PROCESS OF PROVIDING TITANIUM DIOXIDE AND/OR VANADIUM OXIDE

Description

The present invention relates to a process of providing titanium dioxide and/or vanadium oxide from feedstocks comprising minerals and/or slags comprising titanium and/or vanadium.

BACKGROUND TO THE INVENTION

There are abundant reserves of minerals and waste stockpiles which contain valuable constituents that cannot currently be extracted in an economically viable manner. The main reason for this is that the grade of such constituents within feedstocks is too low, resulting in large effluent or by-product generation rates.

Current methods of producing titanium dioxide (e.g. titanium dioxide pigment) are based on two well-established processes: the sulphate and the chloride processes (Zhang, W., Z. Zhu, and C. Y. Cheng, A literature review of titanium metallurgical processes. Hydrometallurgy, 2011. 108(3-4): p. 177-188).

The older sulphate process involves solubilising ilmenite or high grade titanium (>70 wt % TiO₂) slag by dissolving it in concentrated sulphuric acid; obtaining pure TiO₂ by hydrolysis of liquors containing the solubilised TiO₂ (Braun, J. H., A. Baidins, and R. E. Marganski, TiO2 pigment technology: a review. Progress in organic coatings, 1992. 20(2): p. 105-138). The major disadvantage of the sulphate process is that it is only established on two materials (ilmenite and high grade slag). In addition, the consumption of acid is very high for this process.

The newer chloride process requires high-grade TiO₂ feedstocks to be commercially viable. The process involves fluidising feedstock such as ilmenite, slag, synthetic rutile or natural rutile at high temperatures in a stream of chlorine gas to produce a vapour mix of chlorides, including TiCl₄. Selective condensation is utilised to separate TiCl₄ from the impurity chlorides and oxidation of the TiCl₄ subsequently leads to pure TiO₂ by contacting with oxygen at high temperatures. Chlorine gas is recovered in the oxidation treatment (Braun, J. H., A. Baidins, and R. E. Marganski, TiO2pigment technology: a review. Progress in organic coatings, 1992. 20(2): p. 105-138; Mackey, T. S., Upgrading ilmenite into a high-grade synthetic rutile. Jom, 1994. 46(4): p. 59-64; and Rosenbaum, J. B., Titanium technology trends. JOM, 1982. 34(6): p. 76-80).). The major disadvantage of the chloride process is that it does not work on feedstocks which are high in calcium and magnesium oxides due to formation of liquid chlorides. As a result, a prior process of removing magnesium and calcium oxides is applied before chlorination. However, the prior process of removing calcium and magnesium is lengthy and costly.

The range of feedstock suitable for the chloride or sulphate process is presented in Table 1. It can be observed from Table 1 that the sulphate process is more suitable for ilmenite material and high grade TiO₂ slag material.

TABLE 1
Feedstock used in the chloride and sulphate processes (Kogel, J. E.,
Industrial minerals & rocks: commodities, markets, and uses. 2006: SME).

Feedstock	Type	Process	TiO2 (weight %)
Chloride-grade ilmenite	Natural	Chloride	~60
Leucoxene	Natural	Chloride	75-91
Rutile	Natural	Chloride	90-96
Chloride-grade slag	Synthetic	Chloride	85-95
Synthetic rutile	Synthetic	Chloride	90-96
Sulphate-grade ilmenite	Synthetic	Sulphate	44-57
Sulphate-grade slag	Synthetic	Sulphate	70-80

The UpGraded Slag (UGS) process was developed by Rio Tinto plc to remove impurities such as oxides of iron, silica, magnesium and calcium from the sulphate-grade slag (Sorelslag™) to produce a feedstock containing a minimum of 94.5 wt. % TiO₂ which is suitable for the chloride process (Kogel, J. E., Industrial minerals & rocks: commodities, markets, and uses. 2006: SME; Filippou, D. and G. Hudon, Iron removal and recovery in the titanium dioxide feedstock and pigment industries. JOM, 2009. 61(10): p. 36-42; U.S. Pat. Nos. 5,830,420 A; 6,531,110 B1). According to this process, titanium-containing slag is firstly oxidised in air to convert Ti³⁺ to Ti⁴⁺, Fe²⁺ to Fe³⁺ and migrate most of the iron to the periphery of the particles. The air-roasted sample is re-roasted in the presence of a reducing agent such as carbon monoxide to metallise iron. The reductive roasted sample is then leached in hydrochloric acid media to solubilise iron and part of the magnesium, calcium and aluminium. Nonetheless, from the two types of feedstock that were used, titanium dioxide in the material was only upgraded from 33 to 54.2 wt % and 56 to 67.9 wt % thereby representing an upgrade factor of 1.64 and 1.18, respectively. More than 98 wt % of the calcium oxide was removed by this process. However, more than 70 wt % of the magnesium oxide remained in the solid or was unrecovered via the leaching process. In summary, the major disadvantage of the UGS process is that it requires two stages of high temperature roasting (both above 800 degrees Celsius). Titanium does not upgrade to higher than 75 wt % in the final product even after caustic soda leaching. The hydrochloric acid leaching step requires more than 6 hours.

Another known method of extracting titanium from ilmenite mineral is the Becher process (Filippou, D. and G. Hudon, Iron removal and recovery in the titanium dioxide feedstock and pigment industries. JOM, 2009. 61(10): p. 36-42; Becher, R., et al. A new process for upgrading ilmenitic mineral sands, AUSTRALASIAN INST MINING MET PROC. 1965). According to this process, ilmenite concentrates are said to be upgraded into synthetic rutile product containing 92-95 wt % TiO₂. The process involves preferential oxidation of iron to hematite in a rotary kiln at a temperature above 1,000 degrees Celsius. The oxidised iron migrates to the periphery of the particles and is reduced in the presence of coal in the rotary kiln, again at a temperature above 1000 degrees Celsius. The calcined reduced samples are leached in acidic media to dissolve iron thereby leaving upgraded titanium dioxide. One disadvantage of this process is that it is unable to separate vanadium that is associated with titanium. Secondly, the temperature of the two calcinations makes the process expensive. Thirdly, the process does not work on low grade feedstock.

A known modified Becher process involves the use of colemanite, ulexite or borax (U.S. Pat. No. 5,578,109 A) as an additive to remove radionuclides present in titaniferous materials such as ilmenite, reduced ilmenite, altered ilmenite or synthetic rutile to produce synthetic rutile enhanced product. Similarly, this process is unable to recover vanadium from titanium-containing feedstock.

Another known method of recovering titanium and vanadium from ilmenite and low-grade feedstocks is the Leeds University process. The process, as described in U.S. Pat. No. 10,508,320 B2, involves mixing the feedstock with alkaline (e.g. sodium carbonate, sodium hydrogen carbonate, potassium carbonate, potassium hydrogen carbonate etc.) and heating in the presence of a reducing agent. Iron is metallised during the reductive roasting process while sodium forms sodium titanate and water-soluble sodium vanadate. The roast calcine is leached in water to dissolve vanadium and aluminium. The water leach residue is leached in an acidic environment to dissolve iron and sodium oxide. The acid leach residue is baked with sulphuric acid to sulphate titanium which is later leached out and hydrolysed. Even though this process effectively simultaneously recovers both titanium and vanadium, it has the following disadvantages:

• • a) The acid leach stage consumes a lot of acid due to the presence of alkaline salts (sodium titanate, sodium aluminate, sodium chromate etc.). The feed to acid leach step has a pH of more than 9.0;
  • b) The presence of high alkaline salts in the acid solution makes filtration extremely difficult and nearly impossible at times;
  • c) The process has two different leach stages (water and acidic leaching) between the roasting and baking (curing) stages thereby making the capital cost high as leaching tanks are required for both of these leaching stages;
  • d) The presence of coal in the roasting stage increases the operating cost;
  • e) The high addition of alkaline salts not only increases capital and operating costs but also promotes ring formation during the roasting stage.

All other processes that have been developed to recover both titanium and vanadium from titanium-vanadium feedstocks involve roasting in alkaline salt (sodium carbonate, sodium hydrogen carbonate, potassium carbonate, potassium hydrogen carbonate, sodium sulphate etc.). Vanadium is recovered via leaching in alkaline media. The alkaline leach residue is leached in acidic environment to remove iron and other constituents. The acid leach residue is then digested or cured with acid to extract titanium. Disadvantages of these processes include:

• • a) High acid consumption, because most of the excess alkaline that remains after alkaline leaching consumes part of the acid. In general, acid consumption is more than 1,000 kg per ton of feedstock;
  • b) The process has two leaching steps (alkaline and acidic leaching) prior to titanium digestion or curing. The two leaching steps increase capital and running costs;
  • c) Poor filtration, especially during the acid leaching step, due to e.g. the presence of alkaline salts.

It is one object of the present invention to overcome at least some of the disadvantages of the prior art or to provide a commercially useful alternative thereto.

It is a further object of the present invention to provide a process which requires less acid consumption for equivalent or improved titanium and/or vanadium recovery, compared to known processes.

It is a further object of the present invention to provide a more economical titanium and/or vanadium recovery process, compared to known processes.

It is a further object of the present invention to provide a financially more viable titanium and/or vanadium recovery process for lower grade feedstocks, compared to known processes.

It is a further object of the present invention to provide a titanium and/or vanadium recovery process with comparable or increase percentage recoveries, compared to known processes.

It is a further object of the present invention to provide a titanium and/or vanadium recovery process which allows for improved filtration, compared to known processes.

SUMMARY OF THE INVENTION

In a first aspect the present invention provides a process of providing titanium dioxide and/or vanadium oxide, the process comprising:

• • (a) providing a feedstock, the feedstock comprising at least one of
    • (I) a mineral comprising titanium and/or vanadium, and
    • (II) a slag comprising titanium and/or vanadium,
  • (b) adding an alkaline material to the feedstock to provide a modified feedstock, wherein the alkaline material has a pH of less than 9.0,
  • (c) roasting at least a portion of the modified feedstock at a temperature in the range of 600° C. to 900° C. to provide a roasted calcine,
  • (d) optionally leaching at least a portion of the roasted calcine under acidic conditions to provide a pre-cure leach solid and a pre-cure leach liquor,
  • (e) optionally obtaining vanadium oxide from at least a portion of the pre-cure leach liquor via solvent extraction and/or precipitation,
  • (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid to provide a first acidic solid;
  • (g) curing at least a portion of the first acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured first acidic solid,
  • (h) leaching at least a portion of the cured first acidic solid to provide a post-cure leach solid and a post-cure leach liquor,
  • (i) hydrolysing at least a portion of the post-cure leach liquor to provide titanium dioxide, and/or
  • (j) obtaining vanadium oxide from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation.

The present inventors have surprisingly found that the present process allows improved titanium and/or vanadium recovery from feedstocks and/or requires less acid and/or allows easy filtration, compared to known methods. Furthermore, the use of an alkaline material having a pH of less than 9.0 advantageously permits direct acid leaching of roasted calcines. This is in contrast to known methods which use other alkaline salts such as (sodium carbonate, sodium hydrogen carbonate, potassium carbonate, potassium hydrogen carbonate, sodium sulphate etc.), which all have a pH of more than 12.0.

Each aspect or embodiment as defined herein may be combined with any other aspect(s) or embodiment(s) unless clearly indicated to the contrary. In particular any feature indicated as being preferred or advantageous may be combined with any other feature or features indicated as being preferred or advantageous.

Other preferred embodiments of the compounds according to the invention appear throughout the specification and in particular in the examples.

DETAILED DESCRIPTION OF THE INVENTION

Unless otherwise defined herein, scientific and technical terms used in connection with the present invention shall have the meanings that are commonly understood by those of ordinary skill in the art. The meaning and scope of the terms should be clear, however, in the event of any latent ambiguity, definitions provided herein take precedent over any dictionary or extrinsic definition.

As used herein, the terms ‘roasting’ and ‘calcining’ are equivalent. Roasting involves heating at a temperature in the range of 600° C. to 900° C.

As used herein, the terms ‘curing’ and ‘baking’ are equivalent. Curing involves heating at a temperature in the range of 100° C. to 300° C. Curing may be carried out in a rotary drier, for example.

As is known and understood in the art, the term ‘borax’ refers to disodium; 3,7-dioxido-2,4,6,8,9-pentaoxa-1,3,5,7-tetraborabicyclo[3.3.1]nonane;decahydrate, i.e. a hydrate salt of boric acid. The term ‘borax’ is synonymous with the terms sodium borate, sodium borate decahydrate, and sodium tetraborate dechydrate, for example. Borax is commercially available, commonly in powder or granular form. Borax is understood to be weakly alkaline with a pH less than 8.0.

As is known and understood in the art, the term ‘slag’ refers to material (usually waste material) separated from metals during the smelting or refining or minerals/ore.

Unless indicated to the contrary, percentages are by weight.

The present invention provides a process of providing titanium dioxide and/or vanadium oxide, the process comprising:

• • (a) providing a feedstock, the feedstock comprising at least one of
    • (I) a mineral comprising titanium and/or vanadium, and
    • (II) a slag comprising titanium and/or vanadium,
  • (b) adding an alkaline material to the feedstock to provide a modified feedstock, wherein the alkaline material has a pH of less than 9.0,
  • (c) roasting at least a portion of the modified feedstock at a temperature in the range of 600° C. to 900° C. to provide a roasted calcine,
  • (d) optionally leaching at least a portion of the roasted calcine under acidic conditions to provide a pre-cure leach solid and a pre-cure leach liquor,
  • (e) optionally obtaining vanadium oxide from at least a portion of the pre-cure leach liquor via solvent extraction and/or precipitation,
  • (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid to provide a first acidic solid;
  • (g) curing at least a portion of the first acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured first acidic solid,
  • (h) leaching at least a portion of the cured first acidic solid to provide a post-cure leach solid and a post-cure leach liquor,
  • (i) hydrolysing at least a portion of the post-cure leach liquor to provide titanium dioxide, and/or
  • (j) obtaining vanadium oxide from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation.

It is understood that when the feedstock (as (I) a mineral or (II) a slag) comprises titanium, titanium as an element may be in any form. In particular, the titanium may be present in the feedstock in the form of one or more titanium-containing compounds (e.g. titanium dioxide) and/or elemental titanium and/or one or more titanium-containing alloys. For example, in one preferable embodiment, the majority of the titanium in the feedstock is in the form of titanium dioxide.

It is also understood that when the feedstock (as (I) a mineral or (II) a slag) comprises vanadium, vanadium as an element may be in any form. In particular, the vanadium may be present in the feedstock in the form of one or more vanadium-containing compounds (e.g. vanadium oxide) and/or elemental vanadium and/or one or more vanadium-containing alloys. For example, in one preferable embodiment, the majority of the vanadium in the feedstock is in the form of elemental vanadium.

Preferably, the feedstock comprises 0.3 to 5.0 wt % vanadium and/or 5.0 to 85.0 wt % titanium dioxide, based on the total weight of the feedstock. In this context, when the feedstock comprises 0.3 to 5.0 wt % vanadium, the weight percentage refers to the weight percentage in the feedstock of vanadium as an element, which may be in the form of one or more vanadium-containing compounds (e.g. vanadium oxide) and/or elemental vanadium and/or one or more vanadium-containing alloys.

Preferably, the feedstock comprises a slag comprising titanium and/or vanadium, or the feedstock comprises one or more of the group consisting of ilmenite, leucoxene, rutile, and complex calcium magnesium aluminium silicates.

Preferably, the feedstock is subjected to a magnetic separation step before the alkaline material is added in step (b). Such a magnetic separation step is thought to remove at least a portion of the magnetic components present in the feedstock.

Preferably, the feedstock is provided in step (a) in crushed and/or milled form. For example, preferably, the feedstock is provided in step (a) in the form of a powder. More preferably, the feedstock is provided in step (a) in the form a powder, wherein the average diameter of the feedstock particles is less than 500 μm, more preferably less than 300 μm. Commercially available screen sizers may be used to separate out feedstock particles having diameters in specified ranges for further processing.

Preferably, the alkaline material has a pH of less than 8.0. Without wishing to be bound by theory, it is thought that using an alkaline material with a pH of less than 8.0 allows further reduced acid consumption and easier filtration. The roasted calcine produced is also more suitable for direct acid leaching.

Preferably, the alkaline material comprises an alkaline salt. More preferably, the alkaline material consists essentially of, or consists of an alkaline salt.

Preferably, the alkaline material comprises or is selected from borax and/or potassium borate. More preferably, the alkaline material comprises borax. More preferably still, the alkaline material consists essentially of borax, or consists of borax. It is understood that borax is weakly alkaline and has a pH of less than 8.0. Using borax and/or potassium borate is thought to reduce acid consumption and allow easy filtration, as well as permit direct acid leaching of the roasted calcine.

Preferably, the alkaline material is added to the feedstock in an amount of 3 to 20 wt %, based on the total weight of the feedstock. More preferably, the alkaline material is added to the feedstock in an amount of 5 to 15 wt %, based on the total weight of the feedstock. Most preferably, the alkaline material is added to the feedstock in an amount of 10 to 15 wt %, based on the total weight of the feedstock. Without wishing to be bound by theory, it is thought that the consumption of sulphuric acid increases with the amount of alkaline material, e.g. borax, and the recovery of vanadium and/or titanium increases with the amount of alkaline material, e.g. borax.

Preferably, the alkaline material comprises, consists essentially of, or consists of borax, and is added to the feedstock in an amount of 3 to 20 wt %, based on the total weight of the feedstock. More preferably, the alkaline material comprises, consists essentially of, or consists of borax, and is added to the feedstock in an amount of 5 to 15 wt %, based on the total weight of the feedstock. Most preferably, the alkaline material comprises, consists essentially of, or consists of borax, and is added to the feedstock in an amount of 10 to 15 wt %, based on the total weight of the feedstock.

The process comprises (c) roasting at least a portion of the modified feedstock at a temperature in the range of 600° C. to 900° C. to provide a roasted calcine. Preferably, the roasting in step (c) is at a temperature in the range of 600° C. to 800° C. More preferably, the roasting in step (c) is at a temperature in the range of 650° C. to 800° C., or from 700° C. to 800° C. Without wishing to be bound by theory, it is thought that the percentage vanadium recovery and/or the titanium recovery increases with increasing temperature and that consumption of sulphuric acid increases sharply when samples are roasted above 800° C.

Preferably, the roasting in step (c) is for 0.5 hours (30 minutes) to 4 hours. More preferably, the roasting in step (c) is for 1 hour to 4 hours, more preferably for more than 1 hour to 4 hours, or 1.5 hours to 4 hours. Without wishing to be bound by theory, it is thought that roasting for 1 hour or 1.5 hours to 4 hours improves percentage recovery of titanium and/or vanadium.

Preferably, the roasting in step (c) is in an oxidizing atmosphere. For example, the roasting may be in air. Preferably, the roasting is in the absence of coal.

Preferably, the alkaline material comprises, consists essentially of, or consists of borax, and is added to the feedstock in an amount of 5 to 15 wt %, more preferably 8 or 9 or 10 to 15 wt %; and the roasting in step (c) is at a temperature in the range of 700° C. to 800° C., more preferably 700° C. to 800° C. or 750° C. to 790° C.; and the roasting in step (c) is in an oxidizing atmosphere, preferably in the absence of coal. More preferably, the alkaline material comprises, consists essentially of, or consists of borax, and is added to the feedstock in an amount of 10 to 15 wt %; and the roasting in step (c) is at a temperature in the range of 750° C. to 790° C.; and the roasting in step (c) is in an oxidizing atmosphere.

The process comprises (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid to provide a first acidic solid. Preferably, the process comprises adding sulphuric acid to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid to provide a first acidic solid.

Preferably, step (f) further comprises adding water to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid. More preferably, step (f) comprises adding water and sulphuric acid to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid to provide a first acidic solid.

In one preferable embodiment, the process comprises

• • (d) leaching at least a portion of the roasted calcine under acidic conditions to provide a pre-cure leach solid and a pre-cure leach liquor; and
  • (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the pre-cure leach solid to provide a first acidic solid.

More preferably, the process comprises

• • (d) leaching at least a portion of the roasted calcine under acidic conditions to provide a pre-cure leach solid and a pre-cure leach liquor; and
  • (f) adding water and sulphuric acid and/or hydrochloric acid to at least a portion of the pre-cure leach solid to provide a first acidic solid.

Leaching is a technique that is known and well understood in the art. Generally, a solid containing one or more entities of interest (e.g. TiO₂ and/or vanadium) is contacted, e.g. at least partially submerged, in a fluid, e.g. water or an acid. The entities of interest may leach out of the solid into the fluid, from which they may be recovered.

Preferably, the leaching of the at least a portion of the roasted calcine in step (d) is at a pH of 1.0 to 3.0, preferably 1.0 to 2.0. More preferably, the leaching of the at least a portion of the roasted calcine in step (d) is at a pH of 1.0 to 1.8, or 1.2 to 1.6. Most preferably, the leaching of the at least a portion of the roasted calcine in step (d) is at a pH of 1.3 to 1.5. Without wishing to be bound by theory, it is thought that leaching the at least a portion of the roasted calcine in step (d) at a pH of 1.0 to 3.0, 1.0 to 2.0, 1.0 to 1.8, 1.2 to 1.6, or 1.3 to 1.5, leads to optimum recovery of vanadium, optionally as a function of acid consumption.

Preferably, the leaching of the at least a portion of the roasted calcine in step (d) is at a temperature in the range of 15° C. to 25° C., more preferably in the range of 18° C. to 22° C., more preferably at about room temperature (20° C.).

Preferably, the process comprises separating the pre-cure leach solid and the pre-cure leach liquor, preferably via filtration. Techniques for separating solids and liquids after leaching are known and well understood in the art, e.g. conventional filtration.

Preferably, the at least a portion of the roasted calcine prepared in step (c) is not subjected to a water or alkaline leaching step before subsequent process steps. More preferably, when the process comprises (d) leaching at least a portion of the roasted calcine under acidic conditions to provide a pre-cure leach solid and a pre-cure leach liquor, the at least a portion of the roasted calcine prepared in step (c) is not subjected to a water or alkaline leaching step before acidic leaching step (d).

Preferably, the process comprises (e) obtaining vanadium oxide from at least a portion of the pre-cure leach liquor via solvent extraction and/or precipitation. Preferably, the feedstock further comprises aluminium, and the process further comprises obtaining aluminium oxide from at least a portion of the pre-cure leach liquor via precipitation. Preferably, the feedstock further comprises magnesium, and the process further comprises obtaining magnesium oxide from at least a portion of the pre-cure leach liquor via precipitation. Preferably, the feedstock further comprises aluminium and magnesium, and the process further comprises obtaining aluminium oxide and magnesium oxide, in addition to the vanadium oxide, from at least a portion of the pre-cure leach liquor via precipitation.

Preferably, the process further comprises pelletizing the first acidic solid before curing step (g). Pelletizing of solids is known in the art and the skilled person is aware of commercially available equipment that may be used to pelletize the first acidic solid, e.g. a pelletizer. Preferably, the first acidic solid is pelletized into pellets having an average diameter in the range of 1 to 80 mm, more preferably in the range of 5 to 50 mm. Commercially available screen sizers may be used to separate out pellets having diameters in specified ranges for further processing.

The process comprises (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid to provide a first acidic solid; and (g) curing at least a portion of the first acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured first acidic solid. Without wishing to be bound by theory, it is thought that adding sulphuric acid and/or hydrochloric acid to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid to provide a first acidic solid and curing at least a portion of the first acidic solid at a temperature in the range of 100° C. to 300° C. leads to increased overall recovery of vanadium.

Preferably, the curing in step (g) is at a temperature in the range of 100° C. to 250° C., more preferably from 160 to 220° C.

Preferably, the curing in step (g) is for 15 minutes to 5 hours, more preferably for 30 minutes to 3 hours.

Preferably, the process further comprises pulverizing the at least a portion of the cured first acidic solid before leaching step (h). More preferably, the at least a portion of the cured first acidic solid is pulverized into particles having an average diameter of less than 1 mm, preferably 0.01 to 0.9 mm, or 0.03 to 0.7 mm, and more preferably 0.06 mm to 0.5 mm. Pulverizing techniques are known in the art and commercially available equipment may be used. It is understood that pulverizing includes crushing, for example. Pulverizing/crushing may be carried out, for example, by using a road mill, a roller crusher, and/or a ball mill. Commercially available screen sizers may be used to separate out pulverized particles having diameters in specified ranges for further processing. Preferably, the generation of fines in any pulverizing step is substantially avoided, more preferably completely avoided. Without wishing to be bound by theory, it is thought that pulverizing the cured first acidic solid before leaching enables better liberation of vanadium and/or TiO₂ from the cured first acidic solid.

The process comprises (h) leaching at least a portion of the cured first acidic solid to provide a post-cure leach solid and a post-cure leach liquor. Preferably, the leaching in step (h) is in water or sulphuric acid or hydrochloric acid. More preferably, the leaching in step (h) is in water or sulphuric acid. More preferably, the leaching in step (h) is in water. Alternatively, preferably, the leaching in step (h) is in sulphuric acid.

Preferably, the leaching of the at least a portion of the cured first acidic solid in step (h) is at a pH of less than 3.0. More preferably, the leaching of the at least a portion of the cured first acidic solid in step (h) is at a pH of less than 2.0.

Preferably, the leaching of the at least a portion of the cured first acidic solid in step (h) is at a temperature in the range of 15° C. to 25° C., more preferably in the range of 18° C. to 22° C., more preferably at about room temperature (20° C.).

Preferably, the process comprises separating the post-cure leach solid and the post-cure leach liquor, preferably via filtration.

Preferably, the process comprises (i) hydrolysing at least a portion of the post-cure leach liquor to provide titanium dioxide. Hydrolysis as a general concept is known and well understood in the art. For example, hydrolysis to precipitate titanium dioxide from at least a portion of the post-cure leach liquor may be carried out by heating the pregnant solution at a temperature of equal to or greater than 50° C. Preferably, the hydrolysing at least a portion of the post-cure leach liquor in step (i) is at a temperature in the range of 50° C. to 80° C. More preferably, the hydrolysing at least a portion of the post-cure leach liquor in step (i) is at a temperature in the range of 50° C. to 70° C. Without wishing to be bound by theory, hydrolysis at a temperature in the range of 50° C. to 70° C. is thought to lead to titanium dioxide particles having an average diameter of more than 20 μm (e.g. 20 μm to 60 μm), which is thought to be advantageous for any subsequent chlorination steps, in which fine particles can cause problems.

Preferably, the process comprises (j) obtaining vanadium oxide from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation. It is understood that vanadium oxide may be obtained from at least a portion of the post-cure leach liquor after titanium has been extracted from the post-cure leach liquor. Techniques to obtain vanadium oxide via solvent extraction and/or precipitation are known and well understood in the art. For example, the aqueous solution remaining after hydrolysis of titanium may be fed to the solvent extraction process. The initial pH of the liquor may be determined and adjusted via addition of e.g. lime or sodium hydroxide or ammonia solution. An organic solution may be prepared at a strength of about 5%-40% v/v (extractant to diluent), before contacting this organic solution with the aqueous solution for about 10-15 minutes in a ratio of 1:2 to aid the efficient separation of the loaded organic and raffinate. The loaded organic (solvent) may be subjected to a two-stage scrubbing process using about 2 gpl of H₂SO₄ for 10-15 minutes to extract metal ions (e.g. Fe, Al, Ni and Cr) from the loaded organic, and the raffinate will contain the desired metal ions (vanadium).

Preferably, the feedstock further comprises manganese, iron, chromium, and/or nickel, and the process preferably further comprises obtaining manganese, iron, chromium, and/or nickel from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation. It is understood that manganese, iron, chromium, and/or nickel may be obtained from at least a portion of the post-cure leach liquor after titanium and/or vanadium have been extracted from the post-cure leach liquor. Manganese, iron, chromium, and/or nickel, if present, may be extracted from the loaded organic solvent referred to above in relation to the extraction of vanadium.

Preferably, the feedstock further comprises manganese, iron, chromium, and nickel, and the process preferably further comprises obtaining manganese, iron, chromium, and nickel from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation; and the process preferably further comprises melting the obtained manganese, iron, chromium, and nickel to provide an alloy comprising manganese, iron, chromium, and nickel.

Preferably, the process is a process of providing titanium dioxide and vanadium oxide, and the process comprises

• • (i) hydrolysing at least a portion of the post-cure leach liquor to provide titanium dioxide, and
  • (j) obtaining vanadium oxide from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation.

Preferably, the overall process uses less than or equal to 1,000 kg acid per ton of feedstock. More preferably, the overall process uses less than or equal to 900 kg acid per ton of feedstock, or less than or equal to 800 kg acid per ton of feedstock, or less than or equal to 750 kg acid per ton of feedstock. Most preferably, the overall process uses less than or equal to 700 kg acid per ton of feedstock.

Preferably, the process further comprises

• • (f′) adding sulphuric acid and/or hydrochloric acid to at least a portion of the post-cure leach solid to provide a second acidic solid;
  • (g′) curing at least a portion of the second acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured second acidic solid;
  • (h′) leaching at least a portion of the cured second acidic solid to provide a post-cure second leach solid and a post-cure second leach liquor;
  • (i′) hydrolysing at least a portion of the post-cure second leach liquor to provide titanium dioxide, and/or
  • (j′) obtaining vanadium oxide from at least a portion of the post-cure second leach liquor via solvent extraction and/or precipitation.

Without wishing to be bound by theory, it is believed that a higher % recovery of vanadium and/or titanium dioxide can be achieved when the process further comprises steps (f′) to (j′), particularly when the process does not comprise leaching step (d).

Preferably, step (f′) further comprises adding water to at least a portion of the post-cure leach solid. More preferably, step (f′) comprises adding water and sulphuric acid to at least a portion of the post-cure leach solid to provide a second acidic solid.

Preferably, the process further comprises pelletizing the second acidic solid before curing step (g′). Preferably, the second acidic solid is pelletized into pellets having an average diameter in the range of 1 to 80 mm, more preferably in the range of 5 to 50 mm.

Preferably, the curing in step (g′) is at a temperature in the range of 150° C. to 300° C. More preferably, the curing in step (g′) is at a temperature higher than the temperature of the curing in step (g). Preferably, the curing in step (g′) is for 15 minutes to 5 hours, more preferably for 30 minutes to 3 hours.

Preferably, at least a portion of the cured second acidic solid is pulverized before leaching step (h′), preferably into particles having an average diameter of less than 1 mm, preferably 0.01 to 0.9 mm, or 0.03 to 0.7 mm, and more preferably 0.06 mm to 0.5 mm.

Preferably, the leaching in step (h′) is in water or sulphuric acid or hydrochloric acid. More preferably, the leaching in step (h′) is in water or sulphuric acid. More preferably, the leaching in step (h′) is in water. Alternatively, preferably, the leaching in step (h′) is in sulphuric acid.

Preferably, the leaching of the at least a portion of the cured second acidic solid in step (h′) is at a pH of less than 3.0, more preferably, at a pH of less than 2.0.

Preferably, the leaching of the at least a portion of the cured second acidic solid in step (h′) is at a temperature in the range of 15° C. to 25° C., more preferably in the range of 18° C. to 22° C., more preferably at about room temperature (20° C.).

In one embodiment, the process comprises separating the post-cure second leach solid and the post-cure second leach liquor, preferably via filtration.

In one embodiment, the process comprises (i′) hydrolysing at least a portion of the post-cure second leach liquor to provide titanium dioxide. Preferably, the hydrolysing at least a portion of the post-cure second leach liquor in step (i′) is at a temperature in the range of 50° C. to 80° C., more preferably in the range of 50° C. to 70° C.

In one embodiment, the process comprises (j′) obtaining vanadium oxide from at least a portion of the post-cure second leach liquor via solvent extraction and/or precipitation. It is understood that vanadium oxide may be obtained from at least a portion of the post-cure second leach liquor after titanium has been extracted from the post-cure second leach liquor.

In one embodiment, the process is a process of providing titanium dioxide and vanadium oxide, and the process further comprises

• • (f′) adding sulphuric acid and/or hydrochloric acid to at least a portion of the post-cure leach solid to provide a second acidic solid;
  • (g′) curing at least a portion of the second acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured second acidic solid;
  • (h′) leaching at least a portion of the cured second acidic solid to provide a post-cure second leach solid and a post-cure second leach liquor;
  • (i′) hydrolysing at least a portion of the post-cure second leach liquor to provide titanium dioxide, and
  • (j′) obtaining vanadium oxide from at least a portion of the post-cure second leach liquor via solvent extraction and/or precipitation.

In one embodiment, which may be combined with any of the preferred embodiments described above (where applicable), the process is a process of providing titanium dioxide and/or vanadium oxide, the process comprising:

• • (a) providing a feedstock, the feedstock comprising at least one of
    • (I) a mineral comprising titanium and/or vanadium, and
    • (II) a slag comprising titanium and/or vanadium,
  • (b) adding an alkaline material to the feedstock to provide a modified feedstock, wherein the alkaline material has a pH of less than 9.0,
  • (c) roasting at least a portion of the modified feedstock at a temperature in the range of 600° C. to 900° C. to provide a roasted calcine,
  • (d) leaching at least a portion of the roasted calcine under acidic conditions to provide a pre-cure leach solid and a pre-cure leach liquor,
  • (e) optionally obtaining vanadium oxide from at least a portion of the pre-cure leach liquor via solvent extraction and/or precipitation,
  • (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the pre-cure leach solid to provide a first acidic solid;
  • (g) curing at least a portion of the first acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured first acidic solid,
  • (h) leaching at least a portion of the cured first acidic solid to provide a post-cure leach solid and a post-cure leach liquor,
  • (i) hydrolysing at least a portion of the post-cure leach liquor to provide titanium dioxide, and/or
  • (j) obtaining vanadium oxide from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation.

In a further embodiment, which may be combined with any of the preferred embodiments described above (where applicable), the process is a process of providing titanium dioxide and/or vanadium oxide, the process comprising:

• • (a) providing a feedstock, the feedstock comprising at least one of
    • (I) a mineral comprising titanium and/or vanadium, and
    • (II) a slag comprising titanium and/or vanadium,
  • (b) adding an alkaline material to the feedstock to provide a modified feedstock, wherein the alkaline material has a pH of less than 9.0,
  • (c) roasting at least a portion of the modified feedstock at a temperature in the range of 600° C. to 900° C. to provide a roasted calcine,
  • (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the roasted calcine to provide a first acidic solid;
  • (g) curing at least a portion of the first acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured first acidic solid,
  • (h) leaching at least a portion of the cured first acidic solid to provide a post-cure leach solid and a post-cure leach liquor,
  • (i) hydrolysing at least a portion of the post-cure leach liquor to provide titanium dioxide, and/or
  • (j) obtaining vanadium oxide from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation;
  • (f′) adding sulphuric acid and/or hydrochloric acid to at least a portion of the post-cure leach solid to provide a second acidic solid;
  • (g′) curing at least a portion of the second acidic solid at a temperature in the range of 100° C. to 300° C. to provide a cured second acidic solid;
  • (h′) leaching at least a portion of the cured second acidic solid to provide a post-cure second leach solid and a post-cure second leach liquor;
  • (i′) hydrolysing at least a portion of the post-cure second leach liquor to provide titanium dioxide, and/or
  • (j′) obtaining vanadium oxide from at least a portion of the post-cure second leach liquor via solvent extraction and/or precipitation.

When introducing elements of the present disclosure or the preferred embodiments(s) thereof, the articles “a”, “an”, “the” and “said” are intended to mean that there are one or more of the elements. The terms “comprising”, “including” and “having” are intended to be inclusive and mean that there may be additional elements other than the listed elements.

The foregoing detailed description has been provided by way of explanation and illustration, and is not intended to limit the scope of the appended claims. Many variations in the presently preferred embodiments illustrated herein will be apparent to one of ordinary skill in the art, and remain within the scope of the appended claims and their equivalents.

These and other aspects of the invention will now be described with reference to the accompanying Figures, in which:

FIGS. 1 and 2 are scanning electron microscope images of Material A under backscattered imaging.

FIG. 3 is a plot of percentage recovery for vanadium during the initial acid leach step, as a function of wt % borax.

FIG. 4 is a plot showing overall percentage recovery for vanadium for Example 1 as a function of wt % borax.

FIG. 5 is a plot showing overall recovery for vanadium for Example 2 as a function of wt % borax.

FIG. 6 is a plot showing the % recovery for TiO₂ following the acid curing steps for Example 1.

FIG. 7 is a plot showing the % recovery for TiO₂ following the acid curing steps for Example 2.

FIG. 8 is a plot of sulphuric acid consumption per ton of material against roasting temperature in degrees Celsius (° C.).

FIG. 9 is a plot of percentage recovery against acid consumption.

FIG. 10 is a scanning electron microscope image of precipitated TiO₂ via hydrolysis at a lower temperature (less than 70° C.).

FIG. 11 is a scanning electron microscope image of precipitated TiO₂ via hydrolysis at a temperature of 70-80° C.

FIG. 12 is a scanning electron microscope image of precipitated TiO₂ via hydrolysis at a temperature of >80° C., i.e. 90-100° C.

FIG. 13 is a photo of precipitated TiO₂ via hydrolysis at 70-80° C.

EXPERIMENTAL SECTION

Test Methods

pH is measured by a pH meter. Suitable pH meters are commercially available.

Diameters and average diameters of particles are measured by techniques known in the art, dependent on the particle size. For example, the average diameter of Titanium dioxide particles is measured by Scanning Electron Microscopy (SEM). Scanning electron microscope images were obtained using a commercially available scanning electron microscope using backscattered imaging.

Chemical analysis of leach liquors in parts per million (ppm) was performed using known ICP-OES techniques. Examples of analysers that may be used are well known in the art.

Chemical analysis of feedstocks can be performed using known XRF (X-ray fluorescence) and/or ICP-OES techniques.

Chemical Analysis and Mineralogy of Slag Feedstocks

A slag feedstock (Material A) was used for all of the Examples below. Chemical analysis (by XRF and ICP-OES) of this feedstock is shown in Table 2.

TABLE 2
Chemical analysis of Material A

Sample	TiO2	V2O5	Na2O	CaO	SiO2	MgO	Al2O3	Fe	S	Mn	Cr
ID	(wt %)	(wt %)	(wt %)	(wt %)	(wt %)	(wt %)	(wt %)	(wt %)	(wt %)	(wt %)	(wt %)

Material	29.53	0.88	5.84	15.57	20.09	10.57	12.62	2.74	0.48	0.49	0.11
A

For mineralogical examination, a representative portion of Material A was cast in epoxy and re-cast in epoxy for transverse section exposure of the sample block for analysis using a scanning electron microscope. Mineralogical data of Material A is shown in Table 3 from which the following observations can be made:

• • The main phase is the complex Ca—Mg—Al—Si—Ti—V—O, which appears to contain most of the titanium and vanadium;
  • Rutile, which is one of the titanium-bearing minerals, accounts for 15% of the total weight of the material. Based on this, about 50% of the titanium exists in the rutile form while the rest of the titanium is present in the complex calcium magnesium aluminium silicate.

TABLE 3
Mineralogical data for Material A

	Mineral Distribution List	% Weight

	Ca—Mg—Al—Si—Ti—V—O Phase	63.34
	Rutile	15.2
	Al—Ca—Mg—O—V—Ti—Cr—Fe Phases	6.73
	Perovskite	4.18
	Fe-oxides (Hematite)	2.24
	Ca—Si—Mg—Al—O low Ti Phases	2.18
	Rutile-Ca—Al—Mg Phases	1.48
	Olivine	0.98
	Ca—Mg—O—Ti Phases	0.62
	Kassite-Mg Al Si Phases	0.50
	Ilmenite	0.36
	Calcite	0.20
	Quartz	0.14
	Pyrrhotite	0.08
	Kassite	0.07
	Dolomite	0.07
	Chlorite	0.05
	Olivine-Ca—Ti Phases	0.04
	Anhydrite	0.03
	Braunite	0.03
	Others	0.03
	Biotite	0.03
	Actinolite	0.02
	Kaolinite	0.01
	Muscovite	0.01
	K-feldspar	0.01
	[Unclassified]	1.33
	OTHERS	0.03

Scanning electron microscope images of Material A under backscattered imaging are presented in FIGS. 1 and 2. FIGS. 1 and 2 show that rutile and perovskite are embedded in the complex Ca—Mg—Al—Si—Ti—V—O matrix.

EXAMPLES

Example 1

A titanium and vanadium-containing slag feedstock (Material A) was mixed with 5-20 weight % borax and roasted (calcined) at different temperatures in an oxidising atmosphere, in the range of 600-790 degrees Celsius. The roast calcines were leached at room temperature (i.e. about 20° C.) in sulphuric acid media (pH of around 1.8). Up to 55 wt % of the vanadium present in Material A was recovered into solution during leaching. FIG. 3 is a plot of percentage recovery for vanadium during the initial acid leach step. The materials were roasted at 700, 750 and 790° C. with varying wt % amounts of borax in air (in the absence of coal) and leached in sulphuric acid media (pH of around 1.8). The recovery of vanadium was dependent on roasting temperature, as shown by FIG. 3.

Example 2

Example 1 was repeated but roasting was undertaken in a reductive atmosphere and hence coal was added during the roasting step. The roast calcine was leached in acidic environment and the recovery of vanadium into leach solution was similar to the recovery in example one. Therefore, it was concluded that coal has no effect on increasing the recovery of vanadium.

The acid leach residues from Examples 1 and 2 were admixed with water and 400-600 kg sulphuric acid per ton of acid leach residue and cured for 30 minutes to 3 hours at a temperature of between 160° C. to 250° C. The cured materials were highly acidic. They were leached out in water (without addition of acid). Most of the residual vanadium that remained after the previous steps was recovered into the leach solution. Titanium oxysulphate that had formed during curing was also recovered during this leaching step.

A summary of results showing the overall recoveries of vanadium and titanium oxide (V and TiO₂) is shown in Table 4. Further information is provided in FIGS. 4 to 8.

TABLE 4
Results showing overall recovery of vanadium and titanium
dioxide after acid curing the leach residue (i.e.
the overall recovery from the first acid leaching
step and the later acid curing and leaching steps).

Type of experiment	% Recovery of Vanadium	% Recovery TiO2
Oxidative roasting	74-92	64-91.7
(Example 1)
Reductive roasting	70-83	68-80
(Example 2)

FIG. 4 is a plot showing overall percentage recovery for vanadium for Example 1 as a function of wt % borax (the materials were roasted at 700, 750 and 790° C. with varying amounts of borax in the absence of coal and leached in acid media at a pH of around 1.8. Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached). A comparison of FIGS. 3 and 4 shows that mixing the leach residue with water and sulphuric acid followed by curing and leaching increases the overall recovery of vanadium.

FIG. 5 is a plot showing overall recovery for vanadium for Example 2 as a function of wt % borax (the materials were roasted at 700, 750, and 790° C. respectively, with varying amounts of borax in the presence of coal at a coal/borax ratio of 0.8. The roasted samples were leached in acid media at pH of around 1.8. Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached.

FIG. 6 is a plot showing the % recovery for TiO₂ as a function of wt % borax following the acid curing steps for Example 1. The materials were roasted at 700, 750 and 790° C. with varying amounts of borax in air (in the absence of coal). The roasted samples were leached in acid media (pH of around 1.8). Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached.

FIG. 7 is a plot showing the % recovery for TiO₂ as a function of wt % borax following the acid curing steps for Example 2. The materials were roasted at 700, 750 and 790° C. in the presence of coal at a coal/borax ratio of 0.8. The roasted samples were leached in acid media (pH of around 1.8). Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached.

Results showing sulphuric acid consumption in the first acid leaching step (step (d)) against roasting temperature for the first acid leaching step, are shown in FIG. 8. FIG. 8 is a plot of sulphuric acid consumption per ton of feedstock against roasting temperature in degrees Celsius (° C.). FIG. 8 shows that acid consumption is sensitive to both quantity of borax and roasting temperature. In essence, consumption of sulphuric acid increases with increase in roasting temperature and amount of borax. It can be seen that consumption of sulphuric acid generally increases when samples are roasted above 800° C. Therefore, it is thought to be advantageous to roast samples below 800° C.

Effect of pH During Leaching of Roasted Calcine Samples (Derived from Material A)

The effect of pH during leaching of the roasted calcine samples in sulphuric acid media was studied by leaching the samples at different pH and the results are shown in Table 6. It is evident from the results in Table 6 that acid consumption increases with decrease in leaching pH and this is thought to be due to decreased selectivity at lower pH or higher acidity. Nonetheless, the highest recovery of vanadium was achieved at a pH of 1.4 and the weight loss was higher than at any other pH.

TABLE 6
Effect of pH during leaching of the roasted
calcine samples in sulphuric acid media

	% Recovery		Acid consumption
Leaching pH	Vanadium	% Weight loss	(kg/ton ore)
0.8	26.78	16.00	664
1.2	22.30	13.40	233
1.4	39.73	17.90	220
1.6	21.40	14.00	182
1.8	23.38	15.10	171
2.2	19.13	12.10	159

Effect of Amount of Acid Addition

The effect of acid addition per ton of material was studied because acid, in this case sulphuric acid, is one of the major costs in TiO₂ production. The results showing the dependency of recoveries of TiO₂ and vanadium on acid consumption are shown in FIG. 9. FIG. 9 is a plot of percentage recovery against acid consumption. The roasted calcine samples were leached in sulphuric acid at a pH of 1.8, and sulphuric acid was added to the residue. This was subsequently cured at a temperature of 160° C. to 250° C.

FIG. 9 shows that recoveries for both TiO₂ and vanadium are dependent on acid consumption. The recoveries increase as sulphuric acid addition is increased, thought to be due to increased sulphation of TiO₂ and vanadium. FIG. 9 also shows that increasing acid consumption per ton to more than 700 kg per ton of material does not have a significant effect on the recovery of TiO₂ and vanadium.

Based on the results shown in FIG. 9, it can be concluded that the optimum acid consumption is around 700 kg per ton of feedstock material. Recoveries of 68 and 74% for TiO₂ and vanadium, respectively, were achieved. The acid consumption is lower than known processes, in which acid consumption is generally more than 1,000 kg per ton of ore (feedstock).

Effect of Size of Particles

Screen analysis was performed on samples taken from the cured leached residue (post-cure leach solid) to understand the distribution of vanadium and TiO₂ according to particle size after leaching step (h). The results are shown in Table 7. The finest screen size fraction (less than 0.038 mm) had the lowest percentage of vanadium and TiO₂. Table 7 shows that more vanadium and TiO₂ was unobtainable in the coarser size fractions. Put another way, it was more difficult to liberate vanadium and TiO₂ from the larger particles. It is therefore thought to be advantageous to pulverize or crush the cured first acidic solid before leaching.

TABLE 7
Screen analysis of the cured leached residue sample

Particle size (diameter) (mm)	% Mass	% Vanadium	% TiO2

More than 0.075	19.64	0.68	15.76
More than 0.063 to less than	20.68	0.63	16.18
0.075
More than 0.038 to less than	26.83	0.64	12.44
0.063
Less than 0.038	32.85	0.36	7.87

Effect of Pelletizinq Before Curing

Two experiments were carried out to investigate the effect of pelletizing before curing. Samples of first acidic solids were prepared by mixing Material A with borax, roasting, leaching, and adding sulphuric acid and water. The first acidic solid samples (using Material A and according to steps (a) to (f)) were pelletized, having a diameter of 5 to 80 mm. The pellets were cured for 30 minutes to 2 hours at 160-250° C. The cured pellets were then pulverized, taking care to avoid fines generation, and the resulting particles were leached out in an acidic environment. The results showing the recoveries of vanadium and TiO₂ are shown in Table 8. These results can be compared to results shown in FIG. 9, for example. The recoveries of both vanadium and TiO₂ were higher for similar levels of acid consumption, which is thought to be due to increased sulphation when the pellets are being baked. Without being bound by theory, it is thought that sulphur dioxide gas does not escape easily from the pellets, thereby leading to increased sulphation and consequently increased percentage recovery of vanadium and TiO₂.

TABLE 8
Summary of results for samples that were calcined,
leached in acid, and cured via pelletizing

Acid consumption
(kg sulphuric acid/ton	% Recovery

feedstock)	Vanadium	TiO2
500-750	70-78	85-88
600-800	80-85	90-92

Obtaining Vanadium from Pre-Cure Leach Liquor

Previously, vanadium has conventionally been extracted by leaching roasted calcines in water (generally in an alkaline environment due to the make-up of the calcines).

The present inventors roasted samples in air in the presence of borax for durations less than and more than an hour, respectively. The roasted calcine samples were then treated as follows:

• • Leaching the roasted calcines sample in water (alkaline environment) and re-leaching the residue in acidic environment; or
  • Directly leaching the roasted calcine samples in an acidic environment.

The summary of results for the samples is shown in Table 9 which shows that water leaching is not as effective at extracting vanadium from the roasted calcine samples. A majority of vanadium together with boron are extracted when the roasted calcine samples are directly leached out in acidic environment. Similarly, extraction of aluminium is more efficient when the roasted calcine samples are directly leached in an acidic environment. Based on the results in Table 9, it can be concluded that calcined samples should preferably be directly leached out in acidic environment to maximise vanadium/titanium recovery.

TABLE 9
Chemical analysis of the various leach liquors using ICP-OES technique in parts per million
(ppm). The samples had been roasted (calcined) in air in the presence of borax

	Roasting	Al	Ca	Co	Cu	Fe	Mg	Mn
Sample	Time (hr)	(ppm)	(ppm)	(ppm)	(ppm)	(ppm)	(ppm)	(ppm)
Water (alkaline) leach	<1	804	4.2	0.1	2.8	14.4	0.1	1.2
Water (alkaline) leach	>1	787	3.2	0.1	5.8	14.7	0.1	0.8
Water (alkaline) leach followed	<1	678	429	0.2	7.1	335	635	82
by Acid leach at pH = 1.8
Water (alkaline) leach followed	>1	1206	415	0.4	9.8	291	581	61
by Acid leach at pH = 1.8
Direct Acid leach at pH = 1.8	<1	2528	500	0.08	9.3	361	436	61
Direct Acid leach at pH = 1.8	>1	1733	405	0.18	11.4	499	494	91
	Roasting	Ni	Na	K	Cr	V	Ti	B
Sample	Time (hr)	(ppm)	(ppm)	(ppm)	(ppm)	(ppm)	(ppm)	(ppm)
Water (alkaline) leach	<1	0.1	3681	210	0.1	92	0.1	860
Water (alkaline) leach	>1	0.1	3586	262	0.5	82	0.1	838
Water (alkaline) leach followed	<1	0.4	712	17	5.2	47	128	320
by Acid leach at pH = 1.8
Water (alkaline) leach followed	>1	0.5	989	18	11.7	78	122	152
by Acid leach at pH = 1.8
Direct Acid leach at pH = 1.8	<1	0.4	3191	139	11.7	269	192	1566
Direct Acid leach at pH = 1.8	>1	1.1	2879	120	7.2	223	343	1837

Example 3

Two samples of Material A were mixed with borax and then roasted in air at a temperature of 790° C. for (a) 1 hour and (b) 1.5 hours, respectively. The roasted calcine samples (a) and (b) were not subjected to any leaching steps. Instead, the roasted calcine samples (a) and (b) were mixed with sulphuric acid and water, pelletized, and then cured at a temperature of 160° C. to 250° C. for 0.5 to 1.5 hours. The cured samples were leached in water. Titanium dioxide and Vanadium and were recovered from the post-cure leach liquor via hydrolysis and solvent extraction/precipitation. The results are shown in Table 10 from which it can be observed that recoveries were 41-48% for vanadium and 39-44% for titanium.

The leach residues (the post cure leach solid) in Table 10 was again mixed with sulphuric acid and water and pelletized before re-curing at a temperature of 300° C. in order to increase recoveries of vanadium and titanium. The cured samples were leached in water. Titanium dioxide and V₂O₅ and were recovered from the post-cure leach liquor via hydrolysis and solvent extraction/precipitation. The results showing overall recoveries of vanadium and TiO₂ are presented in Table 11. Table 11 shows that 85-90% of vanadium and 89-91% TiO₂ were recovered. The weight losses were very high implying that most of the constituents were extracted into the leach liquor. Table 11 shows that that only about 30 weight % of the waste remains after extraction.

Furthermore, Tables 10 and 11 also indicate that roasting (calcination) should preferably be carried out for more than 1 hour, in order to achieve increased recoveries.

TABLE 10
Leach residue results for samples that were roasted in air, pelletized
with sulphuric acid and cured, followed by leaching

	Roasting
	(calcination)	% Weight	% Recovery	% Recovery
	Time	loss	Vanadium	TiO2

1	hour	30.37	41.96	39.29
1.5	hours	34.00	47.92	44.15

TABLE 11
Leach residue results for the samples that were roasted in air at 790°
C., pelletized and cured at 300° C., followed by leaching

	Roasting
	(calcination)	% Weight	% Recovery	% Recovery
	Time	loss	Vanadium	TiO2

1	hour	67.67	85.55	89.16
1.5	hours	70.33	90.09	91.48

Hydrolysis of Titanium Dioxide (TiO₂)

Titanium dioxide was precipitated from the solutions (post-cure leach liquor) from Examples 1 to 3 via hydrolysis which involved heating the pregnant solution at a temperature of more than 50° C. The particle size of the hydrolysed titanium dioxide is of importance because fine particles can create problems during any subsequent chlorination steps.

It was found that growth of the titanium dioxide particles could be controlled by varying the hydrolysis temperature.

Scanning electron microscopy images for the TiO₂ that was hydrolysed at varying temperature are shown in FIGS. 10-12. When the hydrolysis was performed at less than 70° C., it was found that the particle sizes (average particle diameter) were around 20-60 μm, as shown by FIG. 10. When the hydrolysis was performed at higher than 70° C. to 80° C., it was found that the particle sizes (average particle diameter) were around 3 μm to 20 μm, as shown by FIG. 11. When the hydrolysis was performed at higher than 80° C., it was found that the particle sizes (average particle diameter) were less than 3 μm, as shown by FIG. 12.

A digital image of the TiO₂ precipitates which was produced via hydrolysis at more than 70-80° C. is shown in FIG. 13. As shown in FIG. 13, precipitated TiO₂ had white colour, thought to be due to a low iron content. The digital image in FIG. 13 further shows that the precipitated TiO₂ settled well.

FIG. 10 is a scanning electron microscope image of the precipitated TiO₂ via hydrolysis at lower temperature (less than 70° C.).

FIG. 11 is a scanning electron microscope image of the precipitated TiO₂ via hydrolysis at a temperature of 70-80° C.

FIG. 12 is a scanning electron microscope image of the precipitated TiO₂ via hydrolysis at a temperature of >80, i.e. 90-100° C.

FIG. 13 is a photo of the precipitated TiO₂ via hydrolysis at 70-80 degrees Celsius.

Solvent Extraction

Solvent extraction was carried out on Example 1 to extract vanadium and base metals (Cr, Ni, Co, Mn etc.) that dissolved in solution during leaching after acid curing. The procedure was as follows:

• • The aqueous solution, after hydrolysis of titanium, was fed to the solvent extraction process. This liquor contained vanadium, nickel, iron, chromium, and aluminium.
  • The initial pH of the liquor was determined and adjusted via addition of lime or sodium hydroxide or ammonia solution.
  • An organic solution was prepared at a strength of 5%-40% v/v (extractant to diluent).
  • The aqueous solution was contacted with the organic solution for 10-15 minutes in a ratio of 1:2 to aid the efficient separation of the loaded organic and raffinate.
  • The loaded organic (solvent) is subjected to a two-stage scrubbing process using 2 gpl of H₂SO₄ for 10-15 minutes to extract the metal ions (Fe, Al, Ni and Cr) from the loaded organic.
  • The raffinate contained the desired vanadium metal ions which were recovered and V₂O₅ obtained.

Chemical analysis of base metals extracted from the loaded organic is shown in Table 12. The main components were iron, chromium, nickel, and manganese.

TABLE 12
Chemical analysis of the extracted base metals

	% Mn	% Fe2O3	% Cr2O3	% NiO
	1.64	64.34	16.55	10.20

The base metals iron, chromium, nickel, and manganese were melted to obtain an alloy comprising manganese, iron, chromium, and nickel.