OA16481A

OA16481A - Dissolution and recovery of at least one element Nb or Ta and of at least one other element U or rare earth elements from ores and concentrates.

Info

Publication number: OA16481A
Application number: OA1201300284
Authority: OA
Inventors: Jérôme Agin; Nicolas Durupt; Antoine Greco; Jacques Thiry; Fatima Hammy; Guillaume Laroche
Original assignee: Areva Mines; Eramet
Priority date: 2011-01-06
Filing date: 2012-01-06
Publication date: 2015-10-21

Abstract

The main subject-matter of the present invention is a process for the dissolution of at least one element chosen from niobium and tantalum and at least one other element chosen from uranium and the rare earth elements, advantageously for the dissolution of niobium, tantalum, uranium and rare earth elements, present in an ore or an ore concentrate. Said process comprises : - the roasting of a material, comprising said elements, which material is mixed, dry or/in the presence of water, with an acid roasting agent in order to obtain a calcine; said material consisting of said ore or concentrate or having been obtained from said ore or said concentrate and said acid roasting agent providing for roasting in a sulphate medium; and - the dissolution in an aqueous solution of the calcine obtained in order to obtain a slurry, the liquid fraction of which includes iron, in the ferric state, at a concentration of at least 50 g/l, advantageously of at least 70 g/l and very advantageously of at least 120 g/l.

Description

Titre dissolution and recovery of at least one element Nb or Ta and of at least one other element U or rare earth éléments from ores and concentrâtes.

Abrégé :

The main subject-matter of the présent invention is a process for the dissolution of at least one element chosen from niobium and tantalum and at least one other element chosen from uranium and the rare earth éléments, advantageously for the dissolution of niobium, tantalum, uranium and rare earth éléments, présent in an ore or an ore concentrate. Said process comprises :

-the roasting of a material, comprising said éléments, which material is mixed, dry or/in the presence of water, with an acid roasting agent in order to obtain a calcine; said material consisting of said ore or concentrate or having been obtained from said ore or said concentrate and said acid roasting agent providing for roasting in a sulphate medium; and

- the dissolution in an aqueous solution of the calcine obtained in order to obtain a slurry, the liquid fraction of which includes iron, in the ferrie state, at a concentration of at least 50 g/l, advantageously of at least 70 g/l and very advantageously of at least 120 g/l.

Concentrate or Ore

-----Source ofFe

Wake+jp of sutphur dioxide----«

Water

Gonceritrated of thecaWne in the presence of iron

Ferrie roasting teachale Mb. Ta, U_tRE f FIG.1

Down stream sépara ton Nb, Ta, U, RE

O.A.P.I. - B.P. 887, YAOUNDE (Cameroun) - Tel. (237) 22 20 57 00- Fax: (237) 22 20 57 27- Site web: http:/www.oapi.int - Email: oapi@oapi.int i

Dissolution and recovery of at least one element Nb or Ta and of at least one other element U or rare earth éléments from ores and concentrâtes

The présent invention relates to the treatment of ores and ore concentrâtes for the purpose of quantltatively extractlng, from sald ores or concentrâtes, the typical éléments: niobium (Nb), tantalum (Ta), uranium (U) and rare earth éléments (RE).

A subject-matter of the présent Invention Is more particularly:

- a process for the dissolution of at least one element chosen from niobium and tantalum and of at least one other element chosen from uranium and rare earth éléments présent in an ore or an ore concentrate; and

- a process for the recovery of said dissolved éléments.

The dissolution process of the Invention provides for quantitative dissolution of niobium and/or tantalum and of uranium and/or of at least one rare earth métal (several rare earth éléments are generally présent and are dissolved according to the process of the invention) without the use of hydrofluoric acid, with the use of sulphuric acid under spécifie conditions. Said dissolution process Is particularly advantageous in that It is suitable in particular for treating ores which cannot be beneficiated or can be beneficiated only to a certain extent or with dlfflculty by a physical treatment (for example fine and/or weathered pyrochlore ores, in particular those very rich in iron).

The dissolution process of the invention Is suitable for the dissolution of ail the éléments identlfied above, assuming, very obviously, that they are présent In the ore or ore concentrate treated. It Is possible to speak of the dissolution of the NTURE (N for niobium, T for tantalum, U for uranium and RE for rare earth metal(s)) or of a process for the treatment of NTURE ores or ore concentrâtes.

• In the natural state, niobium and tantalum are often combined in complex minerais, such as pyrochlore, colombite, tantalite, colomblte-tantallte (coltan) and loparite. The ores or ore concentrâtes comprising these minerais are also capable of including uranium and rare earth éléments.

Currently, niobium is mainly produced by a pyrochlore flotatlon benefîclation process, combined with a pyrometallurgical treatment in an aluminothermie furnace (Minerais Engineering, Volume 14, Number 1, January 2001, pp 99-105 (7): Kinetlcs of Pyrochlore Rotation from Araxa Minerai Deposlts by Oliveira J.F., Saraiva S.M., Pimenta J.S. and Ollveira A.P.A.; Mining Magazine, February 1982: Araxa niobium mine). Alternatives to this process, for the production of niobium and tantalum, conslsts essentially of attack processes by the action of hydrofluoric acid on concentrâtes containing chemically refractory minerai phases, such as tantallte and colombite-tantalite (Extractive Metallurgy of Niobium (Chapter 2: Sources and their treatment procedures by C.K. Gupta and A.K. Suri, CRC Press, London, pages 49-62)). Processes which do not resort to hydrofluoric acid, based on the use of sulphuric acid, hâve also been described in the lïterature. They generally require an acid or sulphatlon roasting, combined with dissolution of the niobium and/or tantalum In a highly acidic medium (H2SO4 > 35% welght/welght) (Patent US 3 087 809; Annual General Meeting, Montreal, Aprll, 1964: The production of hlgh-purity niobium oxide from pyrochlore-perovsklte concentrate, by FJ. Kelly and W.A. Gow), or a medium comprising either carboxylic acids (Patent Application CN 1 904 097; Patent US 2 481 584), or ammonium Ions (Tsvetnye Metally/non-ferrous metals, 1986, Vol. 27, No. 11, pages 61-62: Industrial tests and Introduction of the sulphate extraction technology for the processing of low-grade tantalum-nloblum concentrâtes, by A.I. Karpukhin, G.I. Il'ina, V.G. Kharlov, A.I. Usenko and Yu. G. Popov).

• As regards the uranium, It Is generally extracted from the ores or ore concentrâtes by chemical (acid) attack, optionally carried out under pressure.

• As regards the rare earth éléments, they are generally produced from monazite, bastnaesite, xenotlme or loparite ores beneficiated by physical treatment and then attacked chemically. A portion of the current production of rare earth éléments is thus obtaîned by roasting bastnaesite and loparite concentrâtes with sulphuric acid. An alternative process has been described In Patent Application CN 1 721 559 for specifically the recovery of rare earth éléments and thorium. Said process comprises a) a calcination of a rare earth métal ore and/or rare earth métal concentrate In the presence of concentrated sulphuric acid and iron, said iron favourably playing a part, in said calcination, In the

conversion of the rare earth éléments and thorium to substances which are soluble in water or in a dilute acid solution, and b) a leaching with water or with dilute add of the calcine obtained. This document of the prior art Is silent with regard to the treatment of ores or ore concentrâtes in order to extract the niobium and/or the tantalum therefrom. It does not describe a leachate including high concentrations of Iron with reference to the extraction of niobium and/or tantalum.

Patent Application CN 101 492 771 describes a process for the treatment of a Xinganite ore concentrate for the purpose of extractlng lanthanum, béryllium, tantalum and niobium therefrom. Said process comprises the roasting of the concentrate in the presence of sulphuric acid and the dissolution in water of the calcine obtained. This document does not comprise any teachlng with regard to the opportune presence of a high content of dissolved ferrie Iron in the slurry obtained. The examples provided illustrate the presence of dissolved ferrie Iron in said slurry at contents of 11.6 g/1 (Example 1) and 1.1 g/1 (Example 2) respectively (these values were calculated from the ratios by weight shown: Fe/(Nb + Ta)).

In this context, the inventors hâve looked for and found an attack (dissolution) process:

a) which uses conventional reagents, including sulphuric acid (and which thus does not use hydrofiuoric acid); and

b) which is of high performance (the dissolution of the éléments, in particular niobium and/or tantalum, is carried out quantitatively), including when it Is carried out on ores which cannot be beneficiated or can be beneficiated only to a slight extent or with difficulty by a prior physical treatment.

Said process is based on the bénéficiai intervention of the ferrie ions in the dissolution of at least one of the éléments chosen from niobium and tantalum.

According to its first subject-matter, the présent invention thus relates to a process for the dissolution of at least one element chosen from niobium and tantalum and of at least one other element chosen from uranium and rare earth éléments, said éléments being présent in an ore or ore concentrate. Said process is advantageously employed for the dissolution of niobium, tantalum, uranium and rare earth éléments

(several rare earth éléments are generally présent together but the presence of just one rare earth element cannot be ruled out), présent together in the ore or ore concentrate concerned. It is, however, very obvlously understood that it is possible to dissolve only éléments Initially présent in said ore or ore concentrate. It should be noted here, by the way, that the process of the Invention is entlrely capable of being suitable also for the dissolution of other éléments of value présent, such as manganèse, titanium, thorium, and the like.

It should also be noted that the ore or ore concentrate treated according to the Invention can very obviously consist of, respectively, a mixture of ores or a mixture of concentrâtes. Moreover, the starting material treated can also consist of a mixed mixture of ore(s) and concentrate(s), which a person skilled In the art will generally also describe as ore or concentrate according to its contents of éléments of value. It is understood that the invention thus relates to the treatment of a starting material chosen from ores, ore concentrâtes and their mixtures (for the dissolution of the éléments identified above in a first step and the recovery of said éléments in a second step).

The ore concentrate can be of any type. It can in particular be a mlneralurglcal concentrate, resulting from a physical enrichment, or a concentrate resulting from a chemical enrichment (such as a leaching residue).

The dissolution process of the Invention characteristically comprises:

- the roastlng of a material, comprising the searched éléments, which material Is mixed, dry or in the presence of water, with an acidic roastlng agent in order to obtain a calcine; said material consisting of said ore or concentrate or having been obtained from said ore or said concentrate and said acidic roasting agent providing for roastlng In a sulphate medium; and

The process of the invention thus comprises two main stages carried out In succession: a stage of roastlng a material which includes the

desired éléments of value (éléments which it is desired to dissolve) and a stage of dîssolving the calcine obtained on conclusion of said roasting.

The material subjected to the roasting can consist of the ore or ore concentrate concerned. It may also hâve been obtained from said ore or ore concentrate. It may thus be said ore or ore concentrate physically enriched (thus upstream of the roasting) or a leaching residue (said leaching having been carried out, thus upstream of said roasting, on the said ore or ore concentrate, optionally physically enriched).

The material to be roasted can be dry or in the presence of water. It can in particular be in the form of a slurry, of a wet cake or of a dry solid.

It is (advantageously intimately) mixed with the roasting agent. Said roasting agent is an acidic agent (H₂SO₄ and/or SO₃ and/or at least one alkali métal pyrosulphate, and the like) which provides for roasting in a sulphate medium. It has been seen above that the addition of hydrofluoric acid Is not required. The desired éléments (Nb and/or Ta + U and/or rare earth element(s)) are quantitatively attacked during the roasting stage.

Characteristically, In the context of the process of the invention, the calcine obtained on conclusion of the roasting stage is dissolved in an aqueous solution (such an aqueous solution generally consists in water or in a solution containing salts (such as phosphates and/or sulfates of alkaline metals and/or earth alkaline metals and/or transition metals). Such a sait containing solution is advantageously produced in the carried out process) in order to obtaln a slurry, the liquid fraction of which (isolated after carrying out a liquid/solid séparation) includes iron, in the ferrie state, at a concentration of at least 50 g/l. This concentration is advantageously at least 70 g/l. It is very advantageously greater than or equal to 100 g/l, and even 120 g/l, and generally does not exceed the solubility limit (of Iron) in the medium, which is generally 150 g/l, in order to avoid any précipitation of iron. The amount of water used in the dissolution is advantageously limited, very advantageously it is as low as possible, in order to maxîmize the concentration of desired dissolved éléments of value. It is generally between 0.5 I of water/kg of calcine and 5 I of water/kg of calcine. It is advantageously less than 5 I of water/kg. prf

Under such conditions (sulphate medium concentrated in ions,

In particular ferrie Ions), the dissolution of the Nb and/or Ta has proved to be hïghly quantitative.

The iron présent in the slurry can originate from the calcine or hâve been, at least in part, added to the calcine (in the liquid and/or solid form (the iron can more particularly be added in the dissolution aqueous solution), under any valency (in any case, it is, in situ, In the form of ferrie sulphate)). The source of iron can correspond in particular to an iron ore or to an iron concentrate in the form of oxide or sulphide (for example pyrites), a ferrous or ferrie sait (for example FeSO-t, Fe₂(SO4)₃ or FePOq), a solution comprising iron or Iron métal. It is to the crédit of the inventors to hâve demonstrated the critical nature of the presence of the Iron during this stage of dissolution of the calcine.

Said stage of dissolution of the calcine can be carried out according to different alternative forms. The calcine can be cooled in order to be subsequently crushed or ground before dissolution. It can also be used as Is, under hot conditions, on departing from the roastlng furnace. It is generally slurried In the limited amount of aqueous solution at a température between ambient température and the boiling point of the medium, advantageously at a température between 85°C and a température sllghtly below said boiling point.

On the conclusion of the dissolution of the calcine, the slurry or the liquor obtained after carrying out a liquid/solid séparation on said slurry can be diluted (by a limited factor, generally not exceeding 5) without substantial précipitation of the dissolved éléments of value (mainly Nb and/or Ta). Such a dilution Is targeted at preventing any crystalllzation of the dissolved salts, such as ferrie sulphate. Nb and/or Ta présent in the roasting leachate obtained after carrying out a liquid/solid séparation of the above type, can be separated (by various techniques, such as extraction by solvent(s) or sélective précipitations). The resulting solution can advantageously be recycled upstream of the process, upstream of the first roasting stage, at a preliminary leaching stage (see below). Such a recycllng Is doubly opportune in that It Involves, first, the recycling of the acldlty and, secondly, the recycling of unseparated element(s) of value. A séparation of Nb and/or Ta from the roasting leachate has indeed been mentioned above but a person skilled in the art very obviously

understands that It is not necessarlly a 100% séparation. This comment on the term séparation and on the term separated is valid throughout the présent text and appended figures.

Due to the high concentration of iron, the process of the Invention makes It possible to dissolve the desired éléments of value at a restricted free acidity: 0.1 N < [H⁺] << 15 N, preferably:

N <_[H⁺] < 6 N.

It is surprising that, under such acidity conditions, the niobium and/or the tantalum présent remain(s) ln solution. A person skilled ln the art Is actually not unaware that, ln a medium of restricted acidity, niobium and tantalum sulphates are normally respectlvely hydrolysed to nioblc acid (NbjOs.nHaO) (Nouveau traité de chimie minérale (oxyde de niobium pentavàlent hydraté ou oxyde nioblque)) [New Treatise on Inorganic Chemistry (pentavalent niobium oxide hydrate or nlobic oxide)], by P. Pascal and R. Rohmer, Volume XII, p. 454) and tantallc acid (Ta2O5-nH₂O), which are insoluble. The presence of Iron is thus particularly opportune.

Thus, the dissolution of the calcine according to the invention 1s advantageously carried out without any addition of sulphuric acid. It should be noted that the addition of a make-up of sulphuric acid, in order to optimize the level of acidity, cannot be completely ruled out but that, in any case, the dissolution of the calcine according to the Invention is not carried out in a highly acidlc medium (as is, for example, recommended ln Patent US 3 087 809).

The dissolution of the calcine is advantageously carried out (more particularly subséquent to direct roasting of ore (without prior leaching: see below)) with addition of at least one, generally one, reducing agent to said calcine. Such a reducing agent must not be used ln an excessive amount as it is advisable for the iron présent ln the dissolution to remain In the ferrie state. This reducing agent plays a part in particular with reference to the dissolution of the manganèse. Suiphur dioxide (SO₂) is suitable as such a reducing agent.

The roasting stage of the process of the Invention can consist of an acid roasting or a sulphation roasting. The person skilled in the art knows both these types of roasting. The first is a liquld/solid (H₂SO4/material to be roasted) reaction and the second is a gas/solid

(gaseous SO₃ (generally generated in situ from SO₂, H2SO4, at least one alkali métal pyrosulphate, and the llke)/material to be roasted) reaction,

It is intended to provide below some Information regarding these two types of roasting. Said information is in no way limiting.

Acid roastlng

It is generally carried out between 150°C and 400°C, using sulphuric acid, between 100 kg of H2SO4 /t of dry matter to be roasted and 3000 kg/t, preferably between 500 kg/t and 1500 kg/t. The acidlc solution which Imprégnâtes the material to be roasted then comprises between 30% and 98% of H2SO4 according to the moisture content of the material which is roasted with the acid. Its boiling point is generally between 108 and 340°C. This acid roasting can be carried out below or above the boiling point, preferably between 250 and 300°C. The minerai phase carrying the NTURE (at least some of them) is quantitatively attacked, for example 90% of a pyrochlore phase.

Soluble métal sulphates are then formed.

The métal oxides of the minerai phases are attacked by the sulphuric acid according to:

MO_X + x H2SO4 -> M(SO₄)x + x H2O, in particular:

- iron: 2 FeOOH + 3 H₂SO₄ -> Fe₂(SO4)₃ + 4 H₂O

- rare earth éléments (RE): RE₂O₃ + 3 H2SO4 -> RE₂(SO4)₃ + 3 H2O.

According to Ya. G. Goroshchenko (in the Journal de Chimie Inorganique, Vol. I, No. 5, 1956, pp. 903-908), Nb₂Os can form various sulphate compounds on reacting with sulphuric acid.

According to Ya. G. Goroshchenko (see above) and Patent CN 1 904 097, the niobium présent in the pyrochlore is supposed to be attacked during a roasting above 150°C to give Nb₂O₃(SO₄)2.

The attack reaction for the tantalum is assumed to be identical.

Sulphation roasting

It is generally carried out between 400^QC and 700°C, preferably between 450°C and 550°C, using gaseous sulphur trioxide (SO₃), which is formed either by oxidation of sulphur dioxide (SO2) or by décomposition of sulphuric add (H2SO4) or by décomposition of alkali métal pyrosulphate (M2S2O7, where M = Na, K, and the like).

The métal oxides of the minerai phases are sulphated by the action of gaseous SO3 according to:

ΜΟχ + x SO3 (g) => MfSO-ûx, in particular:

iron: 2 FeOOH + 3 SO₃ —> Fe₂(SO4)3 + H₂O

- rare earth éléments (RE): RE₂O₃ + 3 SO₃ -» RE₂(SO4)₃

Soluble métal sulphates are then formed.

It Is assumed that the niobium présent in the pyrochlore is attacked during the sulphation roastlng to give Nb₂O3(SO4)₂.

The attack reaction for the tantalum Is assumed to be identlcal.

The uranium and the rare earth éléments liable to be présent in the material to be roasted (pyrochlore or other minerai phases) can exist In different valency states. On conclusion of the implémentation of the roastlng, they are converted to a form which is soluble In the concentrated ferrie sulphate medium.

The roasting of the process of the Invention, whatever the exact procedure, is advantageously carried out in the presence of iron, In the ferrie form.

The presence of ferrie iron in the roasting (In addition to the presence of ferrie Iron In the dissolution: see above) Improves the yields of attack on the minerai phase carrying the niobium and/or tantalum, in particular pyrochlore. This improvement is generally significant for levels of Iron such that the Fe/(Nb 4- Ta) molar ratio is greater than 2. It is understood that the ratio concerned is read Fe/Nb, Fe/Ta or Fe/(Nb + Ta) according to the presence, together of separately, of the éléments Nb and Ta In the material concerned (one at least of said éléments obviously being présent). Said molar ratio Fe/(Nb + Ta) is advantageously greater than 3.5 and very advantageously greater than 6.

The amount of iron opportunely présent in the roastlng can be présent in the material to be roasted (for example a partial leaching residue (see below) from an iron ore) or be added, at least in part, in the liquid and/or solid form, to said material to be roasted (it is also possible to add the Iron by Introducing a portion of the ultrafines from an ore or ore concentrate, referred to as sûmes, separated beforehand). The source of Iron can thus be in various physical states solid and/or liquid, and in ail the possible valency states of iron (under the conditions of/tff ίο

Implémentation of the roasting, said iron Is In the ferrie state). The source of iron can correspond in particular to an iron ore or to an iron concentrate in the form of oxide or sulphide (for example pyrites), a ferrous or ferrie sait (for example FeSO₄, Fe2(SO₄)₃ or FePO₄), a solution comprising iron or iron métal. Whatever the source of iron, if it brings said Iron under a valency other than ferrie, the Iron is oxldized in situ during the roasting.

It is thus to the crédit of the inventors to hâve demonstrated the opportune presence of ferrie iron, as aid for the dissolution of Nb and Ta, during the dissolution of the calcine and advantageously during the roasting and the dissolution of the calcine, in a context of the treatment of NTURE ore or ore concentrate.

Furthermore, the inventors hâve also noted the opportune presence of phosphate in the roasting. Thus, according to an advantageous alternative embodiment of the process of the invention, the roasting Is carried out in the presence of phosphate, advantageously présent at a PO₄/(Nb + Ta) molar ratio of greater than 2 and very advantageously of greater than 6; said phosphate being already présent in the material to be roasted and/or having been added, at least In part, In the liquid and/or solid form, to said material to be roasted.

It has been seen above that the material to be roasted can consist of an ore (direct roasting), of an ore concentrate (concentrate obtained subséquent to a prior enrlchment: it is also possible to speak of direct roasting, in the context of implémentation of the process of the Invention) or of a material obtained from such ores or ore concentrâtes (reference is made here to indirect roasting, insofar as the process of the Invention encompasses, upstream of the roasting, a treatment of said ore or ore concentrate).

Thus, according to an alternative form, the process of the invention comprises, upstream of the roasting, a physical enrlchment and/or a chemîcal treatment of the ore or ore concentrate to be treated.

Such a physical enrichment can be based on any conventional method for the physical enrichment of a solid material; for example, it can consist of a low-intensity magnetic séparation (In order to remove the magnetic) or of a flotation (of the silica flotation or apatite flotatron type). ₍

Such a chemical treatment advantageously comprises an acid atmospheric leachlng, very advantageously a sulphuric acid atmospheric leachlng. Such a leaching can thus be carried out on an ore, an ore concentrate, a physically enriched ore or a physically enriched ore concentrate.

Such a leaching is advantageous for upstream dissolution of the matrix (of the éléments of the matrix, such as Fe, Al, P, Mn, and the like) by the action of the used acid, very advantageously sulphuric acid. It only partlally attacks the desired éléments of value (NTURE). It makes it possible to reduce the amount of material to partlclpate in the roasting. A significant proportion of the uranium présent can be extracted at this stage (typically 60 to 80%). However, the niobium and/or the tantalum are not quantitatively dissolved by such a leaching when these éléments are within refractory minerai phases, for example within pyrochlore.

Such a leaching is advantageously carried out according to one or other of the two alternative forms specified below.

Specifically, said leaching can be carried out:

- either In a single stage, referred to as cocurrentwlse;

- or In 2 stages countercurrentwise. The first stage Is referred to as mild as the acidity is restrlcted by the addition of the material to be leached (ore or concentrate), which neutrallzes the recycled acidity. The second stage is referred to as aggressive as this is the stage where the acid is introduced.

The leaching agent is very advantageously sulphuric acid (H₂SO₄).

In other words, the dissolution process of the Invention advantageously comprises, upstream of the roasting:

A) a leaching, in a single stage, cocurrentwise, in the presence of (added) sulphuric acid (such a leaching is generally carried out at a température of between 10 and 115°C), followed by a solid/liquid séparation which produces a leachate (to be recovered upstream of the roasting and which includes a portion of the desired éléments) and a leaching residue to be roasted; or

B) a leaching, in two stages, countercurrentwise, comprising a first leaching stage (carried out under mild conditions (without any addition of sulphuric acid) by the action, on the material to be leached, of<

a liquid with recycled residue acidity (see below)), followed by a solld/llquld séparation which produces a leachate (to be recovered upstream of the roasting and which Includes a portion of the desired éléments) and a first leaching residue, and a second leaching stage carried out on said leaching residue, in the presence of sulphurlc acid (under aggresslve conditions), followed by a solid/liquid séparation which produces a second leaching residue to be roasted and a liquid with residue acidity to be advantageously recycled (this second leachate can be recycled as it is or after having separated at least a portion of the éléments of value présent thereln (see below)) to said first (mild) leaching stage.

The leaching prestage is advantageously (with reference to an optlmizatlon of the consumption of acid) carried out according to its alternative form B) above, which thus produces, upstream of the roasting, on the one hand, after the mild leaching, a leachate of restricted residual acidity and, on the other hand, after the aggresslve leaching, a leaching residue to be roasted.

A leaching, in a single stage, cocurrentwlse, in the presence of H₂SO₄, as mentioned in point A) above, can be carried out under a ferrie condition or under a ferrous condition. Each of said ferrie and ferrous conditions Is speclfled below.

Ferrie:

The potential is greater than 500 mV Ag/AgCI, preferabiy greater than or equal to 600 mV/AgCI, so as to lîmit the amount of ferrous Ions in solution.

The leaching agent is sulphuric acid (H₂SO₄).

The gangue, for example a mixture of crandallite and goethite, is leached according to:

Fe(III)OOH + 3 H₂SO₄ -> Fe(III)₂(SO₄)₃ + 4 H₂O

CaAI₃(PO₄)(HPO₄)(OH)₆ + 11 H₂SO₄ —

AI₂(SO₄)₃ + 2 CaSO₄.2H₂O + 8 H₂O + 4 H₃PO₄.

The attack on one mole of Fe in a ferrie medium thus consumes 1.5 mol of H₂SO₄ and that on one mol of candallite consumes 5.5 mol of H₂SO₄. pd

A reducing agent (at least one), such as sulphur dioxide (SO2) or métal (iron, zinc, scrap Iron, and the like) as divided or massive material (métal powder, scraps, and the like), can be Introduced as makeup. It plays a part, a priori, advantageously with reference to the réduction of the manganèse oxides possibly présent in the ore or ore concentrate treated.

- Ferrous:

The potential is less than 500 mV Ag/AgCI and preferably less than or equal to 450 mV Ag/AgCI, so as to llmit the amount of ferrie Ions in solution.

The leaching agent Is sulphuric acid (H₂SO₄). If sulphur dioxide (SO₂) is présent, it performs the double rôle of leaching agent and of reducing agent (see below). The réduction of the medium Is obtained by addition of (at least) one reducing agent, such as sulphur dioxide (SO₂), métal (iron, zinc, scrap iron, and the like) as divided or massive material (métal powder, scraps, and the like). The Iron Is leached in the Fe²⁺ form according to:

Fe(III)OOH + 3 H₂SO₄ -» Fe(III)₂(SO₄)₃ + 4 H₂O (1) Fe(ni)₂(SO₄)₃ + SO₂ + 2 H₂O - 2 Fe(II)SO₄ + 2 H₂SO₄ (4a)

Fe(III)OOH + H₂SO₄ + SO₂ -> 2 Fe(II)SO₄ + 2 H₂O (4)

The réduction (4a) makes it possible to regenerate the acidity, with the resuit that an attack in a ferrous medium on one mole of Fe by H₂SO₄ (in the possible presence of the reducing agent SO₂) consumes one mole of acid (H₂SO₄ and SO₂ combined), that is to say half a mole less than in the ferrie medium.

The ferrie ions présent participate in the dissolution of the niobium or (and, if both are présent) of the tantalum (see above) and, on converting said ferrie Ions to ferrous Ions, the précipitation Is brought about of said niobium or (and) of said tantalum, leached out, dissolved, before complété réduction of said ferrie ions to ferrous Ions. Carrylng out the leaching at a température close to the boiling point, for example above 90°C, makes it possible to further improve the précipitation of the

niobium or (and) of the tantalum as this promûtes the thermal précipitation of these éléments.

The leaching is consequently preferably carried out in a ferrous medium, so as to obtain séparation of Nb or (and) Ta with respect to II or (and) rare earth éléments starting from upstream of the process. This also makes it possible to limlt the consomption of leaching agent for the matrix (for example H₂SO₄ and SO₂ combined).

Thus, the leaching mentioned In point A) above can be carried out according to the two alternative forms below and conslsts of:

Al) a leaching, in a single stage, cocurrentwise, at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mV Ag/AgCI, in the presence of (added) H₂SO₄ and optionally of a make-up of at least one (generally one) reducing agent, such as sulphur dioxide (SO₂) or métal (iron, zinc, scrap Iron, and the like) as dlvided or massive material (métal powder, scraps, and the like), followed by a solld/liquid séparation which produces a ferrie leachate (to be recovered upstream of the roasting and which includes a portion of the desired NTURE éléments) and a leaching residue to be roasted; or

A2) a leaching, in a single stage, cocurrentwise, at a potential of less than 500 mV Ag/AgCI and advantageously of less than or equal to 450 mV Ag/AgCI, in the presence of (added) H₂SO4 and of at least one (generally one) reducing agent, such as sulphur dioxide (SO₂) or métal (iron, zinc, scrap iron, and the like) as divided or massive material (métal powder, scraps, and the like), followed by a solid/liquid séparation which produces a ferrous leachate (to be recovered upstream of the roasting and which essentially Includes uranium and/or rare earth éléments) and a leaching residue enriched in niobium and/or tantalum to be roasted.

In the same way, the leaching, in two stages, countercurrentwise, mentioned in point B) above, can be carried out according to the two alternative forms specified below and consists of:

Bl) a leaching, in two stages, countercurrentwise, comprising a first leaching stage, at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mV Ag/AgCI, in the presence of H₂SO₄ (recycled, advantageously from the second leaching stage) and optionally of a make-up of at least one (generally one) reducing agent, such as sulphur dioxide (SO₂) or métal (Iron, zinc, scrapi iron, and the like) as divided or massive material (métal powder, scraps, and the like), followed by a solld/liquld séparation which produces a ferrie leachate (to be recovered upstream of the roastlng and which Includes a portion of the desired éléments) and a first leaching residue, and a second leaching stage carried out on said first leaching residue, either at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mv Ag/AgCI, in the presence of H2SO4 and optionally of a make-up of at least one (generally one) reducing agent, such as sulphur dloxide (SO₂) or métal (Iron, zinc, scrap iron, and the like) as divided or massive material (métal powder, scraps, and the like), or at a potential of less than 500 mV Ag/AgCI and advantageously of less than or equal to 450 mV Ag/AgCI, In the presence of H2SO4 and of at least one (generally one) reducing agent, such as sulphur dloxide (SO2) or métal (Iron, zinc, scrap iron, and the like) as divided or massive material (métal powder, scraps, and the like), followed by a solld/liquld séparation which produces a second leaching residue to be roasted and a liquid (ferrie or ferrous leachate) with resldual acidlty to be advantageously recycled to said first leaching stage; or

B2) a leaching, in two stages, countercurrentwise, comprising a first leaching stage, at a potential of less than 500 mV Ag/AgCI and advantageously of less than 450 mV Ag/AgCI, in the presence of H2SO4 (recycled, advantageously from the second leaching stage) and of at least one (generally one) reducing agent, such as sulphur dioxlde (SO2) or métal (Iron, zinc, scrap iron, and the like) as divided or massive material (métal powder, scraps, and the like), followed by a solid/llquid séparation which produces a ferrous leachate (to be recovered upstream of the roasting and which essentially includes uranium and rare earth éléments, if they were présent initially) and a first leaching residue, and a second leaching stage carried out on said first leaching residue, either at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mV Ag/AgCI, in the presence of H2SO4 and optionally of a make-up of at least one (generally one) reducing agent, such as sulphur dloxide (SO₂) or métal (iron, zinc, scrap iron, and the like) as divided or massive material (métal powder, scraps, and the like), or at a potential of less than 500 mV Ag/AgCI and advantageously of less than or equal to 450 mV Ag/AgCI, in the presence of H2SO4 and of at least one<^ (generally one) redudng agent, such as sulphur dioxide (SO₂) or métal (iron, zinc, scrap iron, and the like) as divlded or massive material (métal powder, scraps, and the like), followed by a solld/llquid séparation which produces a second leaching residue enriched in niobium and/or tantalum to be roasted and a liquid (ferrie or ferrous leachate) to be advantageousiy recycled to said first leaching stage. Said liquid to be advantageousiy recycled to said first leaching stage can be recycled directly (as it is) or after having separated at least partially the niobium and/or tantalum présent therein (after séparation, at least partially, of the niobium and/or tantalum présent therein). Such a séparation can be of any type. It can consist in particular of a sélective précipitation or of a solvent extraction.

The alternative forms Bl) and B2) above thus comprise a first leaching stage carried out, respectively, under ferrie conditions (Bl)) and under ferrous conditions (B2)) and a second leaching stage carried out under ferrie or ferrous conditions.

It is seen that the countercurrent thus employed between the acid and the material to be leached (ore or concentrate) is particularly advantageous with reference to an optimization of the consumptlon of acid necessary for the dissolution of the entities of the matrix and to an optimization of the attack on the matrix.

It has been understood in the light of the above remarks that the leaching prestage is advantageousiy carried out according to its alternative forms A2) and B2) above which thus produce, upstream of the roasting, on the one hand a ferrous leachate (rlch in uranium and/or rare earth éléments) and, on the other hand, a leaching residue to be roasted enriched In niobium and/or tantalum. Said leaching prestage Is very advantageousiy carried out according to the alternative form B2).

The material to be roasted is advantageousiy ground upstream of the roasting. Such a grinding is intended to improve the performance of the roasting of the material by promoting the accessibility of the reactants to the minerai phases carrying the éléments of value. Grinding Is advantageousiy carried out In order to obtain 100% passing at 30 pm of the mass to be roasted (I.e., exhibiting a particle size < 30 pm), advantageousiy in order to obtain 100% passing at 15 pm of the mass to be roasted (i.e., exhibiting a particle size < 15 pm) and very advantageousiy in order to obtain 100% passing at 10 pm of the mass to^ be roasted (Le., exhibiting a particle size < 10 pm). The grinding concerned Is thus an ultraflne grindlng. It should be noted that, in any case, even in the absence of an ultraflne grindlng, the operation Is generally carried out, in the context of the process of the Invention, on material exhibiting a particle slze of less than 700 microns and advantageously of less than 400 microns (which has generally been subjected to a pregrinding).

A person skilled in the art has understood that the optional stages of the process of the Invention explained above, to be carried out upstream of the roasting, are advantageously carried out in the following order:

- physical enrichment,

- leaching, and

- grlnding, assuming, of course, that at least two of said stages are carried out.

There may also be provided, in the case of the treatment according to the invention of an iron ore or iron ore concentrate, a desliming of said ore or concentrate, targeted at separating the stimes (the ultraflnes) from the other particles. Such a desliming can be carried out by mlcrocycloning. The particle slze cutoff is typically between 5 and 20 pm, preferably between 5 and 10 pm. A portion of the sûmes recovered Is advantageously used downstream in the process of the Invention as source of Iron (in the dissolution of the calcine and/or In the roasting),

As regards the downstream of the dissolution of the calcine according to the invention, an optional dilution by a limited factor (which generally does not exceed 5) In order to prevent a précipitation of the éléments to be dissolved, was mentloned above. It may generally be Indicated that the process of the invention advantageously comprises, downstream of the dissolution of the calcine:

- a solld/ltquid séparation carried out directly on the slurry or after an additional stage of dilution of said slurry by a limited factor (advantageously < 5) in order to prevent the précipitation of the dissolved éléments; said solid/llquid séparation producing a solid attack residue and a roasting leachate including said dissolved éléments, ecf

In the context of this advantageous alternative form, It is opportune to provide for:

- the séparation of niobium and/or tantalum, In solution in said roastlng leachate, and advantageously the recycling of said leachate, from which said niobium and/or tantalum were separated, which is optionally diluted, for the implémentation of a sulphuric acid atmospheric leaching upstream of the roasting (see above).

The séparation of niobium or (and) of the tantalum from a solution In which It (they) are présent was mentioned above. Such a séparation can be carried out in particular by solvent extraction (cf. figures) or by sélective précipitations. The solution purified from Nb and/or Ta Is advantageously recycled. Its acldlty is thus recycled.

It is intended to specify below two preferred embodiments of the process of the Invention. These two embodiments are represented dlagrammatically in the appended Figures 3 and 4.

According to a first embodiment, the dissolution process of the invention comprises:

- a leaching of A) or B) type above (advantageously of Al), A2), Bl) or B2) type above and very advantageously of A2) or B2) type above) and the recovery (downstream) of the leachate and of a leaching residue;

- an optional ultrafine grlnding of said leaching residue;

- the roasting of said optionally ground leaching residue;

- the dissolution of the calcine obtained on conclusion of said roasting, in order to obtain a slurry;

- a liquld/solid séparation, carried directly on said slurry or on said slurry diluted by a limited factor (see above), which produces a solid attack residue and a roasting leachate including the desired éléments of value (dissolved);

- the séparation of the niobium and/or tantalum from said roasting leachate and the recycling of the leachate, from which said niobium and/or tantalum were/was separated (see the above comment with regard to the term separated), which Is optionally diluted, for the implémentation of said leaching.

According to a second embodiment, the dissolution process of the invention comprises:^

- a desliming of an ore or concentrate (comprising iron), which Is optionally physically enrlched, for the recovery, on the one hand, of sûmes including iron and, on the other hand, of the deslimed ore or concentrate (i.e., ore or concentrate freed from at least a portion (generally more than 50% by weight) of Its sûmes);

- a leachlng of A) or B) type above (advantageously of Al), A2), Bl) or B2) type above and very advantageously of A2) or B2) type above) carried out on said deslimed ore or concentrate, optionally with addition of a portion of the sûmes recovered In step 1 above (generally the portion of said recovered sûmes not used below in the roasting), and the downstream recovery of a leachate and of a leaching residue;

- an optional ultrafine grinding of said leaching residue;

- the roasting of said optionally ground leaching residue, carried out with addition of iron via another portion of the sûmes recovered in step 1 above;

- the dissolution of the calcine obtaîned on conclusion of the said roasting, in order to obtain a slurry;

- a liquld/solld séparation, carried directly on said slurry or on said slurry diluted by a limited factor (see above), which produces an attack residue and a roasting leachate Including the desired éléments (dissolved);

- the séparation of the niobium and/or tantalum from said roasting leachate and the recycling of said roasting leachate, from which the niobium and/or the tantalum was/were prevtously separated, which is optionally diluted, for the implémentation of said leaching.

The process described above, for the dissolution of the identified éléments: Nb and/or Ta + U and/or rare earth éléments, is suitable for treating any ore or ore concentrate including said éléments. It is suitable In particular for treating ores and ore concentrâtes, the ore concerned being chosen from minerais of the pyrochlore (of general formula: A₂B₂O6(O,OH,F), with A = U, RE, Na, Ca, Ba, Th or Bi (in particular) and B = Nb, Ta, Tl or Fe (in particular)), euxenite, samarsklte, perovskite or fergusonite groups and their mixtures. Such ores and ore concentrâtes optionally comprise Iron.

It has already been Indicated above that it is particularly advantageous in that it is suitable for treating ores which cannot be

beneficiated or can be beneficiated only to a slight extent or with difficulty by a prior physical treatment.

On conclusion of the implémentation of the dissolution process of the invention described above, the desired éléments of value are thus obtained in solution, in fact présent in at least one solution. Said process is capable in particular resulting In at least one or more solutions including Nb and/or Ta + U and/or rare earth éléments.

In any case, the NTURE présent in the treated material are reencountered in solution. To recover them, It is appropriate to separate them from their solvent. Processes famtliar to a person skilled in the art for obtainlng such séparations: solvent extraction, sélective précipitation, and the like, hâve been mentioned above.

According to its second subject-matter, the présent Invention thus relates to a process for the recovery of at least one element chosen from niobium and tantalum and of at least one other element chosen from uranium and the rare earth éléments, advantageously for the recovery of niobium, tantalum, uranium and rare earth éléments, présent in an ore or an ore concentrate, characterized in that it comprises:

- the dissolution of said éléments (of value) according to the dissolution process described above; and

- the séparation of said éléments (of value).

A person skilled in the art knows various technologies for separatlng these éléments from one another.

It is now intended to describe the invention, without implied limitation, with reference to the appended figures. These figures repeat and contain a great deal of information. This information forms an intégral part of the description of the invention.

Figure 1 gives a diagrammatic représentation of a reference alternative embodiment of the processes of the invention (dissolution, followed by recovery).

Figures 2A1, 2A2, 2B1 and 2B2 give a diagrammatic représentation of alternative forms of the leaching stage optionally carried out upstream of the roasting of the processes of the invention.

Figure 3 gives a diagrammatic représentation of alternative embodiments of the processes of the invention with prior leaching. X

Figure 4 gives a diagrammatic représentation of an alternative form of Figure 3 carried out on an ore with use of sûmes of said ore as source of iron.

Prellmlnary comments:

1) The optional operations (optional stages, optional additions) hâve been indicated, In the figures, in dotted Unes;

2) The material treated Is assumed to Include Nb, Ta, U and RE.

The process of the invention, represented diagrammatically in Figure 1, is carried out directly (= without upstream leachlng) on an ore or ore concentrate including the éléments of value: Nb, Ta, U and RE.

It optionally comprises a stage of ultrafine grinding (in order to obtain a material to be roasted exhibiting a particle size of less than or equal to 30 pm) of the material to be treated (ore or ore concentrate).

It comprises the successive stages of roasting (the two alternative forms of sulphation roasting and acid roasting are represented diagrammatically) and of dissolution of the calcine obtained on conclusion of said roasting. Said dissolution stage is carried out In the presence of iron with a small amount of water (a concentrated dissolution is concerned in order to obtain a slurry including at least 50 g/l of iron). The iron présent (in the ferrie state In said dissolution stage) can orlginate, at ieast In part, from the ore or ore concentrate treated; it can be added, at least In part, to the calcine; It can also be added only, at least in part, to the material to be roasted. SO₂ (reduclng agent) may also optionally be added In the dissolution stage. It is optionally added in a small amount (an optional make-up of SO₂ has been shown), insofar as, as stated above, the iron présent has to be présent in the ferrie state.

The slurry obtained on conclusion of the dissolution stage is directly (according to the alternative form represented) subjected to a liquid/solld séparation. The solid constitutes the attack residue and the liquid or ferrie roasting leachate includes the dissolved éléments of value Nb, Ta, U and RE.

The final stage represented is that of the séparation of said dissolved éléments of value (for their separate recovery).

Figures 2A1 and 2A2 illustrate a first alternative embodiment of an upstream (acid atmospheric) leaching (variant A above), which is limited to an acid attack by H2SO4, followed by a solld/liquid séparation. The amount of material to be roasted is thus reduced. An atmospheric leachate which includes a portion of the éléments of value initially présent In the material to be treated is recovered. The following hâve been more precisely illustrated:

- In Figure 2A1, the use of said cocurrentwise leaching under ferrie conditions. The reducing agent, for example SO₂, is not compulsorlly added (hence the dotted Unes) and, In any case, it is only added as makeup. The yields for extraction of the éléments of value (présent in the material to be leached) vary according to the element concerned. They generally remain low for niobium and tantalum (see Examples a ànd β below);

- In Figure 2A2, the use of said cocurrentwise leaching under ferrous conditions. A reducing agent, for example SO₂, is compulsorlly added. This reducing agent brings about the précipitation of the leached niobium and/or tantalum (assuming that these two éléments were présent in the material to be leached) and a ferrous atmospheric leachate, which essentially includes uranium and rare earth éléments (If they were initially présent in the material to be leached), and a leaching residue enriched In niobium and tantalum (assuming that these two éléments were présent in the material to be leached) are then obtained (see γ below).

Figures 2B1 and 2B2 illustrate a second alternative embodiment of an upstream (acid atmospheric) leaching (variant B) above). The material to be leached is subjected to a first leaching stage which is described as a mild leaching stage, Insofar as it is carried out, without contribution of H₂SO4, by the action on said material of a recycled liquid which exhiblts a residual acidity, due to the use of H₂SÜ4for carrying out a second leaching stage; said recycled liquid being composed of the leachate from said second leaching stage.

Said second leaching stage (which is described as an aggressive leaching stage (due to the use of H2SO4)) is thus carried out on the (solid) leaching residue from the first leaching stage.

Each leaching stage is conventionally followed by a solid/liquid séparation which generates a leachate (liquid) and a residue (solid).

Such a two-stage upstream leachlng generates, like the singlestage leachlng (see Figure 2A), an atmospheric leachate, which includes a portion of the desired éléments of value (in solution), and a leaching residue to be treated according to the invention (roasting + concentrated dissolution in the presence of iron).

The following hâve more specifically been illustrated:

- in Figure 2B1, the implémentation of the countercurrentwise leachlng, with a first mild leaching stage under ferrie conditions (only a make-up of reducing agent, such as SO₂, is fiable to be added), from where a ferrie atmospheric leachate, which includes a portion of the desired NTURE éléments (présent In the material to be leached) is obtained, and a second aggressive leaching stage under ferrie conditions;

- in Figure 2B2, the implémentation of the countercurrentwise leachlng, with a first mild leaching stage under ferrous conditions (see the involvement of the reducing agent), from where a ferrous atmospheric leachate, which includes essentially uranium and/or rare earth éléments (if they were présent inltially in the material to be leached), and a leaching residue enrlched in niobium and/or tantalum are obtained. It is noted that the leachate from the second aggressive leaching carried out under ferrie conditions can be recycled (with its acidity) directly or after having separated Nb, Ta, and the like from It. A total séparation has been represented diagrammatically.

It Is recalled here, by the way, that the second aggressive leachings, represented carried out under ferrie conditions in said Figures 2B1 and 2B2, without addition of a make-up of reducing agent, can altogether, in the context of the invention, also be carried out under ferrie conditions, with addition of such a make-up or under ferrous conditions (tous with obligatory addition of a reducing agent).

Figure 3 illustrâtes alternative embodiments of the process of the Invention with upstream atmospheric leaching and an advantageous alternative embodiment with regard to the downstream treatment of the ferrie roasting leachate obtained.

In said Figure 3, Figure 1 is in fact reencountered, i.e. toe reference stages of the process of the invention (optional ultraflne grlndlng + roasting + concentrated dissolution of the calcine ln the presence of iron + solid/liquid séparation with recovery of the liquid (ferrie roastlng leachate Including in solution the desired éléments of value).

In addition, an upstream leaching (according to Figure 2A1, 2A2, 2B1 or 2B2) has been provided and, downstream, according to one alternative form, the treatment of the leachate in order to recover the Nb and the Ta présent therein in solution has been provided. According to the alternative form represented (a solvent extraction being carried out, said Nb and Ta are recovered in a concentrated stream (it is possible here altogether to provide, as Indlcated above, a recovery In the solid form, after précipitation). According to the represented alternative form, the leachate, from which said Nb and Ta was/were previously separated, is recycled to the upstream leaching.

Figure 4 shows the alternative forms of the process of Figure 3 carried out with an ore comprising iron and with use of said Iron as source of iron In the roasting stage so that the ferrie Iron is présent at the dissolution (and In the roastlng). It should be noted that It Is not ruled out that additional Iron be added in the dissolution and/or In the roastlng. The iron Is recovered from the ore comprising it by desliming and the sûmes are so used as source of Iron.

The atmospheric leaching Is carried out on the deslimed ore, to which can be added a portion of the slimes recovered (the portion not used In the roasting, so that the problem of the control of slimes does not arise).

With reference to Figures 3 and 4, it has been understood that, according to the type of upstream atmospheric leaching carried out, the atmospheric leachate includes a portion of the desired NTURE éléments (a portion of those présent in the material to be leached) (leaching according to Figures 2A1 and 2B1) or essentially uranium and/or rare earth éléments (présent In the material to be leached) (leaching according to Figures 2A2 and 2B2).

With reference to Figures 1, 3 and 4, it has to be noted that a stream of water and a possible stream of source of Iran hâve been shown at the dissolution step. It Is obvious that a single stream of an aqueous solution could be used in place of a stream of water and a stream of source of iron.

It is intended now to illustrate the invention by examples,

EXAMPIES

Ail the examples were carried out with ore pretreated physically by low-lntenslty magnetic séparation (LIMS) and particle size réduction below 315 microns, except in the case of Example β·, where the particle size réduction was carried out to 630 microns.

A - Direct sulphatïon roasting

Example 1: Direct sulphation roasting with ultrafine grlnding

Pyrochlore ore was ground in a vertical mill stirred with balls, so as to obtaln a particle size corresponding to passing at 4 microns of 80% of the weight.

The analysis of the ore is given below:

Ore- ’	*^0	. LU* - ' vn.
Contents	33.8%	0.27%	1100 ppm	0.81%

					^‘sr. ·*.
2.5%	1100 ppm	2.1%	3.7%	3.3%	1.0%

451 g of slurry of ground ore in water at a solid content of 35%, i.e. 160 g of dry ore présent, are introduced into a porcelain crucible. 251 g of 96% sulphuric acid are added, i.e. 1503 kg of hhSCWtonne of ore. 24.3 g of sodium pyrosuiphate (Na₂S₂O7) are also added, i.e. 150 kg/tonne of ore. The mixture is stirred mechanically in order to ensure homogeneity. The Fe/(Nb+Ta) and POVfNb+Ta) molar ratios in the charge at roasting are respectlvely 22 mol/mol and 4.4 mol/mol. The crucible is placed in a kiln at 500°C for 4 hours. The roasted material Is cooled. 298 g of calcine are obtained and the weight loss during the roasting is estimated at 58%, including 30% attributable to loss of water from the ore slurry.

116 g of the calcine are Introduced into a stirred reactor containing 142 ml of water heated to 90°C, I.e. a water/lnitlal ore ratio of 2.4. This slurry is stirred and maintained at température for 2 hours.

The slurry Is centrifuged and the solid centrlfuging deposit is washed. The following results are then obtained:

			ws	æaær
12.72 g	Solid residue (%)	10.0%	0.18%	1.3%	1.4%	0.09%	5.3%
110 ml	Roasting leachate (g/1)	120	0.42	0,44	8.0	0.30	4.4
315 ml	Aqueous wash liquor (g/t)	15.6	0.26	0.70	1.00	0.03	0.57

The leachate contains 120 g/1 of Fe, 8.0 g/1 of Nb, 300 mg/1 of Ta and 440 mg/1 of Ce.

The dissolution yields, estimated by 2 methods of calculation, are given below:

		S'FSvâ				îfSir·
Ljquid/(Uquld + Solid) distribution	79%	93%	85%	62%	87%	80%	50%

From these yields, the pyrochlore minerai phase carrying the niobium and tantalum has been attacked to approximately 85% by sulphation roasting. zxf

B — Direct acid roasting

Example 2; Acid roasting with ultrafine qrindlnq fwithout use of SO? in the dissolution)

Pyrochlore ore was ground in a vertical mill stirred with balls, so as to obtain a particle size corresponding to passing at 18 microns of 100% of the weight.

The analysis of the ore is given below:

.Ore·'^ 7-···	fe ·	Mn		.Λ' U’ ;.	Ce
Contents	29.9%	2.7%	0.15%	1130 ppm	0.65%

7 Nt/	’&Ta .	17TÇ7:.	.P	s
2.3%	900 ppm	2.1%	3.3%	0.55%

207.7 g of slurry of ground ore in water at a solid content of 45% (i.e., 94.2 g of dry ore présent) are introduced into a porcelain cruclble. 147.5 g of 96% sulphuric acid are added, Le. 1504 kg of H₂SO₄/tonne of ore. The attack solution thus comprises 41% of H₂SO₄ and the Fe/(Nb+Ta) and PO₄/(Nb+Ta) molar ratios in the charge in the roasting are respectively 21.4 mol/mol and 4.3 mol/mol. The mixture Is stirred mechanically in order to ensure homogenelty. The cruclble is placed in a kiln at 300°C and maintained for 4 hours. The roasted material is cooled. 180.1 g of calcine are obtained and theweight loss during the roasting is estlmated at 49%.

Ail the calcine is introduced into a stirred reactor filled with 189 ml of water heated to 90°C; the water/'initlal ore ratio is thus 2.0. This slurry is stirred and maintained at température for 2 hours. The slurry is centrifuged and the solid centrifuging deposit Is washed. The following results are then obtained

Amourïts’					^Ce'-		^;z'^?Ta.^z
16.6 g	Solid residue (%)	10.6%	0.08%	92 ppm	0.22%	2.4%	0.14%
154 ml	Roasting leachate (g/l)	135	0.95	0,53	2.4	8.3	0.31
545 ml	Aqueous wash liquor (g/D	9.9	0.09	0.04	0.24	0.25	0.020

Gorfténts .	·' τι·/	Mh
Solid residue (%)	5.3%	3.7%
Roasting leachate (g/l)	4.1	10.4
Aqueous wash liquor (g/l)	0.28	0.7

The dissolution yields are given below:

.... _ς.·„—n -hl. -..-χ J A. : ^^£^7 r ’L. ^J		J;			vca' ►Λ : - : : * L Ί	¹ -I¹ —— rw-			- Mn
Uquld/(Uquid + Solid) distribution	82%	94%	94%	99%	93%	78%	72%	47%	77%

Success is achieved in dissolvlng 8.3 g/l of Nb in a leachate, the free acidity of which is restricted, i.e. [H⁺] = 2.3N. 78% of the niobium and 72% of the tantalum are extracted in solution.

The titanium is partially extracted and the extracting yield for the manganèse is only 76%. dT

Example 3: Acid roasting with ultrafine arinding, with use of SO_?_ In the dissolution

Pyrochlore ore was ground in a vertical mlll stirred with balls, so as to obtain a particle size corresponding to passlng at 10 microns of 100% of the weight.

The analysis of the ore is given below:

				Wrf.il* i~Î
Contents	31.7%	2.8%	0.22%	1090 ppm	0.70%

Nb		rw	p-ntfjF
2.3%	900 ppm	2.1%	3.4%	0.63%

372.1 g of slurry of ground ore In water at a solid content of 31% (i.e., 115 g of dry ore présent) are introduced into a porcelain crucible. 180.9 g of 96% sulphuric acid are added, I.e. 1510 kg of HiSO^tonne of ore. The attack solution thus comprises 31% of H2SO4 and the Fe/(Nb+Ta) and POVtNb+Ta) molar ratios in the charge in the roasting are respectively 22.3 mol/mol and 4.3 mol/mol. The mixture is stirred mechanically in order to ensure homogenelty. The crucible is placed In a klin at 300°C and maintained for 4 hours. The roasted material Is cooled. 229.2 g of calcine are obtained and the weight loss during the roasting is estimated at 58%.

Ail the calcine is introduced into a stirred reactor filled with 230 ml of water heated to 90°C; the water/inltlal ore ratio is thus 2.0. This slurry Is stirred and maintained at a température for 2 hours under reduced bubbling of SO2. The oxidation/reduction potential is then 570 mV Ag/AgCI (at such a potential, the iron présent is predominantiy ferrie). The slurry is centrifuged and the solid centrifuging deposlt is washed. The following are then obtained:

			w		gR
20.9 g	Solid residue (%)	6.5%	0.03%	113 ppm	0.19%	1.6%	600 ppm
222 ml	Roasting leachate (g/l)	142	1.1	0.54	3.0	8.9	0.38
224 ml	Aqueous wash tlquor (3/1)	20	0.18	0.07	0.6	0.73	0.025

Gontentsl \|	iln ΐ	Mh
Solid residue (%)	2.7%	0.54%
Roasting leachate (g/l)	6.6	16.1
Aqueous wash liquor (g/l)	0.82	2.2

The dissolution ylelds are given below:

	^ffassloss-
Uquld/(Liquid + Solid) distribution	82%	96%	98%	98%	95%	87%	88%	74%	97%

Success is achieved in dissolving 8.9 g/l of Nb in a leachate, the free acidity of which Is restricted, i.e. [H⁺] = 2.2N.

From the dissolution yields, the pyrochlore minerai phase carrying the niobium and tantalum was attacked to approximately 90% by virtue of a high ferrie Iron concentration on taking the calcine up in water (142 g/l of Fe), combined with the fine grindlng before roasting and with the presence of Iron in the roasting.

In comparison with Example 2, the use of SO2 on taking up in water drastically improves the dissolution of the manganèse (97%) and the titanium (74%) and improves by approximately +10% the extraction of the niobium and tantalum.

Example 4: Acid roastlng without ultraflne grinding, with use of SO? In the dissolution

The analysis of the ore used Is given below:

iiglig
Contents	33.6%	3.8%	0.28%	1090 ppm	0.79%

3SK
2.8%	1100 ppm	2.5%	3.8%	0.86%

The particle slze of this ore corresponds to 100% passing at 315 microns.

159.9 g of slurry of ground ore in water at a solid content of 50% (i.e., 77.5 g of dry ore présent) are întroduced into a porcelain crucible. 120.8 g of 96% sulphuric acid are added, i.e. 1496 kg of HîSOVtonne of ore. The attack solution thus comprises 43% of H2SO4 and the Fe/(Nb+Ta) and PO₄/(Nb+Ta) molar ratios in the charge in the roasting are respectively 20 mol/mol and 4.1 mol/mol. The mixture Is stirred mechanlcally in order to ensure homogenelty. The crucible is placed in a kiln at 300°C and maintained for 4 hours. The roasted material Is cooled. 138 g of calcine are obtained and the weight loss during the roasting Is estimated at 51%.

Ail the calcine Is Introduced Into a stirred reactor filled with 182 mi of water heated to 90°C; the water/inltlal ore ratio is thus 2.3. This slurry is stirred and maintained at a température for 2 hours under reduced bubbling of SO₂. The oxidation/reduction potential is then 555 mV Ag/AgCI (at such a potential, the iron présent Is predominantly ferrie). The slurry Is centrifuged and the solid centrifuging deposit is washed. The following results are then obtained:

		«« ,-r mis						ââ5ÎS\|\|f
12.4 g	Solid residue (%)	9.3%	0.1%	288 ppm	0.17%	3.9%	600 ppm	5.4%
172 ml	Roastlng leachate (g/0	112	0.81	-	2.1	6.2	0.2	3.4
315 ml	Aqueous wash liquor (g/l)	5.6	0.04	-	0.2	0.05	0.0012	0.12

The dissolution yields are given below:

		HL·?
Uquid/(Uquld + Solid) distribution	84%	95%	93%	96%*	95%	69%	50%	48%	97%

* Extraction yields from the analysis ofthe soiids.

The leachate contalns 6.2 g/l of Nb, 0.2 g/l of Ta, 2.1 g/1 of Ce and 112 g/l of Fe and a free acidlty corresponding to [H⁺] = 3.0N.

It is found that the extraction yields for niobium, tantalum and titanium are significantly poorer than those obtained for Example 3, between -20% for Nb and -40% for Ta, due to the fact that the ore was not ultrafinely ground beforehand, as Is the case in Example 3.

Example 5; Add roastlng .with ultrafins grinding and teQÎuiiQn of the calcine In a less concentrated medium without use of SO?

The calcine used in this test is identical to that produced in Example 3.

AU the calcine Is Introduced into a stirred reactor fllled with

239 ml of water heated to 90°C, i.e. a water^'initial ore ratio = 4.2. The slurry is stirred and maintained at a température for 1 hour. The

oxidatlon/reduction potential is then 1200 mV Ag/AgCI. The slurry is filtered. The following results are then obtained:

Amount	Contentsi§7	w²			W '1 , . - •IB - -	Ti -
14.6 g	Solid residue (%)	11.4%	0.09%	0.31%	4.5%	0.21%	3.9%
135 ml	Filtrate (g/1)	72	0.56	1.53	2.9	0.16	2.5
380 ml	Aqueous wash llquor (g/1)	19	0.15	0.44	0.66	0.04	1.0

The dissolution yields are given below:


Uqukl/(Uquki + Solid) distribution	74%	91%	91%	89%	50%	55%	56%

The dissolution yields for the combined éléments of the matrix (including Fe), on using a water/ore ratio of 4.2, are comparable to those obtained during Example 3, on using a water/ore ratio of 2.2.

The dissolution of the niobium and tantalum is signiflcantly infiuenced by the water/lnltial ore ratio during the mixing of the calcined product with water. A dissolution yield for the niobium of 50% is actually obtained with a water/initial ore ratio of 4.2, whereas 87% of the niobium is dissolved with a water/initial ore ratio of 2.2, as explalned in Exampie 3. The use of a low waterAinitial ore ratio brings about a higher concentration of iron: 142 g/1 of Fe in Example 3 versus 72 g/1 of Fe in Example 5. Example 7, which is presented later, shows that the addition of synthetic iron drastically Improves the dissolution yields for niobium and tanta lum.X

C- Indirect acid roasting

Atmospheric leachings

Example_a: Ferrie cocurrentwise leaching in the presence of a make-up of redudng agent (Figure 2A1)

Pyrochlore ore was ground, so as to obtain a particle size correspondîng to passing at 106 microns of 100% by weight. The analysis of the ore is given below:

		Mn-	- ~ iir* Σ ' :-¾	-Tte“	’X⁵'		X'Sc_Lv·-
Contents	33.6%	3.8%	1160 ppm	0.27%	0.99%	0.36%	170 ppm

			; ΤΛ.-Τ'ίΖ:
2.6%	1030 ppm	2.4%	3.9%	0.9%

500 g of ground dry ore are slurried in a stirred reactor with 566 g of water heated to 90°C. 835 g of 98% sulphuric acid, i.e. 1600 kg of HîSCWtonne of ore, are added to the slurry. The Initial solids content Is thus 26%. Gaseous SO₂ is gradually added to the slurry, so as to maintain a potential of 580 mV Ag/AgCI in order to maximlze the dissolution of the manganèse.

The mixture is mechanically stirred for 24 h while maintainlng the température at 90°C.

The slurry is filtered and the filtration cake is washed on the filter with demineralized water. The following results are then obtained:

Amount	Contents	Fe	TTl i	U	«X	NB	. . Ta	Tl
34.8 g	Solid residue (%)	7.1%	0.5%	70 ppm	1.61%	6.3%	0.4%	6.6%
500 ml	Filtrate (g/l)	67	0.58	0.43	0.22	4.5	0.2	2.7
790 ml	Aqueous wash liquor(g/i)	32	0.29	0. 21	0.30	2.0	0.01	1.2

The dissolution yields are calculated with respect to the distribution between the liquids (filtrate and aqueous wash liquor) and the solid residue:

	oil		ifesS?	Ér‘U.'^		•S
EJqukl/(Uquid + Solid) distribution	75%	92%	52%	89%	36%	35%	21%	30%

The uranium is dissolved to more than 80%. Although the leaching is carried out under hlghly acidic conditions (i.e., an Initial solids content of 26% and an acld/ore ratio of 1600 kg/t), and, contrary to what Is obtained after an acid roasting, the dissolution yields for niobium, tantalum and the rare earth éléments do not exceed 35%. It could be shown, by characterizlng the leaching residue, that the pyrochlore phase Is not destroyed by leaching with sulphuric acid, even if more acid Is used.

Example β: Ferrie cocurrentwise leaching without make-up of reducing agent (Figure 2A1)

Pyrochlore ore was ground, so as to obtain a particle size corresponding to passing at 630 microns of 100% of the weight. The analysis of the ore is given below:

		MhF	Ll-		Ce . ·
Contents	31.3%	3.4%	1150 ppm	0.29%	0.73%

		T1f\ ;
2.8%	1248 ppm	2.1%	3.7%	0.75%

90.2 g of ground dry ore are slurried in a stirred reactor with

400 ml of water heated to 90°C. 94 g of 98% sulphuric acid, i.e. 1022 kg of H2S0₄/tonne of ore, are added to the slurry. The initial solid content is thus 15%. The mixture is mechanically stirred for 24 h while malntalnlng the température at 90°C.

The slurry is filtered and the filtration cake is washed on the filter with deminerallzed water. The following results are then obtained:

Amount^-		S'.Fe.	Th
38 g	Solid residue (%)	22.2%	-	1213 ppm	1.13%	6.3%	2440 ppm
255 ml	Filtrate (0/D	48.7	0.16	0.14	0.19	0.31	0.047	0.41
5400 ml	Aqueous wash liquor (g/l)	0.35	-	0.0023	-	<0.0005	< 0.0005	-

Dtesol^tton yteids si .'· ai i	Weight	‘i-1-	.'s. Th! • — Bk	u •I'	it	Nbi · âi	M	/¹¼
Llquld/(Uquid + Solid) distribution	58%	65%	16%*	54%	7%*	3.6%	14%	5.5%*

* Extraction yields from the analysis ofthe liquids.

The leaching Is carried out under acidic conditions (i.e., an initial solids content of 15% and an acid/ore ratio of 1022 kg/t) which are less acidic than those of the above test a (i.e., an initial solid content of 26% and an acid/ore ratio of 1600 kg/t). The extraction yields for the éléments of value are consequently limited: 3.6% of Nb, 14% of Ta and 7% of Ce. Only the uranium is quantitatively extracted, at 54%.

Example γ: Ferrous cocurrentwise leaching (Figure 2A2)

151 g of pyrochlore ore are ground, so as to obtain a particle size such that 100% of the grains pass through a 71 micron sieve.

The analysis of the ore is as follows:

		Al	W	Th		' Nb,-	'âiTa	/tl
Minerai composition	34.1%	4.2%	4.4%	0.26%	0.86%	2.6%	0.06%	2.3%

The dry ore Is inserted into a reactor containing 623 g of distilled water and 63.5 g of 96% sulphuric acid, an amount of add necessary to observe a ratio of 400 kg of add per tonne of dry ore. The slurry thus obtained has a solid content of 18%.

Gaseous SO2 is added, so as to bring the potential down to 640 mV Ag/AgCI in the first minutes of the test. The flow rate of reducing agent selected for this test is 3 Sl/h.

The mixture is mechanically stirred at 90°C for 6 hours, preferably using a turbine which makes possible good diffusion of the gas, for example of the Rushton type.

At the end of the test, the slurry is centrifuged and the atmospheric leachate is recovered. The centrifuging deposit is washed with a weight of aqueous wash liquor/weight of wet residue ratio of 10. After a second centrifuging, the aqueous wash liquor is separated and the soild Is dried In an oven at 105°C.

The following results are then obtained:


Final residue (% w/w)	32.1%	4.9%	0.40%	0.37%	1.2%	3.9%	0.05%	3.3%
Leachate (g/i)	30	2.7	9.7	0.094	0.23	0.014	<0.01	0.12
Aqueous wash liquor (g/l)	2.2	0.20	0.68	0.004	0.02	0.004	<0.01	0.003

The dissolution yields, calculated with regard to the Liquid/(Liquid + Solid) distribution, are as follows:

							ràitsB
Uq/(Uq + Sol) distribution yields	37%	25%	94%	13%	11%	0.3%	-	2.0%

The weight loss during this leaching Is 35%.

The ferrous cocurrentwise leaching makes it possible to reduce from 85 to 90% of the iron dissolved in the ferrie form into the ferrous form. The réduction of the ferrie Iron into ferrous iron results In a précipitation of the niobium and tantalum which had been dissolved In the ferrie medium (see Example a), as is shown by the low concentrations found In the filtrate, 14 mg/l of Nb and < 10 mg/l of Ta.

A residue more concentrated in niobium/tantalum/rare earth éléments than the ore Is then obtaîned, which residue can, for example, be treated by an acid roasting stage described in the preceding examples.

Example ¢: Ferrous and ferrie countercurrentwi.se leaching (Figure 2B2)

350 g of ore are introduced into a 2 I glass reactor containing

945 g of distllled water and 950 ml of aggressive ferrie leaching filtrate (with an H⁺ concentration of 3.8 mol/l) recycled after liquid/solid séparation.

In this way, a slurry with an Initial solid content of 13% containing 500 kg of sulphurlc acid per tonne of ore treated is obtained.

SO₂ is bubbled Into the slurry and dissolved by mechanical stirring with a Rushton turbine. After reaction at 90°C for 6 hours, the solid residue is separated from the atmospheric leachate by centrifuglng. This is thus a test on ferrous leaching of the type of that presented in Example γ.

On conclusion of the ferrous leaching, aweight loss of 31% is obtained, along with the following results:

Amounts'	Contents	iStFe	AJ	W·	nr	C®
241g	Final residue (%w/w)	26.5%	4.9%	0.4%	0.27%	0.76%	3.2%
1676 ml	Leachate (9/1)	26	14.8	2.4	1.5	4.3	0.23
3821 ml	Aqueous wash liquor (g/D	5.7	3.3	0.56	0.35	1.54	0.056

Dissolution ytelds _;	Weight loss	Ffe·;	rtll S	Mn	g Th	Ce ·· ξί ·	Nb
Liquld/(Llquld + Solid) distribution	31%	59%	49%	90%	18%	11%	1.6%

After centrifuging the ferrous leaching slurry, the filtrate Is recovered in order to be sent to the downstream process in order to extract therefrom the dissolved metals of value.

A portion of the residue (361 g, unwashed) is, for its part, inserted Into a second reactor containing 82 g of distilled water and 160 g of acid, so as to observe a solid content of 24% and an amount of acid of 1036 kg per tonne of dry ferrous leaching residue. The aggressive ferrie leachlng stage Is the only stage where pure concentrated acid Is introduced.

The combined mixture Is mechanlcally stirred for 6 hours and maintained at 90°C. The leaching takes place without SO2.

On conclusion of the aggressive ferrie leaching, a weight loss of 66% Is obtained, along with the foliowing results:

SI*				MÜ
49 g	Final residue (%w/w)	11.5%	6.2%	0.8%	0.48%	2.0%	5.1%
183 ml	Filtrate (g/l)	155	15	4	0.66	1.28	5.0
363 ml	Aqueous wash liquor (g/i)	16	1.37	0.54	0.06	0.16	0.31

	.Weight loss -	j.R·'.· Ÿ J·’	Ü	iüi	ÿ	u	\|FÎ^-7--fo «>*13,: j ’
Uquid/(Liquid + Solid) distribution	66%	77%	37%	59%	25%	15%	27%

The overall weight loss after these two leachings Is (1-31%) x (1-66%) = 77%, which is équivalent to the weight loss obtained after an aggressive ferrie cocurrentwise leaching.

After liquid/solid séparation, the filtrate Is recovered In order to be recycled to the ferrous leaching, whlle the residue (unwashed) is roasted.

It should be noted that these two successive leachings made it possible to increase the niobium content of the solid from 1.8%, initial content of the ore, to 3.2% in the ferrous leaching residue and up to 5.1% in the ferrie leaching residue feedlng the roasting klln.

Counterexamole 6: a. Indirect acid roasting of a concentrate without a source of Iron

b. Dissolution of the calcine without a source of iron

The éléments of value présent in the pyrochlore ore were concentrated by a cocurrentwise leaching as described In Example a.

After washing and drylng, 78 g of residue are obtained, the analysis of which is as follows:

	·' -(Fè·:,¹!®		^; Mn	Th	^.7	Nb		' i i⁵ ΊΊ
Composition of the leaching residue to be roasted (% w/w)	9.9%	4.1%	1.2%	0.31%	1.6%	4.9%	0.19%	5.3%

The cocurrentwise leaching stage has thus made it possible to Increase the niobium content from 2.4 to 5.0%.

The residue Is subsequently ground, in order to obtain a partlcle size such that 100% of grains pass through a 10 micron sieve, and is then roasted as a mixture with 121 g of 96% sulphuric acid and 210 g of distilled water, so as to observe an acid/materlal to be roasted ratio of 1500 kg/t and a solids content of 26% for the slurry particlpating In the roasting. The Fe/(Nb+Ta) molar ratio in the charge in the roasting is 3.3 mol/mol.

After roasting at 300°C for 4 hours, 153 g of calcine are obtained, which product Is introduced into a reactor with 155 g of distilled water at 90°C; the waterf'weight of residue In the charge to be roasted ratio Is thus 2. The combined mixture is maintained at 90°C and mechanlcally stirred for 2 hours. After centrifuging the slurry and washing the centrifuging deposit, the residue is dried In an oven at 105°C.

The following are then obtained:

		-/ai -	Mn		3 Th	.Ce -	NU .	T®	-TT
Final residue (% w/w)	1.7%	0.68%	0.16%	0.01%	0.12%	0.53%	6.2%	0.26%	4.6%
Filtrate (g/i)	26	15	2.4	0.25	1.5	4.3	0.23	0.012	8.1
Aqueous wash llquor (g/1)	5.7	3.3	0.56	0.060	0.35	1.5	0.056	0.012	0.35

i.e., the following extraction yields (calculated with regard to the solid):

	:W'/	Al			Th			• Ta?'	Tl
Yields calculated with regard to analyses of the solids	90%	90%	92%	92%	78%	81%	27%	21%	49%

A filtrate comprising only 26 g/l of Iron does not make it possible to efficiently dissolve the niobium and the tantalum (< 30%).

In contrast, the uranium and the cérium were well dissolved.

Example 7: a. Indirect acid roastlng of a concentrate without a source of Iron

b. Dissolution of the calcine with a source of Iron (synthetic salts)

After washing and drylng, 115 g of residue are obtained, the analysis of which is as follows

		i»		RW				W	. ΊΤ ν Λ
Composition of the leaching residue to be roasted (% w/w)	9.1%	4.5%	0.84%	0.09%	0.49%	1.9%	5.1%	0.17%	6.3%

Analogously to Counterexample 6, the residue Is subsequently ground, in order to obtain a particle size such that 100% of the grains pass through a 10 micron sieve, and Is then roasted as a mixture with 181 g of 96% acid and 321 g of distilled water, so as to observe an acid/material to be roasted ratio of 1510 kg/t and a solids content of 19% for the slurry particlpating in the roasting. The Fe/(Nb+Ta) molar ratio in the charge is 2.9 mol/mol.

After roasting at 300°C for 4 hours, 243 g of calcine are obtained, which product is Introduced Into a reactor with 267 g of distilled water at 90°C; the water/'welght of residue in the charge In the roasting ratio Is thus 2.3. 118 g of synthetic ferrie sulphate Fe₂(SO₄)3 are also added in order to observe a ferrie sulphate/leaching residue before roasting ratio = 1000 kg/t. The combined mixture Is maintained at 90°C and mechanically stirred for 2 hours. After centrifuging the slurry and washing the centrifuging deposlt, the residue is dried In an oven at 105°C.

The following results are then obtained:

				••i»'*·-.VWflkft©					mg
Final residue (% w/w)	2.9%	0.57%	0.07%	0.0045%	0.05%	0.30%	4.4%	0.18%	3.9%
Filtrate (Sfl)	87	12	2.3	0.24	1.2	4.3	9.S	0.35	12
Aqueous wash liquor (9/1)	12	1.5	0.31	0.03	0.18	0.84	0.57	24	1.1

i.e. the following extraction yields (calculated with regard to the distribution):

			e	Hlhg
Yields calculated with regard to the Uq/(Liq+Sol) distribution	96%	94%	96%	97%	95%	92%	58%	56%	67%

The acid roasting of a cocurrentwise leaching residue followed 10 by the dissolution of the calcine in a ferrie medium (iron having being added after the roasting) made it possible to double the extracting yields for niobium and tantalum with respect to a dissolution of the calcine in water (see Counterexample 6). The extraction yields for the uranium and cerlum were also improved (> 90%).

Example g; a. Indirect acid roasting of a concentrate with a source of iron (synthetic salts)

b. Dissolution of the calcine

A cocurrentwlse leachlng was carried out according to the protocol presented in Example a.

The residue obtained (158 g) was washed, dried and then ultrafinely ground, so that 100% of the grains pass through a 10 micron sieve. The chemical analysis of this residue is given below:

						Nbi-
Composition of the
leachlng residue to be roasted (% w/w)	10.1%	3.9%	1.2%	0.48%	1.7%	4.0%	0.11%	4.9%

This residue was subsequently mixed In a porcelain crucible with 248 g of sulphuric acid and 161 g of ferrie sulphate Fe₂(SO₄)3 crystals, in order to observe respective ratios of 1500 and 1000 kg per tonne of leaching residue treated. The Fe/(Nb+Ta) and PO₄/(Nb+Ta) molar ratios in the charge are respectively 16 mol/mol and 2.4 mol/mol.

After addition of 190 g of distilled water and homogenlzation of the combined mixture, the cruclble is inserted into the roasting kiln heated to 300°C for 4 hours.

The resulting calcine is then dissolved in 318 g of water heated to 90°C, for 2 hours, in the presence of SO₂.

The results obtained are as follows:

	Fe	' * AI .	Mrr-	STHÉ		Nb :	Æ- Ta	. r-TÎ·'—
Final residue (%w/w)	1.8%	0.35%	0.04%	0.04%	0.32%	3.3%	0.15%	1.8%
Filtrate (g/0	82	10	1.7	1.1	3.8	10	0.34	10
Aqueous wash liquor (g/D	7.6	0.92	0.17	0.12%	0.66	0.12	0.003	0.73

i.e., the following extraction yields (calculated with regard to the distribution):

	⁷ Fe .		Mrt\	*Th	Cé		.Ta.	7V^:
Yields calculated with regard to (Jq/(Liq+Soi) distribution	98%	97%	98%	97%	93%	77%	71%	87%

The acid roasting of a cocurrentwise leaching residue in the presence of a source of iron makes it possible to increase the extraction yields with respect to Example 7, where the source of ferrie sulphate Is Introduced during the dissolution of the calcine.

Example 9: a. Indirect acid roasting without recvcling and without grlndinq b. Dissolution of the calcine

A pyrochlore ore is treated by a cocurrentwise atmospheric leaching as described in Example a.

At the end of the leaching, the slurry Is centrifuged, so as to recover the filtrate (97 ml) and a centrifuging deposit comprising the solid residue (48.5 g) impregnated with a liquid having the same composition as the filtrate (213 ml). The solid content of the impregnated residue Is thus 12.4%. The total amount of Iron présent in the impregnating liquid Is then 529 kg per tonne of dry leaching residue (i.e., 1890 kg/t of ferrie sulphate). The Fe/(Nb+Ta) and PCWfNb+Ta) molar ratios In the charge are respectlvely 19 mol/mol and 5.5 mol/mol.

The composition of the leachate Is as follows:

			SBOl
Composition of the leadiing residue to be roasted (%w/w)	9.0%	4.7%	1.0%	0.08%	0.47%	2.0%	5.1%	0.20%	5.9%
Composition of the impregnating liquid of the said residue (g/1)	111	16	8.1	0.35	0.54	0.70	3.7	0.20	1.6

86g of 96% concentrated sulphuric acid and 21 g of distilled water are added to this impregnated residue in a porcelain cruclble, so as to obtain an acid/material to be roasted ratio of 1697 kg/t and a solid content of 10% for the slurry particlpating In the roasting.

After manual homogenization, the combined mixture is roasted at 300°C for 4 hours.

The 296 g of calcine obtained are dissolved In 110 g of water,

i.e. at a water/initial residue in the charge to be roasted ratio = 2.3, for 2 hours, in the presence of SO₂, at a constant température of 90°C.

After centrifuging the slurry and washing and then drylng the residue, 16.5 g of attack residue, i.e. a weight loss during the roasting at 66%, and 80 ml of filtrate are obtained. The compositions are as follows:

								W
Final residue (% w/w)	2.0%	0.36%	0.35%	0.01%	0.06%	0.34%	1.1%	0.05%	1.3%
Filtrate (g/0	139	21	19	0.53	1.5	4.7	14	0.52	14
Aqueous wash llquor (g/i)	12	1.7	1.5	0.037	0.11	0.35	0.35	0.024	0.78

The extraction yields, calculated with regard to the distribution, are as follows:


Yields calculated with regard to Uq/(Uq+Sol) distribution	99%	99%	99%	98%	97%	95%	95%	94%	94%

The iron concentration of the roasting leachate Is particularly high (139 g/1) and makes it possible to obtain excellent dissolution of the niobium, tantalum, uranium and cérium.

The two filtrâtes thus obtained, that is to say the atmospheric leaching flltrate and the filtrate from the dissolution of the calcine, are then mixed, so as to obtain a single stream, which is intended to be treated downstream in order to recover the niobium, tantalum, uranium and rare earth éléments.

The composition of the mixture of the two leachates is as follows:

							jsSîSWs.t.·	'Sw'
Mixture of filtrâtes (9/0	136	16	13	0.82	0.95	2.5	7.6	0.33	6.7

The indirect roasting makes it possible to achleve overall extraction yields (leaching and roasting) of 96% for the niobium, 94% for the tantalum, 99% for the uranium and 96% for the cérium.

Exemple 10: a. Indirect acid roasting without recycling and without grindino

b. Dissolution of the calcine

A pyrochlore ore is treated by a ferrous cocurrentwise atmospheric leaching as described ln Example γ.

At the end of the leaching, the slurry is centrifuged, so as to recover the filtrate (425 ml). The centrlfuging deposit is washed by reslurrying before being centrifuged, so as to obtain an aqueous wash liquor and a wet residue. The wet residue Is dried in an oven at 105°C.

The composition of the leachate Is as follows:

	Fe	·' at^::;	-i···; •MO
Composition ofthe atmospheric leachate (g/1)	30	2.7	9.7	0.09	0.23	0.014	0.010	0.12	0.6N

The composition of the residue which feeds the roasting is as follows:xC so

			W					h-?'.U
Composition ofthe leaching residue to be roasted (% w/w)	32%	4.9%	0.40%	0.37%	1.2%	3.4%	0.05%	3.32%

The dry residue (71 g) is roasted, so as to obtain a calcine; the Fe/(Nb+Ta) molar ratio in the charge is 13 mol/mol. The calcine is subsequently taken up in water, the protocol of Example 3 being followed.

After centrifuging the slurry, the following results are obtained:

			Ayr-tfrUtr*·- Μπ<·			.lW¹*·		SO
Final residue (% w/w)	22%	0.66%	0.42%	0.17%	0.51%	5.7%	0.16%	5.5%	-
Leachate (a/1)	115	20	1.5	1.4	4.1	12.5	0.30	9.3	3.7N
Aqueous wash liquor (g/D	3.6	0.65	0.054	0.044	0.17	0.01	0.01	0.14	-

The extraction yields subséquent to the roasting stage, calculated with regard to the distribution, are shown in the following table: 15

		• Al...			”.€6_τ.	--.nW-		Tl·
Yields calculated with regard to Uq/(Uq+Sol) distribution	82%	96%	75%	87%	88%	64%	62%	59%

.¹

The extraction yields are lower than those presented in Example 9 as the Fe concentration at the dissolution step Is lower (115 g/l of Fe Instead of 139 g/l of Fe) and the Fe/(Nb+Ta) ratio during the roasting Is lower. Furthermore, the leaching carried out on the ore Is less aggnesslve than that presented In Example 9, with the resuit that the matrix is not attacked so well, which does not facilitate the roasting stage.

The ferrous leachate originating from the atmospheric leaching (comprising essentially U and RE) can optionally be mixed with the roasting leachate (comprising Nb, Ta, U and RE) diluted two fold with water, so as to obtain a single stream Intended to be treated in the downstream process. It should be noted that, during the dilution, précipitation of the éléments of value 1s not observed; in particular, précipitation of niobium or tantalum Is not observed.

The composition of the mixture Is then as follows:


Mixture of filtrâtes (g/i)	52	7.2	7.5	0.42	1.3	3.3	0.090	2.5	1.4N

Examole 11: a. Indirect acid roasting without recvcllng and with qrlndlno b. Dissolution of the calcine

A pyrochlore ore is treated by a ferrous cocurrentwlse atmospheric leaching as described In Example γ.

At the end of the leaching, the slurry Is centrifuged, so as to recover the filtrate (2745 ml). The centrlfuglng deposlt is washed by reslurrying, before being recentrifuged, so as to obtain an aqueous wash llquor and a wet residue. The wet residue is dried in an oven at 105°C.

The residue is subsequently ultrafinely ground, so as to obtain a particle size such that 100% of the grains pass through a 10 micron sïeve.

The analysis of the leaching residue before roasting corresponds to:

	n
Composition of the leaching residue to be roasted (% w/w)	31%	5.3%	0.41%	0.29%	1.2%	3.3%	0.07%	0.31%

The dry residue Is roasted, so as to obtain a calcine which Is taken up in water, the protocol of Example 3 being followed. The Fe/(Nb+Ta) molar ratio In the charge in the roasting is 15 mol/nrtol.

In this way, the following results are obtained

		SS*-	.rJpi MhisS		iw	ΐ w
Final residue (% w/w)	8.9%	0.97%	0.25%	0.10%	0.47%	4.0%	0.12%	0,39%
Filtrate (9/1)	123	23	1.5	0.33	3.5	12	0.33	10.2
Aqueous wash liquor (9/1)	3.5	0.71	0.05	0.06	0.63	0.02	4.8	0.16

The extraction yields subséquent to the roasting stage, calculated with regard to the distribution, are shown In the following table:

		: AI	ii»·*- TMrr	' Th;.'.;	r'CeJ;	Nb	Ta .<	;T\|·
Yields calculated with regard to Uq/(Uq+Sol) distribution	95%	97%	89%	86%	93%	79%	78%	97%

The ultrafîne grinding stage has made It possible to increase the extraction yields for the niobium and tantalum by 15% with respect to Example 10.

The ferrous leachate originatlng from the atmospheric leaching can optionally be mixed with the roasting leachate diluted by a factor of 1.8, so as to obtain a single stream intended to be treated In the downstream process.

The composition of the mixture is then as follows:

	Fe.	fl Al	Mn	Th	1	Nb	Ta	--..τι
Mixture of filtrâtes (g/D	43	4.7	9.2	0.11	0.54	1.1	0.03	1.1

Example 12: a. Indirect acid roasting

b. Dissolution of the calcine

After a ferrous and ferrie countercurrentwise leaching as described In Example δ, 219 g of an unwashed residue, containing 62 g of dry residue and 157 g of aggressive ferrie leaching filtrate (i.e., 330 kg of ferrie sulphate per tonne of dry leaching residue), are obtained, which product Is inserted into a porcelain crucible and then mixed with 85 g of concentrated sulphuric acid (ratio of 1370 kg of acid per tonne of dry residue). The combined mixture is roasted at 300°C for 4 hours in order to obtain a calcine, which Is dissolved under the conditions described In

Example 8. The Fe/(Nb+Ta) molar ratio in the charge In the roastlng Is mol/mol.

		W	*Mn	Th ·	- · ·. . >H ! · . eu	- Nb A '	•'.W*	Tl
Composition of the leaching residue to be roasted (% w/w)	11.3%	6.6%	0.56%	0.50%	2.0%	6.4%	0.17%	5.8%
Composition of the imprégna ting liquid (g/l)	59	4.0	1.6	0.17	0.36	1.7	0.061	0.86

After centrifuging the slurry obtained on conclusion of the hours of dissolution of the calcine, the ferrie filtrate comprising the majority of the éléments of value can be treated In order to separate the 10 niobium and tantalum, before being recycled at the start of the countercurrentwise leaching.

The residue is washed by reslurrying at ambient température for 20 minutes. The extraction yields obtained are as follows:

Amounte	Contenir	Fe	/ ÀÎ'i	Th.		Nb	Ta- ,
28 g	Residue after roasting (%w/w)	5.2%	0.76%	0.10%	0.38%	4.2%	0.14%
117 ml	Filtrate (g/l)	118	42	2.5	11	30	0.61
138 ml	Aqueous wash liquor (g/l)	15	5.9	0.35	1.7	3.2	0.047

	ÿtelght^		au- .... ; ;	gfjgïSg		Nb t:
UquW/(Uquld + Solid) distribution	55%	92%	96%	92%	93%	77%	66%

The only liquid exit from this example takes place at the end of the ferrous leaching with a 90% ferrous stream which comprises virtually 5 ail the uranium and rare earth éléments.^

Claims

1. A process for the dissolution of at least one element chosen from niobium and tantalum and of at least one other element chosen from uranium and rare earth éléments, advantageously for the dissolution of niobium, tantalum, uranium and rare earth éléments, présent in an ore or an ore concentrate, characterized In that it comprises:

- the roasting of a material, comprising said éléments, which material Is mixed, dry or in the presence of water, with an acidic roasting agent In order to obtaln a calcine; said material consisting of sald ore or concentrate or having been obtained from said ore or said concentrate and sald acidic roasting agent provlding for roasting in a sulphate medium; and

- the dissolution in an aqueous solution of the calcine obtained in order to obtain a slurry, the liquid fraction of which includes Iran, in the ferrie state, at a concentration of at least 50 g/l, advantageously of at least 70 g/l and very advantageously of at least 120 g/l.

2. The process according to claim 1, characterized in that the Iron présent In sald slurry was présent in said calcine and/or was added, at least in part, in the liquid and/or solid form, to sald calcine.

3. The process according to claim 1 or 2, characterized In that the dissolution of the calcine Is carried out without any addition of sulphuric acid.

4. The process according to any one of claims 1 to 3, characterized In that the dissolution of the calcine is carried out with addition of a reducing agent, such as SO₂.

5. The process according to any one of claims 1 to 4, characterized in that the roasting, a solid/liquid reaction, is an acid roasting.

6. The process according to any one of claims 1 to 4, characterized in that the roasting, a gas/solid reaction, is a sulphation roasting.

7. The process according to any one of claims 1 to 6, characterized that said roasting is carried out In the presence of Iron, in the ferrie form, advantageously présent at an Fe/(Nb+Ta) molar ratio of greater than 2, very advantageously of greater than 3.5 and preferably of/ p .

greater than 6; said Iron being already présent in the material to be roasted and/or having been added, at least in part, in the liquid and/or solid form, to said material to be roasted.

8. The process according to any one of daims 1 to 7, characterized In that the roasting is carried out in the presence of phosphate, advantageously présent at a PO₄/(Nb + Ta) molar ratio of greater than 2 and very advantageously of greater than 6; said phosphate being already présent in the material to be roasted and/or having been added, at least in part, In the liquid and/or solid form, to said material to be roasted.

9. The process according to any one of daims 1 to 8, characterized in that it additionally comprises, upstream of said roastlng:

- a physical enrichment and/or a chemical treatment of the ore or concentrate.

10. The process according to any one of daims 1 to 9, characterized in that it additionally comprises, upstream of said roasting:

- an acid leaching, advantageously a sulphuric acid atmospheric leaching; said leaching being carried out on an ore or concentrate, optionally physically enriched.

11. The process according to claim 10, characterized in that it additionally comprises, upstream of the roastlng:

A) a leaching, in a single stage, cocurrentwise, in the presence of sulphuric acid, followed by a solid/liquid séparation which produces a leachate and a leaching residue to be roasted; or

B) a leaching, in two stages, countercurrentwlse, comprising a first leaching stage, followed by a solid/liquid séparation which produces a leachate and a first leaching residue, and a second leaching stage carried out on said first leaching residue, in the presence of sulphuric acid, followed by a solid/liquid séparation which produces a second leaching residue to be roasted and a liquid to be advantageously recyded to said first leaching stage.

12. The process according to daim 10 or 11, characterized in that It additionally comprises, upstream of the roasting:

Al) a leaching, in a single stage, cocurrentwise, at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mV Ag/AgCI, in the presence of H2SO4 and optionally of a make-up < >

of at least one reducing agent, followed by a solid/liquid séparation which produces a ferrie leachate and a leaching residue to be roasted;

A2) a leaching, in a single stage, cocurrentwise, at a potential of less than 500 mV Ag/AgCI and advantageously of less than or equal to 450 mV Ag/AgCI, In the presence of H2SO4 and of at least one reducing agent, followed by a solid/liquid séparation which produces a ferrous leachate and a leaching residue enriched In niobium and/or tantalum to be roasted;

Bl) a leaching, in two stages, countercurrentwise, comprising a first leaching stage, at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mV Ag/AgCI, in the presence of H2SO4 and optionally of a make-up of at least one reducing agent, followed by a solid/liquid séparation which produces a ferrie leachate and a first leaching residue, and a second leaching stage carrîed out on said first leaching residue, either at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mV Ag/AgCI, in the presence of H2SO4 and optionally of a make-up of at least one reducing agent, or at a potential of less than 500 mV Ag/AgCI and advantageously of less than or equal to 450 mV Ag/AgCI, In the presence of H2SO4 and of at least one reducing agent, followed by a solid/liquid séparation which produces a second leaching residue to be roasted and a liquid with resldual acidlty to be advantageously recycled to said first leaching stage;

B2) a leaching, in two stages, countercurrentwise, comprising a first leaching stage, at a potential of less than 500 mV Ag/AgCI and advantageously of less than 450 mV Ag/AgCI, in the presence of H2SO4 and of at least one reducing agent, followed by a solid/liquid séparation which produces a ferrous leachate and a first leaching residue, and a second leaching stage carrîed out on said first leaching residue, either at a potential of greater than 500 mV Ag/AgCI and advantageously of greater than or equal to 600 mV Ag/AgCI, in the presence of H2SO4 and optionally of a make-up of at least one reducing agent, or at a potential of less than 500 mV Ag/AgCI and advantageously of less than or equal to 450 mV Ag/AgCI, In the presence of H2SO4 and of at least one reducing agent, followed by a solid/liquid séparation which produces a second leaching residue enriched in niobium and/or tantalum to be roasted and a liquid togj' * F be advantageously recycled to said first leaching stage; said advantageously recycled liquid being recycled as it is or after séparation, at least partial, from the niobium and/or tantalum présent therein.

13. The process according to claim 12, characterized In that said leaching is of A2) or B2) type, advantageously of B2) type.

14. The process according to any one of claims 1 to 13, characterized in that it additionally comprises, upstream of said roasting:

- an uitrafîne grinding of the material to be roasted, for the production of said material at a particle size < 30 pm, advantageously < 15 pm and very advantageously < 10 pm.

15. The process according to any one of claims 1 to 14, characterized In that it additionally comprises, downstream of said dissolution:

- a solid/iiquid séparation carried out dlrectly on the slurry or after an additional stage of dilution of said slurry by a limited factor, advantageously < 5, in order to prevent the précipitation of the dissolved éléments; said solid/liquid séparation producing a solid attack residue and a roasting leachate Including said dissolved éléments.

16. The process according to claim 15, characterized In that it additionally comprises, downstream of said dissolution:

-the séparation of the niobium and/or tantalum from said roasting leachate and advantageously the recycling of said leachate, from which said niobium and/or tantalum was/were separated, optionally diluted, for the implémentation of a sulphuric acid atmospheric leaching according to any one of claims 10 to 13.

17. The process according to any one of claims 11 to 16; characterized in that it comprises:

- a leaching of Al, A2, B1 or B2 type according to claim 12 and the recovery of a leachate and of a leaching residue;

- an optional uitrafîne grinding of said leaching residue;

- the roasting of said optionally ground leaching residue;

- the dissolution of the calcine obtaîned on conclusion of said roasting, in order to obtain a slurry;

- a llquld/solld séparation, carried out directly on said slurry or on said slurry diluted by a limited factor, which produces a solid attack residue and a roasting leachate including the desired éléments;

4 ♦* *

-the séparation of the niobium and/or tantalum from said roasting leachate and the recycling of the leachate, from which said niobium and/or tantalum was/were previously separated, which Is optionally dlluted, for the Implémentation of said leaching.

18. The process according to any one of claims 11 to 16, characterized in that it comprises:

- a deslimlng of an ore or ore concentrate, which is optionally physlcally enriched, for the recovery, on the one hand, of sûmes including iron and, on the other hand, of the deslimed ore or concentrate;

- a leaching of Al, A2, BI or B2 type according to claim 12 carried out on said deslimed ore or concentrate, optionally with addition of a portion of the sûmes recovered in step 1 above, and the recovery of a leachate and of a leaching residue;

- an optlonal ultrafine grindlng of the leaching residue;

- a liquid/solld séparation, carried directly on said slurry or on said slurry diluted by a limited factor, which produces an attack residue and a roasting leachate Including the desired éléments;

-the séparation of the niobium and/or tantalum from said roasting leachate and the recycling of said roasting leachate, from which said niobium and tantalum was/were previously separated, which Is optionally dlluted, for the implémentation of said leaching.

19. The process according to any one of claims 1 to 18, characterized in that It Is carried out with an ore or ore concentrate; the ore concerned being chosen from the minerais of the pyrochlore, euxenite, samarskite, perovskite and fergusonite groups and their mixtures.

20. A process for the recovery of at least one element chosen from niobium and tantalum and of at least one other element chosen from uranium and the rare earth éléments, advantageously for the recovery of niobium, tantalum, uranium and rare earth éléments, présent In an ore or an ore concentrate, characterized in that it comprises: rtF < ·

- the dissolution of said éléments according to the process of any one of claims 1 to 19; and

- the séparation of said éléments.^

61 pages