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WO2013040694A1 - Procédé de valorisation de minerais et de concentrés de tantale et de niobium avec récupération d'oxydes de manganèse et de terres rares - Google Patents

Procédé de valorisation de minerais et de concentrés de tantale et de niobium avec récupération d'oxydes de manganèse et de terres rares Download PDF

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WO2013040694A1
WO2013040694A1 PCT/CA2012/000890 CA2012000890W WO2013040694A1 WO 2013040694 A1 WO2013040694 A1 WO 2013040694A1 CA 2012000890 W CA2012000890 W CA 2012000890W WO 2013040694 A1 WO2013040694 A1 WO 2013040694A1
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acid
caustic
leaching
alkali
solution
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Francois Cardarelli
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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/12—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic alkaline solutions
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/22—Treatment or purification of solutions, e.g. obtained by leaching by physical processes, e.g. by filtration, by magnetic means, or by thermal decomposition
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00—Obtaining refractory metals
    • C22B34/20—Obtaining niobium, tantalum or vanadium
    • C22B34/24—Obtaining niobium or tantalum
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B47/00—Obtaining manganese
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B59/00—Obtaining rare earth metals
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B60/00—Obtaining metals of atomic number 87 or higher, i.e. radioactive metals
    • C22B60/02—Obtaining thorium, uranium, or other actinides
    • C22B60/0204—Obtaining thorium, uranium, or other actinides obtaining uranium
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B60/00—Obtaining metals of atomic number 87 or higher, i.e. radioactive metals
    • C22B60/02—Obtaining thorium, uranium, or other actinides
    • C22B60/0291—Obtaining thorium, uranium, or other actinides obtaining thorium
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • the present specification broadly relates to a process for upgrading tantalum and niobium ores and concentrates. More specifically, but not exclusively, the present specification relates to a process for upgrading tantalum and niobium ores and concentrates with the recovery of manganese and rare earth oxides and the electrochemical regeneration of reactants.
  • Tantalum and to a lesser extent niobium are two scarce refractory metals of the group VB(5) with a relative abundance in the Earth's crust of two parts per million by weight and twenty parts million by weight respectively. Both metals are used in various civilian and military applications.
  • niobium and tantalum minerals mined today constitute essentially the niobiate and tantalate of iron, manganese and tin. These minerals belong to three major groups; namely the pyrochlore, the columbite and wodginite groups.
  • the pyrochlore group minerals exhibit a cubic crystal lattice structure having the general formula: AB 2 0 6 (0,OH,F) wherein A is at least one of K + , Cs + , Ca 2+ , Sr 2+ , Ba 2+ , Ln 3+ ,Pb 2+ and U 4+ ; and B is at least one of Nb 5+ , Ta 5+ , Ti 4+ , Sn 4+ and W 6+ .
  • the pyrochlore (sensu stricto) [(Na,Ca)2(Nb,Ta,Ti) 2 04(OH,F) H 2 0 is of paramount commercial importance as the main niobium ore. When tantalum fully replaces niobium, the mineral is called microlite.
  • the columbite group minerals exhibit an orthorhombic crystal lattice structure having the general formula AB 2 0 6 wherein A is at least one of Fe 2+ , Mn 2+ and Mg 2+ ; and B is at least one of Nb 5+ and Ta 5+ with complete isomorphic substitutions between the end members.
  • the tantalum-rich end is called tantalite [(Fe,Mn)Ta 2 0 6 )] and the niobium-rich end is named columbite [(Fe,Mn)Nb 2 0 6 ] with a continuous gradation between them yielding the common name columbo-tantalite or coltan in short that is extensively used in trade. Sometimes their names are preceded by the prefixes ferro-, mangano-, or magno- for indicating the predominance of one of the divalent cations.
  • the wodginite group minerals exhibit a monoclinic crystal lattice structure having the general formula ABC 2 0 8 wherein A is at least one of Mn 2+ , Fe 2+ and Li + ; B is at least one of Sn 2+ , Ti 4+ , Fe 3+ and Ta 5+ ; and C is at least one of Nb 5+ and Ta 5+ .
  • A is at least one of Mn 2+ , Fe 2+ and Li +
  • B is at least one of Sn 2+ , Ti 4+ , Fe 3+ and Ta 5+
  • C is at least one of Nb 5+ and Ta 5+ .
  • wodginite (Mn,Fe)(Sn,Ta)(Ta,Nb) 2 0 8 ] is of commercial importance.
  • Tantalum ores containing tantalum minerals such as tantalite, wodginite and euxenite are usually found in acidic igneous host rocks such as granite and pegmatite as well as in weathered, residual or alluvial deposits derived from such parent igneous rocks.
  • silicate minerals such as albite, microcline, beryl, lepidolite, muscovite, spodumene, tourmaline and zircon, phosphates gangue minerals such as amblygonite, triphillyte and apatite as well as other heavy minerals such as cassiterite, wolframite, and samarskite, forming tantalum- and niobium-rich mineral sands.
  • Typical premier tantalum deposits are those of the Tanco pegmatite, Bernic Lake in the province of Manitoba, Canada and the Wodgina Mine in Western Australia.
  • niobium deposits are mainly based on pyrochlore and columbite minerals that are usually found in ultramafic igneous rocks such as anorthosites and carbonatites.
  • Typical premier pyrochlore deposits are those of the CBMM mine in Araxa, Brazil, the Lueshe deposit in the Congo, the Mabounie deposit in Gabon, and the two Canadian deposits of Niocan in Oka and the Niobec mine in Saint-Honore, both located in the province of Quebec, Canada.
  • the ore material typically undergoes on-site beneficiation by means of common mineral dressing techniques such as gravity separation and magnetic separation to remove the silicate gangue minerals.
  • the tantalum-rich concentrate is subsequently shipped to the metallurgical plant for further processing.
  • the tantalum-rich concentrate is subjected to a chemical decomposition step ⁇ i.e. dissolution step), typically performed using mixtures of concentrated hydrofluoric and sulfuric acids (HF- H 2 S0 4 ).
  • HF- H 2 S0 4 concentrated hydrofluoric and sulfuric acids
  • both tantalum and niobium are co-extracted by solvent extraction using methyl-iso-butyl-ketone (MIBK) dissolved in kerosene. Both MIBK and kerosene are highly flammable and noxious organic solvents.
  • MIBK methyl-iso-butyl-ketone
  • This extraction process provides for the selective extraction of both niobium and tantalum from deleterious manganese, iron, titanium, and other metallic impurities that are later discarded along with the original radionuclides in the processing wastes.
  • the niobium is then stripped from the tantalum by contacting the organic phase with dilute hydrochloric acid (HCI).
  • HCI dilute hydrochloric acid
  • tantalum to be transferred into the aqueous phase.
  • Tantalum can be recovered by two routes, depending on the final product required.
  • the tantalum may be precipitated as hydrated tantalum pentoxide or tantalic acid (Ta 2 0 5 .xH 2 0) by adding gaseous ammonia (NH 3 ).
  • the precipitate can then be calcined to yield pure tantalum pentoxide (Ta 2 0 5 ).
  • the tantalum can be crystalized as potassium heptafluorotantalate (K 2 TaF 7 ) by addition of potassium chloride (KCI) or potassium fluoride (KF) to a hot aqueous solution. Pure tantalum metal powder is obtained by sodiothermic reduction using sodium metal.
  • the liquid effluents (mostly waste waters) and solid wastes produced contain significant concentrations of fluoride ions. Since recycling of the fluorides at the extraction plant is not economical, these wastes are typically simply disposed of. The environmental problem is further compounded considering that all of the radionuclides originating from the ore, and that are responsible for its natural radioactivity, also end up in the wastes. These wastes are thus considered as technologically enhanced naturally occurring radioactive materials, (TENORM) that must be concentrated, stabilized and disposed of in accordance with local regulations, often at prohibitive costs.
  • TELM technologically enhanced naturally occurring radioactive materials
  • the present specification broadly relates to a process for upgrading tantalum and niobium ores and concentrates.
  • the present specification relates to a process for upgrading tantalum and niobium ores and concentrates, the process comprising: a) submitting the ore or concentrate to a caustic fusion or an alkali fusion using a melt comprising at least one salt of an alkali metal to produce a solidified melt; b) submitting the solidified melt to an alkaline leaching step to produce an alkaline pregnant leaching solution comprising manganese values; c) recovering the manganese values from the alkaline pregnant leaching solution to produce a substantially manganese-free leach solution; and d) precipitating tantalum and niobium as an insoluble niobiate and tantalate from the manganese-free leach solution.
  • the process further comprises separating the alkaline pregnant leaching solution from insoluble solid residues; recovering rare earth oxides and/or thorium oxides from the insoluble solid residues producing a spent acid filtrate; and recovering uranium from the spent acid filtrate.
  • the process further comprises acid-leaching the insoluble niobiate and tantalate producing a hydrated tantalum and niobium oxides; and calcining or roasting the hydrated tantalum and niobium oxides to yield tantalum metal and niobium oxides.
  • the present specification relates to a process for upgrading tantalum and niobium ores and concentrates with the recovery of manganese and recovery of rare earths oxides and the electrochemical regeneration of reactants.
  • the caustic fusion or alkali fusion comprises using a molten alkali-metal hydroxide of formula MOH, wherein M is selected from the group consisting of Li, Na and K.
  • the molten alkali- metal hydroxide is selected from the group consisting of potassium hydroxide, sodium hydroxide, lithium hydroxide and mixtures thereof.
  • the caustic fusion or alkali fusion is performed using a potassium salt.
  • the caustic fusion or alkali fusion is performed using a sodium salt.
  • FIG. 1 is an illustration of a flowchart in accordance with an embodiment of the present specification.
  • the caustic or alkali fusion is performed using a potassium salt such as potassium hydroxide.
  • the manganese value is recovered as potassium permanganate following alkaline leaching of the melt and electrochemical treatment of the pregnant leach solution.
  • FIG. 2 is an illustration of a flowchart in accordance with an embodiment of the present specification.
  • the caustic or alkali fusion is performed using a potassium salt such as potassium hydroxide.
  • the manganese value is recovered as manganese dioxide by precipitation following alkaline leaching of the melt.
  • FIG. 3 is an illustration of a flowchart in accordance with an embodiment of the present specification.
  • the caustic or alkali fusion is performed using a sodium salt such as sodium hydroxide.
  • the manganese value is recovered as manganese dioxide by precipitation following alkaline leaching of the melt.
  • FIG. 4 is an illustration of a flowchart in accordance with an embodiment of the present specification.
  • the rare earth values are recovered as rare earth oxides and thorium from the air-dried or calcined residues.
  • FIG. 5 is an illustration of the two-compartment electroiyzer and the electrodes and membrane reactions occurring inside the electroiyzer in accordance with an embodiment of the present specification.
  • the electroiyzer is used for the preparation of potassium permanganate along with the concurrent regeneration of the potassium hydroxide lye.
  • FIG. 6 is an image of a commercially available tantalite concentrate
  • concentrate or “ore concentrate” refers to a mineral product obtained after mineral dressing and beneficiation and from which most of the gangue minerals and waste rock have been removed and discarded as tailings. The concentrate is frequently the raw material for further metallurgical and chemical treatments such as roasting, leaching and smelting.
  • upgrading refers to a chemical or electrochemical process, typically part of an upstream metallurgical process, for removing major impurities and other deleterious impurities from an ore or a concentrate in order to improve the grade, that is, the concentration of the valuable(s) metal(s).
  • the present specification broadly relates to a process for upgrading tantalum and niobium ores and concentrates.
  • the process provides for substantially pure tantalum oxide, substantially pure niobium oxide and mixtures of these oxides to be produced.
  • the present specification relates to a pyrometallurgical and hydrometallurgical process for upgrading tantalum and niobium ores and concentrates with the concurrent electrochemical recovery of manganese metal, manganese oxides and other manganese-based materials as co-products and the regeneration of reactants.
  • the process also provides for the recovery of a product comprising the rare earth metal oxides.
  • the process comprises a high temperature caustic fusion or alkali fusion step followed by a hydrometallurgical alkaline leaching step with the concurrent electrochemical recovery of manganese metal or manganese-based materials, non- limiting examples of which include manganese dioxide and potassium permanganate, as co-products of the process.
  • the process further comprises precipitation steps, leaching steps and calcination steps ultimately yielding substantially pure tantalum oxide, substantially pure niobium oxide and mixtures of these oxides.
  • the process comprises steps for recovering the product comprising the rare earth metal oxides as well as steps for regenerating at least some of the reactants such as the alkaline leaching solution and the alkali metal hydroxides.
  • the process comprises a high temperature caustic fusion or alkali fusion step.
  • This high temperature pyrometallurgical step includes subjecting the tantalum or niobium ore or concentrate to a molten caustic fusion or a molten alkali fusion using a melt comprising an alkali-metal hydroxide.
  • the alkali-metal hydroxide has the general formula MOH, wherein M is selected from the group consisting of Li, Na and K. Accordingly, non-limiting examples of alkali-metal hydroxides include potassium hydroxide, sodium hydroxide, lithium hydroxide and mixtures thereof.
  • an alkali- metal peroxide is added to the melt.
  • the alkali-metal peroxide has the general formula M 2 0 2 , wherein M is selected from the group consisting of Li, Na and K.
  • M is selected from the group consisting of Li, Na and K.
  • the alkali- metal hydroxide and peroxide typically have the same M value (i.e. both are either Li, Na or K).
  • the addition of the peroxide ensures that essentially all of the tin content is oxidized to its tetravalent oxidation state (Sn 4+ ).
  • an alkali- metal carbonate of formula M 2 C0 3 or an alkali-metal nitrate of formula MN0 3 are added to the melt, wherein M is selected from the group consisting of Li, Na and K.
  • M is selected from the group consisting of Li, Na and K.
  • the melt comprises more than one alkali-metal hydroxide (alkali-metal hydroxide blend).
  • alkali-metal hydroxide blend the individual alkali-metal hydroxides can be melted individually and then combined, or melted together to produce the melt to be used in the fusion step. This prior melting drives-off all residual moisture and hydration water.
  • the alkali-metal hydroxide blend is allowed to cool prior to be used in the fusion step.
  • the tantalum or niobium ores and concentrates are ground prior to being used in the fusion step.
  • the grinding produces a particle size distribution not higher than 80-mesh. Fine grinding of the ores and concentrates is typically not required in order to avoid dusting.
  • the tantalum or niobium ores and concentrates are dried prior to being processed.
  • the ground product is dried prior to being processed.
  • the ground and dried material is fed directly into the melt once the desired operating temperature is reached and subsequently continuously stirred.
  • the ground and dried material is added on top of a solidified melt comprising at least one alkali-metal hydroxide followed by raising the temperature until melting of the mixture has occurred.
  • the ground and dried material is mixed with at least one alkali-metal hydroxide that has been previously ground followed by raising the temperature until melting of the mixture has occurred.
  • the caustic or alkali fusion step is performed using a dimensioniess mass ratio of ore or concentrate mass (C) to mass of melt material (i.e. mass of molten hydroxide salt) (S) denoted as C:S or C/S ranging from 1 :1 to 1 :20.
  • C ore or concentrate mass
  • S molten hydroxide salt
  • the caustic or alkali fusion step is performed using a dimensioniess mass ratio maintaining a low melt viscosity and allowing for substantially complete dissolution of the products.
  • the caustic or alkali fusion step is performed using a dimensioniess mass ratio ranging from 1 :1 to 1 :15.
  • the caustic or alkali fusion step is performed using a dimensioniess mass ratio ranging from 1 :1 to 1 : 10. In an embodiment of the present specification, the caustic or alkali fusion step is performed using a dimensioniess mass ratio ranging from 1 : 1 to 1 :6.
  • the caustic or alkali fusion step is performed at a temperature of at least the melting point of the alkali metal salt (i.e. alkali metal hydroxide). In cases where the melt comprises more than one alkali-metal hydroxide, the caustic or alkali fusion step is performed at a temperature of at least the eutectic temperature of the alkali mixture.
  • the melt temperature ranges from about 200°C to about 1200°C. In a further embodiment of the present disclosure, the melt temperature ranges from about 300°C to about 1000°C. In a further embodiment of the present disclosure, the melt temperature ranges from about 300°C to about 10O0°C.
  • the melt temperature ranges from about 400°C to about 900°C.
  • the operating temperature is sometimes increased in order to compensate for heat losses and to keep the charge fully liquid.
  • the maximum operating temperatures are typically dictated so as to prevent losses in molten salts by intense evaporation of caustic fumes and by the limited number of corrosion resistant materials commercially available in which to perform the fusion reaction.
  • the caustic or alkali fusion step is performed over a period of time ranging from 5 minutes to 6 hours. In a further embodiment of the present disclosure, the caustic or alkali fusion step is performed over a period of time ranging from 10 minutes to 4 hours. In a further embodiment of the present disclosure, the caustic or alkali fusion step is performed over a period of time ranging from 15 minutes to 2 hours. In a further embodiment of the present disclosure, the caustic or alkali fusion step is performed over a period of time ranging from 15 minutes to 1 hour.
  • the caustic or alkali fusion step is performed either batch wise using a crucible furnace or a muffle furnace or in continuous mode using a rotary kiln or a rotary heart furnace.
  • the caustic or alkali fusion step is performed by means of direct flame heating, gas fired burners, radiant gas heaters, external electrical heaters, Joule heating by immersed AC or DC electrodes, or by induction heating of the crucible used as susceptor.
  • Other suitable heating means for performing the caustic or alkali fusion step are known in the art, and are within the capacity of a skilled technician.
  • the caustic or alkali fusion step is performed using a containment vessel or crucible comprising a construction material capable of withstanding both the high operating temperatures as well as the inherent corrosiveness of the molten alkali-metal hydroxide(s) without contaminating the melt by releasing deleterious metallic impurities.
  • a containment vessel or crucible comprising a construction material capable of withstanding both the high operating temperatures as well as the inherent corrosiveness of the molten alkali-metal hydroxide(s) without contaminating the melt by releasing deleterious metallic impurities.
  • a certain number of materials suitable for use in the fusion step are described in the literature: i) Probst, H.B.; May, C.E.; and McHenry, H.T. - Corrosion resistance of Ni-alloys in molten NaOH. - National Advanced Committee for Aeronautics (NACA) Technical Note No.
  • Suitable materials include metals such as pure nickel (Ni), pure zirconium (Zr), and their alloys (e.g., Duranickel® 201 , Zircadyne® 702); pure iron (Fe), nickel cast iron, cast irons; mild steels; and non-metals such as graphite and carbon-based materials.
  • suitable materials include metallic composite materials comprising inexpensive bulk commercial alloys such as heat resistant stainless steels (e.g., AISI 310); copper-nickel (e.g. , Monel®) and high nickel-alloys (e.g., Hastelloys X® and Inconel® 601 ) coated with an inert, protective and impervious metal lining composed of a highly corrosion resistant pure metal or alloy.
  • heat resistant stainless steels e.g., AISI 310
  • copper-nickel e.g. , Monel®
  • high nickel-alloys e.g., Hastelloys X® and Inconel® 601
  • Protective lining materials include gold (Au), gold-alloys, silver (Ag), silver-alloys, nickel (Ni), nickel-alloys, iron (Fe), iron-alloys, zirconium (Zr), zirconium-alloys, hafnium (Hf), hafnium-alloys and combinations thereof.
  • the protective lining materials can be applied by various techniques such as electroplating, electroless plating, physical or chemical vapor deposition, mechanical cladding, loose lining and explosion bonding.
  • Yet further suitable materials include advanced ceramic materials useful as refractory brick linings, castables and coatings.
  • Non-limiting examples of such ceramic materials include graphite, diamond like carbon (DLC), carbon-carbon composites, silicon carbide (SiC, carborundum®), fused zirconia (Zr0 2 ), fused magnesia (MgO), fused ceria (Ce0 2 ), fused calcia (CaO) and combinations thereof.
  • DLC diamond like carbon
  • SiC silicon carbide
  • Zr0 2 fused zirconia
  • MgO fused magnesia
  • Ce0 2 fused ceria
  • CaO fused calcia
  • the caustic or alkali fusion step and the alkaline leaching step are performed using potassium hydroxide.
  • the caustic or alkali fusion step and the alkaline leaching are performed using sodium hydroxide.
  • Table 1 Corrosion Resistance of Various Materials to Molten KOH.
  • the molten caustic fusion is performed using at least one potassium salt selected from the group consisting of potassium hydroxide, potassium peroxide, potassium carbonate and potassium nitrate.
  • potassium hydroxide selected from the group consisting of potassium hydroxide, potassium peroxide, potassium carbonate and potassium nitrate.
  • Table 2 The chemical reactions and reaction products contained in the melt when potassium hydroxide is used as the salt for the caustic fusion step are illustrated hereinbelow in Table 2. A distinction is made between the soluble and insoluble reaction products.
  • Deleterious impurities such as iron, titanium, zirconium, uranium, rare earth metals and thorium will remain locked as insoluble residues, while substantially all the tantalum, niobium, manganese values and to a lesser extent tin, tungsten, silica, aluminum will be completely dissolved into the melt.
  • a significant amount of water is released as water vapor or as superheated steam as a result of the elevated operation temperatures.
  • the steam release is at the origin of the effervescence or apparent boiling of the melt occurring following the introduction of the ore or concentrate into the melt.
  • the steam evolution ceases once the chemical reactions are completed as is hence a good indicator of the completion of the caustic fusion reaction.
  • the caustic or alkali fusion step is performed over a period of time ranging from 5 minutes to 6 hours. In an embodiment of the present disclosure, the caustic or alkali fusion step is performed over a period of time ranging from 10 minutes to 4 hours. In an embodiment of the present disclosure, the caustic or alkali fusion step is performed over a period of time ranging from 15 minutes to 2 hours.
  • the caustic or alkali fusion step is performed over a period of time ranging from 15 minutes to 1 hour.
  • the time required for the caustic or alkali fusion is typically dependent on the mass ratio between the caustic agent and the ground ore or concentrate.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio of ore or concentrate mass (C) to mass of melt material (i.e. mass of molten potassium hydroxide salt) (S) denoted as C:S or C/S ranging from 1 :1 to 1 :20.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio maintaining a low melt viscosity and allowing for substantially complete dissolution of the products.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio ranging from 1 : 1 to 1 : 15.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio ranging from 1 : 1 to 1 :10.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio ranging from 1 :1 to 1 :6.
  • the melt produced following the caustic or alkali fusion step is subsequently subjected to an alkaline leaching step.
  • the melt produced as a result of the caustic or alkali fusion step is transferred while hot into alkaline water.
  • the transfer may be accomplished by either pouring or scoping the melt into the cold alkaline water.
  • the melt, once cooled and still in the crucible is demoulded onto a hard surface acting as a heat sink and is subsequently transferred into alkaline water.
  • the crucible comprising the melt is cooled after which alkaline water is added to the crucible.
  • the solidified melt is subjected to an alkaline leaching step to produce an alkaline pregnant leach solution.
  • the alkaline leaching step is performed using a dilute aqueous solution of potassium hydroxide (KOH).
  • KOH potassium hydroxide
  • the aqueous potassium hydroxide solution comprises a concentration ranging from 50g/liter to 250g/liter.
  • the temperature of the potassium hydroxide solution ranges from room temperature to the boiling point of the solution.
  • the pH of the potassium hydroxide solution is at least 13.
  • the resulting dark green and strongly alkaline pregnant leach solution contains a suspension of non-dissolved solids and particulates
  • it is subjected to common solid-liquid separation techniques, such as gravity settling, filtration and centrifugation.
  • the solid-liquid separation is carried out while hot.
  • the solid residues obtained following solid-liquid separation consist essentially of insoluble compounds of iron (III), titanium (IV), zirconium (IV), rare earths (i.e. , Y, Sc and lanthanides), and uranides (e.g., uranium and thorium). These solid residues are typically insoluble under the strongly alkaline conditions prevailing in the pregnant leach solution.
  • non-soluble compounds found in the solid residues include mainly ferric oxides and hydroxides [e.g., Fe 2 0 3 , Fe(OH) 3 ], as a result of the decomposition of potassium ferrite; potassium (IV) titanate [K 2 Ti0 3 ]; potassium (IV) zirconate [Zr 2 Ti0 3 ]; potassium (IV) uranate [K 2 U0 4 ], hydroxides of rare earth elements (REEs) [Ln(OH) 3 ], as well as thorium hydroxide [Th(OH) 4 ],
  • REEs rare earth elements
  • the wet filter cake obtained following solid-liquid separation was thoroughly washed with a hot aqueous solution of KOH (5 wt. %).
  • the washed residue was subsequently subjected to oxidative acid leaching producing an acid leachate.
  • the oxidative acid leaching was performed using sulfuric acid and an oxidizing agent selected from the group consisting of potassium peroxodisulfate (K 2 S 2 0 8 ), ammonium peroxodisulfate [(NH 4 ) 2 S 2 0 8 ], hydrogen peroxide (H 2 0 2 ) and Caro's acid (H 2 S0 5 ).
  • This oxidative acid leaching oxidized substantially all the uranium (IV) to its hexavalent oxidation state which is highly soluble.
  • hydrochloric acid is used as the acid with either hydrogen peroxide (H 2 0 2 ), manganese dioxide (Mn0 2 ), or potassium chlorate (KCI0 3 ) as the oxidizing agent.
  • the remaining insoluble solid residue, containing potassium titanate and potassium zirconate is disposed of in accordance with enforced regulations.
  • the pH of the acid leachate is subsequently adjusted to a range from about 0.5 to 2.5. In an embodiment of the present disclosure, the pH of the acid leachate is adjusted to a range from about 1.0 to 2.0.
  • the pH of the acid leachate is adjusted to about 1 .2.
  • a saturated solution of oxalic acid (ca. 10 wt.% H 2 C 2 0 4 ) is added and the solution was allowed to stand overnight resulting in substantially all of the rare earth elements (REEs) along with thorium precipitating out of the solution in the form of insoluble metal oxalates.
  • the precipitate is first rinsed using a dilute oxalic acid solution (2 wt.%) and then with either sulfuric acid or hydrochloric acid (1 wt.%).
  • the residue is then oven dried at 100- 1 10°C and subsequently calcined to yield a product containing rare earth oxides (REOs) and thoria (Th0 2 ).
  • REOs rare earth oxides
  • Th0 2 thoria
  • the remaining liquor contained substantially all of the uranium as uranyl cations (U0 2 2+ ) which could be readily recovered using selective ion exchange resins.
  • the residual liquor contains substantially all of the ferric iron that can be either reduced to ferrous iron and recovered as iron metal by electrowinning or precipitated as ferric hydroxide and neutralized prior to disposal.
  • the calcined product comprises substantially all of the cerium oxidized to cerium(IV) along with other oxidized lanthanides such as Ln 2 0 3 , Pr 6 On, Eu 2 0 3 and Tb 4 0 7 as well as thorium Th0 2 .
  • the calcined product is acid leached with dilute hydrochloric acid at a pH ranging between 2 and 4. The insoluble Ce0 2 is easily recovered by solid-liquid separation.
  • the clear filtrate solution that contains all the thorium and the other lanthanides is then contacted with metallic zinc, aluminum or magnesium or mixtures thereof (e.g., shot, flakes, chunks, or powder) in order to reduce substantially all the europium(lll) to europium(ll).
  • An aqueous solution of barium chloride (BaCI 2 ) is then added to the clear solution and a stoichiometric amount of dilute sulfuric acid is subsequently added in order to co-precipitate the insoluble barium and europium sulfates [(Ba,Eu)S0 4 ].
  • the completion of the europium precipitation is verified by visible spectrophotometry by measuring the absorbance of the supernatant at the characteristic wavelength of 394 nm [i.e., peak of maximum absorption for Eu(ll)].
  • the europium is readily recovered from the wet insoluble (Ba,Eu)S0 4 precipitate by simply oxidizing it with concentrated nitric acid and by precipitating the europium (III) hydroxide from the neutralized solution by adding aqueous ammonia.
  • the cerium- and europium-free liquor that contains the remaining lanthanides and thorium can eventually be treated to separate the remaining lanthanides from the thorium.
  • the clear alkaline pregnant leach solution obtained following solid-liquid separation comprises substantially all of the manganese values, mainly as brilliant-green potassium manganate (K 2 Mn0 4 ) and to a lesser extent blue-green potassium manganite (K 3 Mn0 4 ), both imparting a dark green color to the solution.
  • the pregnant leach solution further contains substantially all of the soluble tantalum and niobium values, mainly as potassium pertantalate and perniobiate of general formula K 8 (Ta,Nb) 6 0 9 and to a lesser extent potassium meta-tantalate and meta-niobiate of general formula K 3 (Ta,Nb) 3 0 9 .
  • the potassium salts are highly soluble under the strongly alkaline conditions prevailing in the pregnant leach solution.
  • the pregnant leach solution also contains tin, silicon and tungsten values as soluble potassium stannate (K 2 Sn0 3 ), potassium metasilicate (K 2 Si0 3 ) and potassium orthotunsgtate or wolframate (K 2 W0 4 ).
  • the alkaline pregnant leach solution is treated by batch or continuous electrochemical process in order to recover the manganese values.
  • the electrochemical process is performed using a divided electrolyzer including an anode compartment equipped with an anode consisting of pure nickel, nickel-plated copper or nickel-plated steel; a separator consisting of a diaphragm cloth or a cation exchange membrane (CEM) that is chemically resistant to both alkaline conditions and oxidizing conditions, non-limiting examples of which include NAFION® N324 or N424, or N438; and a cathode compartment equipped with a cathode consisting of pure iron, mild steel, austenitic stainless steels or nickel-plated copper.
  • CEM cation exchange membrane
  • the pregnant leach solution is typically fed into the anode compartment where the potassium (VI) manganate is oxidized at the anode and converted to the less soluble potassium (VII) permanganate (KMn0 4 ). Any excess potassium cations released during the oxidation reaction migrate and diffuse across the membrane into the cathode compartment where they combine with hydroxyl anions, regenerating potassium hydroxide (KOH) that is subsequently reused for the alkaline leaching step.
  • the pregnant leach solution is at a temperature of at least room temperature when fed into the anode compartment of the electrolyzer.
  • the pregnant leach solution is at 60°C when fed into the anode compartment of the electrolyzer.
  • Table 3 Electrochemical Reactions of Electrosynthesis of KMn0 4
  • the potassium permanganate crystals (KMn0 4 ) are only sparingly soluble in alkaline solutions (ca. 64 g/L at 20°C) and are recovered from the anolyte by crystallization using a multistage vacuum crystallizer.
  • the potassium permanganate product obtained following centrifugation and drying can be packaged and sold as a high value manganese chemical co-product.
  • the potassium hydroxide rich catholyte exiting the cathode compartment of the electrolyzer is further concentrated by thermal evaporation to yield a potassium hydroxide lye having a mass density of about 1 ,514 kg/m 3 .
  • the potassium hydroxide can be crystallized from the lye to produce anhydrous KOH flakes that are recycled back into the caustic fusion process.
  • the steam or water driven-off from the evaporator can optionally be condensed and reused as distilled water in the upstream process.
  • the potassium rich catholyte exiting the cathode compartment of the electrolyzer can be purified by causticization by adding calcium hydroxide (CaO) or spent lime [Ca(OH) 2 ] in order to precipitate calcium carbonate and other minor metallic impurities that can build-up during the recycling of the caustic leaching liquor.
  • CaO calcium hydroxide
  • Ca(OH) 2 spent lime
  • the dark green alkaline pregnant leach solution is simply exposed to air and carbon dioxide (i.e. maturing) for several hours until the potassium manganate(s) are reduced and/or disproportionated, yielding a dense red-brown precipitate of manganese oxides (e.g. , Mn0 2 , Mn 2 0 3 ).
  • the dense precipitate is then separated by solid-liquid separation techniques, washed with water and oven dried. The maturing process constitutes an alternative to the electrochemical recovery of the manganese values.
  • the dried oxide consisting mostly of manganese dioxide (Mn0 2 ), can be sold as a technical manganese dioxide product or as a raw material for the synthesis of chemical (CMD) or electrochemical grade (EMD) manganese dioxides, or even as feed for preparing manganese metal either by electrowinning or smelting.
  • CMD chemical
  • EMD electrochemical grade
  • the clear and manganese-free alkaline pregnant leach solution containing the tantalum and niobium values and to a lesser extent the tin and silicon values is transferred to a plastic-lined steel tank for recovering the tantalum and niobium values.
  • plastic linings include high density polyethylene (HDPE), polypropylene (PP) and polyvinylidene chloride (PVDC).
  • HDPE high density polyethylene
  • PP polypropylene
  • PVDC polyvinylidene chloride
  • the temperature of the solution is raised to a temperature ranging from 30°C to 100°C. A saturated aqueous solution of sodium sulfate or sodium chloride is then added.
  • the total amount of sodium added typically corresponds to the stoichiometric amount necessary to yield hydrated sodium tantalate (Na 8 Ta 6 0i 9 .16H 2 0) and sodium niobiate ( a 8 Nb 6 0i 9 .16H 2 0) respectively, while also providing an excess ranging from 10 to 50%. wt.
  • the pH of the solution is then gradually adjusted to values ranging between 6 and 7 by adding concentrated sulfuric acid or hydrochloric acid.
  • the solubility of sodium tantalate and sodium niobiate decreases as the pH is lowered. In fact, hydrated sodium tantalate starts to precipitate at pH values below 10, whereas hydrated sodium niobiate starts to precipitate at values below 7.
  • Both precipitates quickly settle at the bottom of the reactor and can be recovered by standard solid-liquid separation techniques to yield a wet filter cake.
  • the remaining solution comprising the silicon and tin values, in addition to excess sodium sulfate or sodium chloride, is subsequently used for recovering the tin values by precipitating tin oxide.
  • the wet filter cake comprising the hydrated sodium tantalate and niobiate is transferred to a glass lined autoclave reactor where it is subjected to hot acid leaching using sulfuric acid (30 wt.% H 2 S0 4 ) or hydrochloric acid (20 wt.% HCI) in order to leach out the sodium and any traces of metallic impurities.
  • the leaching is performed at atmospheric pressure.
  • the leaching is performed at pressures in excess of atmospheric.
  • the hot acid leaching yields an insoluble mass comprising hydrated tantalum oxide (Ta 2 0 5 .nH 2 0) and hydrated niobium oxide (Nb 2 0 5 .nH 2 0), also called tantalic and niobic acids.
  • the oxide products are thoroughly washed with a hot and dilute solution of sulfuric acid (or hydrochloric acid), followed by washing with deionized water and solid-liquid separation to provide a bright milky white wet filter cake.
  • the hydrated tantalum and niobium oxides are transferred into a long quartz boat or a large zirconium crucible, dried at 120°C for about 4 hours and calcined in a muffle or rotary kiln to yield a dense white powder of a mixture of pure tantalum and niobium oxides (Ta 2 0 5 + Nb 2 0 5 ).
  • the calcination is performed at temperatures of about 800°C.
  • the tantalum/niobium mass ratio in the oxide product is substantially identical to that found in the original starting material (i.e. ore or concentrate).
  • the brine solution by- produced during the acid leaching of the wet filter cake comprising the hydrated sodium tantalate and niobiate is either concentrated to yield a concentrated brine or evaporated to yield a crystallized salt.
  • the salt comprises either sodium sulfate decahydrate (i.e. Glauber's salt) or sodium chloride (i.e. Rock salt) depending on the type of sodium salt and mineral acid used in the precipitation and acid leaching steps respectively.
  • the concentrated brine or crystallized salts can be reused in the precipitation process and the water driven-off from the evaporator can optionally be condensed and reused as distilled water in the process.
  • the clear and manganese-free alkaline pregnant leach solution containing the tantalum and niobium values and to a lesser extent the tin and silicon values (potassium stannate and silicate) in addition to free potassium hydroxide is evaporated until the potassium niobiate and tantalate have crystallized out.
  • the valuable potassium salts are then removed by common solid-liquid separation, while the concentrated caustic potassium hydroxide lye containing potassium stannate and silicate, can be further purified by adding quicklime or calcium hydroxide in order to precipitate the insoluble calcium stannate and silicate.
  • the purified potassium hydroxide lye is then ready to be recycled into the process.
  • the molten caustic fusion is performed using at least one sodium salt selected from the group consisting of sodium hydroxide, sodium peroxide, sodium carbonate and sodium nitrate.
  • sodium hydroxide sodium peroxide
  • sodium carbonate sodium nitrate.
  • the chemical reactions and reaction products contained in the melt when sodium hydroxide is used as the salt for the caustic fusion step are illustrated hereinbelow in Table 4. A distinction is made between the soluble and insoluble reaction products.
  • the tantalum and niobium values, along with deleterious impurities such as iron, titanium, zirconium, uranium, rare earth metals and thorium will remain locked as insoluble residues, while substantially all the manganese values and to a lesser extent tin, tungsten, silica, aluminum will be completely dissolved into the melt.
  • a significant amount of water is released as water vapor or as superheated steam as a result of the elevated operation temperatures.
  • the steam release is at the origin of the effervescence or apparent boiling of the melt following the introduction of the ore or concentrate into the melt.
  • the steam evolution ceases once the chemical reactions are completed as is hence a good indicator of the completion of the caustic fusion reaction.
  • the caustic or alkali fusion step is performed over a period of time ranging from 5 minutes to 6 hours. In an embodiment of the present disclosure, the caustic or alkali fusion step is performed over a period of time ranging from 10 minutes to 4 hours. In an embodiment of the present disclosure, the caustic or alkali fusion step is performed over a period of time ranging from 15 minutes to 2 hours.
  • the caustic or alkali fusion step is performed over a period of time ranging from 15 minutes to 1 hour.
  • the time required for the caustic or alkali fusion is typically dependent on the mass ratio between the caustic agent and the ground ore or concentrate.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio of ore or concentrate mass (C) to mass of melt material (i.e. mass of molten sodium hydroxide salt) (S) denoted as C:S or C/S ranging from 1 : 1 to 1 :20.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio maintaining a low melt viscosity and allowing for substantially complete dissolution of the products.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio ranging from 1 :1 to 1 : 15.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio ranging from 1 :1 to 1 :10.
  • the caustic or alkali fusion step is performed using a dimensionless mass ratio ranging from 1 :1 to 1 :6.
  • the melt produced following the caustic or alkali fusion step is subsequently subjected to an alkaline leaching step.
  • the melt produced as a result of the caustic or alkali fusion step is transferred while hot into alkaline water.
  • the transfer may be accomplished by either pouring or scoping the melt into the cold alkaline water.
  • the melt, once cooled and still in the crucible is demoulded onto a hard surface acting as a heat sink and is subsequently transferred into alkaline water.
  • the crucible comprising the melt is cooled after which alkaline water is added to the crucible.
  • the solidified melt is subjected to an alkaline leaching step to produce an alkaline pregnant leach solution.
  • the alkaline leaching step is performed using a dilute aqueous solution of sodium hydroxide (NaOH).
  • the aqueous sodium hydroxide solution comprises a concentration ranging from 50g/liter to 250g/liter.
  • the temperature of the sodium hydroxide solution ranges from room temperature to the boiling point of the solution.
  • the pH of the sodium hydroxide solution is at least 13.
  • the resulting blue green and strongly alkaline pregnant leach solution contains a suspension of non-dissolved solids and particulates, and is subjected to common solid-liquid separation techniques, such as gravity settling, filtration and centrifugation.
  • the solid-liquid separation is carried out while hot.
  • the solid residues obtained following solid-liquid separation consist essentially of insoluble compounds of tantalum (V), niobium (V), iron (III), titanium (IV), zirconium (IV), rear earths (i.e. , Y, Sc and lanthanides), and uranides (e.g. , uranium and thorium). These solid residues are typically insoluble under the strongly alkaline conditions prevailing in the pregnant leach solution.
  • non-soluble compounds found in the solid residues include hydrated sodium tantalate (Na 8 Ta 6 0 19 .16H20) and sodium niobiate (Na 8 Nb 6 0 9 .16H 2 0); ferric oxides and hydroxides [e.g. Fe 2 0 3 , Fe(OH) 3 ], as a result of the decomposition of sodium ferrite; sodium (IV) titanate [K 2 Ti0 3 ]; sodium (IV) zirconate [Zr 2 Ti0 3 ]; sodium (IV) uranate [K 2 U0 4 ], hydroxides of rare earth elements (REEs) [Ln(OH) 3 ], as well as thorium hydroxide [Th(OH) 4 ].
  • REEs rare earth elements
  • the wet filter cake obtained following solid-liquid separation was thoroughly washed with a hot aqueous solution of NaOH (5 wt.%).
  • the washed residue was subsequently subjected to oxidative acid leaching producing an acid leachate.
  • the oxidative acid leaching was performed using sulfuric acid and an oxidizing agent selected from the group consisting of sodium peroxodisulfate (Na 2 S 2 0 8 ), ammonium peroxodisulfate [(NH 4 ) 2 S 2 0 8 ], hydrogen peroxide (H 2 0 2 ) and Caro's acid (H 2 S0 5 ).
  • hydrochloric acid is used as the acid with either hydrogen peroxide (H 2 0 2 ), manganese dioxide (Mn0 2 ), or sodium chlorate (NaCI0 3 ) as the oxidizing agent.
  • the insoluble residues remaining following oxidative acid leaching comprise the tantalum and niobium values as hydrated tantalum oxide (Ta 2 0 5 .nH 2 0) and hydrated niobium oxide (Nb 2 0 5 .nH 2 0) also called tantalic and niobic acids.
  • the insoluble residues further comprise, to a lesser extent, hydrated oxides of titanium and zirconium which explains why the "sodium" process provides a final tantalum and niobium oxide product that is generally of a lesser purity than that obtained by the "potassium" process.
  • the hydrated tantalum and niobium oxides are transferred into a long quartz boat or a large zirconium crucible, dried at 120°C for about 4 hours and calcined in a muffle or rotary kiln to yield a dense white fluffy powder comprising a mixture of pure tantalum and niobium oxides (Ta 2 0 5 + Nb 2 0 5 ).
  • the solid further includes oxides of titanium and zirconium.
  • the calcination is performed at temperatures of about 800X.
  • the pH of the acid leachate is adjusted to a range from about 1.0 to 2.0. In an embodiment of the present disclosure, the pH of the acid leachate is adjusted to a range from about 1.0 to 2.5. In an embodiment of the present disclosure, the pH of the acid leachate is adjusted to a range from about 1.2 to 1.5. Following pH adjustment, a saturated solution of oxalic acid (ca. 10 wt.% H 2 C 2 0 4 ) is added and the solution was allowed to stand overnight resulting in substantially all of the rare earth elements (REEs) along with thorium precipitating out of the solution in the form of insoluble metal oxalates.
  • REEs rare earth elements
  • the precipitate is first rinsed using a dilute oxalic acid solution (2 wt.%) and then with either sulfuric acid or hydrochloric acid (1 wt.%).
  • the residue is then oven dried at 100-110°C and subsequently calcined to yield a product containing rare earth oxides (REOs) and thoria (Th0 2 ).
  • REOs rare earth oxides
  • Th0 2 thoria
  • the remaining liquor contained substantially all of the uranium as uranyl cations (U0 2 2+ ) which could be readily recovered using selective ion exchange resins.
  • the residual liquor contains substantially all of the ferric iron that can be either reduced to ferrous iron and recovered as iron metal by electrowinning or precipitated as ferric hydroxide and neutralized prior to disposal.
  • the clear alkaline pregnant leach solution obtained following solid-liquid separation comprises substantially all of the manganese values, mainly as blue-green sodium manganite (Na 3 n0 4 ) and to a lesser extent brilliant-green sodium manganate (Na 2 Mn0 4 ), both imparting a blue-green color to the solution.
  • the pregnant leach solution further contains tin, silicon and tungsten values as soluble sodium stannate (Na 2 Sn0 3 ), sodium metasilicate (Na 2 Si0 3 ) and sodium orthotunsgtate or wolframate (K 2 W0 4 ). These sodium salts are soluble under the strongly alkaline conditions prevailing in the pregnant leach solution.
  • the sodium salts of manganese cannot be oxidized to sodium permanganate because: (i) substantially only sodium manganite (Na 3 Mn0 4 ) is produced during the caustic fusion process with only little sodium manganate (Na 2 Mn0 4 ); (ii) sodium manganite is chemically highly unstable, and; (iii) even if electrolysis were performed, the high solubility of sodium permanganate ⁇ ca. 900 g/L at room temperature) would preclude the harvesting of the crystals from the mother liquor.
  • the clear alkaline pregnant leach solution obtained following solid-liquid separation is simply exposed to air and carbon dioxide ⁇ i.e. maturing) for several hours until the sodium manganate(s) are reduced and/or disproportionated, yielding a dense red-brown precipitate of manganese oxides (e.g., Mn0 2 , Mn 2 0 3 ).
  • the dense precipitate is then separated by solid-liquid separation techniques, washed with water and oven dried.
  • the dried oxide consisting mostly of manganese dioxide (Mn0 2 ), can be sold as a technical manganese dioxide product or as a raw material for the synthesis of chemical (CMD) or electrochemical grade (EMD) manganese dioxides, or even as feed for preparing manganese metal either by electrowinning or smelting.
  • CMD chemical
  • EMD electrochemical grade
  • the clear and manganese-free alkaline pregnant leach solution comprising the silicon and tin values, is subsequently used for recovering the tin values by precipitating tin oxide.
  • the spent acidic solution by-produced during acid leaching of the solid residues can be further concentrated and subjected to electrowinning to recover the iron values and concurrently regenerate the acid values.
  • EXAMPLE 1 MOLTEN KOH CAUSTIC FUSION (SAMPLE A)
  • Example A An aliquot mass (1.2 grams) of an oven dried and ground tantalum concentrate (sample A), having the average chemical composition illustrated in Table 5, was weighed using a precision scale. Separately, potassium hydroxide flakes of technical grade (10 g) were pre-melted inside a 25 mL crucible made of nickel 200 and equipped with a zirconium lid to avoid losses by splashing. The intense heating was provided by means of a blast Bunsen burner supplied with propane gas. Once the residual moisture was driven off and the melt surface became quiet, the heating was stopped and the melt was cooled. The ground aliquot sample was subsequently cautiously poured on top of the solidified melt and heating was resumed to bring the crucible temperature to red-dull heat (ca.
  • the temperature of the solution increased.
  • the temperature of the alkaline solution was subsequently maintained at about 60°C by heating with a hot plate.
  • the resulting strongly alkaline pregnant leach solution (PLS) had a dark green color and contained a suspension of non-dissolved solids and particulates.
  • the PLS was filtered using hardened and ashless filter paper (Whatman #541). The brown solid residue was subsequently washed thoroughly with deionized water, oven dried and then calcined inside a porcelain capsule from CoorsTech to yield 0.219 grams of calcined mass.
  • a saturated aqueous solution of sodium chloride (12 mL) (25.4 wt.% NaCI) was subsequently added.
  • the pH of the resulting solution was gradually increased by cautious dropwise addition of concentrated sulfuric acid (98 wt.% H 2 S0 4 ) until the pH was in a range from about 6-7.
  • concentrated sulfuric acid 98 wt.% H 2 S0 4
  • a milky white, dense and voluminous precipitate composed of hydrated sodium tantalate and niobiate formed that quickly settled at the bottom of the beaker.
  • the precipitate was recovered by filtration using hardened and ashless filter paper (Whatman #541 ) and was subsequently rinsed with dilute hydrochloric acid.
  • the precipitate and filter paper were acid leached under reflux conditions over a period of 30 minutes in order to leach out all of the sodium as well as traces of metallic impurities.
  • the filter paper was digested during this treatment, forming a pulp.
  • the hot acid leaching yielded a dense milky white solid consisting of a hydrated mixture of tantalic and niobic acids that was subsequently thoroughly washed with a hot and dilute solution of hydrochloric acid followed by washing with deionized water.
  • the resulting solid was subsequently recovered by filtration using hardened and ashless filter paper (Whatman #541).
  • the hydrated tantalum and niobium oxides, along with the filter paper, were subsequently transferred into a fused silica crucible and dried at a temperature of 120°C over a period of 1 hour.
  • the product was then calcined at 800°C until all the carbon values were burned off and the fumes ceased.
  • the calcination yielded a dense white powder (0.913 grams).
  • the average chemical composition of the white powder was determined to be 90.5 percent by mass of tantalum pentoxide (Ta 2 0 5 ) and 9.5 percent by mass of niobium oxide (Nb 2 0 5 ) with no detectable traces of other impurities.
  • the overall recovery of the tantalum and niobium values as contained in the original concentrate was determined to be 99 percent by weight (Table 6).
  • EXAMPLE 2 MOLTEN KOH CAUSTIC FUSION (SAMPLE B)
  • the temperature of the solution increased.
  • the temperature of the alkaline solution was subsequently maintained at about 60°C by heating with a hot plate.
  • the resulting strongly alkaline pregnant leach solution (PLS) had a dark green color and contained a suspension of non-dissolved solids and particulates.
  • the PLS was filtered using hardened and ashless filter paper (Whatman #541). The brown solid residue was subsequently washed thoroughly with deionized water, oven dried and then calcined inside a porcelain capsule from CoorsTech to yield 0.520 grams of calcined mass.
  • a saturated aqueous solution of sodium chloride (12 mL) (25.4 wt.% NaCI) was subsequently added.
  • the pH of the resulting solution was gradually increased by cautious dropwise addition of concentrated sulfuric acid (98 wt.% H 2 S0 ) until the pH was in a range from about 6-7.
  • concentrated sulfuric acid 98 wt.% H 2 S0
  • a milky white, dense and voluminous precipitate composed of hydrated sodium tantalate and niobiate formed that quickly settled at the bottom of the beaker.
  • the precipitate was recovered by filtration using hardened and ashless filter paper (Whatman #541) and was subsequently rinsed with dilute hydrochloric acid.
  • the precipitate and filter paper were acid leached under reflux conditions over a period of 30 minutes in order to leach out all of the sodium as well as traces of metallic impurities.
  • the filter paper was digested during this treatment, forming a pulp.
  • the hot acid leaching yielded a dense milky white solid consisting of a hydrated mixture of tantalic and niobic acids that was subsequently thoroughly washed with a hot and dilute solution of hydrochloric acid followed by washing with deionized water.
  • the resulting solid was subsequently recovered by filtration using hardened and ashless filter paper (Whatman #541).
  • the hydrated tantalum and niobium oxides, along with the filter paper, were subsequently transferred into a fused silica crucible and dried at a temperature of 120°C over a period of 1 hour.
  • the product was then calcined at 800°C until all the carbon values were burned off and the fumes ceased.
  • the calcination yielded a dense white powder (0.991 grams).
  • the average chemical composition of the white powder was determined to be 63.1 percent by mass of tantalum pentoxide (Ta 2 0 5 ) and 36.9 percent by mass of niobium oxide (Nb 2 0 5 ).
  • the overall recovery of the tantalum and niobium values as contained in the original concentrate was determined to be 99 percent by weight (Table 6).
  • Example A An aliquot mass (1.0 grams) of an oven dried and ground tantalum concentrate (sample A), having the average chemical composition illustrated in Table 5, was weighed using a precision scale. Separately, sodium hydroxide flakes of technical grade (12.4 g) were pre-melted inside a 25 mL crucible made of nickel 200 and equipped with a zirconium lid to avoid losses by splashing. The intense heating was provided by means of a blast Bunsen burner supplied with propane gas. Once the residual moisture was driven off and the melt surface became quiet, the heating was stopped and the melt was cooled. The ground aliquot sample was subsequently cautiously poured on top of the solidified melt and heating was resumed to bring the crucible temperature to red-dull heat (ca.
  • the temperature of the solution increased.
  • the temperature of the alkaline solution was subsequently maintained at about 60°C by heating with a hot plate.
  • the resulting strongly alkaline pregnant leach solution (PLS) had a blue green color and contained a suspension of non-dissolved solids and particulates.
  • the PLS was filtered using hardened and ashless filter paper (Whatman #541). The brown solid residue was subsequently washed thoroughly with deionized water and put aside for the subsequent recovery of the tantalum and niobium values.
  • the brown solid residue (along with the filter paper), containing the tantalum and niobium values as hydrated sodium tantalate and niobiate, as well as the iron, titanium and zirconium values, was transferred into an atmospheric acid leaching apparatus consisting of a 250-mL Erlenmeyer flask containing 100 mL of hydrochloric acid (20 wt.% HCI) connected to a condenser.
  • the residue and filter paper were acid leached under reflux conditions over a period of 30 minutes in order to leach out all of the iron and sodium.
  • the filter paper was digested during this treatment, forming a pulp.
  • the hot acid leaching yielded a dense solid consisting of a hydrated mixture of tantalic and niobic acids that was subsequently thoroughly washed with a hot and dilute solution of hydrochloric acid followed by washing with deionized water.
  • the resulting solid was subsequently recovered by filtration using hardened and ashless filter paper (Whatman #541).
  • the hydrated tantalum and niobium oxides, along with the filter paper, were subsequently transferred into a fused silica crucible and dried at a temperature of 120°C over a period of 4 hours.
  • the product was then calcined at 800°C until all the carbon values were burned off and the fumes ceased.
  • the calcination yielded a dense off-white powder (0.782 grams).
  • the average chemical composition of the white powder was determined to be 87.7 percent by mass of tantalum pentoxide (Ta 2 0 5 ) and 9.2 percent by mass of niobium oxide (Nb 2 0 6 ) in addition to 3.1 percent by mass of other metal oxides, mainly titanium dioxide and zirconium dioxide tainted by traces of ferric oxide.
  • the overall recovery of the tantalum and niobium values as contained in the original concentrate was determined to be 98.9 percent by weight (Table 6).
  • EXAMPLE 4 MOLTEN KOH CAUSTIC FUSION (SAMPLE C)
  • Example C An aliquot mass (2.44 grams) of an oven dried and ground columbite concentrate (sample C), having the average chemical composition illustrated in Table 5, was weighed using a precision scale. Separately, potassium hydroxide flakes of technical grade (20 g) were pre-melted inside a tall 200 mL crucible made of deep drawn pure zirconium (BJ Scientific). The intense heating was provided by means of a pair of blast Bunsen burners supplied with propane gas. Once the residual moisture was driven off and the melt surface became quiet, the heating was stopped and the melt was cooled. The ground aliquot sample was subsequently cautiously poured on top of the solidified melt and heating was resumed to bring the crucible temperature to red-dull heat (ca.
  • the temperature of the alkaline solution was subsequently maintained at about 60°C by heating the zirconium crucible directly on a hot plate equipped with a magnetic stirrer. A PTFE-coated magnetic bar was subsequently introduced into the crucible to ensure proper agitation and mixing.
  • the resulting strongly alkaline pregnant leach solution (PLS) had a dark green color and contained a suspension of non-dissolved solids and particulates. After cooling to room temperature, the PLS was filtered using hardened and ashless filter paper (Whatman #541). The brown solid residue was subsequently washed thoroughly with aqueous potassium hydroxide, deionized water, oven dried and then calcined inside a porcelain capsule from CoorsTech to yield 1.030 grams of calcined mass.
  • a saturated aqueous solution of sodium chloride (10 mL) (25.4 wt.% NaCI) was subsequently added.
  • the pH of the resulting solution was gradually increased by cautious dropwise addition of concentrated sulfuric acid (98 wt.% H 2 S0 4 ) until the pH was in a range from about 6-7.
  • concentrated sulfuric acid 98 wt.% H 2 S0 4
  • a milky white, dense and voluminous precipitate composed of hydrated sodium tantalate and niobiate formed that quickly settled at the bottom of the beaker.
  • the precipitate was recovered by filtration using hardened and ashless filter paper (Whatman #541) and was subsequently rinsed with water with dilute hydrochloric acid.
  • the precipitate and filter paper were acid leached under reflux conditions over a period of 30 minutes in order to leach out all of the sodium as well as traces of metallic impurities.
  • the filter paper was digested during this treatment, forming a pulp.
  • the hot acid leaching yielded a dense milky white solid consisting of a hydrated mixture of tantalic and niobic acids that was subsequently thoroughly washed with a hot and dilute solution of hydrochloric acid followed by washing with deionized water.
  • the resulting solid was subsequently recovered by filtration using hardened and ashless filter paper (Whatman #541).
  • the hydrated tantalum and niobium oxides, along with the filter paper, were subsequently transferred into a fused silica crucible and dried at a temperature of 120°C over a period of 1 hour.
  • the product was then calcined at 800°C until all the carbon values were burned off and the fumes ceased.
  • the calcination yielded a dense white powder (1.650 grams).
  • the average chemical composition of the white powder was determined to be 19.2 percent by mass of tantalum pentoxide (Ta 2 0 5 ) and 80.8 percent by mass of niobium oxide (Nb 2 0 5 ).
  • the overall recovery of the tantalum and niobium values as contained in the original concentrate was determined to be 97.6 percent by weight (Table 6).
  • Samarskite concentrate (sample D), having the average chemical composition illustrated in Table 5, was weighed using a precision scale. Separately, potassium hydroxide flakes of technical grade (18 g) were pre-melted inside a tall 200 mL crucible made of deep drawn pure zirconium (BJ Scientific). The intense heating was provided by means of a pair of blast Bunsen burners supplied with propane gas. Once the residual moisture was driven off and the melt surface became quiet, the heating was stopped and the melt was cooled. The ground aliquot sample was subsequently cautiously poured on top of the solidified melt and heating was resumed to bring the crucible temperature to red-dull heat (ca. 800°C) over a period of 30 minutes in order to perform the molten caustic fusion.
  • red-dull heat ca. 800°C
  • the temperature of the alkaline solution was subsequently maintained at about 60°C by heating the zirconium crucible directly on a hot plate equipped with a magnetic stirrer. A PTFE-coated magnetic bar was subsequently introduced into the crucible to ensure proper agitation and mixing.
  • the resulting strongly alkaline pregnant leach solution (PLS) had a dark green color and contained a suspension of non-dissolved solids and particulates. After cooling to room temperature, the PLS was filtered using hardened and ashless filter paper (Whatman #541). The brown solid residue was subsequently washed thoroughly with aqueous potassium hydroxide, deionized water, oven dried and then calcined inside a porcelain capsule from CoorsTech to yield 1.005 grams of calcined mass.
  • a saturated aqueous solution of sodium chloride (10 ml_) (25.4 wt.% NaCI) was subsequently added.
  • the pH of the resulting solution was gradually increased by cautious dropwise addition of concentrated sulfuric acid (98 wt.% H 2 S0 4 ) until the pH was in a range from about 6-7.
  • concentrated sulfuric acid 98 wt.% H 2 S0 4
  • a milky white, dense and voluminous precipitate composed of hydrated sodium tantalate and niobiate formed that quickly settled at the bottom of the beaker.
  • the precipitate was recovered by filtration using hardened and ashless filter paper (Whatman #541) and was subsequently rinsed with water with dilute hydrochloric acid.
  • the filter paper was digested during this treatment, forming a pulp.
  • the hot pressure acid leaching yielded a dense milky white solid consisting of a hydrated mixture of tantalic and niobic acids that was subsequently thoroughly washed with a hot and dilute solution of hydrochloric acid followed by washing with deionized water.
  • the resulting solid was subsequently recovered by filtration using hardened and ashless filter paper (Whatman #541).
  • the hydrated tantalum and niobium oxides, along with the filter paper, were subsequently transferred into a tall platinum crucible and dried at a temperature of 120°C over a period of 1 hour.
  • the product was then calcined at 800°C until all the carbon values were burned off and the fumes ceased.
  • the calcination yielded a dense white powder (0.989 grams).
  • the average chemical composition of the white powder was determined to be 22.5 percent by mass of tantalum pentoxide (Ta 2 0 5 ) and 77.5 percent by mass of niobium oxide (Nb 2 0 5 ).
  • the overall recovery of the tantalum and niobium values as contained in the original concentrate was determined to be 98.9 percent by weight (Table 6).
  • EXAMPLE 6 RECOVERY OF RARE EARTH OXIDES AND
  • the isolated insoluble residue consisted essentially of titanium and zirconium oxide.
  • a saturated solution of oxalic acid (ca. 10 wt.% H 2 C 2 0 4 ) was added and the solution was left stand overnight in order to precipitate all the insoluble rare earth values and thorium oxalates.
  • the precipitate was subsequently thoroughly rinsed using a solution containing 2 wt.% oxalic acid and then air dried at 110°C over a period of several hours to yield a product containing the rare earth oxides (REOs) and thoria (Th0 2 ).
  • the air dried rare earth oxide (REOs) product contained the cerium values as cerium(IV) along with other lanthanide hydroxides (Ln(OH) 3 ) and thorium hydroxide (Th(OH) 4 .).
  • the air-dried residue was subsequently acid leached using a diluted hydrochloric acid solution having a pH ranging between 2 and 4 leaving behind an insoluble cerium(IV) hydroxide which was easily recovered by filtration.
  • the clear filtrate solution containing all the thorium and the other lanthanides, was reacted with zinc powder in order to reduce Eu(lll) to Eu(ll).
  • An aqueous solution of barium chloride (5 wt.% BaCI 2 ) was subsequently added to the clear solution followed by a stoichiometric amount of dilute sulfuric acid in order to co-precipitate the insoluble barium and europium sulfates [(Ba,Eu)S0 4 ].
  • the completion of the europium precipitation was verified by visible spectrophotometry by measuring the absorbance of the supernatant at the characteristic wavelength of 394 nm (i.e., peak of maximum absorption for Eu(ll)).
  • the europium was recovered from the wet insoluble (Ba,Eu)S0 4 precipitate by simply oxidizing it with hot concentrated nitric acid and by precipitating the europium (III) hydroxide once the solution was neutralized with ammonia.
  • the remaining cerium- and europium-free liquor contained the remaining lanthanides and most of the thorium.
  • EXAMPLE 7 ELECTROSYNTHESIS OF POTASSIUM
  • a filtered alkaline pregnant leach solution (150 mL), obtained after caustic fusion of 10 grams of sample B and hot alkaline leaching, was charged in the anode compartment of the electrolyzer.
  • the electrolyzer consisted of a rectangular cell divided into two compartments by a cation exchange membrane (CEM) made of NAFION® N324 (E.I. DUPONT DE NEMOURS).
  • CEM cation exchange membrane
  • the anode compartment was equipped with a rectangular plate anode composed of pure nickel grade 201 while the cathode compartment was equipped with a rectangular plate cathode composed of austenitic stainless steel grade AISI 304. Both anode and cathode exhibited exactly the same dimensions (i.e.
  • the catholyte consisted of a diluted alkaline solution of potassium hydroxide containing 1 12 g/L KOH (i.e. 2 mol/L KOH) as measured by acid-base titration.
  • the concentration of potassium permanganate in the anolyte was determined by redox titration using a standard solution of sodium oxalate 0.1 N. Both the anolyte and catholyte were circulated at room temperature with a volume flow rate of 90 mL/min using a L/S peristaltic pump (MASTERFLEX). The electrolysis was performed under galvanostatic mode (i.e. at a constant current of 1.29 A corresponding to an absolute current density of 500 A/m 2 at both electrodes. The average cell voltage measured between the anode and cathode was 2.9 V and the electrolysis was conducted over a period of 10 minutes.
  • the anolyte color turned purple, characteristic of permanganate anions.
  • the final concentration of potassium hydroxide in the catholyte was measured by acid-base titration to be 1 14.9 g/L KOH.
  • the potassium permanganate concentration in the anolyte was measured by redox titration using oxalic acid to be 8.0 g/L KMn0 4 .
  • the cathode and anode current efficiencies were determined to be 97% and 95% respectively.
  • the specific energy consumption was determined to be 1.43 kWh/kg of KOH and 0.52 kWh/kg of KMn0 4 respectively (Table 7).
  • Table 5 Average chemical composition of the tantalum and niobium concentrates (only the principal metal oxides are shown).

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Abstract

L'invention concerne un procédé de valorisation de minerais et de concentrés de tantale et de niobium. Le procédé comprend : a) la soumission du minerai ou du concentré à une fusion caustique ou à une fusion alcaline en utilisant un mélange fondu comprenant au moins un sel d'un métal alcalin pour produire un mélange fondu solidifié ; b) la soumission du mélange fondu solidifié à une étape de lixiviation alcaline pour produire une solution de lixiviation mère alcaline comprenant du manganèse exploitable, c) la récupération du manganèse exploitable dans la solution de lixiviation mère alcaline pour produire une solution de lixiviation sensiblement dépourvue de manganèse et d) la précipitation du tantale et du niobium sous forme de niobiate et de tantalate insolubles dans la solution de lixiviation sans manganèse. Les oxydes de terres rares sont récupérés à partir des résidus insolubles pendant la lixiviation alcaline.
PCT/CA2012/000890 2011-09-23 2012-09-24 Procédé de valorisation de minerais et de concentrés de tantale et de niobium avec récupération d'oxydes de manganèse et de terres rares Ceased WO2013040694A1 (fr)

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Cited By (12)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
WO2015004375A1 (fr) * 2013-07-09 2015-01-15 Eramet Procédé de purification du niobium et/ou du tantale
FR3043696A1 (fr) * 2015-11-17 2017-05-19 Eramet Procede hydrometallurgique pour la separation et la purification du tantale et du niobium
RU2647966C2 (ru) * 2015-10-20 2018-03-21 Акционерное общество "Ульбинский металлургический завод" Устройство для получения порошка тантала
US10508320B2 (en) 2015-03-04 2019-12-17 University Of Leeds Process for recovering primary metal residue from a metal-containing composition
WO2021217148A1 (fr) * 2020-04-24 2021-10-28 Lawrence Livermore National Security, Llc Compositions et procédés d'utilisation de celles-ci pour la séparation de scandium à partir d'un matériau contenant des terres rares
CN114082521A (zh) * 2021-11-24 2022-02-25 贺州久源矿业有限公司 一种从花岗岩风化壳型钾长石综合回收云母的工艺
CN115109931A (zh) * 2022-06-21 2022-09-27 厦门钨业股份有限公司 一种从钨钼废渣中回收多种金属的方法
CN115180637A (zh) * 2022-07-11 2022-10-14 新疆志存新能源材料有限公司 一种利用钨锡尾矿梯次回收碳酸锂及氢氧化锂的方法
CN115992310A (zh) * 2022-11-29 2023-04-21 包头稀土研究院 含独居石的萤石精矿的处理方法
WO2023137523A1 (fr) * 2022-01-20 2023-07-27 Australian National University Récupération de terres rares
CN117531585A (zh) * 2023-11-09 2024-02-09 江阴检验认证有限公司 一种锰矿石样品处理破碎设备及样品检测方法
WO2025096915A1 (fr) * 2023-11-01 2025-05-08 Worcester Polytechnic Institute Extraction alcaline par micro-ondes

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US2728633A (en) * 1948-04-22 1955-12-27 Arden Thomas Victor Production of uranium and thorium compounds
US2950966A (en) * 1958-06-20 1960-08-30 Nat Distillers Chem Corp Recovery of tantalum values
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Cited By (16)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
FR3008425A1 (fr) * 2013-07-09 2015-01-16 Eramet Procede de purification du niobium et/ou du tantale
WO2015004375A1 (fr) * 2013-07-09 2015-01-15 Eramet Procédé de purification du niobium et/ou du tantale
US10508320B2 (en) 2015-03-04 2019-12-17 University Of Leeds Process for recovering primary metal residue from a metal-containing composition
RU2647966C2 (ru) * 2015-10-20 2018-03-21 Акционерное общество "Ульбинский металлургический завод" Устройство для получения порошка тантала
FR3043696A1 (fr) * 2015-11-17 2017-05-19 Eramet Procede hydrometallurgique pour la separation et la purification du tantale et du niobium
WO2017085404A1 (fr) * 2015-11-17 2017-05-26 Eramet Procede hydrometallurgique pour la separation et la purification du tantale et du niobium
WO2021217148A1 (fr) * 2020-04-24 2021-10-28 Lawrence Livermore National Security, Llc Compositions et procédés d'utilisation de celles-ci pour la séparation de scandium à partir d'un matériau contenant des terres rares
CN114082521A (zh) * 2021-11-24 2022-02-25 贺州久源矿业有限公司 一种从花岗岩风化壳型钾长石综合回收云母的工艺
WO2023137523A1 (fr) * 2022-01-20 2023-07-27 Australian National University Récupération de terres rares
CN115109931A (zh) * 2022-06-21 2022-09-27 厦门钨业股份有限公司 一种从钨钼废渣中回收多种金属的方法
CN115109931B (zh) * 2022-06-21 2023-12-08 厦门钨业股份有限公司 一种从钨钼废渣中回收多种金属的方法
CN115180637A (zh) * 2022-07-11 2022-10-14 新疆志存新能源材料有限公司 一种利用钨锡尾矿梯次回收碳酸锂及氢氧化锂的方法
CN115180637B (zh) * 2022-07-11 2023-12-15 新疆志存新能源材料有限公司 一种利用钨锡尾矿梯次回收氢氧化锂的方法
CN115992310A (zh) * 2022-11-29 2023-04-21 包头稀土研究院 含独居石的萤石精矿的处理方法
WO2025096915A1 (fr) * 2023-11-01 2025-05-08 Worcester Polytechnic Institute Extraction alcaline par micro-ondes
CN117531585A (zh) * 2023-11-09 2024-02-09 江阴检验认证有限公司 一种锰矿石样品处理破碎设备及样品检测方法

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