AU2007237306A1 - Production of synthetic rutile - Google Patents

Description

AUSTRALIA

Patents Act 1990 COMPLETE SPECIFICATION Standard Patent Applicant(s): WIMMERA INDUSTRIAL MINERALS PTY LTD Invention Title: PRODUCTION OF SYNTHETIC RUTILE The following statement is a full description of this invention, including the best method for performing it known to me/us: 2 PRODUCTION OF SYNTHETIC

RUTILE

Ci This invention relates to the treatment of titaniferous a ores, for upgrading the titania content thereof.

C 5 In the prior art synthetic rutile has been formed from titaniferous minerals, e.g. ilmenite, via various I techniques. According to the most commonly applied technique, as variously operated in Western Australia, the titaniferous mineral is reduced with coal or char in a rotary kiln, at temperatures in excess of 1100 0 C. In this C process the iron content of the mineral is substantially metallised. Sulphur additions are also made to convert manganese impurities to sulphides. Following reduction the metallised product is cooled, separated from associated char, and then subjected to aqueous aeration for removal of virtually all contained metallic iron as a separable fine iron oxide. The' titaniferous product of separation is treated with 2 5% aqueous sulphuric acid for dissolution of manganese and some residual iron. This step has the effect of substantial removal of a sulphide phase formed in the reduction step by virtue of the addition of sulphur or sulphur compounds. There is no substantial removal of magnesium, aluminium or radionuclides from the product at any point in this process, and iron removal is only effected in the aeration step. After calcination the synthetic rutile contains approximately 92% TiO 2 and 1 2% iron as oxide. The addition of sulphur into the kiln has the negative environmental impact of pkoducing sulphurous kiln exit gases. In the acid leach stage malodorous sulphide gases are commonly produced. The cost of countering these environmental effects is substantial and coupled with the cost of the sulphur additive has been influential in preventing the use of sulphur in at least one synthetic rutile installation.

3 The major use for synthetic rutile is as feedstock for the CI production of white titanium dioxide pigment via the Schloride process. According to this process titania bearing minerals are charged with a carbon source to 0 5 fluidised bed chlorination reactor wherein gaseous titanium tetrachloride is formed. The titanium tetrachloride is Ssubsequently condensed and purified and then is oxidised to Stitanium dioxide for use in pigments. The impurities iron, magnesium, manganese and aluminium in titaniferous feedstocks each have deleterious effects, either in (C1 chlorination, condensation or purification. Where a titaniferous mineral, such as ilmenite, has high levels of magnesium or aluminium it cannot be converted to a synthetic rutile by the Western Australian process which does not remove these impurities. Residual iron and manganese, as well as magnesium and aluminium, result in performance based penalties for synthetic rutile feedstocks to chlorination. The presence of radionuclides may also render the synthetic rutile unsaleable.

In other prior art inventions high degrees of removal of magnesium, manganese, iron and aluminium have been achieved. In one such process the ilmenite is first thermally reduced to substantially complete reduction of its ferric oxide content, normally in a rotary kiln. The cooled reduced product is then leached under 35 psi pressure at 140 1500 with excess 20% hydrochloric acid for removal of iron, magnesium, aluminium and manganese.

The leach liquors are spray roasted for regeneration of hydrogen chloride, which is recirculated to the leaching step.

In other processes the ilmenite undergoes grain refinement by thermal oxidation followed by thermal reduction (either in a fluidised bed or a rotary kiln). The cooled, reduced O4 0 product is then subjected to atmospheric leaching with O excess 20% hydrochloric acid, for removal of the Sdeleterious impurities. Acid regeneration is also Sperformed by spray roasting in this process.

0 In yet another process ilmenite is thermally reduced S(without metallisation) with carbon in a rotary kiln, followed by cooling in a nonoxidising atmosphere. The c cooled, reduced product is leached under 20 30 psi gauge 0 10 pressure at 130 0 C with 10 60% (typically 18 C-I sulphuric acid, in the presence of a seed material which assists hydrolysis of dissolved titania, and consequently assists leaching of impurities. Hydrochloric acid usage in place of sulphuric acid has also been claimed for this process.

The major disadvantage of all processes using hydrochloric acid leaching for the formation of synthetic rutile from iron bearing titaniferous minerals such as ilmenite is the need to operate acid regeneration from the chloride liquors formed in leaching. Such acid regeneration requires combustion of large quantities of fuel to provide the necessary heat. The cost of synthetic rutile production by these methods, which are applicable to more general titaniferous minerals due to the ability to remove impurities, is uncompetitive with the reduction/aeration process operated in Western Australia. The major reason for this cost disadvantage is the formation of large quantities of iron chlorides in the process of impurity removal, placing a consequent heavy load on the acid regeneration system.

The major disadvantage of processes using sulphuric acid for the formation of synthetic rutile from iron bearing titaniferous minerals such as ilmenites is the need to O dispose of aqueous iron sulphates (and other sulphates) Sfrom the liquors formed in leaching, in the absence of a Sby-product use for such liquors. Neutralisation with lime, O 5 producing large quantities of red gypsum, which must be disposed of in managed land based repository, will normally be necessary.

C In one process disclosed in the prior art ilmenite is first metallised by fluidised bed reduction with hydrogen or Ce carbon monoxide, followed by aqueous aeration for metallic iron removal as separable iron oxides. The titaniferous product of aeration is then optionally acid leached for upgrading from 85 to 90 precent titanium dioxide up to about 96 percent titanium dioxide, with removal of residual impurities.

The use of gaseous reductants for metallisation is associated with poor single pass reductant utilisation.

Further, fluidised beds are limited in maximum temperature when applied to ilmenite metallisation by gaseous reductants as bed sintering occurs when the temperature exceeds 800 900 0 C. Metallisation rates at lower temperatures at which effective fluidisation is achieved are low. Consequently, either highly inefficient use of reductant at low intensities (a severe economic penalty) or high pressure processing with reductant recycle is required.

The prior art is silent on the possibility of metallisation of magnesium and manganese rich ilmenites using a solid carbonaceous reductant at temperatures above 900 0 C followed by removal of magnesium, radionuclides and other residual impurities. At temperatures above about 9000C carbothermic metallisation of ilmenite commences to become achievable C under practical conditions. However, as presently Sdisclosed, under many conditions a residual impurity Sbearing titania phase which cannot easily be leached of 0 5 impurities, and certainly cannot be leached of all impurities with sub-azeotropic hydrochloric acid, is 0 formed. In particular, the formation of this phase is ts enhanced by the presence of magnesium in the original eC ilmenite. There has been no prior disclosure of the required thermal processing or leach conditions which encourage particular impurity bearing phases to form and then be effectively leached. The lower temperature fluidised bed metallisation work has allowed successful leaching of impurities due to a structure present in the residual impurity bearing phase which has previously not been reported as exclusively available under practical metallisation conditions at higher temperatures.

If carbo-thermic reduction of ilmenite is not performed at temperatures in excess of 900 0 C the rate of reduction of the iron oxides contained in the ilmenite is too slow.

However, at temperatures above 900 0 C there is an increased propensity for the M 3 0 5 phase to be formed in preference to the M03 phase. This propensity increases with increasing temperature. Furthermore, manganese and magnesium promote the formation of M 3 0 s at temperatures in excess of 900 0 c.

In order to overcome this problem, Australian Patent No.

516155 proposes the addition of sodium chloride and sulphur to ilmenite contaminated with manganese, thus causing the manganese to react with the sulphur to form a sulphide impurity phase rather than stablise the formation of M 3 0,.

However, such a process is not useful if the ilmenite is contaminated with magnesium since magnesium stabilizes the

M

3 0 s phase but it does not preferentially form a sulphide.

Consequently, the process of Patent No. 516155 is only useful for producing synthetic rutile from ilmenites contaminated with manganese.

In summary, existing processes for the formation of synthetic rutile products from titaniferous minerals such as ilmenite will either not be effective in the removal of deleterious impurities in many circumstances or will not be cost effective, due to the need to regenerate large quantities of expensive reagents, dispose of large volumes of leachate liquors or to operate largely impractical and economically unattractive thermal processing schemes. It is the object of the present invention to overcome, or at least alleviate some of these difficulties.

Accordingly, the present invention provides a process for upgrading the titania content of a titaniferous material which process comprises the steps of:reducing the titaniferous material using a solid carbonaceous reductant under temperature conditions which promote the formation of metallic iron, a rutile phase and a separate impurity bearing titaniferous phase to produce a thermally reduced product; (ii) cooling the thermally reduced product in an environment which prevents substantial re- 4 oxidation to produce a cooled reduced product; and (iii) removing iron and other impurity elements from the cooled reduced product by a leaching step or an aeration and leaching step to produce a synthetic rutile.

O 8 0 In the process the iron present in the titaniferous mineral O may be partially reduced to metallic iron by coke, char or coal. The temperature of reduction should preferably be Sabove 900 0 C. For each mineral the optimum temperature of reduction will depend on the level of impurities present and the reductant used. In general, conditions should be Sset such that the predominant titanium bearing phase in the mineral after reduction is rutile or reduced rutile, while C- impurities such as magnesium, manganese and residual (nonmetallised) iron are predominantly concentrated into a CI small amount of a separate titaniferous phase. This separate phase may have either the anosovite/pseudobrookite structure or an ilmenite-like

"M

2 0 3 structure. The latter structure is advantageous in subsequent steps of the process. The structure obtained will determine the nature of the subsequent steps.

Reduction may be carried out in any suitable device including shaft furnaces and rotary kilns. There is no presently available practical and commercial means of achieving metallisation of ilmenite using solid carbonaceous reductant as sole reductant in fluidised beds due to the difficulty of maintaining a heat balance under reducing conditions in the absence of external heating.

The presently preferred apparatus is a rotary kiln charged with solid carbonaceous reductant such as coal, char or coke and preferably operated with a maximum temperature in the range 950 1050 0 C. However, higher temperatures may be operated, especially for ilmenites having low levels of impurities. It is not anticipated that the process would be operated at temperatures significantly in excess of 1200 0 C due to the unavoidable formation of large quantities of non-rutile phases, particularly

M

3 0 5 and a tendency for the mineral to sinter and accrete at higher temperatures.

The formation of appreciable quantities of non-rutile CI phases may result in low selectivity of impurity removal in Sthe final leaching step of the process due to the Sdissolution of solubilised titania. Formation of M, 3 0 in 0 5 preference to M 2 0 3 will result in a need to use more aggressive conditions (higher acid strength, temperature I and time) in leaching, which may be difficult to apply in practice.

The degree of conversion of the titaniferous mineral iron eC content to metal is not a critical part of the process provided that the sought after phases are produced in reduction. The optimum degree of metallisation will be determined primarily by economic consideration in most circumstances. In general, degrees of metallisation in the range 50 95%, depending on mineral composition, will be suitable. A suitable degree of metallisation can be achieved in residence times from 30 minutes to several hours at or above 900 0 C for carbothermic reduction in a rotary kiln. Actual metallisation for a given reduction time and temperature will depend on the nature of the mineral and the nature of the reductant on carbon reactivity).

After reduction and the attainment of the desired degree of metallisation, the material being heated must be cooled almost to room temperature in an environment that prevents substantial re-oxidation. Cooling may be conducted in a cooler which forms an integral part of the reduction unit or in a separate cooling unit in an atmosphere of inert gases or reduction product gases.

Separation of carbonaceous material from minerals may then be performed by a suitable combination of magnetic and size separations, with the carbonaceous component recirculated, 0 as appropriate.

U Metal may be removed from the cooled mineral particles by any suitable means. Aqueous chemical methods are most suitable. Acid leaching using any commercially available acid is effective in removing iron but results in iron Ssalts in solution. The resulting solution will normally Srequire iron precipitation by neutralisation, spray Sroasting to iron oxide for acid recovery, or some other means of further treatment to avoid the need for disposal of environmentally harmful wastes. The most advantageous method for iron metal removal is aqueous aeration, in which air is blown through an agitated aqueous suspension of metallised mineral in the presence of added reagents such as ammonium chloride. Iron metal is converted to iron oxides according to this method. This technique is well known in the prior art. By adjustment of conditions the nature of the iron oxide product of aeration can be altered and its formation as a separable, finely grained suspension can be ensured.

Following aeration, separation of the titaniferous product from the iron oxides can be effected by any suitable method of sizing separation, such as by passage through cyclone separators. The coarser titaniferous product may then be dewatered by any suitable technique or combination of techniques, e.g. thickening and filtration.

The dewatered titaniferous product of aqueous chemical treatment according to the described process contains virtually all of its original magnesium, manganese aluminium and radionuclides and may have substantial quantities of residual iron oxides which were previously not metallised or have adhered during an aeration step. It has been found that leaching with strong mineral acid CI having a concentration in the range from 4 to 50 weight Spercent is effective in the removal of these impurities, Sprovided that appropriate conditions have been used in S 5 reduction.

Acid leaching using strong mineral acids under agitated conditions may be applied to impurity removal. For (C example, both sulphuric and hydrochloric acids have been shown to be effective. Prior to leaching it may be ^C advantageous to grind the titaniferous mineral to assist leaching kinetics, although this step is not essential to the process. Leaching with excess 18 20 wt% HC1 has been found to be particularly advantageous, and is preferred although lower concentration of acid down to 4 wt%) have also been found to be effective.

Acid leaching may be conducted in any suitable batch or continuous leach vessel. For example, heated, agitated vessels or fluidised bed vessels may be used. Typically the leaching temperature will be 80 150 0 C, depending on the leachant. Leaching may be conducted either at atmospheric or at elevated pressures, although a feature of the present invention is the ability to operate the leach step without the need for pressure vessels. Leaching time may be from 15 minutes to 24 hours, depending on the phase assemblage present in the reduced mineral and the desired degree of impurity removal. Greater than 80% removal of each of iron, magnesium, manganese and partial removal of aluminium and radionuclides may be easily achieved by the described process.

At the conclusion of leaching the leach liquor may be separated from the mineral by any suitable means including thickening, washing and filtration. The mineral product is then dried and calcinated for removal of moisture and CI chemically combined water.

SCalcination at temperatures in the range 300 900 0 C has S 5 been found to be effective. The resulting synthetic rutile product will contain greater than 90% Tio 2 and up to 99% STiO 2 depending on the level of impurities in the original titaniferous mineral grains, and the presence of non Stitaniferous grains in the original mineral which are retained through the process.

Additional steps may be incorporated in the process, as desired. For example: The original titaniferous mineral may be agglomerated prior to reduction, with or without regrinding, by any suitable technique. In this manner a size consist which most suits the process dynamics of subsequent steps, e.g.

reduction, may be obtained.

Additives, such as chlorides or oxides (e.g.

MnO 2 may be mixed into the titaniferous mineral prior to reduction in order to redistribute the metallic iron produced via segregation reactions, thereby influencing metallic iron removal, or to encourage the formation of an acid leachable minor impurity bearing phase.

The titaniferous mineral, or admixture containing the titaniferous mineral may be oxidised at elevated temperatures, preferably in excess of 700 0 C, to provide a degree of preheat to the mineral prior to reduction, and to enhance the rate and extent of the reduction reaction.

C Following reduction the cooled, partially Smetallised mineral may be subjected to magnetic Sor other separation procedure for removal of o 5 impurity grains which do not metallise. Grinding prior to such separation may be operated with the Sobjective of liberation of impurity grains from titaniferous grains.

Mineral separations based on magnetic separation, C gravity separation, flotation or any other separation technique may be performed either after removal of metallic iron from the-reduced mineral or after final acid leaching and/or calcination. In this manner impurity grains e.g.

chromite may be removed.

The final titaniferous product may be agglomerated, with or without regrinding, by any suitable technique, to produce a size consist which is suitable to the market for synthetic rutile. After agglomeration the product may be fired at temperatures sufficient to produce sintered bonds, thereby reducing dusting losses in subsequent applications. Firing in this manner may remove the need for product calcination.

Leaching may be conducted either batch-wise or continuously, or in multiple co-current or countercurrent (in relation to solids and liquid flows) stages.

14 CI Within the disclosed process there is great flexibility in Srelation to the degree of iron removal in the first and second stages of aqueous treatment, and therefore the acid S 5 recovery or neutralisation costs. For many titaniferous feeds higher degrees of metallisation in reduction will Scorrespond to greater difficulty of subsequent impurity Sremoval in acid leaching due to the stabilisation of impurities in the less reactive anosovite phase.

Consequently, an optimum balance between leach liquor treatment costs and difficulty of impurity removal may be struck, depending on the economic environment.

Examples: The following examples describe a number of laboratory and pilot scale tests which serve to illustrate the techniques disclosed herein.

Example 1: 300 g of ilmenite in the size range 45 65pm having the composition given in Table 1 was mixed in equal weight proportions with Victorian brown coal char and placed in a I.D. lidded stainless steel pot. This pot was then situated in a 950 0 C muffle furnace for 3.5 hours, after which time it was removed and allowed to cool.

The cooled mineral product was separated from associated char by magnetic separation, and then leached for removal of metallic iron with excess 5% sulphuric acid for minutes.

In this step 89% of the iron content of the reduced mineral was removed into solution. The solids residue was filtered away from the liquor and then leached with refluxing 50 wt% sulphuric acid for 24 hours. After 24 hours of leaching the leached solids contained 0.77% MgO, compared with an Cl initial 2.25% dry basis). However, approximately 15% of the titania was also taken into solution.

Example 2: 1 kg of agglomerated ilmenite (-710 250 pm) having the Icomposition given in Table 1, was mixed in equal weight proportions with Victorian brown coal char 5mm) and heated to 1000 0 C under 0.3m sec 1 nitrogen superficial velocity in a fluidised bed reactor. Upon reaching C-1 temperature an 0.3 sec-1 superficial velocity flow of carbon monoxide fluidising gas was commenced and maintained for a total of 4 hours. At the end of this time the bed was permitted to cool under nitrogen flow and the bed was separated magnetically and by sizing into char and mineral. Chemical analysis indicated that the mineral was 95% metallised.

A 1 g sample of reduced mineral was leached with 5 wt% sulphuric acid to the point of complete removal of metallic iron. The solids residue was then leached with excess boiling 20 wt HCl solution under reflux for 31 hours.

the removal of various elements from the mineral is summarised in the following table: Element Mg Ti Mn Fe Al Removal 96.1 9.0 99.6 97.9 80.3 Example 3: A 2 .6:1(wt basis) Victorian brown coal char (-5mm 0.Smm)/agglomerated ilmenite (Table 1 4mm 250pm) mixture was fed continuously at 18 kg/hr to an inclined 0.4m internal diameter, 5m long rotary kiln. The kiln was fired from the discharge end with a gas burner, and combustibles in the above-bed gas space were combusted by 16 injection of air at controlled rates via air lances at C points along the kiln length. The kiln solids bed Stemperature profile increased uniformly from 200 0 C to 1000 0 C over the length of the kiln from the charge point to C 5 the discharge. Total solids residence time was estimated at 4 hours over this length. The kiln discharge was cooled I to room temperature through a spiral cooler. A 300g sample M of cooled kiln discharge was magnetically separated for pC char removal. A subsample of the magnetic product was 10 analysed by X-ray diffraction, indicating major rutile and metallic iron phases, with minor quantities of the impurity bearing phases anosovite pseudobrookite and ilmenitelike metatitanate occurring in roughly equal proportions.

A further subsample was subjected to analysis for degree of metallisation by measurement of the magnetic attractive force on the sample in a saturating magnetic field against a known calibration. The indicated degree to which iron had been converted to metal was 78.3%.

A further 5g subsample of the magnetic product was subjected to 5 wt% sulphuric acid and 20% hydrochloric acid leaches as described in the previous example. After four hours of the final leach virtually all of the ilmenitelike metatitanate phase had been removed, while most of the anosovite/pseudobrookite phase remained. According to this example the formation of the ilmenite-like metatitanate residual phase is to be encouraged as it is more readily leached, with consequent removal of associated impurities.

Example 4: This example illustrates the thermal reduction step of the process of the invention. Ilmenite of the composition provided in Table 2 and in the size range -2501m 100 pm was treated through the rotary kiln of Example 3 in a g similar manner to that specified above, with the exception Ci that a flat temperature profile, at 950 50oc, was Smaintained over the final 2 metres of kiln length.

0 5 X-ray diffraction analysis confirmed that for this ilmenite at degrees of metallisation in excess of 90% the residual N impurity bearing phase in the product reduced in this Smanner was predominantly metatitanate. It is apparent that M at greater levels of impurities, as for the ilmenite of 10 Table 1, the anosovite-pseudobrookite phase is more Sfavoured, requiring reduction at lower temperatures if the more readily leached metatitanate phase is desired.

Example Two 3 kg batches of -65 45pm ilmenite having the composition recorded in Table 3 were mixed with 1.5kg of -4 1.4mm Victorian brown coal char and placed in a muffle furnace for heating to a final steady state bed temperature of 1000 The first batch was held above 900 0 C (metallisation initiation temperature) for 5 hours, while the second batch was held above 900 0 C for 3 hours. The batches were removed for cooling in air at the end of the heating period.

Magnetic separation was performed on the products of such reduction for removal of char, and the degree of metallisation of contained iron was recorded for the magnetic fraction as follows: metallisation Batch 1 87 Batch 2 a7 SThe metallised minerals were subjected to iron metal c removal by leaching with excess 5 wt% H 2 S0 4 for 90 minutes Sat 80 0 C, before-filtration, washing and drying to recover leached solids. X-ray diffraction analysis indicated the 0 5 following phase distributions: SBatch 1 Batch 2 Rutile 44% 32%

M

3 0O 30%

M

2 0 3 12% 43% The above materials were each subjected to leaching with refluxing excess 20 wt% hydrochloric acid for 6 hours.

Extractions of residual impurities from the already demetallised material were as follows: %Removal Batch 1 Batch 2 Iron 41.4 96.5 Manganese 14.3 88.9 Magnesium 27.8 80.9 Aluminium 17.1 18.0 Titanium extraction was negligible in each case.

X-ray diffraction analysis diffraction analysis of the residues in each case indicated complete removal of the ilmenite, with slight removal of the M 3 0, phase in the case of batch 2, but no M 3 0O removal in the case of batch 1.

Samples of each of the demetallised materials were also subjected to leaching with excess refluxing 50% sulphuric acid for up to 24 hours. Extractions of residual impurities in the leaches after one hour were as follows: %Removal Batch 1 Batch 2

C

N Iron 88.4 95.7 SManganese 81.7 86.0 Magnesium 75.5 85.7 C 5 Aluminium 26.1 24.9 Titanium extraction after one hour in the above cases was Shigh 20%) but hydrolysis of dissolved titania Soccurred over time in the leach to result in losses as low as 0.3%.

X-ray diffraction analysis of the residues in each case indicated virtually complete digestion of both M03 and

M

3 0s residual phases.

Example 6: This is a comparative example illustrating a process for reducing ilmenite at a temperature less than 900 0 C in a fluidised bed using a large excess of hydrogen as reductant. A 5 kg charge of ilmenite (-65 35pm) having the composition provided in Table 4 was fluidised with air at a superficial velocity of 30 cm sec-' (at temperature) in an external heated oxidising fluid bed roast conducted at 900oC for 30 minutes.

The temperature of the fluidised bed was then allowed to fall to 750 0 C and the bed was purged with nitrogen at sec- 1 fluidising velocity for 30 minutes. The fluidising gas was then replaced with hydrogen at a superficial velocity of 64cm sec- 1 Hydrogen reduction continued for 160 minutes, after which time the hydrogen was replaced with a purge of nitrogen and the bed was allowed to cool.

Analysis of the reduced ilmenite product indicated that 76% of its contained iron was metallised. This metallisation g was removed by a 9 minute leach in S wt% H 2 S0 4 at C- The filtered and dried leach residue was then subjected to a further leach with excess 8.7 wt% hydrochloric acid/100 Sgram per litre ferrous chloride leachant, under reflux S 5 conditions. Extraction of residual impurities in the final leach were as follows: 0C %Removal 1 Iron 98.7 C Manganese 99.2 Magnesium 99.0 Aluminium 13.8 X-ray diffraction analysis of demetallised and residue samples indicated that the only residual impurity bearing phase in the demetallised sample was 4203 and that this phase was entirely removed by the final leach.

Although the process described above results in the removal of most of the iron magnesium and manganese, it would not normally be economic because a substantial excess of hydrogen was used during the reduction step.

Example 7: Two kilograms of the ilmenite used in Example 6 (see Table 4) was oxidised in a rotation pot inserted into a laboratory muffle furnace, at 1000oC in the presence of excess air. The oxidised ilmenite was allowed to cool, and then mixed 1:1 (weight basis) with Victorian brown coal char 0.5mm). The mixture was then held for one hour in the absence of air in the rotating pot assembly with the furnace set at 950oC, and then allowed to cool.

Char was separated from the cooled mixture by magnetic separation and screening. The iron content of the separated reduced ilmenite was found to be 79.2% Ci metallised.

SMetallic iron was removed by a 90 minute leach in 5 wt%

H

2

SO

4 at 80 0 C. The filtered and dried leach residue was then subjected to further leaches as follows: SLeach 1: Excess 18.5 wt% HC1 for 6 hours at 104 0

C

Leach 2: Excess 20 wt% H 2

SO

4 for 6 hours at 130 0

C

under pressure.

1 0 Extraction of residual impurities from demetallised samples were as follows: %Removal Leach 1 Leach 2 Iron 77.5 88.1 Manganese 18.9 90.3 Magnesium 81.2 91.7 Aluminium 29.6 45.5 X-ray diffraction analysis of residue samples indicated that the M 2 0 3 (predominant impurity bearing phase) was completely removed by both leaches while Leach 2, with sulphuric acid, also removed most of the M 3 0, phase. Only approximately 4% of the contained titania was dissolved in each of the leaches.

Example 8 Several tonnes of an ilmenite concentrate having the composition indicated in Table 5 were lightly milled (passing 60 microns) and agglomerated with 1% bentonite addition to a size range of 100 microns to 2mm in a small agglomeration plant. The agglomerated ilmenite was then fed to a pilot rotary kiln with excess brown coal char g reductant and heated to a maximum bed temperature of 950oc.

C The kiln discharge was cooled through an Archimedes spiral U mounted on the kiln without reoxidation. After magnetic separation of the kilned agglomerates from the residual 0 5 char it was analytically determined that 80-85% of the iron in the original ilmenite had been converted to iron metal.

SA sample of the metallised ilmenite was then subjected to M leaching with excess 20% hydrochloric acid for the initial 10 removal of metallic iron followed by completion of leaching Sfor the removal of minor impurities, including radioactive impurities. The final product of leaching is also indicated in Table 5. Clearly there has been substantial removal of iron, magnesium, manganese, aluminium, uranium and thorium. Radiometric analysis of material treated in an effectively identical manner indicated that 70% of the radioactivity in the original material (measured in Becquerels per gram) had also been removed.

Example 9 Agglomerates of a titaniferous concentrate previously milled to pass a 35 micron screen and having the composition indicated in Table 6 were prepared in a laboratory agglomeration facility. The agglomerates, which were sized to pass a 2 mm screen,.were formed with the addition of 1% sodium silicate, 10% water and micronised sodium carbonate added to result in a weight ratio of silica to soda in a subsequent roasted material of 2.4:1.

The agglomerates were reduced with 5% Victorian brown coal char addition at 1000oc in a rotating pot reactor under a flow of nitrogen to prevent air ingress for 30 minutes. At the end of this time the pot was rapidly cooled by application of water sprays onto its external surface.

Residual char was removed from the product of reduction by 23 magnetic separation and the magnetic product so formed was C sent for further treatment. This product was found by Schemical and X-ray diffraction analysis to contain amongst Sother phases metallic iron, rutile, an ilmenite like phase S 5 (M 2 0 3 and a minor glassy phase discovered to contain approximately 20% TiO 2 and containing virtually all of the QD original silica and much of the original thorium. (The silica was originally in the form of quartz inclusions in the titaniferous grains of the concentrate.).

Approximately 2.5% of the contained iron in the reduced Sproduct was as iron metal.

The reduced product was crushed to pass a 75 micron screen aperture and subjected to a one hour leach with 30% sodium silicate (silica to soda 2.4:1 by weight) under reflux for decomposition of the glassy phase. After solid/liquid separation and washing the leach residue was leached with a slight excess of 20% hydrochloric acid under reflux for one hour. The combined leach treatments removed over 90% of the contained iron, magnesium, aluminium and manganese, over 70% of the contained silica and alumina, and approximately 80% of the contained thorium and radioactivity. The acid leach was effective in the removal of the greatest proportion of these impurities with the exception of alumina and silica which were primarily removed in the sodium silicate leach. Removal of silicon and aluminium, and proportional removal of thorium was enhanced by the presence of the glassy phase.

Table 1: Compodition of Ilmenite in Examples 1 3 FeO Fe 20 3 2 Cr 2

O

3 S i0 2 A1 2 0 3 Cao MgO MnO

V

2 0 5 ZrO 2 P 2 0 5 Wt% 9.68 25.3 53.4 0.62 1.60 1.94 0.06 1.48 1.23 0.25 0.17 0.46 Table 2: Composition of Ilmenite in Example 4 Wt%6 FeO Fe 2

O

3 Tib 2 Cr 2

O

3 S i0 2 A1 2 0 3 CaO HgO

MAO

V

2 0S ZrO 2

P

2 0S 23.2 16.8 53.8 0.05 o0.68 0.84 0.26 0.34 1.50 0.14 0.07 0.06 Table 3: Composition of Ilmenite in Example FeO Fe 2 0 3 T1O 2 Cr 2 03 Si0 2 A1 2 0 3 CaO HgO Mno

V

2 05 Zr0 2

P

2 05 Wt%6 10.5 23.6 51.4 1.01 1.23 1.22 0.11 1.60 1.19 0.25 0.73 1.55 Table 4: Composition of Ilmenite in Example 6 FeO Fe 2 0 3 Ti0 2 Cr 2

O

3 S i0 2 Al1 2 0 3 CaO Ego MnO

V

2 0 5 ZrO 2

P

2 0S 1*01 Wt% 8.76 26.2 57.3 0.54 1.16 0.65 0.05 1.40 1.30 0.25 0.15 0.05 0.71 Table U Concentrates and Product in Example 8.

Wt% Concentrates Product Ti0 2 56.1 95.0 Fe as FeO 35.4 0.40 IDSi0 2 0.95 2.1 A1 2 0 3 0.73 0.46 Hg0 1.33 0.13 Cr 2 0 3 0.34 0.57 MnO 1.45 0.078

V

2 0 5 0.24 0.15 ZrO 2 0.11 0.2-1

U

3 0 8 10 6 ThO 2 105 32

Claims (9)

1. A process for upgrading the titania content of a titaniferous material which process comprises the steps of:- reducing the titaniferous material using a solid carbonaceous reductant under temperature conditions which promote the formation of C metallic iron, a rutile phase and a separate impurity bearing titaniferous phase to produce a thermally reduced product; (ii) cooling the thermally reduced product in an environment which prevents substantial re- oxidation to produce a cooled reduced product; and (iii) removing iron and other impurity elements from the cooled reduced product by a leaching step.

2. A process according to claim 1, wherein the leaching step is preceded by an aeration step for removal of metallic iron.

3. A process according to claim 1, wherein the leaching step comprises two stages

4. A process according to claim 1 wherein the minor impurity bearing phase is a metatitanate (M 2 0 3 or an anosovite/pseudobrookite like phase (M 3 A process according to any one of the preceding claims wherein step comprises reducing the titaniferous ore or concentrate with a solid carbonaceous reductant at a temperature in a range from 900 0 C to 1200oc.

6. A process according to Claim 1, wherein step (i) M 5 is performed at a temperature in a range from 900 0 C to 1050 0 C. ID

7. A process according to claim 1, wherein the minor Simpurity bearing phase is anosovite or pseudobrookite and step (iii) includes leaching in excess hot sulphuric acid whilst agitating the hot sulphuric acid for a period of from 15 minutes to 24 hours.

8. A process according to claim 7, wherein the acid has an initial concentration of up to 50 wt%.

9. A process according to claim 1, wherein the solid carbonaceous reductant is sub-bituminous or lignitic coal or char derived therefrom. A process according to claim 1, wherein the rutile phase includes reduced rutiles.

11. An upgraded titaniferous material produced according to the process of any one of the preceding claims.