitate) can be washed using concentrated nitric acid. The acid solution used for the washing step typically has an acid concentration ranging from 45 to 65% by weight, preferably from 50 to 60% by weight. Washing removes any residual traces of the liquid phase. The calcium nitrate tetrahydrate crystals are then isolated. Separation can be performed by vacuum filtration. Step (g) allows the TRs to precipitate and concentrate. Step (g) Step (g) is a pH adjustment step for the liquid phase obtained at the end of step (f) to a value ranging from 1 to 1.8 by adding a strong base. Adjusting the pH allows for the selective precipitation of the TRs contained in the liquid phase. The resulting solid residue (precipitate) and liquid phase are collected separately. Separation can be carried out by centrifugation (e.g., 3000 rpm for 15 min). The strong base can be a solution of ammonium hydroxide, sodium hydroxide, potassium hydroxide, sodium carbonate, potassium carbonate, or mixtures thereof, preferably a solution of ammonium hydroxide or sodium hydroxide. The concentration of the strong base solution typically ranges from 4 to 7 mol / L, preferably from 4 to 6 mol / L. The pH is preferably adjusted to a value of 1 to 1.4. The duration of step (g) generally varies from 10 to 60 min, preferably from 30 to 60 min. Step (h) is a drying step of the solid residue (precipitate) obtained at the end of step (g). The dried solid residue can then be ground and sieved. The conditions of step (h) are identical to those of step (b). The solid residue is enriched in TRs. The process according to the invention makes it possible to obtain solid residues typically having rare earth contents ranging from 4% to 6% by weight. Optional step for processing the liquid phase obtained during step (g) The process may further include a step of adjusting the pH of the liquid phase obtained at the end of step (g) to a pH ranging from 6 to 10, preferably from 7 to 10, by adding a strong base, the strong base being preferably a solution of ammonium hydroxide or sodium hydroxide. This step allows the precipitation of residual and undesirable metallic elements present in the liquid phase, such as Fe, Al, etc. It allows obtaining a purified solution of ammonium nitrate or sodium nitrate depending on the strong base used. The same strong base solution used in step (g) can be used to adjust the pH of the solid residue obtained at the end of step (h), also noted The dried solid residue obtained at the end of step (h) can be ground, possibly sieved, and undergo the following treatment: - washing of the ground solid residue, possibly sieved with a polar solvent and collection of the solid residue obtained and the liquid phase obtained, then drying and grinding of the solid residue to obtain a solid product containing rare earths, or calcination of the ground solid residue, possibly sieved at a temperature ranging from 700°C to 1000°C to obtain a solid product containing rare earths. The ground solid residue, possibly sieved, can be washed with a polar solvent. Advantageously, the polar solvent is water. Advantageously, the mass ratio of polar solvent to solid residue is 5 to 15, preferably 10 to 15. The washing time typically varies from 15 minutes to 60 minutes, preferably from 45 minutes to 60 minutes. The washed solid residue is collected. Then the solid residue is dried, ground and possibly sieved to obtain a solid product containing rare earth elements. The conditions of the drying and grinding stage are as described in step (b). Alternatively, the crushed solid residue, possibly sieved, can be calcined at a temperature ranging from 700°C to 1000°C to obtain a solid product containing soils rare. The calcination stage advantageously eliminates traces of organic matter and residual water, thus increasing the rare earth content of the solid residue. The solid residue can be calcined at a temperature of 800 to 900°C. Advantageously, the calcination time is 30 to 120 min, preferably 45 to 60 min. A minimum temperature of 700°C allows for the calcination of rare-earth-enriched insolubles in the solid residue, whereas applying a temperature above 1000°C represents an unnecessary additional cost. A calcination temperature ranging from 700°C to 1000°C advantageously achieves complete calcination of the rare-earth-enriched insolubles in the solid residue within 30 minutes to 2 hours. This optional step of treating the solid residue obtained at the end of step (h) allows the said residue to be enriched in TRs. By implementing the optional step (i), the process according to the invention can yield solid residues typically having rare earth contents ranging from 5% to 25% by weight. This TR content advantageously ranges from 10% to 25% by weight when step (i) includes a washing step of the solid residue obtained at the end of step (h), or from 8% to 15% by weight when step (i) includes calcination of the solid residue obtained at the end of step (h). In addition to rare earths, the process according to the invention makes it possible to isolate phosphorus, fluorine, silicon, calcium and sulfur contained in PG and BPs, thus opening the way to the valorization of these elements. The proposed process thus makes it possible to valorize by-products in the form of sodium fluorosilicates, divalent alkali sulfates, calcium nitrate tetrahydrate, or sodium nitrate. These by-products are isolated. They can be purified and obtained in solid form or in solution. In the process according to the invention, the separation of a liquid phase and a solid residue can be achieved by any separation technique known to those skilled in the art, in particular, by vacuum filtration or by centrifugation. Centrifugation may be preferred when it is desired to recover the liquid phase. Vacuum filtration may be preferred when it is desired to recover the solid phase. Unless otherwise stated, the steps of the process are advantageously carried out at room temperature, particularly for cost reasons. Figure 1 is a schematic representation of an example of a process of the present invention implemented using phosphoric sludge. Those skilled in the art will be able to adapt the described steps to the treatment of phosphogypsum. With reference to Figure 1, the illustrated process comprises 3 main steps: • A step (A) intended to reduce the content of soluble impurities in the BPs. This step A includes: - separation of the acid 52% P2O5(F1) and the sludge, generally by centrifugation (1), typically carried out at 3000 rpm for 10 min; - a multiple washing of the sludge (B1) by a polar protic aqueous solution at room temperature, in order to recover the P2O5, in a reactor (2); - a separation of the sludge (B2) and the filtrates F2, F3, F4, F5 and F6, typically by vacuum filtration (3); - washing of the sludge (B2) with a polar protic aqueous solution previously heated in a reactor (4); - a separation of the sludge (B3) and the filtrate (F7), typically by vacuum filtration (5); - a drying of the sludge (B3), for example at 80 °C, followed by grinding and sieving, typically to less than 250 mm (6); - cooling of the filtrate (F7) to room temperature in a tank (22); - separation of the crystals (C1) and the filtrate (F8), typically by filtration (23) under vacuum; - precipitation of crystals (C2) by attack of the filtrate (F8) with a sodium-based salt of the type NaOH, NaCl, NaF, NaL or Na2CO3 in a reactor (24); - separation of crystals (C2) and filtrate (F9) by filtration (25), typically under vacuum. • A step (B) of concentration and leaching of the TRs with recovery of the components of the sludge and the reagents used. This step B includes: - the preparation of a basic solution of the type NaOH, KOH, LiOH, RbOH, or mixtures thereof in a reactor (7); - an introduction of the BPs (B4) and the basic solution into a reactor (8); - treatment of the mixture at a temperature generally above 35°C, typically for 5 min at an agitation speed of 500 rpm; - a separation of the filtrate (F10) and the residue (R1), typically by vacuum filtration (8); - neutralization of the filtrate (F10), for example with sulfuric acid, in a reactor (26); - evaporation of the neutralized filtrate (F11) in an oven (27), generally at 80°C; - a drying, typically at 80°C, followed by grinding (9) of the residue (R1), to prepare the residue (R2); - preparation of a nitric acid solution in a reactor (10); - an introduction of the residue (R2) and the nitric acid solution into a reactor (10); - treatment of the mixture by mechanical agitation, typically at a speed of 500 rpm for 5 min; - separation of the leachate (F12) and residue (R3) by centrifugation (11), typically at 3000 rpm for 15 min; - placing the leachate (F12) in a refrigerator (12) set at -10 °C; - inoculation, for example using calcium nitrate tetrahydrate, of the leachate placed in the refrigerator; - separation of the leachate (F13) and the crystals (C4) by filtration (13), typically under vacuum; - washing the crystals (C4) with nitric acid, for example at 50% by weight (28); - a separation of the filtrate (F14) and the crystals (C5) by filtration (24), for example under vacuum; • A step (C) of precipitation and concentration of TRs and purification of leachate (F15). This step (C) includes: - an adjustment of the pH of the filtrate (F13) to pH=1.2 by a basic solution of NaOH, KOH, NasCOs, K2CO3 OR NH4OH in a reactor (14); - separation of the leachate (F15) and the precipitate (P1) by centrifugation (15), typically at 3000 rpm for 15 min; - an adjustment of the pH of the filtrate (F15) by the solution of NaOH, KOH, NasCOs, K2CO3OR NH4OH in a reactor (30); - separation of the leachate (F16) and the precipitate (P6), generally by centrifugation (31), typically at 3000 rpm for 15 min; - a drying and grinding (16) of the precipitate (P1) to prepare a solid (P2); - a calcination of the solid (P2) in a furnace (21), generally for one (1 h) hour at a temperature of 900 °C; - washing of the precipitate (P2) in a reactor (18), generally under agitation at 300 rpm with distilled water; - separation of the leachate (F17) and the precipitate (P3), typically by centrifugation (19) at 3000 rpm for 15 min; - Drying and grinding (20) of the precipitate (P3) to prepare a solid (P4). As previously described, initially (step (A)), the BPs are separated by centrifugation to obtain a solution (F1) containing 52% by weight of P2O5 and a residue rich in TRs (B1) of a gypsum-like nature (CaSC) mixed with sodium fluorosilicate (NasSiFe) and residual phosphorus pentoxide (P2O5). The sludge (B1) is cascade-washed with a polar solvent to recover more P2O5 and minimize the concentration of impurities while concentrating the TRs. For example, 5 washes can be carried out with water at room temperature under stirring at 300 rpm and a contact time of 5 min. A final wash can be carried out with hot water at a temperature of 80 to 90°C, agitated at 300 rpm, with a contact time of 10 minutes. The sludge (B3) is dried, ground, and screened to prepare sludge (B4).Next, the hot filtrate is placed in a tank to promote cooling of the filtrate (F7) to room temperature. The purpose of cooling the filtrate is to induce the crystallization / precipitation of sodium fluorosilicate (Na2SiF6) crystals (C1). Sodium chloride (NaCl) or sodium carbonate (Na2CO3) is then added to the residual filtrate (F8), which is saturated with fluorine, silicates, and sodium. The addition of the aforementioned salts promotes the precipitation of sodium fluorosilicate (Na2SiFe) crystals (C2). Then, to carry out step (B), a basic solution of sodium hydroxide (NaOH), potassium hydroxide (KOH), lithium hydroxide (LiOH), or rubidium hydroxide (RbOH) is prepared. The solution is cooled to a temperature ranging from 33 to 38°C to ensure high sulfate conversion, as sulfates are the major component of the sludge (B4). The residues (B4), rich in TRs, are added to the basic solution to convert the calcium sulfates into double alkali sulfate (X₂SO₄: X = Na, K, Li, or Rb) (F10) and calcium hydroxide (R1), which is very rich in TRs and other metallic elements such as Al, Sr, Ba, Mg, etc. The filtrate (F10) is neutralized with a dilute sulfuric acid (H₂SO₄) solution. The purpose of neutralization is to promote the conversion of the residual base (NaOH, KOH, LiOH, or RbOH) contained in the filtrate into double alkali sulfate (X2SO4: X=Na, K, Li, or Rb), so as to to avoid crystallization and mixing of the two products during drying. This neutralization step (26) allows obtaining sodium sulfate crystals (C3). Next, the residues (R2) previously dried, ground and sieved to less than 400 pm undergo solubilization in a nitric acid solution, for the dissolution of TRs, calcium and other metallic elements present in the residues (R2). After solubilization, the filtrate, consisting mainly of solubilized calcium and nitrates, undergoes crystallization in a refrigerator to prepare calcium nitrate tetrahydrate (Ca(NO3)2.4H2O) crystals (C4) and a leachate (F13) rich in triglycerides and metallic elements. After collection of the leachate (F13), the crystals (C4) are washed. This step is carried out by pouring concentrated nitric acid over the crystals (C4) to remove residual traces of leachate (F13). This facilitates the purification of the final calcium nitrate tetrahydrate (Ca(NO3)2.4H2O) crystals (C5). Finally, in step (C), the leachate (F13) undergoes pH adjustment with a solution of sodium hydroxide (NaOH) or ammonium hydroxide (NH4OH). The pH is controlled to precipitate the TRs. Next, the precipitate (P2), previously dried and ground, undergoes calcination or washing with water to further concentrate the TRs and obtain a washed precipitate (P4) or a calcined precipitate (P5). Finally, the leachate (F15), resulting from the pre-neutralization of the filtrate (F13) at a pH of 1.2, undergoes post-neutralization at a defined pH to precipitate the residual metals in filtrate in order to further purify the liquid phase (F16) based on sodium nitrate or ammonium nitrate. REFERENCES

[0001] G. Allaedini and P. Zhang, “T reatment of phosphoric acid sludge for rare earths recovery II: effect of sonication and flocculant solution temperature on settling rate”, Separation Science and Technology, vol. 54, no. 11, p. 1842-1852, July 2019, doi: 10.1080 / 01496395.2018.1536715 [2] A. Mahrou et al., “Rare Earth Elements Distribution During Phosphoric Acid Production”, Mining, Metallurgy & Exploration, vol. 39, no. 1, p. 161-167, Feb. 2022, doi: 10.1007 / S42461-021-00513-9 [3] G. G. Jang, A. Ladshaw, J. K. Keum, P. Zhang, et C. Tsouris, « Continuous-Flow Centrifugal Solid / Liquid Separation for the Recovery of Rare-Earth Elements Containing Particles from Phosphoric Acid Sludge », Ind. Eng. Chem. Res., vol. 59, n° 50, p. 21901-21913, déc. 2020, doi: 10.1021 / acs.iecr.0c04128 [4] WG2020067856A1 [5] Z. Wei et Z. Deng, « Research hotspots and trends of comprehensive utilization of phosphogypsum: Bibliometric analysis », Journal of Environmental Radioactivity, vol. 242, p. 106778, févr. 2022, doi: 10.1016 / j.jenvrad.2O21 .106778. [6] CN11 1560521 A [7] CN112408452A [8] CN110918251A [9] US10988387B2

[0010] RU2739409C1

[0011] RU2258036C 1

[0012] EP0419318A1

[0013] S. Al-Thyabat and P. Zhang, “REE extraction from phosphoric acid, phosphoric acid sludge, and phosphogypsum,” Mineral Processing and Extractive Metallurgy, vol. 124, no. 3, p. 143-150, Sep. 2015, doi: 10.1 179 / 1743285515Y.0000000002

[0014] SA Stern, JT Mullhaupt, and WB Kay, “The Physicochemical Properties of Pure Nitric Acid. », ACS Publications, May 1, 2002. https: / / pubs.acs.org / doi / pdf / 10.1021 / cr60204a004 EXAMPLES The following examples of implementation of a process according to the invention are given by way of illustration and should not be interpreted as limiting the scope of the invention. To characterize the various raw materials and the resulting residues and solutions, physical and chemical analyses were performed. X-ray crystallography (XRD) was used to track the phase profiles in the solids. Scanning electron microscopy (SEM) was used to study their topography and morphology. Energy-dispersive spectroscopy (EDS) has been used for elemental chemical analysis of solids, inductively coupled plasma mass spectrometry (ICP-MS) for elemental analysis of rare earths in solids and liquids, and atomic absorption spectroscopy (AAS) for elemental analysis of other elements in solids and liquids. The references indicated in the following (e.g., B1, F1...) are made with reference to Figure 1. Unless otherwise stated, the percentages shown in the examples are mass percentages. Example 1: Valorization of phosphoric sludge and preparation of a solid concentrate rich in rare earths by precipitation with soda. The BPs are subjected to a centrifugation step for 10 min with a rotation speed of 3000 rpm to obtain a solid residue (B1) and a liquid phase (F1). Next, the solid residue (B1) is added to a volume of water with a liquid / solid mass ratio of 1 and mechanically stirred at 300 rpm for 5 minutes at atmospheric pressure and room temperature. After vacuum filtration and isolation of the liquid phase (F2), the resulting solid residue is added to a volume of water with a liquid / solid mass ratio of 2 and mechanically stirred at 300 rpm for 5 minutes at atmospheric pressure and room temperature. The liquid phase (F3) and the solid phase are then separated by vacuum filtration, and the washing process is repeated three times. This yields the liquid phases (F4) to (F6) and a solid residue (B2). Vacuum filtration of all mixtures is carried out until a solid moisture content of 25 to 30% by weight is obtained. The resulting solid (B2) is then added to a water solution preheated to 83°C and mechanically stirred at 300 rpm for 10 minutes. The solid (B3) and liquid (F7) are then separated by vacuum filtration, with the volume of water corresponding to 10 times the volume of dried solid. This step aims to purify the sludge by removing the sodium fluorosilicate phase. The filtrates separated from the solid after each wash are analyzed to determine the concentration of P2O5. Figure 2 represents the evolution of the concentration of phosphorus pentoxide as a function of the number of cold washes. The results of analyses show that the concentration of P2O5 decreases from 52.58% by weight to 0% after the cold wash sequence, which implies the total recovery of P2O5. The solid residue (B3) is dried, ground, and sieved through a 250 µm sieve to obtain the solid residue (B4). The results of elemental analyses in the phosphate rock samples (B1) from the phosphate processing plant, Jorf Lasfar / Morocco, and in the solids (B2) and (B4) are detailed in Table 1. The residue (B4) obtained is four times more concentrated in trace elements (TEs) than the initial phosphate rock samples. The mineralogical composition of the phosphate rock samples (B1) and (B4) (Table 1) shows that the increase in the concentration of trace elements in the phosphate rock samples is accompanied by an increase in the concentration of calcium, sulfates, etc., and a decrease in the concentration of phosphorus pentoxide, fluorides, and silicates. The ICP results show that TRs are concentrated in the cold pretreated BPs (B2). The solid is rich in TRs and other metallic elements and the filtrates are rich in P2O5. Table 1: Mineralogical composition of different samples of phosphoric sludge The main phases of the BPs are determined by XRD. The results in Figure 3 show that the pretreated sludge (B2) is a mixture of two phases (CaSC and NasSiFe), while the sludge (B4) contains only the gypsum phase (CaSC). Hot washing therefore removed the sodium fluorosilicates (NasSiFe). Step (c): Basic leaching of the sludge (B4) is carried out as follows: 197.5 g of the sludge (B4) is poured into a volume of 500 mL of NaOH previously prepared by solubilizing 116.15 g of NaOH and cooling the solution to 35 °C. The mixture is stirred at a speed of 500 rpm by a mechanical stirrer at various times, preferably for 5 min. Step (d): After the vacuum filtration step, the residue (R1) undergoes drying and grinding to less than 250 µm to obtain a solid residue (R2). The residue (R2) is analyzed by ICP, XRD, and chemical analysis to determine its chemical composition. Table 2 represents the chemical composition of the residues (R2). The results show that the residue (R2) is rich in TRs with a concentration exceeding 6500 ppm. Furthermore, the sulfate conversion yield and TR concentration are 97.9% and 99.8%, respectively. Table 2: Mineralogical composition of calcium hydroxides (R2) XRD analysis of residue (R2) shows that the BPs have been completely transformed into calcium hydroxide, which confirms the high conversion of calcium sulfates (figure 4). Tl The residue R2 is subjected to acid leaching. 200 g of residue (R2) are slowly added to 490 mL of 50% (w / w) nitric acid solution. The mixture is mechanically stirred at a speed of 500 rpm for a variable duration, preferably 5 min. After separation by centrifugation at 3000 rpm for 15 min, the filtrate (F12) is seeded with a few grains of calcium nitrate tetrahydrate and cooled to -10 °C. Next, the calcium nitrate crystals formed (C4) are separated by vacuum filtration to obtain a filtrate (F13). The acid leaching yield (corresponding to the percentage of solid residue solubilized in step (e)) and the leachate concentration yield for TRs are 77.6% and 84.5%, respectively. The results of TRs and other metal analyses in the different matrices are shown in Table 3. The crystals (C4) are washed with 50 wt% nitric acid to prepare purified calcium nitrate tetrahydrate crystals (C5). The analytical results of the different fractions are presented in Table 3. Table 3: Mineralogical composition of the filtrate (F12), (F13) and the crystals (C5) Step (a): The filtrate (F13), consisting of a concentrate of TRs dissolved in a medium containing predominantly nitrates, undergoes pH adjustment. A 150 mL volume of filtrate is treated with 5 mol / L sodium hydroxide to adjust the pH to 1.2. The mixture is stirred at 300 rpm by a magnetic stirrer for one (1) hour. After recovery of the precipitate (P1) by centrifugation at 3000 rpm for 15 min, it is dried and ground to give the solid residue (P2) containing the TRs. The mass content of TRs in the solid residue P2 is 4.8%. The filtrate (F15) is treated to adjust the pH to 8.78 (F16) with sodium hydroxide, which allows residual metals to precipitate and yields a purified sodium nitrate solution. The solid residue (P2) is washed with water at an agitation speed of 300 rpm for a period of one (1) hour for a liquid-solid mass ratio of 10, then the residual solid (P3) is isolated by centrifugation at 3000 rpm for 15 min. The solid (P3) undergoes drying at 80°C followed by grinding to prepare a solid (P4). According to one variant of the process, the solid residue P2 undergoes a calcination step at 900°C for one hour to obtain the residue (P5). The results of analyses in TRs and other mineral components are shown in Table 4. The ICP analysis results shown in Table 4 demonstrate the effectiveness of the cascade method in selectively precipitating TRs in precipitate (P1) and purifying the filtrate (F14). Precipitates (P4) and (P5) are advantageously rich in TRs, reaching mass concentrations of 19% and 9.4%, respectively. Table 4: Mineralogical composition of precipitates from the filtrate of BPs neutralized with sodium hydroxide / HG: out of range Additional steps in valorizing by-products As described above, the process may include steps enabling the valorization of by-products, for example to obtain purified calcium nitrate tetrahydrate crystals (see step (f) above) or to obtain a purified sodium nitrate solution (see step (h) above). Other by-products can also be valorized. - treatment of the liquid phase obtained at the end of step (a) to valorize sodium fluorosilicates The filtrate from the hot wash (F7) obtained in step (a) is cooled to room temperature to ensure supersaturation of the medium with sodium fluorosilicates, which promotes their precipitation. The solid phase is separated from the liquid phase (F8) by vacuum filtration, then dried, ground, and analyzed by XRD. The analysis (Figure 5) shows that the solid residue contains only sodium fluorosilicates (C1). A 1 L volume of filtrate saturated with sodium fluorosilicates (F8) is used to precipitate residual sodium fluorosilicates from the above step, by adding a 50 mL volume of a 3 mol / L NaCl solution. According to a variation of the process, precipitation by adding 30 mL of a 1.5 mol / L Na2CO3 solution can be used to precipitate NasSiFe. The reaction time is 1 h and the stirring speed is 500 rpm. X-ray diffraction analysis (Figure 6) shows that the precipitate contains only the sodium fluorosilicate (C2) phase. Chemical analyses determined the purity of the products formed. The NasSiFe resulting from supersaturation and the addition of NaCl is pure (>99%), however, the NasSiFe resulting from the addition of Na2CO3 is of lower quality (>60%). - treatment of the liquid phase obtained at the end of step (c) to valorize the sodium sulfate The filtrate (F10) obtained at the end of step (d) is neutralized with a H2SC>4 solution to a pH ranging from 6 to 6.5 to remove traces of residual sodium hydroxide (NaOH). The solution is neutralized, dried at 80 °C, and the solid residue is ground. XRD analysis (Figure 7) shows that the residue obtained is of the sodium sulfate (C3) type. The elemental analysis shown in Table 5 indicates that the sodium sulfate (C3) crystals are slightly contaminated with metallic elements. The purity of the product obtained is 99.7%. Crystals (C3) Table 5: Mineralogical composition of sodium sulfates Valorization of phosphoric sludge and preparation of a solid concentrate rich in TRs by precipitation with ammonium hydroxide solution. The process for treating BPs is identical to that of example 1 except that the strong base used in step (g) is in example 2 an ammonium hydroxide solution. Step (g) A 150 mL volume of filtrate (F13) is neutralized with a 5 mol / L ammonium hydroxide solution to adjust the pH to 1.2. The mixture is stirred at 300 rpm using a magnetic stirrer for one (1) hour. After recovering the precipitate (P1) by centrifugation at 3000 rpm for 15 min, the filtrate is neutralized again to pH 7.45 using the previously prepared ammonium hydroxide solution. The solid is then separated from the filtrate by centrifugation. Step (h) and step (i) The solid residue (P2) obtained after drying, grinding and sieving of the precipitate (P1) undergoes an optional washing step or an optional calcination step. The solid residue (P2) has a mass content of TRs of 4.3%. The washing is carried out with water at an agitation speed of 300 rpm for a duration of one (1) hour for a liquid-solid ratio of 10. Then the solid (P3) hyper-concentrated in TRs was recovered by centrifugation at 3000 rpm for 15 min and then underwent drying and grinding to obtain the TRs concentrate (P4). The calcination of precipitate (P2) is carried out at 900 °C for one hour, and the resulting precipitate is designated (P5). The results of TR analyses and other mineral components are shown in Table 6. The concentration of TRs in precipitates P4 and P5, after implementation of the optional step (i), is 19.2% and 12.3%, respectively. Table 6: Mineralogical composition of precipitates from the filtrate of BPs neutralized with ammonium hydroxide Example 3: Valorization of phosphoric sludge and preparation of a solid concentrate rich in TRs by precipitation with sodium hydroxide. In this example, we implement the same BP treatment process described in example 1, except that the number of cold washes in step a) is 3 in this example. The protocol consists of pretreating the BPs by centrifugation at 3000 rpm for 10 min. The resulting solid phase (B1) is then washed with a volume of water corresponding to a liquid-solid ratio of 1 for 5 min at atmospheric pressure and ambient temperature. After vacuum filtration, the solid phase, containing a moisture content ranging from 25 to 30%, is added to a volume of water corresponding to a liquid-solid ratio of 2 under mechanical agitation for 5 min at atmospheric pressure and ambient temperature. The liquid phase is then separated from the solid phase by vacuum filtration, and the final washing is repeated once. The mineralogical composition of the solid residue (P2) obtained at the end of step h) and of the solid residue (P4) obtained at the end of step i) according to the variant of the process with washing of the solid residue P2 is detailed in Table 7. The solid residue (P2) according to the invention has a mass content of TRs of 5.4%. By implementing the optional step (i), the TRs content of the solid product is 21.3%. Table 7: Mineralogical composition of precipitates from the filtrate of BPs neutralized with sodium hydroxide 5 Example 4: Valorization of phosphoric sludge and preparation of a solid concentrate rich in TRs by precipitation with ammonium hydroxide solution. In this example, we implement the same BP treatment process described in example 2, except that the number of cold washes in step a) is 3 in this example. The 3 wash steps are as described in example 3. The mineralogical composition of the solid residue (P2) obtained at the end of step h) and of the solid residue (P4) obtained at the end of step i) according to the variant of the process with washing of the solid residue P2 is detailed in Table 8. 5 The solid residue (P2) according to the invention has a mass content of TRs of 4.6%. By implementing the optional step (i), the concentration of TRs in the P4 precipitate is 18.66%. Table 8: Mineralogical composition of precipitates from the filtrate of BPs neutralized with ammonium hydroxide

Claims

CLAIMS 1. Process for treating phosphogypsum or phosphoric sludge allowing the recovery of rare earths, the process comprising the following successive stages: (a) treatment of phosphogypsum or phosphoric sludge respectively according to the following steps (a1) or (a2): (a1) washing the phosphogypsum with a polar solvent and collecting the solid residue obtained; or (a2) separating the phosphoric sludge into a liquid phase and a solid residue and collecting the obtained solid residue and the obtained liquid phase separately; washing the obtained solid residue after the treatment with a polar solvent at room temperature and collecting the obtained solid residue and the obtained liquid phase separately; and washing the obtained solid residue after the washing with a polar solvent at a hot temperature and collecting the obtained solid residue and the obtained liquid phase; (b) drying, grinding and sieving the solid residue obtained at the end of step (a); (c) basic leaching of the solid residue obtained at the end of step (b) with a strong mineral base, and collection of the solid residue obtained and the liquid phase obtained; (d) drying, grinding and sieving the solid residue obtained at the end of step (c); (e) acid leaching of the solid residue obtained at the end of step (d) with nitric acid and collection of the solid residue obtained and the liquid phase obtained; (f) precipitation of calcium nitrate crystals from the liquid phase obtained at the end of step (e) and collection of the solid residue and the liquid phase obtained; (g) adjusting the pH of the liquid phase obtained at the end of step (f) to a value ranging from 1 to 1.8 by adding a strong base and collecting the solid residue obtained and the liquid phase obtained; (h) drying the solid residue obtained at the end of step (g) to obtain a solid product containing rare earths.

2. Process according to the preceding claim in which the polar solvent of step (a) is water, methanol, ethanol, acetone, an ether or their mixtures, preferably water.

3. Method according to claim 1 or 2 in which the strong mineral base of step (c) is a solution of sodium hydroxide, potassium hydroxide, lithium hydroxide or rubidium hydroxide, preferably a solution of sodium hydroxide.

4. Process according to any one of claims 1 to 3 in which step (c) is carried out at a temperature ranging from 30 to 50°C, preferably from 32 to 37°C.

5. Process according to any one of claims 1 to 4 in which the strong base of step (g) is a solution of ammonium hydroxide, sodium hydroxide, potassium hydroxide, sodium carbonate, carbonate, potassium or mixtures thereof, preferably a solution of ammonium hydroxide or sodium hydroxide.

6. Method according to any one of claims 1 to 5 further comprising a step of precipitation of the fluorosilicate salts contained in the liquid phase obtained at the end of step (a2) and of collection of the salts obtained, preferably by cooling and addition of sodium chloride (NaCl) or sodium carbonate.

7. Method according to any one of claims 1 to 6 further comprising a step of neutralizing the liquid phase obtained at the end of step (c), preferably by adding sulfuric acid to the liquid phase obtained at the end of step (c) to obtain a pH ranging from 5 to 8, preferably from 5 to 7, and collecting the sodium sulfate crystals obtained.

8. Method according to any one of claims 1 to 7 in which the precipitation of the calcium nitrate crystals in step (f) is carried out by cooling and optionally seeding the liquid phase.

9. Method according to any one of claims 1 to 8 further comprising a step of washing with nitric acid the solid residue obtained at the end of step (f) and collecting the solid residue obtained comprising crystals of calcium nitrate tetrahydrate.

10. Method according to any one of claims 1 to 9 further comprising a step of adjusting the pH of the liquid phase obtained at the end of step (g) to a pH varying from 6 to 10 by adding a strong base, the strong base preferably being a solution of ammonium hydroxide or sodium hydroxide, to obtain a solution of ammonium or sodium nitrate. 1 1. Method according to any one of claims 1 to 10 further comprising a step (i) comprising grinding, optionally sieving the dried solid residue obtained at the end of step (h) and: washing the ground solid residue, optionally sieved with a polar solvent and collecting the solid residue obtained and the liquid phase obtained, then drying and grinding the solid residue to obtain a solid product containing rare earths, or calcining the ground solid residue, optionally sieved at a temperature ranging from 700°C to 1000°C to obtain a solid product containing rare earths.

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