Method for recovering an arsenic-enriched concentrate from a radiologically contaminated ore

Abstract

The present invention relates to a method for recovering a concentrate enriched with arsenic and other optional metals from a uranium-bearing ore comprising radium and arsenic, by flotation.

Description

[0001] DESCRIPTION

[0002] TITLE: METHOD FOR RECOVERING AN ARSENIC-ENRICHED CONCENTRATE FROM RADIOLOGICALLY CONTAMINATED ORE

[0003] TECHNICAL FIELD

[0004] The present invention relates to the field of mineral deposit processing, and in particular uranium deposits.

[0005] Uranium ores can be contaminated by the presence of other elements such as arsenic, and possibly other metals such as radium, nickel, cobalt and / or copper.

[0006] To separate these elements, ore processing generally involves, firstly, an oxidative leaching process using sulfuric acid to dissolve the uranium and some of the arsenic (referred to below as leached arsenic). This results in a leachate containing these elements and a mine tailings also containing arsenic (referred to below as unleached arsenic) and possibly other elements, such as radium, nickel, cobalt, and / or copper. The leachate is then subjected to solvent extraction of the uranium, the resulting organic phase being a uranium-enriched phase and the remaining aqueous phase containing the leached arsenic. The mine tailings containing the unleached arsenic are then subjected to successive washings with industrial water in countercurrent clarifiers.

[0007] The aqueous phase containing leached arsenic, as well as the washed mine leaching residues containing unleached arsenic, are subjected, before storage, to neutralization typically consisting of the addition of lime and ferric iron (Fe 3+ ) in the form of ferric sulfate (Fe2(SO4)3) in excess of arsenic. This treatment is illustrated in Figure 1.

[0008] The neutralization of leached arsenic differs from that of unleached arsenic: Ferric sulfate is added at an Fe / As ratio of 3 to neutralize leached arsenic, while for unleached arsenic, the Fe / As ratio is approximately 14. Neutralizing unleached arsenic therefore requires about five times more ferric iron than neutralizing arsenic in solution and thus represents a significant cost factor in ore processing.

[0009] Furthermore, the depletion of uranium deposits de facto induces an increase in the arsenic content of the ore, and an increase in the financial impact associated with the increased consumption of oxidizing reagents to compensate for the increase in arsenic content.

[0010] This financial impact can currently be estimated at several million dollars per year and per deposit.

[0011] It is therefore necessary to provide an alternative treatment route for leaching residues for the management of unleached arsenic, in order to reduce the consumption of ferric sulfate and lime.

[0012] Furthermore, lowering ore processing costs would eventually make new deposits accessible that are currently considered economically unattractive, due, among other things, to their high arsenic content.

[0013] PREVIOUS STATE OF THE ART

[0014] US patent 4,406,864 describes a process for recovering uranium from an ore containing iron, arsenic, and silicon using flotation, and specifically mentions the recovery of nickel salts from arsenic by flotation. However, the flotation is carried out after leaching steps on a leaching residue that typically remains low in arsenic. Therefore, the arsenic recovery yield is not satisfactory.

[0015] US patent 10,822,673 B1 describes a process for remediating arsenic from a lead concentrate employing ozone treatment and a flotation step.

[0016] One aim of the invention is therefore to optimize the costs of uranium production by limiting the costs associated with the treatment of unleached arsenic contained in mining residues after leaching.

[0017] Furthermore, the treatment and recovery of arsenic and, where applicable, valuable metals from uranium mining tailings must include a radiological decontamination step. The presence of radium and uranium hinders the recovery of these concentrates by third parties.

[0018] Another objective is therefore to collect a concentrated product rich in arsenic and other potential metals, which can then be more easily decontaminated radiologically for reuse. Another aim of the invention is also to recover the arsenic in order to reduce the environmental footprint of the ore processing.

[0019] DESCRIPTION OF THE INVENTION

[0020] These goals are achieved by the invention which proposes a process for concentrating arsenic in a radiologically decontaminated fraction in order to extract the arsenic and to be able to recover the critical metals associated with the arsenic.

[0021] To this end, the invention relates to a process for recovering an arsenic-enriched concentrate from an ore comprising uranium, radium, and arsenic, said process comprising:

[0022] Oxidative leaching A of said ore;

[0023] The separation B of a uranium-enriched L1 leachate and a radium- and arsenic-enriched R1 leaching residue;

[0024] Washing C of residue R1 leading to residue R2;

[0025] The treatment of residue R2 by particle size separation D leading to a fraction F1 of fine particles enriched in radium and a fraction F2 of coarse particles enriched in arsenic;

[0026] Collection E of fraction F2;

[0027] The flotation treatment F of fraction F2 leads to a floated product enriched in arsenic and a sunken product; and

[0028] The collection G of the flotation product obtained, which flotation product consists of the arsenic-enriched concentrate.

[0029] It was thus discovered that particle size separation and flotation of the resulting coarse particle fraction substantially enriched the sample with arsenic, while simultaneously decontaminating it of radium, which is predominantly found in the fine particle fraction. The process according to the invention therefore optimizes the recovery of a decontaminated, arsenic-enriched concentrate.

[0030] The ore

[0031] The ore is typically a uranium ore, mined for uranium extraction. Typically, this uranium ore contains 0.1 to 10% arsenic by mass, specifically 1 to 5% by mass relative to the total mass of the ore. The ore is also radiologically contaminated by radium, particularly radium-226.

[0032] The ore may also include one or more other metals, including valuable metallic elements such as cobalt, copper, and nickel.

[0033] Leaching

[0034] The leaching stage refers to the treatment of the ore, generally in powder form, to solubilize the metallic elements for extraction. This leaching typically involves treating the ore with an aqueous solution of acid, particularly sulfuric acid, often under agitation (dynamic leaching).

[0035] Typically, uranium ore is crushed and then ground by mechanical processes, then leached by oxidative attack in an acidic environment, under agitation in tanks, for several hours.

[0036] Leaching leads to an L1 leachate, enriched in uranium, and an R1 residue enriched in arsenic and radium, and possibly including other metallic elements such as nickel, cobalt and / or copper.

[0037] The L1 leachate can typically be extracted using organic solutions to extract uranium, then be precipitated and calcined to obtain a uranium concentrate (yellowcake) in the form of, for example, magnesium uranate.

[0038] Washing

[0039] The R1 residue is then washed, typically to recover the uranium carried along in the leaching residues. This washing process results in a residue R2. Residue R2 typically takes the form of a pulp or sludge, thicker than R1 and washed of its impregnating agent.

[0040] Typically, it contains the same arsenic content as R1.

[0041] According to one embodiment, residue R2 comprises an arsenic content of between 0.1 and 15%, in particular between 0.1 and 5%, and even more specifically from 0.1 to 0.5% (by mass).

[0042] This washing step can typically be carried out by decantation, in particular by counter-current decantation.

[0043] The term "counter-current settling" as used here refers to settling carried out in one or more successive cycles. The number of cycles can be adjusted according to the nature of the residue R1 and / or the desired degree of purity of the residue R2. Thus, settling can, for example, be carried out using 1 to 15 washing cycles, particularly between 3 and 10 cycles, advantageously using 6 cycles.

[0044] The settling is said to be counter-current in that the lower phase of cycle i is collected to be used in the following cycle i+1, while the supernatant (upper phase) of cycle i+1 is recovered to be reinjected into cycle i.

[0045] Particle size separation

[0046] The R2 residue then undergoes particle size separation, aimed at separating the population of the smallest particles, called "fines", from the coarse particles.

[0047] The term "particle size separation" used here is not limited to separation on a sample in solid form, and also refers to the separation of particles present in the form of sludge or pulp.

[0048] This particle size separation can typically be achieved by desliming (also called deslimmage or "desliming"). Desliming involves using a classifier (such as a hydrocyclone) to separate the finest particles from the coarser particles.

[0049] The hydrocyclone is widely used to efficiently separate pulp particles according to their size and density. Thanks to centrifugal force, fine particles cluster towards the center of the hydrocyclone and are ejected into the overflow, while larger particles are collected in the underflow. The principle of hydrocycloning relies on the tangential injection of the mixture under pressure: the centrifugal force applied to this mixture induces separation between the particles in the cylindrical and conical sections of the hydrocyclone. The coarsest and densest particles, constituting fraction F2, are discharged into the underflow, while the finest and least dense particles, constituting fraction F1, are discharged into the overflow.

[0050] Advantageously, hydrocyclone separation offers great flexibility of adjustment, a small footprint, low energy consumption and allows optimization according to the specific characteristics of the pulp.

[0051] Typically, several successive cycles on hydrocyclone can be conducted, if necessary, for example between 1 and 5 cycles, typically 2 cycles.

[0052] In one embodiment, the hydrocyclone parameters are adapted to achieve a particle size reduction of 20 µm. Typically, these parameters include the opening of the underflow outlets (spigot in mm) and overflow outlets (vortex finder in mm), which can be adjusted according to other parameters of the instrument used, such as the hydrocyclone body angle and / or the injection pressure. For example, the underflow outlet opening can be between 3 and 5 mm, particularly between 4 and 5 mm, for an overflow outlet opening of 8 mm, an angle of 5°, and an injection pressure between 1 and 5 bar, particularly between 2 and 3 bar.

[0053] According to the invention, particle size separation, in particular by hydrocyclone, allows separation between fine particles depleted in arsenic and comprising radium and coarse particles depleted in radium and comprising arsenic and possibly at least one of the valuable metallic elements, such as copper, cobalt and / or nickel.

[0054] As used here, the terms "fine particles" and "coarse particles" refer to the particle populations separated by the particle size distribution as defined here.

[0055] Thus, according to one embodiment, the F1 fraction of fine particles corresponds to a population of particles whose diameter D90 (by number) is less than or equal to 20 pm, which means, in other words, that 90% by number of particles have a diameter less than 20 pm and, conversely, 10% by number of particles have a diameter greater than 20 pm. More specifically, for particles with a diameter less than or equal to 20 pm, this can range from 5 to 20 pm and, even more specifically, from 10 to 20 pm.

[0056] According to one embodiment, the fraction F2 of coarse particles corresponds to a population of particles whose diameter D10 (by number) is greater than 20 pm, which means, in other words, that 90% by number of particles have a diameter greater than 20 pm and, conversely, 10% by number of particles have a diameter less than 20 pm, in particular less than 15 pm, particularly less than 10 pm.

[0057] Particle size and diameter D10 can be measured by a laser diffraction particle size analyzer.

[0058] The particle size separation step may be followed by one or more dilution steps of the F2 and / or F1 fractions, aimed at diluting the percentage of solids from the fractions of each cyclone cycle. Typically, the dilution aims to achieve a concentration between 5 and 35%. Thus, according to one embodiment, the process may include diluting the F2 fraction after the F2 fraction collection step at the outlet of the particle size separation.

[0059] Flotation

[0060] The term "flotation" refers to frothing flotation, generally used as a mineral separation technique in the mining and metallurgical industry.

[0061] It relies on the ability of air bubbles injected into an aqueous suspension containing mineral particles to collect the mineral particles at the air / water interface and make them float on the surface of the aqueous suspension, thus forming a foam of recoverable minerals.

[0062] More specifically, the foam flotation process is based on the surface properties of mineral particles. Mineral particles have electrically charged surfaces, and this charge can be positive or negative depending on the pH of the solution. Air bubbles injected into the aqueous suspension adhere to the surfaces of the electrically charged mineral particles. The adhesion of the air bubbles to the surfaces of the mineral particles is enhanced using chemical reagents called collectors, which are amphiphilic and have an affinity for the surfaces of the targeted mineral particles.

[0063] Thus, according to one embodiment, collectors can be added to the aqueous suspension to increase the hydrophobicity of the mineral particle surfaces. The air bubbles adhere more readily to the hydrophobic surfaces of the mineral particles, forming a mineral foam that floats on the surface of the aqueous suspension.

[0064] Other chemical reagents can also be used to control the size and stability of air bubbles in the aqueous suspension.

[0065] In particular, foaming agents are generally added to the aqueous suspension to increase the stability of the foam.

[0066] pH regulators can also be used to maintain the pH of the aqueous suspension at an optimal value for the flotation process.

[0067] The foam flotation process typically comprises the following steps: In a conditioning step, the F2 fraction is brought into contact with a flotation agent, also called a collector, and a foaming agent. The contact time varies and typically ranges from approximately 5 to 15 minutes. The resulting mixture is then transferred to a cell for the flotation step.

[0068] During this step, the mixture is agitated and mixed with an airflow to promote foam formation and increase the surface area for exchange between the air and the pulp. The arsenic concentrate is then recovered by scraping the surface of the foam, and the arsenic-depleted fraction is collected under the overflow.

[0069] In a reconditioning step, the arsenic-depleted fraction recovered from the underflow is directed into an agitated conditioning tank where collector and foaming agent are added again.

[0070] The sequence of these steps can be carried out successively, over several cycles.

[0071] The recovered mineral particles can then be washed and dried before being transported for further processing.

[0072] The term "floating product" here refers to the floating material containing mineral particles, present on the surface of the flotation cell. This material is generally in the form of a foam.

[0073] More specifically, the flotation step according to the invention may use one or both of the following reagents:

[0074] As a collector, we can mention Potassium Amyl Xanthate (PAX or AERO 350) generally used at slightly acidic pH (around pH 4), or the AEROPHI NE 3418A, 3302 and MX-5160 collectors, marketed by Solvay.

[0075] The typical concentration of the collector is between 50 and 400 g per tonne, specifically between 100 and 300 g / tonne, relative to the total mass of fraction F2.

[0076] As a potential pH regulator, bases can be used if an increase in the pH of the mixture being treated is desired, which is generally very acidic after previous treatment steps, such as leaching. For example, the pH of fraction F2 is typically close to 1. Bases also allow for intensifying or reducing the water-repellent effect of the collector on the surfaces of the mineral particles. This allows flotation collectors to act selectively on certain mineral particles.

[0077] Lime or sodium hydroxide can be cited in particular as pH regulators.

[0078] Typically, the pH during the flotation stage is below 9, specifically below 6, particularly below 5, and most commonly between 1 and 4.5. The type of foaming agent is not critical. Examples of foaming agents include OREPREP X-13, AEROFROTH 65, and MIBC, marketed by Solvay. Some foaming agents may have collecting properties that can improve flotation yields. The typical concentration of foaming agent can range from 1 to 300 g per tonne, specifically from 10 to 100 g per tonne of fraction F2.

[0079] Activators and / or depressants can be used to activate certain mineral forms at the expense of others, in order to improve the selectivity of element recovery. Copper sulfate as an activator or calcium hydroxide Ca(OH)2 as a depressant can be considered to improve yields and / or minimize reagent consumption.

[0080] According to one embodiment, the flotation step includes mixing the F2 fraction with one or more agents selected from foaming agents, collecting agents, activating agents, depressant agents, pH regulators and mixtures thereof, in aqueous medium, and blowing air bubbles into the mixture.

[0081] The flotation step can be carried out in one or more cycles. Thus, in the case of (i) cycles, it can be carried out in i successive flotation reactors, each pass being conducted as described above, it being understood that for each reactor, the floated product is collected giving rise to a residual mixture.

[0082] According to one embodiment, the flotation treatment F comprises successive passage through i flotation reactors, such that i is an integer between 2 and 4, in particular 2, and such that the residual mixture obtained at reactor (i-1) is fed into reactor (i), and the flotation product obtained for each flotation reactor is collected.

[0083] According to a preferred embodiment, the process according to the invention comprises: Oxidative leaching A of an ore comprising uranium, radium, arsenic, and possibly comprising valuable metallic elements such as cobalt, nickel and / or copper, by sulfuric acid;

[0084] The separation B of the uranium-enriched L1 leachate and of a radium- and arsenic-enriched R1 leachate residue possibly containing valuable metallic elements such as cobalt, nickel and / or copper; The solvent extraction of said L1 leachate, leading to a uranium-rich organic phase;

[0085] The C washing by counter-current settling comprising 6 successive cycles of residue R1 leading to residue R2;

[0086] The treatment of residue R2 by desliming D by hydrocyclone leading to a fraction F1 of fine particles of diameter D90 (in number) less than or equal to 20 pm enriched in radium and a fraction F2 of coarse particles of diameter D10 greater than 20 pm enriched in arsenic, and possibly including valuable metallic elements, such as cobalt, nickel and / or copper and possibly including radium;

[0087] Neutralization H of fraction F1 and of said lower phase with ferric sulfate (Fe2(SO4)s) in excess relative to arsenic;

[0088] Collection E of fraction F2;

[0089] The flotation treatment F comprises 3 flotation cycles of the F2 fraction in the presence of a foaming agent and a collecting agent; and

[0090] The collection G of the floated product obtained enriched in arsenic and, where applicable, in nickel, copper and / or cobalt.

[0091] Advantageously, the flotation step of the F2 fraction of the coarse particles leads to a floated product, which is a sulfide concentrate containing arsenic and possibly one of the aforementioned elements (nickel, cobalt and / or copper), and possibly radium.

[0092] The said concentrate is enriched in arsenic, and depleted in radium, the terms "enriched" and "depleted" being understood in relation to the starting ore and / or in relation to products R1, R2 and F2.

[0093] Typically, the flotation product comprises at least 70% (by mass), including at least 80% of the arsenic present in the original ore, while the residual flotation mixture comprises less than 20% (by mass), including less than 10% of the arsenic present in the F2 fraction.

[0094] A typical composition of the floated product includes (by mass relative to the total mass of floated product):

[0095] - 10 to 30% arsenic;

[0096] - from 20 Bq / g to 800 Bq / g of radium;

[0097] - 0.01 to 0.05% uranium;

[0098] - 0 to 15% nickel; - 0 to 20% copper; and

[0099] - Less than 5% cobalt.

[0100] Advantageously, the process according to the invention makes it possible to recover at least 60%, preferably at least 70%, in particular more than 80% of the arsenic present in the ore.

[0101] If necessary, the recovered floated product can then be recovered for the metallic elements it contains (Cobalt, nickel, copper).

[0102] It can also be stored or buried.

[0103] If the concentration of radium in this concentrate is too high to allow its downgrading and exploitation by a refiner, it is possible to partially decontaminate it (i) by selectively dissolving the radium and then precipitating the radium or (ii) by removing the remaining fine particles carried in the flotation product by a sieving and / or hydrocyclone step.

[0104] Additional steps

[0105] The process of the invention may also include one or more additional processing steps, typically implemented in ore processing, such as precipitation, filtration, calcination, washing, drying and / or packaging.

[0106] The process may also include one or more treatment steps for the fractions and residues generated, with a view to their recovery and / or residue disposal. This may notably concern the L1 leachate generated by the leaching step, the F1 fine particle fraction generated by the particle size separation step, and / or the residual mixture after flotation.

[0107] Typically, fraction F1 can be neutralized before its removal. According to one embodiment, the process further comprises the neutralization H of fraction F1 with ferric sulfate (Fe2(SC>4)3) in excess relative to arsenic, in particular such that the ratio (by mass) of iron to arsenic is between 10 and 20, preferably between 12 and 16.

[0108] The L1 leachate, enriched in uranium during leaching, is used in the uranium industry. Thus, the L1 leachate can typically be extracted using one or more solvents or organic solutions to separate a uranium-enriched upper phase from a residue (lower phase).

[0109] The upper phase can be subjected to precipitation and calcination to obtain a uranium concentrate ("yellowcake"), for example, in the form of magnesium uranate.

[0110] The residue from solvent extraction (lower phase) can be neutralized before being removed.

[0111] Thus, according to one embodiment, the process according to the invention may further comprise:

[0112] - collection of leachate L1;

[0113] - solvent extraction of said L1 leachate, leading to an upper phase enriched in uranium and a lower phase.

[0114] Typically, said lower phase can be neutralized with ferric sulfate (Fe2(SC>4)3) in excess relative to arsenic, in particular such that the ratio (by mass) of iron to arsenic is between 1 and 5, typically 3.

[0115] Thus, advantageously, said lower phase can be added to the neutralization step H of fraction F1.

[0116] The residual mixture from flotation can also be recovered for neutralization.

[0117] According to one embodiment, the process according to the invention can therefore include collecting said residual flotation mixture and neutralizing it with ferric sulfate (Fe2(SC>4)3) in excess relative to arsenic, in particular such that the ratio (by mass) of iron to arsenic is between 10 and 14, preferably between 1 and 13

[0118] The invention also relates to a product that can be obtained by a process according to the invention, comprising (by mass relative to the total mass of floated product): - from 10 to 30% arsenic;

[0119] - from 20 Bq / g to 800 Bq / g of radium;

[0120] - 0.01 to 0.05% uranium;

[0121] - 0 to 15% nickel;

[0122] - 0 to 20% copper; and

[0123] - Less than 5% cobalt.

[0124] BRIEF DESCRIPTION OF FIGURES [Fig. 1] Figure 1 is a diagram illustrating the conventionally used uranium ore processing method, by oxidative leaching and involving neutralization of the aqueous phase including leached arsenic as well as washed leaching mine residues including unleached arsenic.

[0125] [Fig. 2] Figure 2 is a diagram illustrating an illustrative embodiment of the process according to the invention.

[0126] [Fig. 3] Figure 3 represents the results of an embodiment of the process according to the invention described in the examples below, in terms of mass yield (%) represented by the curve, and arsenic recovery yield represented by the histograms.

[0127] [Fig. 4] Figure 4 represents the results of an embodiment of the process according to the invention described in the examples below, in terms of mass yield (%), and arsenic recovery yield, including arsenic from the coarse fractions and from exhaustion.

[0128] DETAILED DESCRIPTION OF SPECIFIC IMPLEMENTATION METHODS

[0129] The invention will become clearer upon reading the following description, given solely by way of non-limiting example, and made with reference to Figure 2, as well as in comparison with the conventionally used treatment process illustrated in Figure 1. An ore 1, typically a uranium ore, comprising arsenic is subjected to oxidative leaching A with sulfuric acid and separation B of a leachate L1 and a residue R1.

[0130] The uranium-rich L1 leachate continues the uranium extraction pathway, by extraction I with an organic solution, giving rise to an upper phase 6 comprising uranium and a lower phase 5 potentially comprising arsenic.

[0131] Phase 6 can then be processed by precipitation and calcination to obtain a uranium concentrate (yellowcake), for example, in the form of magnesium uranate. Phase 5 is intended for disposal and storage in a waste management facility. Prior to this, it can undergo hydrogen neutralization with ferric sulfate (Fe2(SC>4)3), typically in an Fe / As ratio of approximately 3.

[0132] The leaching residue R1 enriched in arsenic is then washed with industrial water, for example, by several counter-current settling cycles C (for example 6 cycles), at the end of which a residue R2 is obtained.

[0133] Typically, and as illustrated in Figure 1, mine tailings containing unleached arsenic R2 are subjected to neutralization H, typically consisting of the addition of a mixture 3 of lime and ferric iron (Fe 3+) in the form of ferric sulfate (Fe2(SC>4)3) in excess of arsenic: Ferric sulfate is added in an Fe / As ratio of 14. Then, the resulting residues can be disposed of and stored in a waste management facility.

[0134] According to the invention, and as illustrated in Figure 2, the residue R2, which is generally in the form of a pulp, will be subjected to a desliming D by hydrocyclone, allowing the separation E by centrifugal force of a fraction F1 of fine particles in overflow, and a fraction F2 of coarse particles in underflow, the fraction F1 being depleted in arsenic, while the fraction F2 is enriched in arsenic.

[0135] Advantageously, a particle size cut-off at 20 pm allows the separation of an F2 fraction comprising approximately 90% of the committed arsenic.

[0136] Fraction F1 is intended to be disposed of and stored in a waste management facility. Prior to this, it may be neutralized with a mixture of ferric sulfate (Fe2(SC>4)3) and lime, typically in an Fe / As ratio of approximately 14.

[0137] Fraction F2 then undergoes separation F by flotation through successive passages in several flotation reactors, for example, 3 cycles, as shown here, in the presence of a foam-collecting agent and possibly a foaming agent. At the surface of each flotation reactor, the floated product is recovered by scraping G: the sum of the floated products leads to the resulting concentrate 2, enriched in arsenic, which can be recovered for any metals it may contain (not shown here).

[0138] The residual mixture 4 from the last flotation reactor is collected: residue 4 is intended for disposal and storage in a waste management facility. Prior to this, it may undergo H₂ neutralization with ferric sulfate (Fe₂(SC>4)₃) (not shown here).

[0139] EXAMPLES

[0140] 1. Treatment of thickened mining tailings by hydrocyclone:

[0141] 1.1 Characteristics of mining residues before hydrocyclone treatment The mining residues considered are those from the treatment by counter-current washing (6 cycles), designated R2 at a rate of 12 kg and presenting an arsenic content of approximately 0.2%.

[0142] These residues were analyzed by Scanning Electron Microscopy (SEM) coupled with Energy Dispersive Systems (EDS) to determine their chemical composition at different measurement points. The mineral analysis results indicate that arsenic, nickel, cobalt, and copper are primarily bound to the sulfide in the following mineral forms:

[0143] Gersdorffite: NiAsS

[0144] Cobaltite: CoAsS

[0145] Chacolcite: CuS2

[0146] 1.2 Treatment of residues by hydrocyclone - Results

[0147] The objective of the study is to treat a mine tailings pulp by hydrocyclone (residue R2) in order to separate arsenic from radium.

[0148] The treatment is carried out in two hydrocyclone cycles (marketed by Salters Cyclone). After separation and at the end of the first cycle, the coarse fraction at the underside of the cyclone is recovered, diluted, and returned to the hydrocyclone for a second cycle. The purpose of this second cycle is to refine the particle size separation between the coarse and fine fractions to around 20 µm.

[0149] The protocol is as follows:

[0150] Cyclone cycle 1

[0151] A sample of residue R2 is taken for analysis for U, As, Co, Ni, Fe and Ra. The average solids content of the pulp is measured by measuring pulp density and solids density.

[0152] The particle size distribution of the residue is measured by laser granulometry.

[0153] The pulp is diluted by adding industrial water to achieve a maximum solids content of 20%. The hydrocyclone pump is then activated to obtain a pressure of 2 to 3 bar at the cyclone inlet. Once this pressure is reached, the overflow and underflow are collected in separate tanks.

[0154] A representative sample is taken from the fine and coarse fractions for the characterization of particle size and solids content.

[0155] The net mass of the fine and coarse fractions of the residue is measured.

[0156] A sample of residues from the coarse fraction is taken for analysis of U, As, Co, Ni, Fe and Ra.

[0157] Cyclonic cycle 2: The particle size distribution can be refined by performing a second particle size separation on the coarse fraction of the residues.

[0158] The coarse fraction is diluted to achieve a solids content of between 15 and 20%.

[0159] The procedure from cycle 1 is then repeated.

[0160] A representative sample is taken from the fine and coarse fractions for the characterization of particle size and solids content.

[0161] The net mass of the fine and coarse fractions of the residue is measured.

[0162] A sample of residues from the coarse fraction is taken for analysis of U, As, Co, Ni, Fe and Ra.

[0163] The results presented below in Table 1 were obtained after 2 successive cycles on hydrocyclone.

[0164] [Table 1]

[0165]

[0166]

[0167] SB: No cyclone body

[0168] The particle size distribution of the initial residue R2 and of the fractions F1 and F2 after 2 separation cycles is analyzed for each opening of the sub-flow.

[0169] The results demonstrate the achievement of a particle size cut-off between 5 and 20 µm. The results obtained after hydrocyclone treatment demonstrate that the target was met. After treatment, the coarse fraction obtained represented approximately 60% of the total mass of the initial residue.

[0170] The arsenic content between the initial and final products increased from 2.83% to between 3.72% and 5.34%, depending on the test. During these tests, the arsenic recovery rate for the coarse fraction averaged 86%. Therefore, approximately 90% of the arsenic is found in fraction F2, which represents 50% to 60% of the initial residue mass.

[0171] Furthermore, the analysis of radium in the different fractions produced confirms that 226Ra is mainly contained in the fine fraction of the residue with a recovery of 90% of the radium in about 40% of the residue mass.

[0172] These tests were repeated numerous times on different batches of R2 residue. The repeatability of the results demonstrated the robustness of this process. This particle size reduction allows for a satisfactory arsenic recovery yield (approximately 90% by mass) compared to the mass loss of the residue (approximately 50% of mass recovered).

[0173] Several kg of R2 pulp were thus descaled using a hydrocyclone with a 4.5 mm underflow opening to validate the robustness and reproducibility of the particle size separation. The results are summarized in Table 2.

[0174] [Table 2]

[0175]

[0176] A production campaign yielded 7 batches of delimed fraction F2 by processing approximately 130 kg of residue R2 with a solids content of 23%. Analysis of the particle size distributions summarized in Table 2 demonstrates very good reproducibility.

[0177] 2. Treatment of residues by flotation:

[0178] 2.1 Context

[0179] The mineralogical and elemental analysis of the R2 mining residues and the F2 fraction obtained after the hydrocyclone particle size separation stage indicates that arsenic, nickel, cobalt, and copper are mainly bound to the sulfide in the following mineral forms:

[0180] Gersdorffite: NiAsS

[0181] Cobaltite: CoAsS

[0182] Chacolcite: CuS2

[0183] 2.2 Methodology

[0184] In a conditioning step, the coarse particle fraction F2 is brought into contact with a flotation agent, also called a collector, and a foaming agent. The contact time varies and typically ranges from approximately 5 to 15 minutes.

[0185] The resulting mixture is then transferred into a cell for the flotation step.

[0186] During this step, the mixture is agitated and mixed with an airflow to promote foam formation and increase the surface area for exchange between the air and the pulp. The arsenic concentrate is then recovered by scraping the surface of the foam, and the arsenic-depleted fraction is collected under the overflow.

[0187] In a reconditioning step, the arsenic-depleted fraction recovered from the underflow is directed into an agitated conditioning tank where collector and foaming agent are added again.

[0188] Collector:

[0189] Product name: Potassium Amyl Xanthate (PAX) or Aerophine 3418A Concentration: 100 to 200 g / t of fraction F2

[0190] Conditioning time: 5 to 10 minutes

[0191] Foaming:

[0192] Designation: Aerofroth 65 or MIBC

[0193] Concentration: 25 to 100 g / t of fraction F2

[0194] Operating conditions:

[0195] Initial percentages: from 10 to 50%

[0196] pH: not controlled but measured between pH 1 and 3

[0197] Three successive flotation cells: the product not floated by cell 1 is introduced and reprocessed in cell 2 and then in cell 3

[0198] Residence time: 10 minutes per flotation cell

[0199] The operating conditions of the tests are summarized in Tables 3 and 4: [Table 3]

[0200]

[0201]

[0202] [Table 4]

[0203]

[0204] An illustrative protocol is detailed below:

[0205] Settings:

[0206] - Conditioning time: 5 minutes

[0207] - Retention time for roughing, exhaustion 1 and 2: 10 minutes

[0208] - Collector and concentration: AEROPHI NE 3418A and 150 g / t

[0209] - Foaming agent and concentration: AEROFROTH 65 and 50 g / t

[0210] Preparation and packaging:

[0211] A 10 g / L solution of AEROPHINE 3418A is prepared. The solid content of the pulp is adjusted to a maximum of 25% by adding industrial water. The flotation cell is filled to the level preceding the recovery slope. The pulp is agitated with the residue to float in the flotation cell. The redox potential (Eh) and pH are characterized using a glass electrode and Ag / AgCl electrodes. A 10 g / L solution of Aerophine 3418A is added to the pulp to achieve a concentration of 150 g / t of fraction F2. The mixture is left under agitation for 5 minutes.

[0212] An Aerofroth 65 solution is added to achieve a concentration of 50 g / t of residue (ore).

[0213] Roughing stage:

[0214] The compressed air flow rate is adjusted to optimize bubble formation (ideally large and well-filled bubbles). The floating fraction is regularly collected until the surface of the bubbles clears, indicating a lack of floating material or a missing collector. The foam is then filtered, washed, and dried for 24 hours at 150°C. ère stage :

[0215] 75 g / t of Aerophine 3418A at 10 g / L are added to the non-floating paste. The compressed air flow rate is adjusted to optimize bubble formation. The floating fraction is collected regularly until the bubble surface clears, indicating a lack of floating material or insufficient collector. The foam is then filtered, washed, and dried for 24 hours at 150°C.

[0216] 2 ème stage :

[0217] 40 g / t of Aerophine 3418A at 10 g / L are added to the non-floating paste. The compressed air flow rate is adjusted to optimize bubble formation. The floating fraction is collected regularly until the bubble surface clears, indicating insufficient floating material or a lack of collector. The foam product is then filtered, washed, and dried for 24 hours at 150°C. The dry residue is filtered, washed, and weighed separately. Each fraction of the foam and residue is analyzed (As, Co, Ni, and Ra).

[0218] 2.3 Results

[0219] In this case, several flotation steps are coupled, thus allowing the recovery of nearly 100% of the arsenic contained in the coarse fraction F2 produced by the hydrocyclone in a mass fraction representing approximately 20% of the mass of residue introduced.

[0220] The results are presented in Figures 3 and 4.

[0221] The different parameters tested on the residue with low arsenic content indicate: ■ Conducting the particle size separation step and cutting at 20 pm upstream of the flotation step significantly improves arsenic recovery.

[0222] ■ The product obtained is a dark grey powder typical of flotation products with a high sulfide content.

[0223] ■ Increasing the collector concentration (from 150 g / t to 300 g / t) improved the arsenic flotation yield by 15% according to tests 15 and 16. A similar yield was observed for test 1 after the first depletion stage.

[0224] ■ The type (Aerofroth 65 vs MIBC) and concentration of the foaming agent (from 200 to 20 g / t) do not significantly affect flotation efficiency according to tests 15 and 17. The observed performance differences can be explained by the heterogeneity of the samples used for tests 15 and 17 and by the variability of operating conditions related to the experimenter.

[0225] ■ Increasing the solids content (from 17% w to 25% w) does not significantly affect flotation yields according to tests 17 and 18. This observation indicates that the collector is not limiting for flotation yield.

[0226] ■ Finally, decreasing the pH of the pulp from pH 6 to pH 4.5 increases flotation performance when using potassium amyl xanthate by increasing the flotation yield from 75% to 62% of the arsenic recovered.

[0227] At the end of the test, the concentrate obtained has on average the following composition:

[0228] [Table 5]

[0229]

[0230] 3. Conclusion Considering a particle size cut at 20 pm by hydrocyclone and a flotation step, an overall arsenic recovery yield of 85% is obtained for 5 to 10% of the mass of the initial residue introduced into the process.

[0231] Implementing this entire process allows for the removal of 85% of the unleached arsenic.

[0232] Ultimately, this concentrate can be sold and its value derived from its constituent elements. If necessary, the radium, the main contributor to the activity of this concentrate, can be selectively leached with a hydrochloric acid solution until the target radiological specifications are met.

Claims

24 DEMANDS 1. A process for recovering an arsenic-enriched concentrate from an ore comprising uranium, radium and arsenic, said process comprising: Oxidative leaching A of said ore; The separation B of a uranium-enriched L1 leachate and a radium- and arsenic-enriched R1 leaching residue; Washing C of residue R1 leading to residue R2; The treatment of residue R2 by particle size separation D leading to a fraction F1 of fine particles enriched in radium and a fraction F2 of coarse particles enriched in arsenic; Collection E of fraction F2; The flotation treatment F of fraction F2 leads to a floated product enriched in arsenic and a sunken product; and The collection G of the resulting floated product.

2. A process according to claim 1 wherein the flotation step comprises mixing fraction F2 with one or more agents selected from foaming agents, collecting agents, activating agents, depressant agents, pH regulators and mixtures thereof, in aqueous medium, and blowing air bubbles into the mixture.

3. A process according to claim 2 wherein the collecting agents are present in a concentration of between 50 g / t and 400 g / t, preferably between 100 and 300 g / t, relative to the total mass of fraction F2 4. A process according to any one of the preceding claims such that during the flotation step F, the pH is less than 9, preferably less than 6, in particular between 1 and 4.

5.

5. A process according to any one of the preceding claims wherein the flotation treatment F comprises successive passage through i flotation reactors, where i is an integer between 2 and 4, in particular 2, and where the residual mixture obtained at reactor (i-1) is fed into reactor (i), and the flotated product obtained for each flotation reactor is collected.

6. A process according to any one of the preceding claims wherein the particle size separation step D is carried out by desliming, preferably by hydrocyclone treatment.

7. A process according to any one of the preceding claims wherein the residue R2 comprises an arsenic content of between 0.1 and 15% (by mass), more specifically from 0.1 to 5% and even more specifically from 0.1 to 0.5%.

8. A method according to any one of the preceding claims wherein the fraction F2 corresponds to a population of particles whose diameter D10 (in number) is greater than 20 pm.

9. A method according to any one of the preceding claims, wherein it includes the dilution of fraction F2 after the step of collecting fraction F2.

10. A process according to any one of the preceding claims further comprising the neutralization H of fraction F1 with ferric sulfate (Fe2(SC>4)3) in excess relative to arsenic, in particular such that the ratio (by mass) of iron to arsenic is between 10 and 20, preferably between 12 and 16.

11. A process according to any one of the preceding claims further comprising collecting the residual mixture from flotation and neutralizing it with ferric sulfate (Fe2(SC>4)3) in excess relative to arsenic, in particular such that the ratio (by mass) of iron to arsenic is between 10 and 14, preferably between 11 and 13.

12. A method according to any one of the preceding claims, further comprising: - collection of leachate L1; - solvent extraction of said L1 leachate, leading to an upper phase enriched in uranium and a lower phase.

13. A process according to claim 12, wherein the lower phase is neutralized with ferric sulfate (Fe2(SC>4)3) in excess relative to arsenic, in particular wherein the ratio (by mass) of iron to arsenic is between 1 and 5.

14. Process any one of the preceding claims such that the ore further comprises one or more metallic elements, such as cobalt, copper, nickel.

15. A process according to any one of the preceding claims allowing the separation of at least 60%, preferably at least 70%, in particular more than 80% of the arsenic present in said ore.

16. A method according to any one of the preceding claims comprising: Oxidative leaching A of an ore comprising uranium, radium, arsenic, cobalt, nickel and copper by sulfuric acid; The separation B of the uranium-enriched L1 leachate and of a radium- and arsenic-enriched R1 leaching residue possibly including valuable metallic elements such as cobalt, nickel and / or copper; Solvent extraction of said L1 leachate, leading to an organic phase rich in uranium; The C wash by counter-current settling comprising 6 successive cycles of residue R1 leading to residue R2; The treatment of residue R2 by desliming D by hydrocyclone leading to a fraction F1 of fine particles of diameter D90 (in number) less than or equal to 20 pm enriched in radium and a fraction F2 of coarse particles of diameter D10 greater than 20 pm enriched in arsenic, and possibly including valuable metallic elements, such as cobalt, nickel and / or copper and possibly including radium; Neutralization H of fraction F1 and of said lower phase with ferric sulfate (Fe2(SC>4)3) in excess relative to arsenic; Collection E of fraction F2; The flotation treatment F comprises 3 flotation cycles of the F2 fraction in the presence of a foaming agent and a collecting agent; and The collection G of the obtained float product enriched in arsenic, nickel, copper and cobalt.

17. Product obtainable by the process according to any one of the preceding claims, said product comprising (by mass relative to the total mass of floated product): - 10 to 30% arsenic; 27 - from 20 Bq / g to 800 Bq / g of radium; - 0.01 to 0.05% uranium; - 0 to 15% nickel; - 0 to 20% copper; and - Less than 5% cobalt.

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