Such metals include Titanium (Ti), Zirconium (Zr), Hafnium (Hf), Vanadium (V), Niobium (Nb) , Tantalum (Ta) , Chromium (Cr), Molybdenum (Mo), Tungsten ( ) , Manganese (Mn), Cobalt (Co), Nickel (Ni), Copper (Cu) , Zinc (Zn), Cadmium (Cd) , Magnesium (Mg), the Rare Earths (Sc, Y and La to Yb), Thorium (Th), Uranium (U) and Lead (Pb) .

Applications of these oxides generally requires their isolation in a reasonably pure form from their ores, before further processing. There are a wide variety of pyrometallurgical and hydrometallurgical processes for extraction of metal oxides. The details of any extraction process depend very much on the particular element, the composition and physical properties of the ore body and the availability of local resources among other factors.

Tungsten, for example, principally occurs in nature as the minerals Scheelite and Wolframite. Industrial processes known in the prior art for the extraction of tungsten oxide from its ores include: (a) roasting the concentrate and digesting it in sodium carbonate under high pressures and temperatures

(>1.2 Mpa, 200°C), to afford sodium tungstate

(Na-WO.). Addition of Al and Mg sulphates removes impurities, then the Na-WO. is converted into the ammonium form by ion-exchange.

(b) acid leaching with hydrochloric acid to produce tungstic acid (H ₂ WO.) and digestion in aqueous ammonia to give an ammonium tungstate solution. Addition of MgO and activated carbon removes

impurities, which are filtered off.

In either case the ammonium tungstate can be calcined to produce pure WO_.

Niobium and Tangalum usually occur together in the minerals pyrochlore, columbite, tantalite and in tin refinery slags. Processing of the Nb/Ta concentrates to the oxide forms include:

(a) Nb concentrates are usually converted directly to the metallic form or to a carbide by a variety of reduction methods (e.g. aluminothermic, carbo- thermic) . The products may then be burnt to the oxide form.

(b) Direct attach of the concentrate with hot 70-80% hydrofluoric acid and then solvent extraction. (c) Dissolution in cone, sulphuric acid and 300— 00 ^β C and subsequent hydrolysis of the sulphato complex to produce niobic acid.

(d) Fusion with NaOH at 500-800°C in an iron crucible, subsequent leaching with water to leave an insoluble niobic acid.

(e) Reaction of a mixture of the ore and carbon with chlorine gas at 500-1000°C to yield volatile NbCl-. which is fractionated from other volatile chlorides. Hydrolysis affords niobic acid. It can be seen that from these two cases that such processes generally require complex and expensive processing plant and equipment to withstand the reaction conditions encountered and the number of necessary steps. Moreover, considerable waste material is generated and needs to be disposed of.

Accordingly, it is an object of the present invention to overcome, or at least alleviate, one or more of the difficulties or deficiencies related to the prior art. Accordingly, in a first aspect, there is provided a process for producing a metal oxide product, which process includes providing a metal-containing material; and

a leaching composition including a source of ammonium ions; a source of carbonate ions; and water; contacting the metal-containing material with the leaching composition to form an aqueous slurry for a time sufficient such that a metal complex is produced; subjecting the metal-containing material to a size reduction step prior to, or simultaneously with, the leaching step such that the average diameter of the particles is below approximately 100 micron; and isolating a leaching solution containing the metal complex so formed.

The. present invention may provide a general method for the extraction of various metal oxides from their ores whilst reducing or eliminating the complexities attendant to prior art processes.

Accordingly, it will be understood that the metal oxides within the metal-containing material react with aqueous solutions of a leaching agent derived from ammonium carbonate, bicarbonate, carbamate or related compounds.

The metal-containing material may be a mineral or mineral ore, concentrate or a processing waste or by-product, such as a slag or spent shale. The metal- containing material may be a metal compound, complex or other chemical combination, and be present in the material in a chemical combination such as an oxide, a silicate, sulphide, other chalcogenide, phosphate, carbonate, sulphate, borate, a mixed metal oxide or combinations of these.

The metal-containing material may include one or more of metals selected from Groups 4, 5 and 6 of the Periodic Table, as well as other electropositive and oxophilic metals. The metal-containing material may include one or more of the following metals: Vanadium (V), Niobium (nb) , Tantalum (Ta) , Chromium (Cr) , Molybdenum (Mo), Tungsten (W), and the Rare Earths (Sc, Y and La to Yb) . The metal-containing material may additionally

include metals selected from Titanium (Ti), Zirconium (Zr), Hafnium (Hf) , Manganese (Mn) , Cobalt (Co), Nickel (Ni), Copper (Cu) , Zinc (Zn), Cadmium (Cd) , Magnesium (Mg), Thorium (Th) , Uranium (U) and Lead (Pb) . The metal-containing material may contain varying amounts of impurities, either in solid solution or as a discrete amorphous or crystalline phase.

The material may be converted into another phase, more reactive than the original phase, before leaching. Accordingly, the process may include a preliminary step of providing a metal-containing material, and an alkaline earth metal salt, contacting the metal-containing material with the metal salt at elevated temperatures to form an alkaline earth metallate.

Preferably the preliminary reaction is conducted at a temperature of approximately 500°C to 1600°C, preferably approximately 750 ^β C to 1200°C. The preliminary reaction may continue for from approximately 1 to 24 hours.

The heating step may be performed in air, in an inert gas or under a reducing atmosphere.

The alkaline earth metal salt may be a carbonate, hydroxycarbonate, oxide or hydroxide. The alkaline earth metal salt may be a salt of barium or calcium, barium is preferred.

For example, where barium hydroxide is used, for example with iron chromite, the reaction sequence may be idealised by the following chemical equation 2FeCr ₂ 0 ₄ + 4Ba(OH) ₂ + 3.50 ₂ - 4BaCr0 ₄ + Fe ₂ Og + 4H ₂ 0

In another example, impure v ^" ₂ ° ₅ may be converted into a reactive form by reaction with barium carbonate in a reaction typified by the following idealised equation: γ ₂ 0 ₅ + BaCC - BaV ₂ O _g + C0 ₂ The particle size of the metal-containing material is selected to improve the efficiency of the process according to the present invention. Relatively small particles, generally below approximately 100 microns in average diameter and preferably below approximately 50

microns have been found to be suitable. Thus, as stated above, the process includes the step of subjecting the metal-containing material to a size reduction step.

The size reduction step may include a crushing and/or grinding step. The size reduction step may include a milling process. The milling process may continue over an extended period. The milling process may continue for from approximately 4 to 48 hours, preferably approximately 5 to 36 hours, more preferably from approximately 12 to 24 hours. Attrition milling is preferred.

The size reduction step may be conducted prior to, or simultaneously with, the leaching step. The size reduction may be preferably achieved in situ by performing attrition milling and leaching simultaneously. The source of ammonium ions may be selected from ammonium compounds such as ammonium carbonate, ammonium bicarbonate, ammonium carbamate, or free ammonia or mixtures thereof.

The source of carbonate ions may similarly be ammonium compounds such as ammonium carbonate, ammonium bicarbonate, ammonium carbamate, or free carbon dioxide or mixtures thereof.

The source of ammonium ions and/or the source of carbonate ions may be provided as pure solids or as mixtures, or may be generated in situ. For example, reactions may be generated with appropriate proportions of ammonia and carbon dioxide with water.

The source of ammonium ions and the source of carbonate ions may be present in any suitable relative amounts. The ratio of contained NH ₃ to C0 ₂ may be in the range of approximately 100:1 to 1:100 by weight.

The source of ammonium ions and the source of carbonate ions may be present in the aqueous slurry in any suitable amounts. The components of the leaching composition may be present in the aqueous slurry such that the concentration thereof is sufficiently high to insure rapid reaction with the metal-containing material and/or to provide at least sufficient stability of the dissolved metal complex. The ratio of the leaching solution to the

metal-containing compound may be such that the atomic ratio of C0 ₂ to metal is at least approximately 1.

The process according to the present invention may be conducted at any suitable temperatures. Temperatures in the range of 0°C to approximately 200°C, preferably 30°C to 60°C, may be used. The temperature should be preferably high enough to ensure rapid reaction with the metal-containing material, but low enough to ensure that decomposition of the dissolved metal complex is minimised. The process may be conducted in any suitable conventional manner, including utilising batch processing in opened or closed vessels including autoclaves or continuous processing. The process according to the present invention may be conducted under pressures ranging from approximately atmospheric to approximately 100 MPa above atmospheric.

The process according to the present invention may continue for a time sufficient for a metal-containing complex to be formed. The leaching composition utilised in the process according to the present invention may optionally include secondary components. Secondary components which alter the thermal stability or solubility of the metal-containing complex in the aqueous solution and/or preferentially separate any metal phase formed from impurity phase or phases, may be used.

Secondary components which may be included in the leaching solution may be selected from the free acids, salts or esters or phosphoric, phosphorous, sulphuric, sulphurous, citric, oxalic, benzoic acetic or higher carboxylic, alkyl- or arylphosphonic or -sulphonic acids or the halides (F ^~ , Cl ^" , Br ^" , I~) , pseudohalides (CN ^~ , OCN ^~ , SCN ^" ) or combinations thereof. The secondary components may be added as pure solids or liquids, as diluted solutions or generated in situ, such as by reaction of free acids with NH_ or the major leachant components. The secondary components may be present in the leaching composition in amounts of 0 to approximately 95% by weight.

The leaching solution containing the metal-complex may be separated from the aqueous slurry in any suitable manner. The metal-complex may be separated from any unreacted material or from insoluble reaction products, by filtration, flotation, centrifugation or by sedimentation. The residue may be recycled and reacted further with fresh leaching composition with the same or a different composition under the same or different reaction conditions as used in the original leaching reaction. Accordingly in a preferred aspect of the present invention the process may further include separating the metal leach product from the leaching solution and recycling the residue of the separation step to the leaching step. In a further preferred aspect the process may further include contacting the residue with an alkaline earth metal salt at elevated temperatures to form an alkaline earth metal product; and recycling the alkaline earth metal product to the leaching step.

For example the residue may be contacted with barium carbonate at a temperature in the range of approximately 500°C to 1600°C. In a preferred aspect of the present invention, the process of producing a metal product may further include subjecting the metal complex to a decomposition step.

A metal-containing product may be isolated from the leaching solution by any suitable means, such as by thermal decomposition in solution or of a metal species isolated from the solution, or by removal of the unreacted leachant components. Decomposition of the metal-containing species may result in formation of free gaseous H ₃ , C0 ₂ or other components of the original leaching solution; these may be returned to the reaction process. After removal of some or all of the soluble metal-containing species, the leaching solution may be recycled for use with a fresh charge of metal-containing material.

In the following description the invention will be more fully described with reference to the accompanying example. It should be understood, however, that the following description is illustrative only and should not be taken in any way as a restriction on the generality of the invention described above.

EXAWPfrE

Extraction of Nb-O _r from BaNb-O

BaNb ₂ 0 _fi (10g) was attrition milled in an aqueous solution of commercial "ammonium carbonate" (80g in 200 ml water) at 18°C for 24 hours in a reaction system open to the air with partially stabilized zirconia (PSZ) balls (1 mm dia., 900g) . A light pink suspension was formed. This suspension was subsequently centrifuged and the supernatant filtered to remove any remaining solids.

This filtrate was evaporated to dryness and the off white solid formed heated to 800°C for 4 hours. Nb-O _c was obtained as a fine light yellow powder. The material was identified as Nb ₂ 0 ₅ by powder X-ray diffraction and by ΞDAX elemental analysis. The leaching process was repeated twice more with the residues from centrifugation. Overall

63% of the niobium (4.0 g) was extracted.

EXAMPLE 2 Extraction of Mo0 ₃ from MoS ₂ ^MoS 2 (209) ^was attrition milled in an aqueous solution of commercial "ammonium carbonate" (80 g in 200 ml water) at 22°C for 16 hours in a reaction system open to the air with PSZ balls (1 mm dia., 900 g) . The resulting black suspension was centrifuged and the very pale blue supernatant liquid filtered. This solution was evaporated to dryness yielding a blue-black solid. Heating the solid to 500°C for 4 hours gave o0 ₃ as a pale yellow solid. The overall yield of Mo extracted was 2.5% (0.4 g) .

EXAMPLE 3 Extraction of Cr ₂ Q ₃ from BaCrO,

BaCroO. (20g) was attrition milled in an aqueous solution of commercial "ammonium carbonate" (80 g in 200 ml water) at 18°C for 24 hours in a reaction system open to the air with PSZ balls (1 mm dia., 900 g) . The resulting

yellow suspension was centrifuged and the yellow supernatant liquid filtered. The solution was slowly allowed to evaporate to dryness at room temperature to afford a black solid, which was subsequently heated to 600°C for 1 hour giving Cr ₂ 0 ₃ as a dark green solid (characterised by powder XRD) . The overall yield or Cr extracted was 3%.

In a separate experiment the yellow filtrate was added to a mixture of methanol/2-propanone (1/1, 50 ml) to afford (NH ₄ ) ₂ Cr ₂ 0~ as an orange solid, characterised by powder XRD.

EXAMPLE 4 Extraction of V ₂ Q ₅ from Mσ(VQ ₃ ) ₂ Mg(V0 ₃ ) ₂ (20g) was attrition milled in an aqueous solution of commercial "ammonium carbonate" (80 g in 200 ml water) at 18°C for 24 hours in a reaction system open to the air with PSZ balls (1 mm dia., 900g) . The resulting suspension was centrifuged and the very pale yellow supernatant liquid filtered. This solution was slowly evaporated to dryness yielding NH.V0 ₃ (2.6g) as a pale yellow solid, characterised by powder XRD. Upon heating to 700°C the material was converted to orange V ₂ O _e . The overall yield of V extracted was 16%.

SSAMP g 5 Extraction of Mo0 ₃ from BaMoQ ₄

BaMoC (10g) was attrition milled in an aqueous solution of commercial "ammonium carbonate" (80g in 200 ml water) at 18°C for 24 hours in a reaction system open to the air with PSZ balls (1 mm dia., 900g) . The resulting suspension was centrifuged and the pale green supernatant liquid filtered. This supernatant was slowly evaporated to dryness yielding (NH ₄ ) _g Mo_0 ₂₄ (4.4g) as a very pale green solid, characterised by powder XRD. Upon heating to 660°C (1 h) the material was converted to yellow o0 ₃ (3.6g, 74% extraction), characterised by powder XRD.

EXAMPLE 6 Extraction of WO- from CaWO, Scheelite, CaWO., (20g) was attrition milled in

an aqueous solution of commercial "ammonium carbonate" (80g in 150 ml water) at 200°C for 20 hours in a reaction system open to the air with PSZ ballw (1 mm dia., 900g) . The resulting suspension was centrifuged and the supernatant liquid evaporated to dryness. Upon heating to 700°C (1 h) the material was converted to W0 ₃ . The leaching process was repeated 3 times with the residues of centrifugation to afford a total of 14.4g W0 ₃ (91% W extraction), characterised by powder XRD. EXAMPLE 6A (Comparative)

In a separate experiment CaW0 ₄ (20g) was stirred in an aqueous solution of commercial "ammonium carbonate" (80g in 150 ml water) at 25°C for 22 hours in a closed reaction system in the absence of any milling media. The resulting suspension was centrifuged, and the supernatant was filtered and evaporated to dryness. The solid residue obtained was calcined in air (700°C, 1 hour) to afford a total of 0.15g of an off-white powder (W0 ₃ , <1% W extraction) . This example shows the importance of milling in the extraction process.

EXAMPLE 7 Extraction of CeQ ₂ from Ce Q ₄ Synthetic monazite CeP0 ₄ .H ₂ 0 (10g) was attrition milled in an aqueous solution of commercial "ammonium carbonate" (60g in 150 ml water) with partially-stabilised zirconia (PSZ) balls (1 mm diam., lOOOg) at 25°C for 24 hours in a closed reaction system, the resulting suspension was centrifuged, and the supernatant filtered and evaporated to dryness. The extraction was repeated twice, and the resulting cream-colored solid was calcined in air (800°C, 4 hours) to give an off-white solid identified by XRD as CeO- (3.39g, 46% Ce extraction).

EXAMPLE 7A (Comparative) In a separate experiment, synthetic monazite

CeP0 ₄ .H ₂ 0 (lOg) was stirred in an aqueous solution of commercial "ammonium carbonate" (60 g in 150 ml water) at 25°C for 24 hours in a closed reaction system, in the absence of milling media. The work-up was the same as

outlined in Example 7 and yielded < 0.lg of solid (approximately 1% Ce recovery). This example also shows the importance of milling in the extraction process.

EXAMPLE 8 (Comparative) Extraction of Cr ₂ Q ₃ from an iron chromite

The procedure of Example 1 was repeated, except that 10 g of a lateritic iron chromite concentrate from New Caledonia was used instead of Very little reaction occurred, with less 0.1 g of solid residue obtained from the extraction process.

EXAMPLE 9 Modified extraction of r ₂ 0 ₃ from iron chromite

The general procedure of Example 8 was repeated, except that before reaction the iron chromite (5g) was intimately mixed with Ba(OH) ₂ (15g) and fired to 850°C in air for 16 h. A portion (4.2g) of the resultant material was attrition milled in an aqueous solution of commercial "ammonium carbonate" (80g in 200 ml water) with partially-stabilised zirconia (PSZ) balls (1 mm diam. , 900g) at 23 ^β C for 16 h in a reaction system open to the air. The brown suspension was centrifuged, and the supernatant filtered to remove any remaining solids. This filtrate was evaporated to dryness and the yellow brown solid, identified as ( H _{4 2} Cr ₂ 0 _? by X-ray powder diffraction. Calcination of this product in air (500°C, lh) have Cr ₂ 0 ₃ (0.82g). A total proportion of 63% of the chromium in the starting concentrate was extracted in this manner.

Finally, it is to be understood that various other modifications and/or alterations may be made without departing from the spirit of the present invention as outlined herein.