WO1995007366A1

WO1995007366A1 - Upgrading titaniferous materials

Info

Publication number: WO1995007366A1
Application number: PCT/AU1994/000528
Authority: WO
Inventors: Ross Alexander Mcclelland; Michael John Hollitt
Original assignee: Technological Resources Pty Ltd.
Priority date: 1993-09-07
Filing date: 1994-09-07
Publication date: 1995-03-16
Also published as: NO960917L; NO317932B1; EP0717783A1; ZA946864B; NO960917D0; EP0717783A4; CA2171185A1; CN1042349C; CN1134730A; JPH09504828A

Abstract

A process for upgrading a titaniferous material by removal of impurities from the titaniferous material is disclosed. The process comprises alternate leaching of the titaniferous material in a caustic leach and a pressure sulphuric acid leach.

Description

UPGRADING TITANIFEROUS MATERIALS

This invention relates to the removal of impurities from naturally occurring and synthetic titaniferous materials. The invention is particularly suited to the enhancement of titaniferous materials used in the production of titanium metal and titanium dioxide pigments by means of industrial chlorination systems.

Embodiments of the present invention have the common features of the use of caustic leaching and pressure sulphuric acid leaching for the upgrading of titaniferous materials, e.g. titaniferous slags, derived from hard rock ilmenites. Additional steps may be employed as will be described below.

In industrial chlorination processes titanium dioxide bearing feedstocks are fed with coke to chlorinators of various designs (fluidised bed, shaft, molten salt), operated to a maximum temperature in the range 700 - 1200°C. The most common type of industrial chlorinator is of the fluidised bed design. Gaseous chlorine is passed through the titania and carbon bearing charge, converting titanium dioxide to titanium tetrachloride gas, which is then removed in the exit gas stream and condensed to liquid titanium tetrachloride for further purification and processing.

The chlorination process as conducted in industrial chlorinators is well suited to the conversion of pure titanium dioxide feedstocks to titanium tetrachloride. However, most other inputs (i.e. impurities in feedstocks) cause difficulties which greatly complicate either the chlorination process itself or the subsequent stages of condensation and purification and disposal of waste. The attached table provides an indication of the types of problems encountered. In addition, each unit of inputs which does not enter products contributes substantially to the generation of wastes for treatment and disposal. Some inputs (e.g. particular metals, radioactives) result in waste classifications which may require specialist disposal in monitored repositories.

Preferred inputs to chlorination are therefore high grade materials, with the mineral rutile (at 95-96% Ti0₂) the most suitable of present feeds. Shortages of rutile have led to the development of other feedstocks formed by upgrading naturally occurring ilmenite (at 40-60% Ti0₂), such as titaniferous slag (approximately 86% Ti0₂) and synthetic rutile (variously 92-95% Ti0₂) . These upgrading processes have had iron removal as a primary focus, but have extended to removal of magnesium, manganese and alkali earth impurities, as well as some aluminium.

Elemental Chlorination Condensation Purification Input

Fe, Mn Cons me sSolid/liquid chlorine, chlorides c o k e ,f o u 1 increasesductwork, gas volumes make sludges

Alkali & Def luidise a 1 k a 1 i fluid beds e a r t h d e t o metals l i q u i d chlorides , c o n s u m e chlorine, coke

Al ConsumesC a u s e s C a u s e s chlorine, corrosion corrosion, makes coke sludges

Si Accumulates Can May require i ne ncourage distillat ion chlorinator, d u c t from product r e d u c i n gb lockage . c a m p a i g n Condenses in life. part with

Consumes t i t anium c o e , tetrachloride chlorine

Must be removed, by chemical treatment and distillation

Th, Ra Accumulates i n chlorinator brickwork , radioactive; c a u s e s disposal difficulties

In the prior art synthetic rutile has been formed from titaniferous minerals, e.g. ilmenite, via various 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°C. In this process the iron content of the mineral is substantially metallised. Sulphur additions are also made to convert manganese impurities partially 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. There is no substantial chemical removal of alkali metals or alkaline earths, aluminium, silicon, vanadium or radionuclides in this process as disclosed or operated. Further, iron and manganese removal is incomplete.

Recent disclosures have provided a process which operates reduction at lower temperatures and provides for hydrochloric acid leaching after the aqueous aeration and iron oxide separation steps. According to these disclosures the process is effective in removing iron, manganese, alkali and alkaline earth impurities, a substantial proportion of aluminium inputs and some vanadium as well as thorium. The process may be operated as a retrofit on existing kiln based installations. However, the process is ineffective in full vanadium removal and has little chemical impact on silicon.

In another prior art invention relatively high degrees of removal of magnesium, manganese, iron and aluminium have been achieved. In one such process ilmenite is first thermally reduced to substantially complete reduction of its ferric oxide content (i.e. without substantial metallisation), normally in a rotary kiln. The cooled, reduced product is then leached under 35 psi pressure at 140-150°C 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 product is then subjected to atmospheric leaching with excess 20% hydrochloric acid, for removal of the deleterious impurities. Acid regeneration is also performed by spray roasting in this process.

In all of the above mentioned hydrochloric acid leaching based processes impurity removal is similar. Vanadium, aluminium and silicon removal is not fully effective.

In yet another process ilmenite is thermally reduced (without metallisation) with carbon in a rotary kiln, followed by cooling in a non-oxidising atmosphere. The cooled, reduced product is leached under 20 - 30 psi gauge pressure at 130°C with 10 - 60% (typically 18 - 25%) 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 been claimed for this process. Under such circumstances similar impurity removal to that achieved with other hydrochloric acid based systems is to be expected. Where sulphuric acid is used radioactivity removal will not be complete.

A commonly adopted method for upgrading of ilmenite to higher grade products is to smelt ilmenite at temperatures in excess of 1500°C with coke addition in an electric furnace, producing a molten titaniferous slag (for casting and crushing) and a pig iron product. Of the problem impurities only iron is removed in this manner, and then only incompletely as a result of compositional limitations of the process.

In another process titaniferous ore is roasted with alkali metal compounds, followed by leaching with a strong acid other than sulphuric acid (Australian Patent No. AU-B- 70976/87). According to this disclosure substantial removal of various impurities is achieved, with "substantial '' defined to mean greater than 10%. In the context of the present invention such poor removal of impurities, especially of thorium and uranium, would not represent an effective process. No specific phase structure after roasting is indicated for this process but it is evident from analytical results provided (where product analyses, unlike feed analyses do not sum to 100% and analyses for the alkali metal added are not given) that there may have been significant retention of the additive in the final product. Under the conditions given it is herein disclosed that it is to be expected that alkali ferric titanate compounds which are not amenable to subsequent acid leaching will form. The consequent retention of alkali will render the final product unsuitable as a feedstock for the chloride pigment process.

In yet another process a titaniferous ore is treated by alternate leaching with an aqueous solution of alkali metal compound and an aqueous solution of a non-sulphuric mineral acid (US Patent No. 5,085,837). The process is specifically limited to ores and concentrates and does not contemplate prior processing aimed at artificially altering phase structures. Consequently the process requires the application of excessive reagent and harsh processing conditions to be even partially effective and is unlikely to be economically implemented to produce a feedstock for the chloride pigment process.

A wide range of potential feedstocks is available for upgrading to high titania content materials suited to chlorination. Examples of primary titania sources which cannot be satisfactorily upgraded by prior art processes for the purposes of production of a material suited to chlorination include hard rock (non detrital) ilmenites, siliceous leucoxenes, many primary (unweathered) ilmenites and large anatase resources. Many such secondary sources (e.g. titania bearing slags) also exist.

In particular, for titaniferous materials containing elevated levels of silica, alumina and magnesia, such as titaniferous slags derived from hard rock ilmenite sources, none of the previously disclosed upgrading methods is effective for the production of a feedstock for the commercial chloride pigment processing route. The combination of silica which cannot be removed economically by the previously identified techniques and alumina and magnesia which together assist in the formation during thermal processing of pseudobrookite - anosovite type phases which are not amenable to leaching with hydrochloric acid under commercially realistic conditions limits the use of such materials to sulphate pigment process feedstocks. Since the pigment process expected to supply all growth in pigment demand is the chloride process such a limitation is a severe constraint.

A large portion of the world's identified titania reserves is in the form of hard rock ilmenites.

Clearly there is a considerable incentive to discover methods for upgrading of such titaniferous materials which can economically produce high grade products which are suitable as feedstocks to the chloride pigment process.

The present invention provides a combination of processing steps which may be incorporated into more general processes for the upgrading of titaniferous materials, rendering such processes applicable to the treatment of a wider range of feeds and producing higher quality products than would otherwise be achievable.

Accordingly, the present invention provides a process for upgrading a titaniferous material by removal of impurities, which process involves alternately leaching of the material in a caustic leach and a pressure sulphuric acid leach.

In a particular embodiment the present invention ensures that caustic leaching can be conducted economically and effectively despite the need for the use of excess caustic in the leach by circulation of caustic leach liquors after solid/liquid separation through a caustic regeneration step using lime addition to precipitate complex aluminosilicates and regenerate caustic solution. The complex aluminosilicates are then separated from the regenerated caustic solution which is recycled to the leach.

The treatment of titaniferous materials containing both alumina and silica in such a manner has not previously been disclosed, and it is herein revealed that only under specific operating conditions can such a process be operated without precipitation of complex aluminosilicates in the caustic leach.

It has been surprisingly discovered that by limiting the concentrations of silica, alumina, titania and other impurities in caustic leach liquor, i.e. by leaching at low slurry densities and recirculating leach liquors through caustic regeneration, the complex aluminosilicates otherwise formed in the caustic leach can frequently be avoided.

It has also been surprisingly discovered that complex aluminosilicates formed in the caustic leach can actually be removed in the subsequent acid leach along with other impurities. This is a particularly surprising outcome as under most circumstances silica in titaniferous materials cannot be removed by acid leaching.

Consequently, in a further embodiment it is possible to operate a simple process involving a two stage treatment in which complex aluminosilicates are formed in a first stage and consumed by acid leaching in a second stage, wherein silica removal is achieved in the acid leaching stage along with the other benefits of acid leaching in more general upgrading.

In particular it is revealed that the ease of formation in caustic leaching and removal in acid leaching of complex aluminosilicates depends on the caustic to silica ratio in the leach liquor (which determines whether the aluminosilicates are of the sodalite type or in another form) , with high caustic to silica ratios allowing greater ease of removal. Thus, the circulation of caustic leach liquors through a caustic leach and caustic regeneration by lime (which keeps the caustic to silica ratio high) followed by pressure sulphuric acid leaching is under many circumstances a most effective means of upgrading titaniferous materials, especially titaniferous materials derived from hard rock ilmenite.

It has been discovered that the process of the invention can remove iron, magnesium, aluminium, silicon, calcium, magnesium, manganese, phosphorus, chromium and vanadium, which impurities form an almost comprehensive list of impurities in hard rock ilmenite sources of titania.

Additional steps may be incorporated in the process, as desired. For example:

(1) The titaniferous material may be roasted in any suitable device and to any temperature under reducing or oxidising conditions prior to leaching. Such roasting may be conducted in order to enhance the response of the material to the leaching steps or to reduce the production of sulphur dioxide in the leach by oxidation of any trivalent titania in the titaniferous material.

(2) Additives may be made to the titaniferous material prior to such a roasting step in order to enhance the response of the material to the leaching steps, or for any other purpose.

(3) The titaniferous material may be preground prior to roasting or leaching in order to enhance reaction rates or in preparation for agglomeration steps which are improved by generation of a broad particle size distribution in the material to be agglomerated.

(4) An agglomeration step via which additives are incorporated into the titaniferous material prior to roasting may be operated.

(5) Physical separation of material (e.g. magnetic separation of final product in order to selectively remove and recycle iron rich material) for further upgrading.

(6) The final titaniferous product may be agglomerated 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 removing from dusting losses in fluidised bed chlorinators.

(7) Irrespective of final product agglomeration the final product may be calcined in order to remove volatile matter (e.g. water, sulphur dioxide and sulphur trioxide) .

(8) A caustic solution bleed or caustic solution evaporation step (for wash water removal) may be operated.

(9) The sulphuric acid leach exit liquor may be neutralised to produce solid sulphates and hydroxides for disposal.

(10) The sulphuric acid leach exit liquor may be treated for regeneration of sulphuric acid from the aqueous sulphate solutions formed in the process.

(11) Other leach steps, filtration steps and washing steps may be incorporated into the process as desired. For example, a hydrochloric acid leach may; be conducted to assist in the removal of trace levels of radioactivity. Pressure filtration of the complex aluminosilicate precipitated in caustic recovery may be operated to assist solid/liquid separation.

(12) Flocculants and other aids may be used to assist solid/liquid separation.

Examples

The following examples describe a number of laboratory tests which serve to illustrate the techniques described herein.

Example 1

This example is to demonstrate the ineffectiveness of treatments found to be effective for upgrading other titaniferous materials on materials such as titaniferous slags produced from hard rock ilmenites.

Commercial titaniferous slag having the composition indicated in Table 1 was subjected to oxidation roasting in air at 750°C for 30 minutes, followed by reduction roasting in a 1:3 hydrogen to carbon dioxide (volumetric basis) gas mixture at 680°C for one hour. The cooled product of this thermal treatment contained no ferric iron and no trivalent titania. The phase composition of the material was indicated by X-ray diffraction as pseudobrookite.

The thermally treated material was leached in refluxing 10% caustic soda solution at 10% slurry density. After filtration and washing the solid residue had a composition as indicated in Table 2.

It is clear that caustic leach had no appreciable effect on the silica or alumina contents of the material.

The residue of the caustic leach was subjected to a leach with refluxing 20% hydrochloric acid at 30% slurry density for 6 hours. After filtration and washing the solid residue had the composition which is also indicated in Table 2.

Clearly a roast/leach process using 10% caustic soda at 10% slurry density and 20% hydrochloric acid at 30% slurry density as leachants is almost totally ineffective in upgrading the slag.

Example 2

The treatment indicated in Example 1 was repeated with the exception that the caustic leach was conducted under pressure at 165°C.

The compositions of the caustic and acid leached products are indicated in Table 3. It is clear that the caustic leach had no appreciable effect on the silica or alumina contents of the material. The acid leach, however, despite being largely ineffective in producing an upgrade which might be suitable for the chloride pigment process did have a substantial effect on the silica and alumina contents. There was no such effect on a sample of slag submitted directly to hydrochloric acid leaching.

Clearly the pressure caustic leach had altered the state of the silica to allow its subsequent removal in hydrochloric acid leaching but had not resulted in direct removal. Investigations revealed the production of a complex aluminosilicate precipitate in the caustic leach. The caustic leach had been conducted under conditions in which silica could be leached but was not soluble.

The results of this example combined with the results of subsequent examples in which effective caustic leaching is demonstrated illustrate the dependence of the removal of silica and alumina in caustic leaching on the leach conditions.

Example 3

A sample of the slag whose composition is indicated in Table 1 was mixed with 2% borax, formed into pellets and subjected to reduction roasting in a 19:1 hydrogen to carbon dioxide (volumetric basis) gas mixture at 1000°C for 2 hours. The phase composition of the cooled product of this thermal treatment was indicated by X-ray diffraction as pseudobrookite.

A sample of the thermally treated material was leached in refluxing 10% caustic soda solution at 5% slurry density. After filtration and washing the solid residue had a composition as indicated in Table 4. It is clear that the caustic leach was highly effective in the removal of silica, despite the much poorer performance of a leach conducted at 10% slurry density in Example 1, in which complex aluminosilicates were formed.

The residue of the caustic leach was subjected to a pressure leach at 150°C with 20% sulphuric acid at 5% slurry density for 6 hours. After filtration and washing the solid residue had the composition which is also indicated in Table 4.

Clearly the combined effects of a low slurry density caustic leach and a subsequent pressure sulphuric acid leach (which is capable of decomposing pseudobrookite) were to substantially upgrade the slag to a very high grade product which is suitable in composition as a chloride pigment process feedstock.

The leach liquor from the above caustic leach was preserved and after analysis was treated with micronised lime at the weight ratio of 1.3 units of lime per unit of dissolved silica. The resulting complex aluminosilicate precipitate and any excess lime were removed by filtration and the "regenerated" caustic solution was preserved for reuse in leaching.

A further sample of the thermally treated material was leached with the regenerated caustic solution under the same conditions as indicated above. There was no difference of any consequence between the results of the leach with fresh caustic and the results of the leach with regenerated caustic.

Example 4

This example is to demonstrate the ineffectiveness of acid leaching alone in the removal of silica from titaniferous materials such as titaniferous slags produced from hard rock ilmenites.

Commercial titaniferous slag having the composition shown in Table 1 was subjected to roasting for two hours in an atmosphere of 1:19 (volumetric basis) of hydrogen to carbon dioxide at 1000°C. After cooling in the roasting atmosphere the roasted slag was pressure leached at 135°C in 20% sulphuric acid at 25% w/w slurry density for six hours.

The composition of the leach residue is given in Table 5. Such direct acid leach treatment of a roasted titaniferous material may be anticipated to result in little improvement of product quality by leaching, and no removal of Si0₂.

Example 5 A sample of slag to which no addition of additive had been made and which was not subjected to any thermal treatment was treated by the same leaching steps as indicated in Example 3.

The composition of the final product was as recorded in Table 6. Substantial removal of impurities have been achieved without thermal treatment.

Table 1: Composition of Titaniferous Slag

Used In Example 1 - 4

wt%

Ti0₂ 78.9

FeO 8.94

MgO 4.73

MnO 0.25

Cr₂0₃ 0.16

V₂0₅ 0.56

A1₂0₃ 3.14

Si0₂ 2.71

Zr0₂ 0.05

CaO 0.42

Table 2: Composition of Products in Example 1

wt% Caustic Leach Acid Leach

Ti0₂ 78.6 80.8

FeO 9.22 7.4

MgO 4.71 4.69

MnO 0.24 0.23

Cr₂0₃ 0.16 0.16 v₂o₅ 0.59 0.59

A1₂0₃ 3.09 3.06

Si0₂ 2.94 2.86

Zr0₂ 0.05 0.04

CaO 0.37 0.16 Table 3: Composition of Products in Example 2

wt% Caustic Leach Acid Leach

Ti0₂ 78.4 82.7

FeO 9.13 7.66

MgO 4.76 4.81

MnO 0.25 0.23

Cr₂0₃ 0.16 0.16 v₂o₅ 0.58 0.60

A1₂0₃ 3.08 2.73

Si0₂ 3.13 0.96

Zr0₂ 0.05 0.04

CaO 0.40 0.13

Table 4: Composition of Products in Example 3

wt% Caustic Leach Acid Leach

Ti0₂ 81.3 97.9

FeO 9.56 0.89

MgO 4.96 0.44

MnO 0.27 0.02

Cr₂0₃ 0.20 0.12 v₂o₅ 0.57 0.12

A1₂0₃ 1.75 0.23

Si0₂ 0.73 0.09

Zr0₂ 0.05 0.06

CaO 0.45 0.003 Table 5: Composition of Product in Example 4

wt% Acid Leach

Ti0₂ 84.93

FeO 6.09

MgO 2.92

MnO 0.16

Cr₂0₃ 0.16

V₂0₅ 0.60

Al₂0₃ 1.33

Si0₂ 3.15

Zr0₂ 0.06

CaO 0.03

Table 6: Composition of Product in Example 5.

wt% Acid Leach

Ti0₂ 92.1

FeO 2.98

MgO 1.21

MnO 0.08

Cr₂0₃ 0.16

V₂0₅ 0.18

A1₂0₃ 0.60

Si0₂ 0.71

Zr0₂ 0.06

CaO 0.003

Claims

CLAIMS :

1. A process for upgrading a titaniferous material by removal of impurities from the titaniferous material, which process comprises alternately leaching of the material in a caustic leach and a pressure sulphuric acid leach.

2. The process defined in claim 1, wherein the caustic leach is carried out before the acid leach.

3. The process defined in claim 1, wherein the acid leach is carried out before the caustic leach.

4. The process defined in any one of the preceding claims, wherein the acid leach is carried out at a temperature of at least 100°C.

5. The process defined in claim 4, wherein the acid leach is carried out at a temperature of at least 135°C.

6. The process defined in any one of the preceding claims, which comprises controlling the conditions of the caustic leach to precipitate complex aluminosilicates in a controlled and desired manner.

7. The process defined in any one of claims 1 to 5, which comprises carrying out the caustic leach under conditions of low slurry density to avoid precipitation of complex aluminosilicates.

8. The process defined in claim 7, wherein the slurry density is less than 10%.

9. The process defined in claim 7 or claim 8, which further comprises separating the leach residue from the caustic leach liquor and regenerating the caustic leach liquor by precipitating complex aluminosilicates.

10. The process defined in claim 9, which further comprises separating the regenerated caustic leach liquor from the precipitated complex aluminosilicates and recycling the regenerated leach liquor to the caustic leach.

11. The process defined in any one of the preceding claims, wherein the titaniferous material comprises a titaniferous slag derived from hard rock ilmenites.

12. A process for upgrading a titaniferous material derived from hard rock ilmenites by removal of impurities from the material, which process comprises:

(a) leaching the material in a caustic leach under conditions selected to precipitate complex aluminosilicates;

(b) separating the caustic leach residue from the caustic leach liquor; and

(c) leaching the caustic leach residue from step (b) in a pressure sulphuric acid leach to remove the complex aluminosilicates and other impurities to form the upgraded material.

13. A process for upgrading a titaniferous material derived from hard rock ilmenites by removal of impurities from the material, which process comprises:

(a) leaching the material in a caustic leach;

(b) separating the caustic leach residue from the caustic leach liquor; (c) leaching the caustic leach residue from step (b) in a pressure sulphuric acid leach to remove the complex aluminosilicates and other impurities to form the upgraded material;

(d) regenerating the caustic leach liquor from step (b) by precipitating complex aluminosilicates to form a regenerated caustic leach liquor suited to the effectiveness of steps (a) and (c); and

(e) recycling the regenerated caustic leach liquor to step (a) .