IL39102A - The beneficiation of ilmenite ores - Google Patents

Description

Improvements in and relating to the beneficiation of ¾jnenite ores LAPOHTB INDUSTRIES LIMITED Cs 57127 This invention relates to a process for the beneficiation of ilmenite ores* and is especially concerned with the production of a suitable feed material for the manufacture of titanium tetrachloride by chlorination under fluidised-bed reaction conditions.

In the production of tita&iura tetrachloride "by fluidised-bed chlorination, the overall efficiency of the process is impaired when some of the particles present i the chlorinator are so small that they are entrained in the fluidising gas and are carried out of the reactor either unreacted or incompletely reacted· This can either arise from th© presence in the starting material of particles which are too email, or from breakdown of the' starting material under the fluidised-bed reaction conditions to yield fines. Inadequate particle sise in the starting material can result from grinding the ore material before it is beneficiated, and is also observed when* following undesirable dissolution of titanium in a leaching operation, it proves necessary to effect some re-precipitation.

A number of previously proposed processes for beneficiating titaniferous ore materials involve oxidising the ore to convert its iron content entirely to the ferric state, reducing the oxidised material so as to convert its iron content, either wholly or in part, to metallic iron, and finally leaching the reduced product* In a recent proposal, it has been suggested that the reduction should not bo carried bejrond the stage in which metallic iron -constitutes 20$ by weight of the reduced product.

The present invention is based in part on the observation that the tendency for breakdown of the beneficiated product in a fluidised-bed to yield fines is associated with the presence of metallic iron in the reduced material prior to the leaching stage. It has "been found that the product obtained by leaching material that contains metallic iron contains cavities having diameters of about 10 microns or more, and it is believed that the presence of these cavities is responsible for the undesirable breakdown characteristics of such beneficiates under fluidised-bed reaction conditions.

The present invention provides a process for beneficiating an ilmenite ore, which comprises forming from the llmenite a material of which at least 20$ by weight has a crystal structure of the without forming any substantial quantity of metallic iron and pseudobrookite type, educing the said material/under such conditions that not less than 4$ of the iron content of the reduced ma erial^ (calculated as Fe) is in the ferric state, and leaching the reduced material so obtained to yield a beneficiated titaniferous material* It has been found that, by ensuring that at least 4$ of the iron content of the material remains in the ferric state, no detectable metallic iron is formed under typical reduction conditions* As a result of the formation of material having a structure of the pseudobrookite crystallographic type (hereinafter termed merely "pseudobrookite" for convenience), the reduced material formed subsequently is very readily leachable, making it possible to obtain a beneficiated product having a high titanium dioxide content, with very little loss of titanium values in the leaching step. Typically, the process of the invention leads to a product having a titanium dioxide content of 9 -93 » with 95-98$ efficiency of recovery of titanium values from the reduced materialo In the case of certain ilmenites, for example, Western Australian ilmenites, even better results have been obtained.

Advantageously, the formation of pseudobrooki e is effected b sub ectin the ilmenite ore to oxidisin conditions. Thus pseudobrooki e formation may be brought about by contacting ^the ilmenite with oxygen or, preferably, an oxygen-containing gas, at an elevated temperature. Conveniently, the pseudobrookite is formed by heating the ilmenite ore in ai e Under oxidising conditions, pseudobrookite can be formed from all types of ilmenite ores. In the case of weathered ilmenites, however, the desired extent of pseudobrookite formation c an, if desired, be brought about under substantially inert conditions, ~ although in practice it will again generally be more convenient simply to heat the ilmenite in air. Conditions are "substantially inert" for this purpose if they are^such that not more than 1/15 by weight of the iron content of the ore is reduced. Preferably, non-reducing conditions are employed. The following expression is to be used in calculating the fraction of tbe total iron content that has been reduced: where X^ is the percentage by weight of ferrous iron (calculated as FeO) in the starting material; Xg is the percentage by weight of ferrous iron (calculated as FeO) in the reduced material; and is the total iron content of the starting material (expressed as a percentage by weight and calculated as Fe) .

The expression "weathered ilmenite" is intended to include beach sand ilmenites, alluvial ilmenites, and leucoxenes. The use of inert conditions leads to better results with relatively highly weathered ores such as, for example, Quilon ilmenites, than with less weathered ores such as, for example, Western Australian ilmenites.

Pseudobrookite cannot be formed from non-weathered massive ilmenites, for example, Norwegian and Maclntyre ilmenites, merely by heating under inert conditions.

The rate of formation of pseudobrookite is very strongly ° affected to a small extent by the presence of artificially added materials in admixture with the ilaenite ore, or by the presence in the ore of atypically large quantities of impurities The formation o f pseudobrookite may be ef ected in the presence a compound of lithiumt sodium9. inegnesi rap zirconium* or s.ine? preferably in the presence of a lithium compound» Such compound oerve to promote the formation of pseudobrookite e d enable that st relatively low temperatures. If desired, the promoters may instead be used to reduce the time required for the formation of pseudobrookite at a given temperature, or for a combination of these purposes* Whilst it is only necessary to form 20$ by weight of pseudo brookite in order to bring about an improvement in the leacha-bility of the reduced material obtained subsequently, the proportion of pseudobrookite formed is advantageously at least 30 (and preferably at least 50$¾) , estimated by X-ray analysis. It will generally be found to be desirable to carry out the first stage of the process under such conditions that the theoretical maximum proportion of pseudobrookite is approached. Typically, when the pseudobrookite is formed by oxidation, the theoretical maximum proportion will be in the range offrom about 60 to about 70$ by weight, although it may be as high as 75$ in some cases.

The pseudobrookite formed from ilmenite will have a composition in the range depending on the composition of the original ilmenite and on the extent of oxidation occurring in the first stage of the process. If oxidation of the iron content of the original ilmenite is substantially complete, the pseudobrookite will have the composition TiFe20^. The crystal structure of that material is well-known and is described, for example, by Pauling in Z. ristall, volume 73 (1930), pages 97 to 112. Other materials in the composition range given above have closely similar crystal structures (see Akimoto et al, Nature volume 179 (1957), pages 37 and 38).

The pseudobrookite content of a given sample may be measured by comparing the X-ray diffraction pattern of the sample with patterns obtained using standard samples contairing known ro ortions of the same kind of seudobrookite.

Akimoto e al describe a method of preparing different pseudo-brookites within the composition range by heating a mixture of ferric oxide, titanium dioxide, and metallic iron to 1150°C. and thereafter quenching the hot mixture. The total iron and titanium contents of the standard pseudobrookite-containing sample should be in the same ratio by weight as in the original ilmenite. This can be achieved by mixing the synthetic pseudobrookite with appropriate amounts of TiC>2 and/or FegO^. It should be noted that, using the foregoing method, it will in general be found that there will be no appreciable detection unless the pseudobrookite content is about 5# b weight or above* ■ The formation of pseudobrookite may be carried out in any suitable apparatus, for example, in a rotar kiln, which may be operated either co-current or counter-current* The reduction reaction may be carried out at a temperature in the range of from 800° to 1000°C, advantageously from 850° to 950°C. The use of temperatures in excess of 1000°C. leads to increasin difficulties in preventing excessive reduction* Preferably, the material containing pseudobrookite is raised to the desired reduction temperature rapidly. The rate of attainment of the reduction temperature is dependent principally on the design of the reduction apparatus, bu it is preferable to transfer the material containing pseudobrookite into the reduction zone while the material is still hot, so lessening the time taken for the material to reach the desired reduction temperature.

The reduction step should be carried out in an oxygen-deficient atmosphere, say, about io of oxygen by volume, and the reducing agent should be present in excess, taking into account the oxygen derived from the reactants and any other oxygen that may be present in the reaction apparatus. As is explained hereinbefore it is essential that the reduced material should contain no detectable metallic iron, and this leads in turn to the requirement that at least 4S& of the iron content of the reduced material must remain in the ferric state* Although the proportion of trivalent iron in the reduced material can exceed 4%, a relatively lengthy leaching operation is required at high proportions of ferric iron.

Also, pseudobrookite itself is not reddily leachable, and the presence of pseudobrookite in the reduced material leads to a leached product^having inferior titanium dioxide content* Accordingly, the reduction conditions are advantageously such that the reduction product contains substantially no detectable pseudobrookite (which implies that not more than 5 by weight of pseudobrookite is present)* The reduction product then consists predominantly of ilraenite and rutile. In principle, it is desirable to eliminate all of the pseudobrookite, but broadly similar results are obtained with pseudobrookite contents of up to about 5 by weight.

In general, optimum titanium dioxide content in the leached material is obtained if the reduction reaction is not continued after the pseudirookite has been eliminated. It is believed that the .teaching characteristics of the reduced material are related to its porosity and crystallinit . It is also believed that the material possesses the optimum properties in terms of porosity and crystallinity when the pseudobrookite phase has just disappeared. It will frequently be the case, however, that the ferric iron content of the material is so high at that stage that the required leaching time will b c n nt r m ct b . ferric state when elimination of pseudobrookite is substantially complete varies according to the type of ore in question, being about 18 in the case of Western Australian ilmenite, about 40 in the case of Norwegian ilmenite, and about 11$ in the case of Ojuilon ilmenite. It will usually be found advantageous, however, to decrease the proportion of ferric iron below these maximum figures. In the case of Western Australian ilmenite, the reduction is advantageously continued until a proportion of from 8 to 12$ of the iron content of the material is in the ferric state. More generally, if the reduction stage is carried out under such conditions that not more than 10$, and preferably at least 8$, of the iron content of the reduction product is in the ferric state, good results will be obtained.

When the desired extent of the reduction has been achieved, the temperature of the reduced material should be quickly lowered, advantageously by discharging the hot reduced material into a cooling zone. If the temperature of the reduced material remains high after the reduction is complete, especially if its temperature is then relatively high, its leaching characteristics will suffer. This deterioration is believed to be associated with changes in rosity and crystallinity that take place if the material is maintained at a high temperature with no significant chemical reaction occurring. Similar changes (which involve increasing pore size and development of crystallinity) do in fact take place to some extent if the material is maintained hot for any length of time after elimination of the pseudobrookite, but their effect will generally be compensated by the concomitant removal of ferric iron, provided that reduction still continues while the material is hot.

The association between leachability and prosity has already been mentioned. It has been found that the porosity of the reduced product in the size range 0.04 to 2 microns is a useful guide to its leachability. The percentage porosity of a material is given by the expression: pore volume χ 1Q(^g apparent volume The pore size distribution in a material may be measured by the well-known technique of mercury porosimetry.

Generally, the reduced product will be readily leachable if it has at least i porosity in the size range 0.04 to 2 microns. Desirably, the porosity in that range . is at least 8J&. In principle, materials having even higher porosities, say of the order of 10?o or more, will be more readil leachable* As pointed out hereinbefore, however, difficulties can arise in endeavouring to combine high porosity with low ferric iron content.

If the temperature of the material is not quickly lowered when the reduction is complete, the pore size distribution will in general not satisfy the conditions laid down in the preceding paragraph.

If the material is to be discharged from the reduction apparatus when the pseudobrookite has been e?.liminated, then it is preferable to attain a relatively high temperature at that point. If, on the other hand, the reduction is to be continued after the elimination of pseudobrookite, it will generally be advisable to- operate with a relatively lower maximum temperature, although it is again preferable that elimination of pseudobrookite should take place at or near the maximum temperature employed.

If the material is to remain in the apparatus after elimination of pseudobrookite, it must be emphasised that the desirable crystallinity and porosity will be lost if the temperature remains high with no reduction occurrin * The reductio stage of the process may be effected using a gaseous reducing agent, for example, hydrogen or a hydrocarbon gas such as methane. A preferred gaseous reducing agent is carbon monoxide. Typically, when the reducing agent comprises carbon monoxide, the reduction may be effected at about 800°C, as compared with about 900°C. for methane reduction. Temperatures as low as 700°C. may be employed when the reducing agent comprises hydrogen. Reduction with a gaseous agent may be effected under fluidised-bed reaction conditions.

The reduction may instead be effected by contacting the pseudobrookite-containing material with a solid carbonaceous reducing agent at an elevated temperature. Although the use of a gaseous reductant^has the advantage that no solid/solid separation is needed after the reduction, the reaction will commonly proceed more rapidly when a gaseous agent for example, methane) is used, and especially careful control is then required in order to ensjire that at least 4$ by weight of the iron content of the reduced material is in the ferric state.

It will be appreciated that the reducing agent may b e used in admixture with other substances which may assist in the reduction process. For example, hydrogen may be used in admixture with steam.

Any solid carbonaceous reducing agent may be used, but there should be present, preferably contained in the solid material, a quantity of material that is volatile under the conditions of he reaction. The presence of volatiles is essential if the reaction is to proceed at an adequate ra.te. If used alone, pure carbon containing no volatile matter, for example, calcined petroleum coke, is extremely inefficient in bringing about the desired degree of reduction. The exact volatile content of the''reducing agent is not critical, however, and substances having widely differing volatile contents produce similar effects in the reduction reaction, althoigh the rate of reaction will not be the same in ea.ch case. In eneral more rapid is the rate of reduction at a given temperature.

Examples of solid carbonaceous reducing agents which may be used include green coke, semi-coke, lignite,. and the coal/coke mixture known as reject char, which is a material obtained in the low-temperature carbonisation of coal.

Advantageously, the reducing agent is coal, for example, anthracite. Any coal may be used, provided that it is non-caking under the conditions of the reaction. The tendency of certain coals to form calces may be overcome by oxidising the coal for a period of from about ¾ hour to aboutl-g- hours at a temperature in the range of from 200° - 400°C, and such pre-oxidised coal may be used in the process of the present invention.

With regard to the particle size of a solid carbonaceous reducing agent used in the reduction step, the limits are not-critical. The most important factor governing the maximum particle size is the desirability of the surface area of the particles being sufficiently high to give reasonable reaction efficiency. If the reaction mixture is to be exposed to a gas stream, the use of very small particles (which might of course be entrained in the stream) should be avoided. Material having a suitable particle size can readily be obtained by grinding.

When an excess of a solid reducing agent is used, the reduced material may be separated by magnetic techniques from the unconsumed reducing agent.

Reduction with a solid reductant may be effected in any suitable apparatus. For example, a rotary kiln may be employedj and may be operated either co-current or counter current.

Advantageously, the conditions in the reduction zone of the apparatus are such that the residence time of the material at the reduction temperature does not exceed 2 hours, and preferably does not exceed l£ hours. Especially good results are obtained if the residence time at the desired temperature is from 1 to 1½ hours.

By way of example, the temperature profile in a gas-fired kiln using coal as reductant may lie within the limits given in the following Table.

Time (minutes) 0 25 3 50 Temperature 500*70 700+50 800+60 900+70 Time 60 70 85 100 110 Temperature 910+50 900+50 850+40 800+30 750+20 In a process operated in accordance with the above Table, the reduction reaction will continue after elimination of pseudobrookite.

It will sometimes be found advantageous to conduct the reduction in two stages, a first stage at a temperature of up to about 850°C. followed by a second stage at 900-950°C. there being a .airly sharp rise in the temperature profile between the two stages.

There is a risk that a certain amount of re-oxidation will take place as a result of atmospheric air coming into contact with the reduced material while it is cooling, and it will generally be found advisable to take precautions to prevent any significant re-oxidation taking place, as such re-oxidation impairs the leaching characteristics of the material to some extent. For example, when a solid carbon- aceous reducing agent is used, the reduction product is most preferably cooled out of contact with the atmosphere before being separated from any unconsumed reducing agent* The material obtained by the process steps described hereinbefore is readily leachable and a wide variety of leaching techniques may therefore be employed. Typically, a single-stage treatment with a 25$ excess (over the stoichiometric quantity required in respect of the iron values only) , of 17½-20$ w/w hydrochloric acid for 3 to 4 hours at a temperature of from 100° to 108°G. leads to a product having a titanium dioxide content of at least 90$ with at least 95$ recovery of titanium values, and with substantially no degradation of particle size. Although other mineral acids may be used, hydrochloric acid is the preferred leaching agent, and may be recovered from the residual leach liquor by a thermal decomposition process, for example, by the process disclosed in British Patent Specification No. 793,700.

It will generally be found that the reduced material is sufficiently reactive in leaching that a single-stage leaching treatment at normal atmospheric pressure is all that is required. Accordingly, it will usually be unnecessary to resort to more severe leaching conditions and systems, for example, multi-stage leaching, continuous co-current, counter-current, or cross-current leaching systems, or the use of superatmospheric pressure. The reduced material will frequently be found to be sufficiently reactive that hydrochloric acid having a strength as low as 15$ will be adequate. Preferably, the strength of the hydrochloric acid does not exceed 20$ w/w.

It will generally be found that a single-stage acid leaching treatment (involving no replenishment of used acid) will not only remove almost all of the iron values, but will also remove significant quantities of impurity elements such as, for example, manganese, chromium, and vanadium. In any leaching operation in which the acid is not continually replenished, the strength of the leaching acid will decrease steadily as the leaching operation progresses* In this connection, it has been found that the rate of leaching at any given instant is a function not only of the prevailing acid strength at that instant, but also of the initial acid strength.

Advantageously, the reduced material is leached with an acidic solution that contains some ferrous iron initially, for example a solution of ferrous chloride in hydrochloric acid. In general, such a solution will produce similar results in terms of titanium dioxide yield and content to those obtained by using the same quantity of acid alone, bu will produce such results in a shorter time.

A solution containing ferrous iron may be obtained in a variety of ways. For example, if a quantity of leached material is washed with hydrochloric acid solution, the wash liquor will contain some ferrous iron in solution, and that wash liquor may thenbe used as leachant. Instead, the acid leaching solution may contain residual leach liquor. In such a process, the residual leach liquor will eventually contain such a high concentration of ferrous iron as to be unsuitable for recycle. If a hydrochloric acid leaching solution is employed, the acid may be recovered from the residual leach liquor by a thermal decomposition process, for example, by the process described in British Patent Specificati No. 793*700, and such a recovery process will be especially valuable when the residual leach liquor is unsuitable for direct recycle.

When a solid reducing agent is used, it is preferable to separate the reduction product from any unconsumed reducing agent before the leaching treatment is carried out. Advantageously the desired separation is effected by mesjis of a magnetic separation technique. Such a separation process will also reduce the quantity of any gangue material that may be present, for example, silica sand.

The beneficiated titaniferous material resulting from the leaching step does not break down to yield fine particles under fluidised-hed chlorination conditions, at least to any substantial extent, and the chlorination reaction can therefore produce a high yield. Accordingly, the present invention also provides a process for the production of titanium tetrachloride which comprises a forming from an ilmenite ore/material of which at least 20$£ by weight has a crystal structure of the pseudobrookite type, reducing the said material under such conditions that not less than 4$ of tb iron content of the reduced material (calculated as Fe) is in the ferric state, leaching the reduced material so obtained to yieLd a beneficiated titaniferous material, and chlorinating the beneficiated material under fluidised-bed reaction conditions.

The beneficiated material contains a very small proportion of material that is relatively resistant to leaching. Such material does tend to yield fines on breakdown, but has a crushing strength which is sufficiently high that the particles do not in practice break down to any substantial extent under the chlorination conditions o Desirably, the material to be chlorinated has a mean particle diameter of at least 100 microns, with substantially none of the particles having a diameter of less than 65 microns. It is a feature of the beneficiation process according to the present invention that the mean particle diameter of the beneficiated material is not substantially different from that of the ilmenite ore starting material. Accordingly, the particle size distribution of the ilmenite is advantageously . such that the mean weight diameter is at least 100 microns, and that there are substantially no particles having a diameter of less than 65 microns.

After being leached, the material may be calcined to a temperature of from 600° to 700°C. Before being calcined, the leached material will be washed to remove residual leach liquor.

Such calcination of the leached material enhances its reactivity (measured as described hereinafter) in subsequent chlorination under fluidised-bed reaction conditions.

The reactivity of the calcined material under fluidised-bed chlorination conditions is measured for the purposes of the present invention by evaluating the ratio CO/CO+COg) for the exhaust gases from the chlorinator. The value of this ratio provides a measure of the exothermicity of the chlorination reaction, and a low value of the ratio is taken to indicate a highly reactive material. It should be noted that, in practice, the value for the ratio is taken as the mean of a number of readings made during the chlorination process.

Typically, the value of the C0/(C0+C02) ratio obtained in chlorinating natural rutile is of the order of 0.20 -0.24* The value of the ratio for the product of the present invention may be as low as 0.25» or even lower in some cases, indicating that the reactivity of the material is approaching that of natural rutile, and may be of closely similar reactivity in some cases. The carbon monoxide and carbon dioxide contents of the exhaust gases from the chlorinator may be determined by gas-liquid chromatography.

The calcinatio may be effected in any suitable apparatus^ and may be continued for a period of from { to 2 hours, preferably for about hour.

Chlorination of the product may be effected, for example, in a fluidised-bed chlorinator having a diameter of 15 cm. and constructed of silica. The chlorine flow rate in such a reactor may be about 35 l/min* and the reaction temperature may be maintained at 925 + 25°C. with a bed having a depth of 1.3 m. By way of example, a leached beneficiate was calcined at 600°C. and then chlorinated under the conditions described in this paragraph. The CO/CCOj-COg) ratio in the exhaust gases was found to be 0.25» By comparison, when' a sample of the same leached beneficiate was chlorinated under the same conditions, but without being calcined at 600° - 700°C. prior to chlorination, the value for the ratio was 0.3 · A ratio of 0.3 was also obtained when, in another comparison experiment, the beneficiate was calcined at a temperature in excess of 800°C. before being chlorinated.

In the chlorination of material obtained in accordance with the present invention, the chlorine utilisation is generally of the order of 100 .

Calcination in accordance with the present invention is effective in thoroughly removing water from the leached beneficiate.

The invention also provides a reactive titaniferous material obtained by the process of the invention and suitable for use as a feed material in the production of titanium tetrachloride by fluidised-bed chlorination.

The invention further provides a beneficiated titaniferous material obtained by a beneficiation process according to the nv .

The beneficiated material has a prous structure, substantially all of the pores being below 2 microns in diameter* The structure is generally free from large cavities and it is believed that this contributes to the desirable behaviour of the material under fluidised-bed reaction conditbns .

The following Examples illustrate the invention: Example 1 2 Kg» of a sample of Western Australian ilmenite was brought up to 1000°C. and was roasted in an oxidising atmosphere in a gas-fired furnace at that temperature for a period of one hour. The ilmenite was contained in shallow trays and the co tents of the trays were raked at frequent intervals throughout the oxidising treatment. Initially, the ilmenite ore was found by wet analysis to contain 54«6?S titanium calculated as TiOg, 17«0?¾ ferric iron calculated as and 25·7?° ferrous iron calculated as FeO, each percentage being by weight. The particle size distribution of the ilmenite ore was such that 99-8 by weight of the material had a particle size in the range of from 75 to 200 microns, the mean weight diameter being 160 microns. After the oxidising roast, no FeO could be detected on wet analysis, and X-ray examination showed the structure of the oxidised material to comprise 65 by weight pseudobrookite with some rutile* 100 Gr. of the oxidised material was mixed intimately with 50 g. powdered anthracite, and the resulting mixture was maintained for 100 minutes in a gas-fired furnace at 1000°C. The mixture was contained in a fire-clay crucible and was prevented from coming into contact with the furnace atmosphere.

The reduction stage was repeated with two further 100 g. portions of the oxidised material, using 50 g. powdered non-caking high-volatile bituminous coal as the reducing agent for one portion, and 50 g. of pre-oxidised (200°C, ·½- hour) medium-volatile caking coal as the reducing agent for the other portion.

On completion of the reduction stages, the crucibles were allowed to cool with the contents out of contact with air, and the contents were then separated magnetically into ore and coal char fractions. Wet analysis of the three ore fractions showed that they each contained in excess of 35*0$ by weight ferrous iron calculated as PeO and about 4«0$ by weight ferric iron calculated as FegO^, that is to say, about 9$ of the iron content was in the ferric state. Bone of the ore fractions contained any detectable metallic iron; X-ray analysis showed that each ore fraction contained only ilmenite and rutile.

Each of the ore fractions was then leached in boilin 18$ w/w hydrochloric acid for 4 hours, the quantity of acid used in each case being 25$ in excess of the stoichiometric requirement calculated with respect to iron values only. The beneficiates so obtained were washed and dried, and the titanium dioxide content of the dried material, and the efficiency of recovery of titanium, were determined in each case.

The results obtained are set out in the following table.

TiO content Efficiency of Reducing Agent of beneficiate TiOg recovery anthracite 91·7$ 98$ high- olatile coal 92. $ 96$ medium-volatile coal 92.7$ 95$ The particle size of the beneficiated material was substantially the same as that of the initial ilmenite sample, and the grains showed a porou structure, generally free from ( large cavities.

Example 2 Western Australian ilmenite was fed to a gas-fired rotary kiln, Oo55m. in diameter and 7«3m» in length, at a rate of 45 kg. per hour, and was oxidised in the kiln at 900°C. by means of atmospheric air. The residence time of the ilmenite in the kiln was about 2 hours, and X-ray examination of the oxidised material showed it to comprise 35 by weight pseudo-brookite together with rutile and haematite.

The oxidised material was mixed with 30$ by weight of kibbled anthracite and heated in the kiln to 950°C. over a period of one hour. During this period, the oxygen content of the exit gases from the kiln varied between 0.5 and 1$ by volume. The material was then discharged from the kiln and cooled out of contact with the atmosphere. After cooling, the reduced material was . separated magnetically from the unconsumed anthracite. On wet analysis, the ore material was found to contain 38$ by weight ferrous iron calculated as FeO and about 2| by weight ferric iron calculated as E^O^, with no detectable metallic iron, that is to say, ·6$ of the iron content of the material was in the ferric state.

The reduced material was then leached under the same conditions as described in Example 1, and the resulting beneficiated material was found to contain 9 .6$ by weight of Ti02» The efficiency of recovery of titanium in the leaching step was 98$.

No large cavities could be detected within the grains of the beneficiated material, which had a porous structure.

Example "5 The process described in Example 1 was repeated using a quantity of Quilon ilmenite containing 60.3$ by iight of titanium (calculated as TiOg and using a quantity of Norw¾Lan ilmenite having a titanium content of 44·6 (calculated as Ti02) · Before being oxidised, the Norwegian ilmenite was sieved in order to remove particles smaller than 170 mesh (British Standard).

In each case, X-ray analysis of the oxidation product showed that it consisted predominantly of pseudobrookite with some rutile, and X-ray analysis of the reduction product showed the presence of ilmenite and a small proportion of rutile, with no pseudobrookite* The reducing agent used in each case was non-caking high-volatile bituminous coal, and wet analysis of the reduced ore material showed that its ferric iron content was 7«5 in the case of the Quilon ilmenite, and 7 » 6 ° in the case of the Norweigan ilmenite (each percentage being by weight and based on the total iron content of the material) . Neither of the reduction products contained any detectable metallic iron.

The titanium dioxide contents of the final beneficiatfis , and efficiencies of recovery of titanium dioxide in the leaching step, were as follows; Ilmenite iO^ content iO^ recovery Quilon 92 o5 95f° Norwegian 9 4 Although the crushing strengths of the beneficiated products were relatively low (about 6.4 x 10 g./cm ), breakdown of the material was not accompanied by the formation of fines, and the material was therefore suitable for chlorination under fluidised-bed reaction conditions. The beneficiated material had a porous structure, generally free from large cavities.

Example 4 A sample of Western Australian ilmenite was oxidised as described in Example 1. 100 G. of the oxidised material was reduced in a two- stage process as follows: 100 g. was mixed intimately with 50 g. powdered anthracite, and the resulting mixture was maintained at 850°C. for 40 minutes in a gas-fired furnace and then at 925°C. for a further 40 minutes. The mixture was contained in a fire-clay crucible and was prevented from coming into ^contact with the furnace atmosphere.

On completion of the reduction stage, the crucible was allowed to cool with the contents out of contact with air, and the contents were then separated into ore and coal char fractions.

The ore fraction was leached for 4 hours as described in Example 1, and the leached material was washed first with 20 w/w hydrochloric acid to yield a wash liquor comprising solution of ferrous iron in hydrochloric acid, and then with water. The washed material was calcined at 650°C. in shallow trays in a gas-fired furnace, thereby producing a dry material suitable for fluidieed-bed chlorination. The titanium dioxide content of the material was 91· 7$· An additional quantity of reduced material was prepared as described above, and was lee,ched using the ferrous iron-containing wash liquor that had been obtained by washing the previous sample with 20$ w/w hydrochloric acid. In this case, however, leaching was continued for 3 hours only, after which time the leached material was washed with water and calcined as before. It was found that the titanium dioxide content of the resulting beneficiate was 91«7 , demonstrating that the leaching efficiency of the solution of ferrous iron in hydrochloric acid was greater than that of hydrochloric acid alone.

In the claims which follow, the term "pseudobrookite" has the meaning assigned hereinbefore, that is to say, "material having a structure of the pseudobrookite cr stallographic type".

Claims (1)

1. · A process for beneficiating an ilmenite ore, which comprises forming from the ilmenite a material of which at least 20 by weight comprises pseudobrookite, reducing the without forming any substantial quantity of metallic iron and said material/under such conditions that not less than 4$ of the iron content of the reduced material (calculated as Pe) is in the ferric state, and leaching the reduced material so obtained to yield a beneficiated titaniferous material* 2. A process as claimed in claim 1, wherein the formation of pseudobrookite is effected by heating the ilmenite under oxidising conditions. , 3· A process as claimed in claim 2, wherein the formation of pseudobrookite is effected by heating the ilmenite in the presence of oxygen or an oxygen-containing 4· A process as claimed in claim 1, wherein the ilmenite is a weathered ilmenite and the formation of pseudobrookite is effected by heating the ilmenite under "substantially inert conditions" (as hereinbefore defined). 5» A process as claimed in any one of claims 1 to 4, wherein the formation of pseudobrookite is effected by maintaining the ilmenite at a temperature of at least 900°C. 6. A process as claimed in claim 5, wherein the temperature is at least 925°C. 7· A process as claimed in claim 6, wherein the temperature is 950°C. 8. A process as claimed in any one of claims 1 to 7, wherein the temperature of the ore material does not exceed 1000°C. during the formation of pseudobrookite therefrom. 9· A process as claimed in any one of claims 1 to 8, wherein the formation of pseudobrookite is effected in the presence of a compound of lithium, sodium, magnesium, zirconium, or zinc. 10· A process as claimed in claim 9» in which the said compound is a lithium compound- 11. A process as claimed in any one of claims 1 to 10, in which a material comprising at least 30$ by weight of pseudobrookite is formed from the ilmenite. 12. A process as claimed in claim 11, wherein the proportion of pseudobrookite is at least 50 by weight. 13· A process as claimed in> claim .12, wherein the proportion of pseudobrookite is in the range of from 60 to 70 by weight, and the pseudobrookite is formed under oxidising conditions. 14» A process as claimed in any one of claims 1 to 10, wherein the pseudobrookite content of the material formed from the ilmenite is substantially equal to the theoretical maximum corresponding to that ilmenite. 15· A process as claimed in any one of claims 1 to 14» wherein the pseudobrookite content of the reduced material does not exceed by weight. 16. A process as claimed in any one of claims 1 to 14» wherein substantially no pseudobrookite can be detected in the reduced material. 17» A process as claimed in claim 15 or claim 16, wherein the temperature profile in the reduction stage is such that the maximum temperature reached in that stage is attained during the elimination of the pseudobrookite. 18. A process as claimed in any one of claims 1 to 17» wherein the ilmenite is a Western Australian ilmenite and the reduction conditions are such that not more than ^ 18 "by weight of the iron content of the reduced material is in the ferric state* 19 · A process as claimed in claim 18, wherein from 8 to 12°/o "by weight of the iron content of the reduced material is in the ferric state. 20 » A process as claimed in any one of claims 1 to 17 » wherein the ilmenite is a Quiloii ilmenite and the reduction conditions are such that not more than 11$ by weight of the iron content of the reduced material is in the ferric state. 21. A process as claimed in> any one of claims 1 to 3» or in any one of* claims 5 to 17 except when dependent on claim 4 , wherein the ilmenite is a Norwegian ilmenite and the reduction conditions are such that not more than 405¾ by weight of the ferric iron content of the reduced material is in the ferric state. 22. A process as claimed in any one of claims 1 to 21 , wherein the reduction conditions are such that not more than lO o by weight of the iron content of the reduced material is in the ferric state. 2J . A process as claimed in claim 22 , wherein at least 8°/o by weight of the iron content of the reduced material is in the ferric state. 24 » A process as claimed in any one of claims 1 to 23 t wherein the temperature of the reduced material is quickly lowered when the desired extent of reduction has been achieved. 25 · A process as claimed in any one of claims 1 to 24 » wherein the pore size distribution in the reduced product is such that it has at least 8$ porosity in the size range 0·04 to 2 microns. 26» A process as claimed in any one of claims 1 to 25» wherein substantially no re-oxidation of the reduced material takes place. 27» A process as claimed in any one of claims 1 to 26, wherein the reduction reaction is carried out at a temperature in the range of from 800° to 1000°C. 28. A process as claimed in claim 27» wherei 'the reduction temperature is in the range of from 850° to 950°C. 29· A process as claimed in any one of claims 1 to 28» wherein the pseudobrookite-containing material is transferred to the reduction apparatus without any substantial cooling taking pl^ace*. 30. A process as claimed in any one of claims 1 to 29» wherein the reduction is effected using a gaseous reducing agent. 31» A process as claimed in claim 30, wherein the gaseous reducing agent comprises hydrogen. 3 » A process as claimed in claim 30, wherein the gaseous reducing agent comprises a hydrocarbon gas. 33· A process as claimed in claim 30, wherein the gaseous reducing agent comprises carbon monoxide. 34» A process as claimed in any one of claims 30 to 33, wherein the reduction is effected under fluidised-bed reaction conditions. 35» A process as claimed in claim 31 except when appended to claim 27 or claim 28, wherein the reduction reaction is carried out at a temperature of from 700° to 800°C. 36 o A process as claimed in any one of claims 1 to 29» wherein the reduction is effected by contacting the pseudobrookite-containing material at an elevated temperature with a solid carbonaceous reducing agent in the presence of volatile material. 37» A process as claimed in claim 36» wherein the volatile material is contained initially in the solid carbonaceous reducing agent* 38.. A process as claimed in claim 37» wherein the solid reducing agent comprises green coke, semi-coke, lignite, or reject chare 39· A process as claimed in claim 37» wherein the solid reducing agent comprises a coal that does not cake under the reduction conditions* 40· A process as claimed in any one of claims 36 to 39» wherein excess reducing agent is employed, and a magnetic separation technique is employed to isolate the reduced material from unconsumed solid reducing agent. 41· A process as claimed in claim 40, wherein the said separation is effected prior to leaching. 42. A process as claimed in any one of claims 36 to 41, wherein the conditions in the reduction zone of the reduction apparatus are such that the residence time of the material at the reduction temperature does not exceed 2 hours. 43» A process as claimed in claim 42, wherein the said residence time does not exceed 1½ hours. 44· A process as claimed in claim 43» wherein the residence time is in the range of from 1·£ to 1¾- hours. 45· A process as claimed in any one of claims 6 to 44, wherein the reduction is effected in a rotary kiln. 6· A process as claimed in claim 45» wherein the reducing agent is coal and the temperature profile in the kiln is within the limits set out in the Table herein. 47· A process as claimed in any one of claims 36 to 45 except when dependent on claim 28, wherein the reduction is effected in two stages : a first stage at a temperature of up to 850°C. followed immedia ely by a second stage at a temperature of 900° to 950°C. 48· A process as claimed in any one of claims 1 to 47 , wherein the reduced material is leached with hydrochloric acid* 49· A process as claimed in claim 48, wherein leaching is effected as s, single-stage process at a temperature of from 100° to 108°C. using a 25°/° excess of hydrochloric acid having a strength^ of from 17¾- to 20 w/w, the duration of the leaching process being from 3 to 4 hours- 50. A process as claimed in claim 48, wherein the strength of the acid is not in excess of 20?ό w/w. 51· A process as claimed in any one of claims 1 to 48 or claim 50, wherein the reduced material is leached with an acidic solution that contains some ferrous iron initially 52. A process as claimed in claim 51, wherein the leachant comprises the liquor obtained by washing with hydrochloric acid a quantity of previously leached materialo 53· A process as claimed in any one of claims 1 to 52, wherein the reduced material is acid leached in a single-stage treatment involving no replenishment of used acid. 54· A process as claimed in any one of claims 1 to 53? wherein the reduced material is leached with a hydrochloric acid solution, and the acid is recovered from the residual leach liquor by a thermal decomposition process and is recycled to the leaching stage* 55· A process for the beneficiation of an ilmenite ore,^ conducted substantially as described in any one of the Examples herein* 56 · A process as claimed in any one of claims 1 to 55» wherein the beneficiated titaniferous material resulting from the leaching step is calcined to a temperature of from 600° to 700°C. 57· A process as claimed in claim 56, wherein the calcination is continued for a period of from { to 2 hours o 58. A process as claimed in claim 57» wherein the calcination is continued for a period of 1 hour. 59· A process for the production of titanium tetrachloride, wherein- a titaniferous material prepared by a process as claimed in any one of claims 1 to 55 is chlorinated under fluidised-bed reaction conditions. 60. A process as claimed in claim 59» wherein, prior to being chlorinated, the titaniferous material is calcined as specified in any one of claims 56 to 58· 61. A process as claimed in claim 59» or claim 60, wherein the ilmenite ore starting material is a material having a mean weight particle diameter of at least 100 microns, and containing subs a tially no particles having a diameter of less than 65 microns. 62. A beneficiated titaniferous material whenever prepared by a process as claimed in any one of claims 1 to 55· 63· A reactive titaniferous material whenever prepared by a process as claimed in any one of claims 56 to 8· 64» Titanium tetrachloride whenever prepared by a process as claimed in any one of claims 59 to 61.