Process for recovering titanium dioxide

Abstract

The present invention relates to a process for the recovery of titanium dioxide from a titanium-containing material, the process comprising the steps of: leaching the titanium-containing material in a first leaching step at atmospheric pressure and a temperature of from 70°C to 97°C with a first leaching agent to produce a first leach solution comprising undissolved first leach solids containing titanium content and a first leach liquor, the first leaching agent comprising hydrochloric acid at a concentration of less than 23% w / w; separating the first leach liquor from the undissolved first leach solids; leaching the first leach solids in a second leaching step at atmospheric pressure and a temperature of from 60°C to 80°C with a second leaching agent in the presence of a Fe powder reducing agent to produce a second leach solution comprising undissolved second leach solids and a second leach liquor comprising leached titanium content and iron content, the second leaching agent comprising a mixed chloride solution comprising less than 23% w / w hydrochloric acid and additional chloride selected from an alkali metal chloride, magnesium chloride and calcium chloride or mixtures thereof; separating the second leach liquor from the undissolved second leach solids; then separating the titanium dioxide and iron content from the second leach liquor by precipitation and regenerating the second leaching agent for recycle to the second leaching step.

Description

[0001] Priority Cross Reference

[0002] This application claims priority from Australian Provisional Patent Application No. 2020901698, filed 26 May 2020, the contents of which should be understood as incorporated by reference into the present specification. TECHNICAL FIELD

[0003] The present application relates generally to a process for the recovery of titanium dioxide from a titanium-containing material. The present application is particularly applicable to the recovery of titanium dioxide from titanium-containing ores or ore concentrates, and will hereinafter be conveniently disclosed in relation to this exemplary application. However, it will be appreciated that the present application is not limited to this application and can be used to recover titanium dioxide from a variety of sources, including other ore bodies containing titanium minerals, vanadium associated with titanium minerals such as titanomagnetite, vanadium-containing minerals, and titanium-containing leach residues and slags. BACKGROUND

[0004] The following background discussion of the present application is included to facilitate an understanding of the present application. However, it should be appreciated that this discussion is not an acknowledgement or admission that any of the material referred to was published, known or part of the common general knowledge as at the priority date of the application.

[0005] Titanium is the ninth most abundant element, making up approximately 0.6% of the Earth’s crust. A variety of titanium-containing minerals exist in nature, including ilmenite (FeO-TiO2or TiFeO3), rutile (TiO2) and leucoxene (Fe2O3-nTiO2). Ilmenite is reported to contain between 40% and 65% TiO2, which is approximately 91% of the world’s demand for titanium minerals. In 2019, the world ilmenite production reached approximately 7 million tonnes. In addition to titanium, titanium-containing minerals typically contain other valuable metals, the identity and quantity of which can vary greatly depending on the source of the ore. These titanium-containing ores can include one or more of vanadium, aluminium, manganese, magnesium, molybdenum, chromium, copper, lead, nickel, zinc, zirconium, niobium and tantalum. These titanium-containing ores also typically contain varying amounts of Fe2O3and gangue material, typically silicates, alumina, lime and magnesia.

[0006] Titanium-containing ores can be leached or beneficiated to produce a concentrate, with beneficiation being employed if the titanium content of the ore is low. Methods for the recovery of titanium dioxide from ilmenite and other titanium-containing ores are known. These methods mostly involve digesting the ore in a mineral acid (e.g. hydrochloric acid (chloride process) or sulphuric acid (sulphate process)) to remove at least the titanium values from the ore. In many such processes, the purity of the titanium dioxide obtained can be about 90 to 95%, thus requiring further purification procedures to produce a high quality, pigment grade product.

[0007] The sulphate route is carried out through hydrometallurgical route using ilmenite or low grade titania slag (72 to 87% Ti02) as feed material, where the product quality is still poor and the process generates large amount of waste. In contrast, the chloride route traditionally processes only high grade synthetic rutile (90 to 95% Ti02), natural rutile (95% Ti02) or high grade titania slag (>90% Ti02) through a complex process to produce a purer product with relatively less waste generation.

[0008] Hydrometallurgical processing of ilmenite ores with hydrochloric acid has been the main focus of research in recent years. A number of processes have been proposed, such as using i) direct leaching, ii) leaching in the presence of an oxidizing agent, iii) leaching in the presence of a reducing agent such as iron powder, and iv) leaching after pre-oxidation of the concentrate at high temperatures. The leaching agent is typically HCl based, either containing high concentration of HCl (30 to 40 wt%) or optionally adding chloride species such as MgCh which have been found to enhance the direct leaching process.

[0009] An important issue for any HCl leaching route of ores / minerals / concentrates is the cost of HCl, hence the need to regenerate HCl from the treatment liquor to ensure the process is economically viable. To regenerate HCl from the treatment liquor, either thermal hydrolysis (e.g. International Patent Publication No. WO 2014 / 125275 Al) or high temperature hydrolysis (e.g. International Patent Publication No. WO 2011 / 094858) technology is used. Both processes are energy intensive, requiring high temperatures: 400°C to 800°C for thermal hydrolysis decomposition, where the metal chloride salts decompose to metal oxides; and 170 to 180°C for high temperature hydrolysis to enable a hydrothermal reaction to precipitate the remaining metals as metal oxides, such as iron as hematite. Both processes also require expensive reactor construction materials due to the highly corrosive gaseous HCl produced at these temperatures.

[0010] Another issue with the HCl leaching route is the valuable metal recovery process. In many processes, expensive solvent extraction (SX) technology (e.g. US Patent No. 7803336) is used to separate the valuable metals including titanium and iron. The adoption of solvent extraction technology in the process to recover Fe and Ti is a capital intensive option, which can hinder the successful commercialization of the process.

[0011] An example of a chloride process is taught in Canadian Patent Publication CA2878744, which includes among many methods a process for recovering titanium dioxide and valuable metals from a titanium-containing material using a two-stage chloride-based leaching process. The titanium-containing material can for example be selected from a titanium-containing ore or a recycled industrial titanium-containing material, such as a slag, a red mud or a fly ash. The first leaching stage uses a HC1-based lixiviant comprising HC1 at a concentration of 25 to 45 wt%, at a temperature of 125 to 225 °C on a titanium-containing material comprising Ti, Si and a first metal, to produce a first leach liquor of the first metal and a solid comprising Si and Ti. The leach liquor and solid are separated using a custom recovery process for recovering the first metal from the first leach liquor. The solid comprising Si and Ti from the first stage is subjected to a second leach using a lixiviant comprising less than 20% w / w HC1, in the presence of a chloride (MgCl2or ZnCl2), at a temperature of less than 85 °C, to produce a second leach liquor comprising TiCl4. The titanium is recovered as TiO2by heating, solvent extraction and subsequent formation of titanium dioxide from the solvent extraction, or reaction with water, oxygen and / or a base to cause precipitation of TiO2. The HC1 is regenerated from the leach liquor. Recovery of hematite (Fe2O3) from FeCl3that can be present in the ore leach liquor can also be achieved using high temperature hydrolysis at 160 to 175 °C, producing HC1 for recycle to the leaching stage.

[0012] While CA2878744 provides a general chloride leach, the conditions of the first leach stage provide non-ideal, capital-intensive conditions for recovery of a number of important valuable metals, particularly the high concentration of HC1 lixiviant and the high temperature. Furthermore, the second leach stage is conducted under conditions that require the value of the resulting metals in the lixiviant, including any iron content, to be recovered in an energy-intensive manner.

[0013] International Patent Publication WO2015 / 131266 teaches another chloride method relating to the recovery of high-grade synthetic rutile (95 to 98% TiO2) from low-grade ores containing less than 12% TiO2. Similar to previous patent disclosures, this method comprises a two-stage leaching process in 35 to 40% by weight HCl at an acid-to-ore ratio of 2 to 2.5. The first stage leaching is carried out at 60 to 70°C with 80% ground ore of -200 mesh particle size. The second stage leaching is carried out at 75 to 80°C with the residue from the first stage leaching. The leachates from the two stages are combined after solid-liquid separation and boiled to distill off unreacted HCl until the dissolved titanium is hydrolyzed and most of the ferric chloride precipitates as hydrate. After filtering the slurry of hydrolyzed titanium containing ferric chloride crystals, the crystals are dissolved in a minimal amount of dilute hydrochloric acid, leaving insoluble TiO(OH)2, which is then calcined to obtain a TiO2 product of 95 to 98%. The Ti-free liquid obtained after the Ti hydrolysis step is further processed by solvent extraction or selective precipitation to recover V and Cr, respectively. The HCl leaching agent is regenerated using a spray reactor and subjected to high-temperature hydrolysis in a micro-oxidizing atmosphere to produce iron oxide and HCl, which are used to recover iron and hydrochloric acid. However, many valuable metal recovery methods, including iron recovery and HCl regeneration stages, remain energy-intensive.

[0014] Therefore, it is desirable to provide an improved or at least alternative method for recovering titanium / titanium dioxide from titanium-containing materials such as titanium-containing ores or concentrates. Summary of the Invention

[0015] A first aspect of the present invention provides a method for recovering titanium dioxide from titanium-containing materials, the method comprising the following steps:

[0016] In the first extraction step, the titanium-containing material is extracted with a first extractant at atmospheric pressure and a temperature of 70°C to 97°C to produce a first extraction solution, which contains undissolved first extraction solids and a first extraction liquid containing titanium contents, preferably substantially all the titanium contents of the titanium-containing material, and the first extractant contains hydrochloric acid with a concentration of less than 23% w / w.

[0017] Separate the first extract from the undissolved first extract solid;

[0018] In the second extraction step, the first extraction solid is extracted with a second extractant in the presence of Fe powder reducing agent at atmospheric pressure and a temperature of 60°C to 80°C to produce a second extraction solution containing undissolved second extraction solid and a second extract containing extracted titanium and iron contents. The second extractant contains a mixed chloride solution containing less than 23% w / w hydrochloric acid and additional chlorides selected from alkali metal chlorides, magnesium chloride and calcium chloride or mixtures thereof.

[0019] separating the second leach liquor from the undissolved second leach solids;

[0020] precipitating titanium dioxide from the second leach liquor by adding heated or boiling water under an inert gas or nitrogen atmosphere to raise the temperature of the second leach liquor to 85°C to 100°C, thereby producing a treated second leach liquor and a titanium dioxide containing solid;

[0021] separating the titanium dioxide containing solid from the treated second leach liquor;

[0022] precipitating iron content from the treated second leach liquor by adding a neutralizing agent and an oxidizing agent to the treated second leach liquor at a temperature of 70°C to 90°C to raise the pH of the second leach liquor to 4 to 8, thereby producing an iron removed slurry comprising an iron removed second leach liquor and an iron precipitated solid;

[0023] separating the iron removed second leach liquor from the iron precipitated solid; and

[0024] regenerating the second leaching agent for recycling to the second leaching step,

[0025] thereby recovering titanium as titanium dioxide from the second leach solution.

[0026] Dual stream leaching process

[0027] The process of the present invention involves a dual stream leaching process using hydrochloric acid leaching followed by mixed hydrochloric acid and calcium chloride leaching which selectively leaches impurities and valuable metals other than titanium (such as vanadium and aluminum) from the titanium bearing material in a first leaching step and then selectively leaches the titanium content from the titanium bearing material in a second leaching step which can then be recovered. This dual leaching process strategy results in a more efficient leaching process specifically targeting titanium in the second leaching step compared to the previous single step leaching process where titanium is dissolved from the ore material along with the impurities in a single stage leaching.

[0028] Importantly, the second leaching step is conducted using metallic Fe powder under a reducing atmosphere. The reducing conditions provide the advantage of higher Ti extraction of Ti minerals (e.g., ilmenite, rutile, pseudorutile, and anatase, etc.) from the first leach solids and thus from the feed Ti-bearing material. In this leaching step, iron powder is added to assist in the dissolution of the Ti-bearing minerals. The iron content is removed by iron precipitation from the ferrous chloride solution and oxidizing agent (e.g., alkali metal peroxides, alkali metal perchlorates, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, alkali metal chlorates, chlorine gas, alkali metal hypochlorites, hydrogen peroxide, perchloric acid, oxygen-containing gases such as air or oxygen, other non-sulfur containing oxidizing agents, or mixtures thereof) and neutralizing agent (e.g., limestone, lime, or MgO) to form typically one or more of goethite (a-FeOOH), akageneite (β-FeOOH), hematite (Fe203), magnetite (Fe304), or mixtures thereof, or preferably only as magnetite. The iron removal process of the present invention provides significant advantages over conventional iron removal techniques, which typically focus on more expensive (higher operating cost) removal techniques such as hot hydrolysis of hydrous ferrous chloride or high temperature hydrolysis of ferric chloride to hematite (Fe203).

[0029] It will be appreciated that the "Ti-bearing material" can be any material, including a material that contains a Ti-bearing substance, which comprises one or more of:

[0030] a. a Ti-bearing ore material, including a Ti-bearing ore or ore body, a concentrate thereof, a modified ore thereof, and a tailings thereof, and mixtures thereof;

[0031] b. an ore body that contains Ti-bearing minerals such as ilmenite, rutile, and / or leucoxene;

[0032] c. vanadium, Ti-bearing minerals associated with Ti-bearing minerals such as titanomagnetite;

[0033] d. a titanomagnetite ore or ore body, a concentrate thereof, a modified ore thereof, and a tailings thereof, and mixtures thereof;

[0034] e. Ti-bearing leach residues and slags; or

[0035] f. mineral processing residues.

[0036] In preferred embodiments, the Ti-bearing material is a Ti-bearing ore, a Ti-bearing ore concentrate, a modified tailings of a Ti-bearing ore, or mixtures thereof. In some embodiments, the Ti-bearing material is a Ti-bearing ore, a concentrate thereof, a modified ore thereof, and a tailings thereof, or mixtures thereof. In these embodiments, the Ti and Fe values are leached from the Ti-bearing ore material. In various embodiments, the Ti-bearing material includes ilmenite. However, it will be appreciated that the Ti-bearing material can include other Ti-bearing minerals, including (but not limited to) rutile, pseudorutile, anatase, and / or leucoxene.

[0037] The titanium-containing ore material can be ore itself, but is preferably a concentrate thereof. Techniques for processing titanium-containing ores, such as ilmenite ores, to form concentrates or beneficiating the ores are known in the art and include the use of gravity or magnetic separation steps. The process preferably operates using ore concentrates. In other embodiments, the ore can have been subjected to a smelting step in the presence of carbon and / or fluxing agents, after which the slag is separated from the smelting process and subjected to a leaching step. Thus, the ore can be in the form of a matte, such as a converter matte or a bath matte. The ore can also be in the form of a roasted and / or reduced titanium-containing concentrate or other intermediate, all of which include the above-mentioned matte and are referred to herein as modified ore. The ore can also be in the form of a titanium-containing ore tailings. It will be appreciated that the expression "ore" also includes any other form of ore and mixtures of various forms of ore can be used. The process of the present invention can operate without pre-treatment of the titanium-containing ore. In particular, the process can operate with or without roasting or reducing the ore.

[0038] Pre-treatment of the ore, such as oxidation and / or reduction of the ore, is generally not required prior to leaching. The process operates with relatively low concentrations of hydrochloric acid, in particular with hydrochloric acid having a concentration of less than 23% w / w (weight by weight). As pre-treatment of the ore is not required and the leaching step produces a solution of titanium value, the process can be described as a direct process for leaching and recovering titanium. The process of the present invention is considered to be environmentally friendly, requiring no extensive pre-treatment procedures.

[0039] First leach treatment stream

[0040] The process of the present invention is a two-stream process for treating titanium-containing material, wherein each treatment stream can be operated independently, with separate dissolution of the valuable metals and regeneration of the leaching agent.

[0041] The first leaching step primarily separates any hydrochloric acid-soluble impurities and valuable metals, such as vanadium, aluminium and iron, that can be present in the titanium-containing material, leaving the titanium content substantially in the first leaching solids. The leaching step is conducted at atmospheric (ambient) pressure, i.e. the leaching step does not need to be conducted under pressure. The leaching is conducted under conditions such that the titanium leached from the titanium-containing ore material is substantially retained in the titanium-containing material (solids), i.e. the titanium does not leach into solution. In this step, the leaching conditions are selected to leach a substantial portion of the vanadium and aluminium content of the titanium-containing material into solution. Thus, no extraction and recovery steps for titanium are required in the associated processing stream. To achieve this, the first leaching step is conducted using a first leaching agent comprising less than 23% w / w HCI solution, preferably 20% to 22% w / w. The temperature of the leaching is between 70°C and 97°C, preferably between 85°C and 97°C.

[0042] The first leaching step can be carried out as a co-current step, a counter-current step or otherwise continuously, or the leaching step can be carried out as a batch step.

[0043] A solution enriched in valuable metals is obtained in the first leaching step (first leachate). The residue (un-dissolved first leaching solids) can be in the form of a suspension. The leaching mixture is fed to a solid / liquid separation step to effect separation of the first leachate from the first leaching solids (e.g. leach residue and other gangue). Techniques for such separation are known in the art, for example using pressure or vacuum filters, counter-current decantation, thickeners or centrifuges.

[0044] The titanium-bearing material can also contain one or more additional valuable metals, such as iron, vanadium, manganese, magnesium or aluminium. Other trace elements, substances or impurities can also be present. Accordingly, the method of the present application can include a step of removing and recovering any iron, vanadium, manganese, magnesium or aluminium from the leachate in this first leaching process stream. In these embodiments, the first leachate is subjected to a step of recovering at least one valuable metal therefrom.

[0045] In exemplary embodiments, the valuable metals in the titanium-bearing material include at least vanadium and / or aluminium. In such embodiments, the method further includes a vanadium and / or aluminium removal step comprising:

[0046] adding a neutralising agent, preferably at least one of limestone, lime or MgO, to the first leachate under an inert gas or nitrogen atmosphere at a temperature of 50°C to 80°C to raise the pH of the liquid to 3 to 6, thereby precipitating vanadium and aluminium to produce a V / Al-removed slurry; and

[0047] separating the V / Al-removed slurry into a liquid portion comprising a V / Al-removed liquid and a solid portion comprising V / Al-precipitated solids.

[0048] Accordingly, the separation of vanadium and / or aluminium from the first leachate is a precipitation technique that results from the addition of limestone, lime or MgO to raise the pH of the liquid. In contrast, most prior art methods use more expensive solvent extraction techniques to separate vanadium.

[0049] The method steps are preferably carried out under an inert gas or nitrogen atmosphere, preferably under a nitrogen blanket, to prevent oxidation of ferrous iron to ferric iron and to hinder precipitation of any iron content (in the form of ferric iron) that can be present in the first leachate. The vanadium and / or aluminium precipitation is preferably carried out prior to removal / recovery of other valuable metals that can be present in the first leachate.

[0050] Recovery of vanadium and aluminium from the V / Al precipitation solids can be carried out by any suitable method known in the art, for example using an ammonia or HC1 solution leach followed by a precipitation and optional calcination step. Details of these process steps will be described in more detail later in the specification.

[0051] As mentioned above, the titanium-containing material can include an iron content (i.e. one of the valuable metals), for example, the titanium-containing material is a titanium-containing ore or concentrate thereof. Alternatively or additionally, the first leach liquor can include an iron content of Fe added at some point in the first leach process stream. In such embodiments, Fe powder is added as Fe(III) is present in the leach liquor. The Fe powder reduces the Fe(III) to Fe(II) and the Fe(II) does not precipitate due to the nitrogen blanket and remains in solution during the V / Al reduction. In these embodiments, the process can further include an iron removal step comprising:

[0052] adding a neutralising agent and an oxidising agent to the first leach liquor at a temperature of 70 to 90 °C to raise the pH of the liquor to 4 to 7, thereby precipitating iron to produce an iron-removed slurry; and

[0053] separating the iron-removed slurry into a liquid portion comprising an iron-removed liquor and a solid portion comprising iron precipitation solids.

[0054] The neutralising agent can include any suitable neutralising substance or compound, and preferably includes at least one of limestone, lime or MgO.

[0055] The oxidising agent can include one of an alkali metal peroxide, an alkali metal perchlorate, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, an alkali metal chlorate, chlorine gas, an alkali metal hypochlorite, hydrogen peroxide, perchloric acid, an oxygen-containing gas such as air or oxygen, other non-sulphur containing oxidising agents, or mixtures thereof. A preferred oxidising agent is H2O2 or an oxygen-containing gas, for example oxygen or air, etc. The most preferred oxidising agent is oxygen or air.

[0056] The iron removal solids can include one or more of magnetite, goethite, hematite and tetragonal goethite. However, magnetite is the preferred form of the iron removal solids. Thus, in exemplary embodiments, the iron is substantially precipitated as magnetite, preferably only as magnetite. The iron precipitate (preferably predominantly magnetite) can be used to produce Fe powder, for example by reacting the precipitated magnetite with carbon / charcoal / coke / coal at high temperature of ~800 to 1000 °C. The resulting Fe powder can be recycled for use in the process, for example in the second leach process stream or the optional reduction step of the first leach process stream.

[0057] The iron removal step is preferably carried out after the vanadium and / or aluminium removal step(s) throughout the first leach process stream.

[0058] The titanium-containing material can also include manganese and / or magnesium (i.e. one of the valuable metals). Alternatively or additionally, the first leach liquor can include a manganese and / or magnesium content derived from the addition of Mg or Mn, for example MgO, at a point in the first leach treatment stream. In these embodiments, the method further comprises a manganese and / or magnesium removal step comprising:

[0059] adding to the iron removal liquor a neutralising agent, preferably lime, and an oxidising agent, preferably H2O2 or an oxygen-containing gas, more preferably air, to raise the pH of the liquor to 9 to 10, thereby causing the Mg and / or Mn to precipitate to produce a Mg / Mn-removed slurry; and

[0060] separating the Mg / Mn-removed slurry into a liquid portion comprising a Mg / Mn removal liquor and a solid portion comprising precipitated Mg and / or Mn solids.

[0061] The precipitation step is carried out in the presence of an oxidising agent, which can be an oxidising agent, preferably H2O2 or an oxygen-containing gas such as oxygen, air or the like for oxidising Mn(II) to Mn(IV). In this step, it is preferred to use lime as the neutralising agent. The precipitated Mg and / or Mn solids will typically comprise a mixture of Mg(OH)2and Mn-oxides / hydroxides. The Mg / Mn removal liquor will be primarily a chloride solution, for example calcium chloride in the case where limestone and / or lime is used in the preceding step. The manganese and / or magnesium removal step is preferably carried out after the iron removal step throughout the first leach treatment stream.

[0062] To aid process economics, it is preferred to regenerate and recycle the first leachant to the first leaching step. In these embodiments, the method further comprises:

[0063] regenerating the first leachant and recycling the first leachant to the first leaching step.

[0064] In embodiments, the first leachant is regenerated by:

[0065] concentrating the chloride content of the Mg / Mn removal liquor by removal of water, preferably boiling and / or evaporation, to produce an evaporation liquor;

[0066] reacting the evaporation liquor with at least 98% w / w sulphuric acid at atmospheric conditions at a temperature of 30°C to 90°C, preferably 80°C to 85°C, to produce 20 to 22% w / w hydrochloric acid and a solid precipitate,

[0067] separating the precipitated solids from the hydrochloric acid liquor; and

[0068] recycling the hydrochloric acid liquor to the first leaching step.

[0069] The composition of the chloride content will depend on the composition of the additives to the first leach process stream. In many cases, the chloride content will comprise a calcium chloride solution / liquid. Thus, the evaporation liquor will comprise a calcium chloride liquid. In such embodiments, the evaporation liquor is reacted with concentrated sulphuric acid (98% w / w) in a stoichiometric ratio of calcium chloride to sulphuric acid to produce HCl and a precipitate comprising at least one of gypsum, hemihydrate or anhydrite. Furthermore, in such embodiments, the reaction between the evaporation liquor and concentrated sulphuric acid is preferably carried out at a temperature range of 80°C to 85°C, thereby precipitating only anhydrite.

[0070] The first leach process stream preferably comprises a plurality of process methods prior to the valuable metal recovery step, more particularly prior to the vanadium and / or aluminium precipitation step. In these embodiments, the method further comprises the following steps prior to the precipitation of vanadium and / or aluminium from the first leach liquor:

[0071] neutralising at least some of the free acid (HCl) in the first leach liquor by adding at least one of a feed titanium-containing material (preferably a titanium ore concentrate), limestone, lime or MgO to the first leach liquor to produce a first liquid neutralisation slurry comprising neutralised leach solids; and

[0072] separating the first liquid neutralisation slurry into a solid portion comprising the neutralised leach solids and a liquid portion comprising the neutralised first leach liquor.

[0073] It will be appreciated that other neutralising agents can also be used, such as sodium hydroxide and the like. In some embodiments, the first leach liquor neutralisation stage solids or leach solids (when using an ore) are fed to the first leach stage.

[0074] The first leach process stream preferably comprises the following steps after the neutralisation step:

[0075] reducing the neutralised first leach liquor by adding metallic iron, preferably iron powder, to convert the ferric chloride in the first leach liquor to ferrous chloride at 45°C to 75°C; and

[0076] separating the reduced first leach liquor into a liquid portion comprising the reduced liquor and a solid portion comprising any unreacted solid iron powder.

[0077] The reduction is preferably carried out under an inert gas or nitrogen atmosphere, preferably under a nitrogen blanket, and achieves an oxidation reduction potential (ORP) of the liquid of less than 100 mV.

[0078] Second leach treatment stream

[0079] The second leaching step and the associated second leaching process flow is essentially focused on efficient titanium recovery in the form of titanium dioxide (rutile or anatase). The specific recovery steps and conditions depend on the composition of the titanium-containing material and thus on the method required to recover the titanium dioxide and regenerate the leaching agent used in the second leaching step.

[0080] The second leaching step is performed at atmospheric (ambient) pressure, i.e. there is no need to perform the leaching step under pressure. The conditions under which the leaching is performed are such that the titanium leached from the titanium-containing ore material is leached into solution and remains in solution during the leaching process, i.e. the titanium does not precipitate as, for example, titanium dioxide. In particular, the leaching is performed at a temperature of less than or equal to 80 °C, typically between 60 °C and 80 °C, and most preferably at a temperature in the range of 70 °C to 80 °C. The leaching is performed with a second leaching agent in the presence of a Fe powder reducing agent. The second leaching step is preferably performed for 2 to 6 h, and in some embodiments for 4 to 6 h.

[0081] The second leaching step can be performed as a co-current step, a counter-current step or otherwise continuously, or the leaching step can be performed as a batch step.

[0082] The second leaching agent comprises a mixed chloride solution comprising less than 23% w / w hydrochloric acid and an additional chloride selected from an alkali metal chloride, magnesium chloride and calcium chloride or mixtures thereof. In embodiments, the mixed chloride solution comprising 20% to 22% w / w HC1 and an additional chloride has a total chloride ion concentration of 400 to 550 g / L (calculated based on the amount of chloride and hydrochloric acid in the leaching agent solution). It will be appreciated that the metal chloride / HCI (metal to hydrochloric acid) ratio in the leaching is preferably adjusted based on, for example, the specific ore being leached and the temperature to optimise the leaching. The upper limit of the chloride ion concentration can depend on the ions present in the leaching solution, especially due to the leaching of the ore and the complexes formed thereby.

[0083] In the second leaching step, the additional chloride is selected from an alkali metal chloride, magnesium chloride and calcium chloride or mixtures thereof. In exemplary embodiments, the chloride is preferably calcium chloride, whereby the hydrochloric acid is regenerated and the mixed chloride solution containing the hydrochloric acid and unreacted calcium chloride is recycled in the process. However, it will be appreciated that other chlorides, such as magnesium chloride, can equally be used.

[0084] In a particularly preferred embodiment of the application, the chloride is derived from calcium chloride and hydrochloric acid, and the 400 to 550 g / L chloride ion concentration is calculated based on the amount of calcium chloride and hydrochloric acid in the leachant solution. In various embodiments, the amount of hydrochloric acid is in the range of 255 to 280 g / L and the amount of calcium chloride is in the range of 300 to 400 g / L. For the sake of clarity, a concentration of 255 to 280 g / L HC1 results in -20 to 22% w / w HC1 in the HC1 and CaCl2 mixed chloride solution, with a liquid SG of reasonably high -1.3. If there is water in the HC1, 20 to 22% w / w HC1 would be -220 g / L to 240 g / L HC1.

[0085] A solution enriched in valuable metals (second leach liquor) is obtained in the second leaching step. The residue (un-dissolved second leaching solids) can be in the form of a suspension. The leach mixture is fed to a solid / liquid separation step to effect separation of the second leach liquor from the second leaching solids (e.g. leach residue and other gangue). Techniques for such separation are known, for example using pressure or vacuum filters, counter-current decantation, thickener or centrifuge.

[0086] For example, where the titanium-bearing material is a titanium-bearing ore or concentrate thereof, the first leaching solids will typically contain a substantial portion of the titanium-bearing minerals, such as ilmenite and any precipitated rutile, pseudorutile, anatase, etc., as well as gangue minerals that were not dissolved during the first leaching process. The leaching conditions in this second leaching step can be adjusted to assist in the dissolution of each of these titanium-bearing minerals. Thus, in some embodiments, the second leaching step comprises two leaching regimes, including:

[0087] a first leaching regime in a mixed chloride solution (without the addition of any iron powder); and

[0088] a second leaching regime in a mixed chloride solution with the addition of iron powder.

[0089] The second leaching reaction can be conducted as a two-stage reaction. Here, the first leaching regime and the second leaching regime of the second leaching step can be conducted as a continuous leaching step in (i) the same leaching stage / vessel; or (ii) separate leaching stages / vessels. In various embodiments, the first leaching regime is conducted for a duration of 1 to 2 h and the second leaching regime is conducted for a duration of 1 to 4 h. In some embodiments, the first leaching regime is conducted for a duration of 1 to 2 h and the second leaching regime is conducted for a duration of 2 to 4 h.

[0090] The first leach process is used to dissolve most of the ilmenite mineral. The second leach process has Fe to dissolve the remaining unreacted ilmenite and other Ti-containing minerals or precipitated solids, such as rutile, pseudorutile or anatase, etc., in a reducing atmosphere. The iron powder will also reduce the ferric iron present in the leach liquor to ferrous iron during the leach reaction. If desired, an additional small amount of fresh second leach agent (mixed chloride solution) can be added in the second stage second leach to stabilize the dissolved metals and additional iron from the added Fe powder.

[0091] After leaching, the second leach process stream includes a method of recovering the dissolved titanium and iron from the leach liquor and regenerating the leach agent HCl and CaCl2mixed solution.

[0092] The iron content of the treated second leach liquor is removed by adding a neutralizing agent and an oxidizing agent to the V / Al removal liquor at a temperature of 70°C to 90°C to raise the pH of the liquor to 4 to 8 to cause the iron to precipitate to produce an iron removed slurry; and separating the iron removed slurry into a liquid portion containing the iron removal liquor and a solid portion containing the iron precipitated solids. The neutralizing agent can include any suitable neutralizing substance or compound, and preferably includes at least one of limestone, lime or MgO. The oxidizing agent can include one of an alkali metal peroxide, an alkali metal perchlorate, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, an alkali metal chlorate, chlorine gas, an alkali metal hypochlorite, hydrogen peroxide, perchloric acid, an oxygen-containing gas such as air or oxygen, other non-sulfur containing oxidizing agents, or mixtures thereof. The preferred oxidizing agent is H2O2or an oxygen-containing gas, such as oxygen or air, etc. The most preferred oxidizing agent is oxygen or air. The iron removal solids can include one or more of magnetite, goethite, hematite and tetragogoethite. However, magnetite is the preferred form of the iron removal solids. Also, in exemplary embodiments, the iron is substantially precipitated as magnetite, preferably only as magnetite.

[0093] The iron precipitate, preferably primarily magnetite, can be used to produce Fe powder, such as by reacting the precipitated magnetite with carbon / charcoal / coke / coal at high temperatures of ~800 to 1000°C. The resulting Fe powder can be recycled for use in the process, such as the second leach process stream or the optional reduction step of the first leach process stream.

[0094] In some embodiments, the titanium content of the second leach liquor can include Ti(III) content. To recover titanium dioxide, it is preferred to convert any Ti(III) content to Ti(IV) prior to the titanium dioxide precipitation step. Accordingly, in these embodiments, the method further includes the step of:

[0095] introducing an oxidizing agent into the second leach solution prior to the titanium dioxide precipitation step to oxidize any Ti(III) content to Ti(IV) by controlling the redox potential of the second leach solution to be within 100 to 200 mV,

[0096] wherein the oxidizing agent is selected from air, oxygen, alkali metal peroxides, alkali metal perchlorates, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, alkali metal chlorates, chlorine gas, alkali metal hypochlorites, hydrogen peroxide, perchloric acid, other non-sulfur containing oxidizing agents, or mixtures thereof.

[0097] Examples of alkali metal peroxides are sodium peroxide and potassium peroxide. Examples of alkali metal perchlorates are sodium perchlorate and potassium perchlorate. Ammonium perchlorate, magnesium perchlorate, and magnesium chlorate can also be used. Examples of alkali metal chlorates are sodium chlorate and potassium chlorate. One example of an alkali metal hypochlorite is sodium hypochlorite. Other oxidizing agents are non-sulfur containing oxidizing agents; the presence of sulfur in the oxidizing agent should be avoided. Preferred oxidizing agents are selected from the group consisting of air, oxygen, chlorine gas, sodium chlorate, sodium perchlorate, hydrogen peroxide, perchloric acid, and mixtures thereof. In exemplary embodiments, the oxidizing agent comprises hydrogen peroxide, and in some embodiments, dilute hydrogen peroxide.

[0098] Thereafter, the titanium dioxide can be recovered using a precipitation step in which heated or boiling water is added to the second leach solution under an inert gas or nitrogen atmosphere to raise the temperature of the second leach solution to 85°C to 100°C, thereby producing a treated second leach solution and a titanium dioxide containing solid. The titanium dioxide precipitation step preferably includes hydrolyzing the Ti(IV) content of the second leach solution to precipitate as a titanium dioxide (Ti02) solid. The reaction is conducted under an inert gas or nitrogen atmosphere, such as a nitrogen blanket, to prevent the oxidation of ferrous iron to ferric iron and the precipitation of ferric iron with the Ti02during the washing stage of the Ti02. The Ti(IV) hydrolysis will release HCl in solution. In embodiments, the HCl can be partially neutralized by the addition of at least one of limestone, lime, or a MgO slurry to maximize Ti02recovery.

[0099] The ferrous solution can be optionally V / Al removed according to the same procedures described for the first leach treatment stream, as a function of the V and / or Al concentration in the Ti02precipitate solution. In these embodiments, the second leach treatment stream further includes the steps of:

[0100] adding a neutralizing agent, preferably at least one of limestone, lime, or MgO to the treated second leach solution under an inert gas or nitrogen atmosphere at a temperature of 50°C to 80°C to raise the pH of the liquid to 3 to 6, thereby precipitating vanadium and aluminum to produce a V / Al removed slurry; and

[0101] The V / Al removal slurry is separated into a liquid portion comprising a treated second leachate depleted in V / Al and a solid portion comprising V / Al precipitated solids.

[0102] Likewise, the V / Al removal is carried out under an inert gas or nitrogen atmosphere, preferably under a nitrogen blanket, to prevent oxidation of ferrous iron to ferric iron and precipitation of ferric iron in this step.

[0103] As previously mentioned, for the first leach treatment stream, the recovery of vanadium and aluminium from the V / Al precipitated solids can be carried out by any suitable method known in the art, for example using an ammonia or HC1 solution leach followed by a precipitation and optional calcination step. Details of these process steps will be described in more detail later in the specification.

[0104] Where applicable, the Mg / Mn removal can also be applied to the second leach stream. In these embodiments, the second treatment stream further comprises the steps of:

[0105] A neutralising agent, preferably lime, and an oxidising agent, preferably H2O2 or an oxygen containing gas, more preferably air, are added to the iron removal liquor at a temperature of 60 to 90 °C to raise the pH of the liquor to 9 to 10, thereby causing Mg and / or Mn to precipitate to produce a Mg / Mn removal slurry; and

[0106] The Mg / Mn removal slurry is separated into a liquid portion comprising a Mg / Mn removal liquor and a solid portion comprising precipitated Mg and / or Mn solids.

[0107] In this step, lime is preferably used as the neutralising agent. An oxidising agent, preferably H2O2 or an oxygen containing gas such as air, is added to assist in the oxidation of Mn(II) to Mn(IV). The precipitated Mg and / or Mn solids will typically comprise a mixture of Mg(OH)2and Mn-oxides / hydroxides. The Mg / Mn removal liquor will primarily be a calcium chloride solution.

[0108] To aid process economics, the second leachant is regenerated and recycled to the second leaching step. The regeneration of the second leachant for recycling to the second leaching step preferably comprises:

[0109] The chloride content of the treated second leachate is concentrated by removal of water, preferably by boiling and / or evaporation, to produce a concentrated chloride solution (in some cases, at a concentration below its saturation concentration);

[0110] The evaporated liquor is reacted with at least 98% w / w sulphuric acid at atmospheric conditions at a temperature of 30 to 90 °C, preferably 80 to 85 °C, to produce a mixed chloride solution having 20 to 22% w / w hydrochloric acid, additional chloride content in solution, and a solid precipitate,

[0111] separating the precipitated solids from the mixed chloride solution; and

[0112] recycling the mixed chloride solution to the second leach step.

[0113] The composition of the chloride content will depend on the composition of the additives that have been fed to the second leach process stream. In many cases, the chloride content will comprise a calcium chloride solution / liquid. Thus, the evaporation liquor will comprise a calcium chloride liquid. In such embodiments, the evaporation liquor comprises a calcium chloride liquid in some cases at a concentration below its saturation concentration, and the evaporation liquor is reacted with concentrated sulfuric acid (98% w / w) in a stoichiometric ratio of calcium chloride to sulfuric acid to produce HC1 and a precipitate comprising at least one of gypsum, hemihydrate or an anhydrite compound.

[0114] In the present invention, the focus of the HC1 regeneration step is typically 20% to 22% w / w HC1 resulting from the reaction of a CaCl2 solution at a temperature preferably above 75°C with concentrated H2SO4 to produce primarily anhydrite calcium sulfate or as a mixture of anhydrite, hemihydrate and dihydrate calcium sulfate. In embodiments where the chloride content comprises calcium chloride, the reaction between the evaporation liquor and concentrated sulfuric acid is preferably carried out in a temperature range of 80°C to 85°C so that only anhydrite is precipitated.

[0115] In some embodiments, the treatment stages of the first leach liquor treatment step and the second leach liquor treatment step can be combined for at least one of a vanadium and / or aluminum removal step, an iron removal step, or a manganese and / or magnesium removal step. In some embodiments, all or common stages of the first leach treatment stream and the second leach treatment stream, including V / Al removal, Fe removal and Mg / Mn removal, are combined to be carried out in one common treatment stream to reduce capital investment and operating costs.

[0116] It will be appreciated that the solid and liquid elements in the process can be separated using any suitable method. Techniques for such separations are known, for example using pressure or vacuum filters, reverse decantation, thickeners or centrifuges.

[0117] In some embodiments, the process of the present invention includes a sulfuric acid production plant that produces sulfuric acid from elemental sulfur. This additional process can provide a significant energy credit for the power and heat required for the various steps of the process.

[0118] In some embodiments, the neutralizing agent in the various process steps / stages comprises MgO. In these embodiments, the process typically further comprises a Mg removal step in which lime is used to precipitate Mg(OH)2, and a MgO regeneration stage in which the Mg(OH)2 is calcined, preferably at 300 to 400°C, to regenerate MgO that is recycled as the neutralizing agent in the process.

[0119] A second aspect of the present invention provides a processing system for recovering titanium dioxide from a titanium-containing material, the system comprising the steps of:

[0120] a first leach vessel for leaching the titanium-containing material with a first leach agent in a first leach step at atmospheric pressure and a temperature of 70°C to 97°C to produce a first leach solution comprising unsolved first leach solids containing a titanium content and a first leach liquor, the first leach agent comprising hydrochloric acid at a concentration of less than 23% w / w;

[0121] a first solid-liquid separator for separating the first leach liquor from the unsolved first leach solids;

[0122] a second leach vessel for leaching the first leach solids with a second leach agent and a Fe powder reductant additive in a second leach step at atmospheric pressure and a temperature of 60°C to 80°C to produce a second leach solution comprising unsolved second leach solids and a second leach liquor comprising leached titanium content and iron content, the second leach agent comprising a mixed chloride solution comprising less than 23% w / w hydrochloric acid and an additional chloride selected from an alkali metal chloride, magnesium chloride and calcium chloride or mixtures thereof;

[0123] a second solid-liquid separator for separating the second leach liquor from the unsolved second leach solids;

[0124] a first precipitation vessel for precipitating titanium dioxide from the second leach liquor by adding heated or boiling water to raise the temperature of the second leach liquor to 85°C to 100°C under an inert gas or nitrogen atmosphere to produce a treated second leach liquor and a titanium dioxide containing solid;

[0125] separating the titanium dioxide containing solid from the treated second leach liquor;

[0126] a second precipitation vessel for precipitating the iron content from the treated second leach liquor by adding a neutralising agent and an oxidising agent to the treated second leach liquor to raise the pH of the second leach liquor to 4 to 8 at a temperature of 70°C to 90°C to produce an iron removed slurry comprising an iron removed second leach liquor and an iron precipitate solid;

[0127] a third solid-liquid separator for separating the iron removed second leach liquor from the iron precipitate solid; and

[0128] a regeneration stage for regenerating the second leach agent for recycling to the second leach step,

[0129] thereby recovering the titanium as titanium dioxide from the second leach solution.

[0130] In this second aspect, the neutralizing agent added to the second precipitation vessel can comprise any suitable neutralizing substance or compound, and preferably comprises at least one of limestone, lime or MgO. Furthermore, the oxidizing agent in the second precipitation vessel can comprise one of an alkali metal peroxide, an alkali metal perchlorate, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, an alkali metal chlorate, chlorine gas, an alkali metal hypochlorite, hydrogen peroxide, perchloric acid, an oxygen-containing gas such as air or oxygen, other non-sulfur containing oxidizing agents, or mixtures thereof. Preferred oxidizing agents are H2O2 or oxygen-containing gases, e.g. oxygen or air, etc. The most preferred oxidizing agent is oxygen or air.

[0131] It will be appreciated that the treatment system of the second aspect of the present application can carry out the method of the first aspect of the present application. Features and additional method steps / phases taught in relation to the first aspect of the present application are equally applicable to this second aspect of the present application.

[0132] The third aspect of the present application provides an apparatus comprising the method of the first aspect of the present application.

[0133] The fourth aspect of the present application also provides titanium dioxide produced by the method of the first aspect of the present application.

[0134] Some advantages of this two-stage leaching method are as follows:

[0135] i) The method of the present application is based on atmospheric precipitation technology at less than 100°C, which means that the capital investment is lower compared to methods with high temperature extraction methods and / or solvent extraction method steps.

[0136] ii) In this method, the reactor design criteria do not require specialized building materials. Standard glass fiber and / or high-density polyethylene (HDPE) and / or polypropylene (PP) tanks can be used to meet the requirements of the reactor / apparatus. Compared to the prior art of thermal or high-temperature hydrolysis technology, the leaching agent regeneration in the present application is a simpler method in which the energy requirement is low and the building material is also not important (i.e. materials resistant to high-temperature operation and high-temperature corrosion are not required).

[0137] iii) The entire method is operated with low or lower concentrations of hydrochloric acid in the concentration range of 20 to 22% w / w HC1.

[0138] iv) The HC1 required in this method is regenerated from the treatment solution containing calcium chloride using sulfuric acid under atmospheric conditions. BRIEF DESCRIPTION OF DRAWINGS

[0139] The present application will now be described with reference to the accompanying drawings which illustrate particular preferred embodiments of the present application, in which:

[0140] Figure 1is a general flow chart showing the method steps of a preferred embodiment of the method of the present application.

[0141] Figure 2 is a general flow chart showing the method steps of another embodiment of the method of the present application, which is Figure 1 an improved process flow chart of the method shown.

[0142] Figure 3 is a general flow chart showing the method steps of another embodiment of the method of the present application, which is Figure 2 an improved process flow chart of the method shown, including combined method steps.

[0143] Figure 4 is a general flow chart showing the method steps of another embodiment of the method of the present application, which is Figure 3 an improved process flow chart of the method shown, including two-stage second leaching steps.

[0144] Figure 5 A graph is provided illustrating the extraction of Fe and Mg in a one-stage leaching test with 21% w / w and 17.5% w / w HCI.

[0145] Figure 6 A graph is provided illustrating the extraction of V and Al in a one-stage leaching test with 21% w / w and 17.5% w / w HCI.

[0146] Figure 7 A graph is provided illustrating the concentration of Ti during one-stage leaching with 21% w / w and 17.5% w / w HCI.

[0147] Figure 8 A graph is provided illustrating the online pH and ORP versus time for an Fe(III) reduction test with 1.17 times stoichiometric Fe sand added at 70°C.

[0148] Figure 9 A graph is provided illustrating the Fe precipitation behavior in a Ti02precipitation neutralization liquor using limestone as a neutralizing agent, airflow > 5 L / min at 80°C.

[0149] Figure 10 A graph is provided illustrating the effect of temperature on the extraction of Fe, V and Ti in a titanomagnetite concentrate under leaching conditions of 20% w / w slurry density, 20.1% w / w HCI concentration and 4 h.

[0150] Figure 11 A graph is provided illustrating the extraction of metals from a titanomagnetite concentrate with 20.4% w / w slurry density and 19.8% HCI solution at 85°C for 2 h.

[0151] Figure 12A graph illustrating the Ti analysis of the extract from a secondary extraction experiment using the residue from a primary extraction is provided. Detailed Implementation

[0152] The method of this invention relates to the recovery of titanium dioxide from titanium-containing materials. "Titanium-containing materials" can be any material, including materials containing titanium substances, such as titanium-containing ores. Titanium can be found in a variety of titanium-containing minerals, including ilmenite (FeO·TiO2 or TiFeO3), rutile (TiO2), anatase (TiO2), and / or leucoxene (Fe2O3·nTiO2). Such titanium-containing materials may also typically include iron, vanadium, aluminum, and manganese, which may also be dissolved in the leaching agent applied in the leaching step. Titanium-containing materials can be titanium-containing ore materials, including titanium-containing ores or ore bodies, their concentrates, their modified ores, and their tailings and mixtures thereof. Titanium-containing materials can also be materials comprising vanadium associated with titanium minerals such as titanomagnetite, vanadium-containing minerals, and titanium-containing leaching residues and slag. However, it should be understood that the invention is not limited to any of those materials and may include other materials containing titanium or titanium-containing substances.

[0153] The method described in this invention is a two-stream leaching method for processing titanium-containing ores, concentrates, etc. Figure 1 And as specified in the related instructions (stream-1 and stream-2), where each stream can be operated independently with its own dissolved valuable metal recovery and HCl leaching agent regeneration. This document describes a method for recovering valuable metals, including titanium, vanadium, aluminum, and iron, from titanium-containing feedstocks by leaching with hydrochloric acid (HCl) and leaching with a mixed solution of hydrochloric acid and additional chloride. The entire method is described as operating through two leaching treatment streams at reduced HCl concentrations (below 23% w / w HCl).

[0154] Figure 1 A general flow chart of one embodiment of the method of the present invention is shown, illustrating a two-step leaching method 100 for recovering titanium dioxide from titanium-bearing ore or concentrate 101. As described above, titanium-bearing ore or concentrate 101 comprises titanium, and in this case includes additional valuable metals such as iron, vanadium, manganese, magnesium, and aluminum. The described and illustrated methods have been adapted to recover the titanium contents and each of those valuable metals. It should be understood that different method steps may be used depending on the valuable metal composition of a particular titanium-bearing material. Therefore, the method of the present invention may include, but should not be strictly limited to, the following steps:

[0155] Stream-1 (first leach treatment stream):

[0156] The method for the first extraction treatment stream 110 (stream-1) is as follows:

[0157] i) A first leach step 120 of the titanium-containing material (in this embodiment, titanium-containing ore 101) is conducted at 70 to 97 °C, preferably at 85 to 97 °C in a first leachant solution comprising 20% to 22% w / w HC1 to leach impurities including vanadium, leaving the titanium values in the first leach solids (solid leach residue). Following the leach process, the first leach slurry is subjected to a solid-liquid separation (part of step 120) to separate the first leach solids 125 and the first leach liquor 126. The first leach solids 125 will contain most of the Ti-containing minerals, such as ilmenite and any precipitated rutile, pseudorutile, anatase, etc., as well as gangue minerals that were not dissolved during the first leach process. The first leach solids 125 are further treated in the second leach treatment stream 210 (stream-2), which will be described in more detail below.

[0158] ii) The excess free acid (HC1) remaining in the first leach liquor 126 after the first leach is neutralised in a neutralisation stage 130, preferably at a temperature of 70 to 97 °C. Here, the first leach liquor 126 is fed to a neutralisation vessel, and a neutralising agent 131 is added to minimise the free acid concentration in the first leach liquor to less than 5 g / L (pH < 0.5). The neutralising agent 131 is preferably the feed Ti-containing ore or concentrate to minimise the free acid concentration in the first leach liquor 126. However, another neutralising agent such as limestone, lime or MgO can be used, but it is to be appreciated that the reagent consumption can be high to achieve the required pH. However, the initial neutralising agent 131 is preferably the feed Ti-containing ore or concentrate, followed by a small amount of another neutralising agent such as limestone, lime or MgO to achieve the required pH in the first leach liquor. The first liquor neutralisation slurry is then subjected to a solid-liquid separation to provide: solids 135, which are fed to the first leach stage 120; and an acid neutralised first leach liquor 136, which is supplied to the Fe powder reduction stage 140.

[0159] iii) In the Fe powder reduction stage 140, the acid neutralised first leach liquor 136 is reduced by the addition of metallic Fe powder 141 at 45 to 75 °C under a nitrogen blanket to convert the ferric chloride present in the liquor to ferrous chloride. This step is conducted to achieve an oxidation reduction potential (ORP) of the liquor of less than 100 mV. The resulting reduced first leach liquor is then subjected to a solid-liquid separation to remove any unreacted solid Fe powder 145 and obtain a reduced liquor 146. The unreacted solid Fe powder 145 can be recycled for use in the reduction stage 140.

[0160] iv) The reduced liquor 146 is then fed to a vanadium and aluminium removal stage 150, wherein the pH of the reduced liquor 146 is raised to ~3 to 6 at 50 to 80 °C by the addition of limestone or lime as neutralizing agent 151, thereby precipitating vanadium and aluminium from the reduced liquor 146 under a blanket of nitrogen. The resulting vanadium (V) and aluminium (Al) removal slurry is then subjected to solid-liquid separation to separate the precipitated solids 155 and V / Al removal liquor 156.

[0161] v) The recovery of V as vanadium pentoxide (V205) from the V / Al precipitated solids 155 can be carried out by any suitable recovery method known in the art. In a preferred embodiment, V as vanadium pentoxide (V205) can be recovered from the V / Al precipitated solids 155 in a recovery process (not shown) by leaching the precipitate in an ammonia solution at higher temperatures to dissolve V as ammonium metavanadate, followed by solid-liquid separation of the slurry to separate the ammonium metavanadate solution and Al rich un-dissolved solids. The ammonium metavanadate is precipitated by cooling the hot solution and the slurry is subjected to solid-liquid separation by filtration. The ammonium metavanadate solids are calcined at above 250 °C to produce the V205product. Alternatively, V can be recovered as vanadium pentoxide (V205) from the V / Al precipitated solids 155 in an alternative recovery process (not shown) by leaching the precipitate in an HC1 solution at ~50 °C to 80 °C to dissolve V and Al, followed by solid-liquid separation to obtain a clean liquid. The prepared V and Al leach liquor can be treated with an organic solvent Cyanex 372 to extract V into the organic solvent leaving Al in the raffinate. The V loaded organic is stripped with an HC1 solution to obtain a V rich strip liquor, and the regenerated organic is recycled to the extraction stage after washing. The V strip liquor is further treated with ammonia to precipitate V as ammonium metavanadate and the slurry is subjected to solid-liquid separation by filtration. The ammonium metavanadate precipitate is washed and calcined at above 250 °C to produce the V205product.

[0162] vi) The V / Al removal liquor 156 is fed to an iron recovery stage 160, wherein the solution pH is changed to ~4 to 7 by the addition of limestone or lime as neutralizing agent 161 in the presence of air 162 (oxidizing agent for the precipitation reaction), thereby precipitating iron from the V / Al removal liquor 156 at a temperature of 70 to 90 °C. The precipitated iron removal solids 165 comprise mainly magnetite, goethite, hematite and tetragonal goethite. However, magnetite is the most preferred precipitate for this stage, and therefore conditions are preferably optimized to substantially precipitate magnetite. The resulting iron removal slurry is then subjected to solid-liquid separation to separate the precipitated Fe solids 165 and Fe removal liquor 166.

[0163] vii) The iron removal liquor 166 is then fed to a Mg and Mn removal stage 170, where magnesium and / or manganese are precipitated from the liquor at pH ~ 9 to 10, temperature of 60 to 90 °C using lime as neutralizing agent 171 and air 172 as oxidizing agent, typically used to oxidize Mn(II) to Mn(IV). The precipitated Mg / Mn removal solids 175 will comprise a mixture of Mg(OH)2and Mn-oxides / hydroxides. The resulting Mg and Mn removal slurry is then subjected to solid-liquid separation to separate the precipitated solids 175 and the Mg / Mn removal liquor 176, which will be predominantly calcium chloride solution.

[0164] viii) The Mg / Mn removal liquor 176, which is predominantly calcium chloride, is fed to an evaporation stage 180 to evaporate the water content to obtain a suitable calcium chloride concentration prior to the subsequent HCl regeneration stage 190. Evaporation is typically achieved by heating / boiling the Mg / Mn removal liquor 176 by the addition of heat 181.

[0165] ix) The evaporated calcium chloride liquor 186 is reacted with concentrated sulfuric acid (98% w / w) 191 in a regeneration stage 190 at the stoichiometric ratio of calcium chloride to sulfuric acid to produce 20 to 22% w / w hydrochloric acid and precipitate calcium as gypsum, hemihydrate, anhydrite compounds or a mixture of these compounds. The reaction can be carried out at atmospheric conditions in the temperature range of 30 to 90 °C. The reaction between the evaporated liquor 186 and the concentrated sulfuric acid 191 is preferably carried out in the temperature range of 80 to 85 °C, thereby precipitating only anhydrite. The resulting regenerated hydrochloric acid slurry is then subjected to solid-liquid separation to separate the precipitated solids 195 and the hydrochloric acid liquor 196, which is recycled back to the first leaching stage 120 for use as the first leaching agent solution.

[0166] Stream-2 (second leach treatment stream):

[0167] The method for the second leaching treatment stream 210 (stream-2) is as follows:

[0168] i) the second leaching stage 220 of the first leached solids 125 from the first leaching stage 120 is performed using a mixed chloride second leaching agent solution of 20% to 22% w / w HC1 and calcium chloride (CaCl2) solution with a total chloride concentration of 400 to 550 g / L at 60 °C to 80 °C, preferably at 70 to 80 °C, for a duration of 4 to 6 h with the addition of Fe powder 221. The Fe powder is added 221 to provide a reducing atmosphere to achieve higher Ti extraction from the first leached solids 125 and to aid in the dissolution of Ti minerals such as ilmenite, rutile, pseudorutile, and anatase, among others. Following the leaching process, the resulting second leached slurry is subjected to solid-liquid separation (part of step 220) to separate the second leached solids 225 and the second leached liquid 226. The second stage second leached slurry / solids 225 exits the process as a tailings. The second leached liquid 226 is fed to subsequent process steps of the second leaching process stream 210.

[0169] In some embodiments, the Fe powder 221 is added throughout the 4 to 6 hour leaching process. In other embodiments, the second leaching stage 220 is performed as two separate leaching campaigns. In these embodiments, the first leaching campaign (initial second leaching) will be performed in the mixed chloride solution without the addition of any Fe powder 221 for 1 to 2 h to dissolve a majority of the ilmenite minerals. This is followed by a second leaching campaign that includes continued leaching in the mixed chloride solution with the addition of Fe powder 221 for an additional 2 to 4 h to dissolve the remaining unreacted ilmenite and other Ti-containing minerals, such as rutile, pseudorutile, anatase, etc., from the first leached solids 125 under a reducing atmosphere. The Fe powder 221 will also reduce the trivalent iron present in the leach liquor to divalent iron during the leaching reaction.

[0170] In other embodiments, for example as Figure 4As shown (and described in more detail below), the second leach stage 220C is completed as a two-stage reaction, where the first stage second leach (222C) includes dissolution of most of the ilmenite mineral without addition of Fe powder 221C, and the second stage second leach (223C) includes reaction of the first stage second leach solids with Fe powder to dissolve the remaining unreacted ilmenite mineral and other Ti-bearing mineral phases. If desired, an additional small amount of fresh second leachant solution (20-22% w / w HC1 and calcium chloride (CaCl2) solution mixed chloride solution) can be added in the second stage second leach (223C) to stabilise the dissolved metals and additional iron from the added Fe powder 221C. The resulting second leach slurry is subjected to a solid-liquid separation to separate the Ti-rich second leach liquor 227C and second leach solids 225C. The second stage second leach slurry / solids 225C exit the process as tailings. The first stage second leach (222C) preferably includes a solid / liquid separation stage to separate the first stage second leach slurry / solids 224C and the first stage second leach liquor 226C, where the first stage second leach slurry / solids 224C are fed to the second stage second leach 223C and the first stage second leach liquor 226C is fed to the subsequent oxidation stage 230C to be combined with the second stage second leach liquor 227C, producing a combined second leach liquor, which is treated as the Ti-rich second leach liquor 226 (as Figure 1

[0171] ii) treating the Ti-rich second leach liquor 226 with dilute H2O2 231 in an oxidation stage 230 to oxidise any Ti(III) content of the liquor to Ti(IV) by controlling the oxidation-reduction potential of the liquor to be within 100-200 mV, producing an oxidised Ti(IV) liquor 236.

[0172] iii) Ti is then recovered from the oxidised Ti(IV) liquor 236 (as titania) by adding heated / boiling water 241 to the oxidised Ti(IV) liquor 236 to hydrolyse the Ti(IV) and thereby precipitate this content as titania (TiO2) solids. The reaction is preferably conducted under an inert atmosphere such as a nitrogen blanket to prevent oxidation of ferrous iron to ferric iron in the liquor and thereby prevent unwanted precipitation of ferric iron with TiO2 in the washing stage of the TiO2. The resulting Ti(IV) hydrolysis will release HC1 in solution. In some embodiments, this additional HC1 can be partially neutralised by the addition of a limestone / lime slurry to maximise TiO2 recovery Figure 1 The resulting TiO2 slurry is then subjected to a solid-liquid separation to separate the TiO2 solids 245 and the ferrous-containing liquor 246.

[0173] iv) as​Figure 2 As shown, depending on the V and Al concentrations in the Ti02precipitate, the ferrous liquor 246 can be subjected to a V / Al removal stage 250, wherein the pH of the ferrous liquor 246 is raised to ~3 to 6 at 50 to 80 °C by the addition of limestone or lime as neutralizing agent 251, thereby precipitating vanadium and aluminum under a blanket of nitrogen gas. The resulting removed vanadium (V) and aluminum (Al) slurry is then subjected to a solid-liquid separation to separate the precipitated solids 255 and V / Al removal liquor 256. Vanadium (V) and aluminum (Al) can be recovered from the V / Al precipitated solids by employing the same procedure as described for the first leach liquor.

[0174] If the V and Al concentrations in the Ti02precipitate are high, the ferrous liquor 246 of the second leach treatment stream 210 can be sent to / combined with the V / Al precipitation stage 150 of the first leach treatment stream 110 (as shown in Figure 3 and 4 ), to simplify / optimize the method of application of these treatment stage operations when needed, as the further downstream steps are the same (see below) prior to the HC1 regeneration stage.

[0175] v) Feeding the ferrous liquor 246 to an iron recovery stage 260, wherein the solution pH is changed to ~4 to 7 by the addition of limestone or lime as neutralizing agent 261 in the presence of air 262 (oxidizing agent for the precipitation reaction), thereby precipitating iron from the liquor at a temperature of 70 to 90 °C. Again, the precipitated iron removal solids 265 mainly comprise magnetite, goethite, hematite and tetragonal goethite. However, magnetite is the most preferred precipitate of this stage. The conditions are preferably optimized to substantially precipitate magnetite. The resulting iron removed slurry is then subjected to a solid-liquid separation to separate the precipitated Fe solids 265 and the Fe removal liquor 266, which is mainly a calcium chloride solution.

[0176] vi) Feeding the calcium chloride solution 266 to an evaporation stage 270 to evaporate the water content to obtain a suitable concentration of calcium chloride prior to the subsequent HC1 regeneration stage 290. Evaporation is usually achieved by heating / boiling the calcium chloride solution 266 by the addition of heat 271. In this evaporation stage 270, water is partially evaporated to produce a concentrated calcium chloride solution 276 having the concentration of calcium chloride required for the second leachant generation.

[0177] vii) In case the liquor contains magnesium and / or manganese content, the second leach treatment stream 210 can include a Mg and Mn removal stage 280. Here, the content is drained from the treatment line into a drain line 279 (as shown in Figure 1 and 2 ), or using a dedicated stage (as shown in Figure 3 and 4). In each case, the liquid is fed to the Mg and Mn removal stage 280, where magnesium and / or manganese is precipitated from the liquid at pH ~ 9 to 10, temperature of 60 to 90 °C using lime as neutralizing agent 281 and air 282 as oxidizing agent, which is commonly used to oxidize Mn(II) to Mn(IV). The precipitated Mg / Mn removal solids 285 will comprise a mixture of Mg(OH)2and Mn-oxides / hydroxides. The resulting Mg and Mn removal slurry is then subjected to solid-liquid separation to separate the precipitated solids 285 and the Mg / Mn removal liquid 286, which will be predominantly a calcium chloride solution.

[0178] viii) The concentrated calcium chloride solution 276 is reacted with concentrated sulfuric acid (98% w / w) 291 in a regeneration stage 290 at a desired stoichiometric ratio of calcium chloride to sulfuric acid in the liquid to regenerate the second leaching agent (equivalent to 20% to 22% w / w hydrochloric acid solution, with the remaining calcium chloride left in solution) and simultaneously precipitate a gypsum, hemihydrate or anhydrite compound or a mixture of these compounds. The reaction can be carried out at a temperature range of 30 °C to 90 °C. The reaction can be carried out at atmospheric conditions at a temperature range of 30 °C to 90 °C. The reaction between the concentrated calcium chloride concentrated solution 276 and the concentrated sulfuric acid 291 is preferably carried out at a temperature range of 80 °C to 85 °C, so that only anhydrite is precipitated. The resulting mixed chloride regeneration slurry is then subjected to solid-liquid separation to separate the precipitated solids and the regenerated second leaching agent solution 296. The regenerated second leaching agent solution 296 is recycled back to the second leaching stage 220.

[0179] It will be appreciated that the neutralizing agent 151, 161, 251, 261 used in the Al / V removal stage 150, 250 and the Fe removal step 160, 260 in stream-1 110 and stream-2 210 of the process 100 can be limestone or lime (as discussed), and / or in other embodiments is MgO. Of these neutralizing agents, limestone is the preferred neutralizing agent 151, 161, 251, 261 as it is a low cost reagent.

[0180] Figure 1Limestone or lime is exemplified for use as the neutralizing agent 151, 161, 251, 261. However, it should be understood that when MgO is added for neutralization in these steps 150, 250, 160, 260, it will form MgCh in the associated liquid. Thus, in the case of using MgO in the process 100, the treated liquid will comprise a MgCh containing solution for stream-1 100 and a mixed solution containing CaCh and MgCh for stream-2 210 (when CaCh is used as the additional chloride in the second leaching agent, it can be a MgCh containing solution for stream-2 210, where MgCh is used as the additional chloride in the second leaching agent). Any Mg content will be removed using the Mn / Mg removal steps 170, 280 described in the process 100. However, in these steps, the MgO will need to be regenerated from the Mg removal steps 170, 280 and the regenerated solids will be recycled back to the neutralization steps 150, 250, 160, 260.

[0181] When using MgO as the neutralizing agent, the following process steps are required after the iron removal step:

[0182] If Mn is present in the liquid obtained after the Fe removal step, lime will be used as the neutralizing agent for the Mn removal and Mg removal steps 170, 280, respectively.

[0183] a) First, the Mn removal step will be performed by adding lime to precipitate Mn as an oxide / hydroxide or mixture using the Fe removal liquid 166, 276 in the presence of an oxidizing agent (e.g. air, oxygen, H2O2) at a pH lower than 9. A solid-liquid separation will be performed to obtain a Mn removal liquid and a Mn rich precipitate.

[0184] b) The Mg removal can then be performed after Mn removal by adding lime to precipitate Mg as Mg(OH)2 using the Mn removal liquid at a pH 9 to 10 and this liquid will mainly have CaCh (for HCl regeneration). The Mg(OH)2 obtained after the solid-liquid separation will be calcined at ~300 to 400 °C to regenerate MgO for recycling.

[0185] c) In multiple embodiments, when Mn is not present in the liquid obtained after the iron removal step, only a single Mg removal step will be required as described above.

[0186] When MgCl2 is used as an additional chloride in the second leaching agent, Ti leaching is performed in the second leaching stage 220 using a second leaching agent comprising a mixture of HCl and MgCl2 solutions. Although not shown, it should be understood that the liquid after TiO2 precipitation can be used for iron removal with MgO to produce magnetite (instead of high-temperature FeCl3 / FeCl2 hydrolysis to produce hematite) and the desired MgCl2 solution. As described above, the desired amount of Mg(OH)2 can be precipitated from the MgCl2 solution using lime to obtain a CaCl2 + MgCl2 solution, wherein the CaCl2 concentration should be equal to or higher for regeneration using 20% ​​to 22% w / w HCl with 98% H2SO4. In this case, MgO and the HCl + MgCl2 solution are also regenerated.

[0187] like Figure 2 and 3 As shown, the common stages of the first leaching process stream 110, the second leaching process stream 210, such as V / Al removal 350, Fe removal 360, and Mg / Mn removal 370, can be combined into a single common process stream 310 to reduce capital investment and operating costs. Here, the liquids from the first leaching process stream 110 and the second leaching process stream 210 are combined before V / Al removal 350 and separated into separate process streams before evaporation stages 180 and 270. However, it should be understood that the evaporation stage 180 after Mg / Mn removal may not require the generation of 20% to 22% w / w HCl for the first leaching process stream 110 because the CaCl2 concentration may be sufficiently high due to the mixing of the process liquids before the V / Al precipitation step 250. Therefore, in some embodiments, only a single evaporation stage may be required.

[0188] As described above, in some embodiments, the second extraction stage 220 can be modified to reduce the addition of Fe powder 221 and increase overall Ti extraction by performing the second extraction in two stages. For example... Figure 4As shown, the second stage leaching 220C can include two leaching stages: i) a second leach (SL) 222C; and ii) a reductive second leach (RSL) 223C. The SL 222C is performed without a reducing agent, and the RSL 223C is performed with the addition of Fe powder. The two-stage second leaching step 220C can allow for the elimination / minimization of H2O2 required for Ti(III) oxidation, as the SL leach liquor 226C from the SL 222C containing Fe(III) will oxidize the Ti(III) present in the RSL leach liquor 227C. However, the ratio of the SL leach liquor 226C to the RSL leach liquor 227C needs to be properly adjusted so that only Ti(III) oxidation occurs in the RSL leach liquor 227C, otherwise, if too much RSL leach liquor 227C is added, the Ti(IV) present in the SL leach liquor 226C can be reduced.

[0189] The stages can be performed in suitable process vessels suitable for the leaching, precipitation, boiling and mixing process steps. As previously mentioned, in this process, the reactor design criteria do not require specialized construction materials. Standard glass fiber and / or high-density polyethylene (HDPE) and / or polypropylene (PP) tanks can be used to meet the reactor / equipment requirements. The leachant regeneration in the present invention is a simpler process compared to the prior art hydrothermal or high temperature hydrolysis technologies, in which the energy requirements are low and the construction materials are not critical (i.e. no high temperature and corrosion resistant materials are required).

[0190] It is understood that the solid / liquid separation of all stages can be operated using any suitable method and method equipment. Techniques for such separation are known, for example, using pressure or vacuum filters, countercurrent decantation, thickeners or centrifuges. In a particular embodiment, the solid / liquid separation can be operated using a thickener operation. Washing stages will only be applicable to the solids from the flow loops, for example: i) final leach solids from the second leach; ii) TiO2precipitate; iii) V / Al precipitate; iv) Fe precipitate; v) Mg / Mn removal solids; vi) gypsum solids. Intermediate solids washing from one stage to another within the process is not essential, as the stages of the process should be able to accommodate any entrained liquids with the intermediate solids.

[0191] The products of the process are a high grade titanium dioxide product, and one or more additional valuable metals selected from vanadium, aluminum, iron, magnesium or manganese.

[0192] Example

[0193] Aspects of the dual flow process of the present invention are illustrated by the following examples:

[0194] Example 1 - Recovery of titanium dioxide from a Ti ore concentrate containing ilmenite.

[0195] 1. Experimental methods

[0196] Developed such Figure 1 The experimental process flow diagram shown is for testing a method for recovering titanium and other valuable metals. This method can operate on Ti ore (see composition below) from Western Australia at low HCl concentrations (~20% to 22% w / w HCl) and can also regenerate HCl at low temperatures (<100°C) under atmospheric conditions. As described above, the proposed flow diagram has two main processing streams: stream-1 is studied in a ~20% to 22% w / w HCl system, and stream-2 is studied in a mixed HCl + CaCl2 system with ~20% to 22% w / w HCl and a CaCl2 concentration of ~300 g / L. The different stages covered by the studies of these two streams are:

[0197] • Processing Stream-1: Primary extraction, neutralization of the extract, reduction of the neutralized solution, V / Al removal, Fe removal, Mg / Mn removal, evaporation of the Mg / Mn removal solution (using CaCl2 in the concentrated liquid for HCl regeneration), and hydrochloric acid regeneration.

[0198] • Processing Stream-2: Secondary leaching of the residue from the primary leaching of Stream-1, TiO2 recovery, Fe removal, Mg / Mn removal, evaporation of the Mg / Mn removal solution (to concentrate CaCl2 in the liquid) and hydrochloric acid regeneration.

[0199] 2. Methods and materials

[0200] 2.1 Materials

[0201] In this study, approximately 10 kg of Ti ore concentrate from Western Australia and approximately 1 kg of Fe powder (Fe gravel 120) were used. The concentrate was thoroughly homogenized, and subsamples were collected for analysis. All chemicals used in this study, such as HCl, H₂SO₄, FeCl₃, CaCO₃, and Ca(OH)₂, were laboratory reagent grade.

[0202] 2.2 Experimental Procedure

[0203] 2.2.1 Pre-extraction experiments using dilute HCl, H2SO4, and FeCl3

[0204] Pre-leaching test work was conducted in 0.5 L glass reactors using 5% w / w HC1, 5% w / w H2SO4 and ~150 g / L of FeCl3 solution at ~65 °C and ~20% w / w pulp density for 2 h. The concentrate and the prepared HC1, H2SO4, FeCl3 solution were fed to the reactor and heated in a water bath at 65 °C for 2 h. The final pulp was filtered and the liquid was analyzed for the desired elements by ICP-OES.

[0205] 2.2.2 Treatment Stream-1

[0206] 2.2.2.1 Primary Leaching

[0207] Primary leaching tests were conducted in 2 L and 5 L glass reactors using 17% to 21% w / w HC1 at 95 °C to 97 °C and 20% w / w pulp density for 1 to 4 h. The reactors were fitted with a glass lid connected to a condenser. For the first two tests, the required amount of HC1 solution was placed in the 2 L reactor and the concentrate was added to the reactor at 50 °C to 60 °C. Once the reaction temperature was reached (~95 °C), samples were collected and the reaction was continued for 4 hours with hourly sampling. The samples were filtered, the solids were first top washed with ~15% HC1 solution and then re-slurried / washed with deionized (DI) water. The reactor pulp was filtered at the end of the reaction; the solids were thoroughly washed and dried in an oven at 60 °C.

[0208] Three pilot leaching tests were conducted in a 5 L reactor for 2 h. The required amount of concentrate material and HC1 solution was placed in the reactor and heated to the test temperature. At the end of the test, samples were collected and filtered in a pressure filter. The solids were washed similar to that described for the initial tests. The pilot pulp was filtered in a pressure filter and the liquid was stored in a gas tight bottle. The wet cake was re-slurried with ~2 times the cake volume of ~15% HC1 solution and then the first washed cake was second re-slurried with ~2 times DI water (deionized water). Representative wet cake samples were collected from the second washed cake for moisture determination and chemical analysis. The washed wet cake was stored in a sealed bag for the reduction leaching test work. The solids, final liquid and wash liquid were analyzed for Fe, V, Ti, Al, Mn, Ca, Mg and Si. The collected samples and free acid in the final liquid were analyzed. The final liquid and the second washed cake from the three pilot leaching tests were homogenized separately and stored in a gas tight container. The homogenized liquid was used for further processing and the cake was used for Stream-2 leaching test work.

[0209] Primary leaching tests were also conducted in a 2 L reactor using regenerated HC1 from Treatment Stream-1 following the same conditions and procedures as the pilot leaching tests. The test was conducted for 2 h and no samples were collected.

[0210] 2.2.2.2 Neutralization of primary leach liquor

[0211] The free acid analysis report showed that the acid concentration in the homogenized primary leach liquor was very high (~ 140 g / L). The majority of the free acid in the leach liquor was neutralized by adding Ti concentrate material in a 5 L reactor. The final slurry was filtered and the liquid was stored for further processing.

[0212] The Ti concentrate neutralized liquor was further treated with limestone to neutralize the remaining free acid, resulting in < 5 g / L free acid in the liquid.

[0213] 2.2.2.3 Iron reduction of neutralized leach liquor

[0214] Reduction trials were performed on the concentrate / limestone neutralized leach liquor in 0.5 L and 5 L glass reactors (equipped with pH and ORP probes) at 70 °C. A calculated amount of Fe gravel 120 was slowly added to the reactor containing the concentrate / limestone neutralized liquor for Fe(III) reduction. The trials were performed under a nitrogen blanket during the addition of the Fe gravel. The online pH and ORP (Oxidation Reduction Potential) were recorded continuously until the ORP of the liquid was found to be negative and relatively stable. The slurry was filtered in a filter press and the solids were re-slurried / washed with water and dried in an oven. The final liquid was stored in a gas tight bottle under a nitrogen blanket for further trial work. Both the solids and the liquid were submitted for analysis.

[0215] 2.2.2.4 Removal of aluminum and vanadium from reduced liquor

[0216] The Al / V removal trials were performed on the reduced liquor in 0.5 L and 5 L glass reactors (equipped with pH and ORP probes) at 70 °C by increasing the pH of the reduced liquor with limestone. Limestone slurry was slowly added to the reactor at 70 °C and the online pH and ORP were measured. The trials were performed under a nitrogen blanket to prevent ferrous oxidation. At the end of the trial, the slurry was filtered in a filter press and the solids were re-slurried / washed with water and dried in an oven. The final liquid was stored in a gas tight bottle under a nitrogen blanket for further trial work. Both the solids and the liquid were submitted for analysis.

[0217] 2.2.2.5 Removal of iron from Al / V removal liquor

[0218] Iron removal tests were conducted in a 2 L glass reactor equipped with pH and ORP probes, a thermometer, an air sparge tube, and a condenser. The test solution was heated to the set temperature (80 °C) under a nitrogen blanket to prevent Fe(II) oxidation. Initially, lime or limestone slurry was added to raise the pH of the reactor to a target precipitation pH of ~4.2 to 5.0, and then air sparging was initiated at a flow rate of ~2.0 to 5.0 L / min. The pH of the reactor was maintained by continuous addition of limestone slurry. Samples were collected before air addition was initiated, and then periodically. The collected samples were immediately filtered, the wet cake was thoroughly rinsed with DI water, and dried in an oven at ~60 °C. The iron concentration in the filtrate was determined by analyzing the ferrous concentration using the standard dichromate method. Based on the ferrous analysis, the retention time of the iron removal test was determined. Typically, the test was run for 3.5 to 5.0 h.

[0219] At the end of the test, the slurry was filtered using a pressure filter. The cake was washed with DI water by re-slurrying, and the washed solids were dried in an oven. The solids and liquids were submitted for chemical analysis.

[0220] 2.2.2.6 Removal of magnesium and manganese from the iron removal liquor

[0221] Magnesium and manganese removal was conducted at 60 °C in a 5 L reactor equipped with pH and ORP probes using the iron removal liquor. Dry lime was added slowly to the reactor at 60 °C to raise the pH of the liquor to ~9, and then a calculated amount of 7.5% w / w H2O2 was added for Mn oxidation. The final slurry was filtered, the solids were washed by re-slurrying, and then dried at 60 °C. The solids and liquids were submitted for analysis.

[0222] 2.2.2.7 Regeneration of hydrochloric acid from the Mg / Mn removal liquor

[0223] Prior to the HCl regeneration test work, the Mg / Mn removal liquor was evaporated in a 5 L beaker using a hot plate to achieve the required Ca concentration in the liquor so that >20% w / w HCl could be produced during the HCl regeneration reaction.

[0224] HCl regeneration test work was conducted in 0.5 L and 1 L reactors at 80 to 85 °C by using evaporation to add a calculated amount of 98% w / w H2SO4. Initially, the solution was heated to ~60 to 70 °C and addition of H2SO4 was started. The acid was added slowly / dropped and the temperature rise of the reactor slurry was recorded. Once the reactor slurry reached ~80 to 85 °C, the addition of acid was controlled to maintain the reactor temperature. At the end of the reaction, the final slurry was filtered in a filter press and the filter cake was re-slurried / washed twice with about one volume of filter cake of gypsum saturated water. The solids were dried at ~45 °C. The final liquor was stored for recycle leaching of feed concentrate material. The acid concentration in the final liquor was determined using standard titration analysis. The solids, final liquor, and wash liquor were submitted for elemental analysis.

[0225] 2.2.3 Process Stream - 2

[0226] 2.2.3.1 Secondary Leaching of Primary Leach Residue

[0227] Secondary leaching tests were conducted in 2 L and 5 L glass reactors at 75 °C to 80 °C for 4 to 6 h in HCl-CaCl2 mixed solution using primary leach residue in the absence and presence of Fe Grit 120 reducing agent. The required amount of primary leach wet cake and HCl-CaCl2 solution (with required concentrations of HCl and CaCl2) were placed in the reactor to obtain a slurry density of ~4.9 to 8.8% w / v. The reactor was fitted with a condenser, thermometer, and ORP probe and placed in a hot water bath. The reaction was continued for 1 to 3 h at the test temperature after which ~1.3 to 2 g of Fe Grit 120 was added manually at fixed intervals of ~5 to 10 minutes until the end of the reaction. The online ORP of the reaction was recorded during the leaching of Fe grit addition. Samples were collected at 1 h intervals and filtered immediately in a filter press. The solids were initially re-slurried / washed with 15% w / w HCl and then with DI water. The final slurry was processed similarly to the collected samples. The liquor (filtrate) was diluted immediately for analysis as it was found to crystallize on storage at ambient temperature.

[0228] Two large sample secondary leaching tests were conducted without sampling in a 5 L reactor with Fe grit addition to generate leach liquor for further processing. At the end of the reaction, the slurry was filtered in a filter press and the liquor was stored in air tight bottles at ~60 °C to prevent iron crystallization. The wet cake was initially re-slurried / washed with ~2 volumes of filter cake of 15% w / w HCl and then with DI water to generate wash data. The second wash cake was dried at 60 °C. The diluted final liquor, wash liquor, and solids were submitted for analysis. The leach liquor from both tests was used for TiO2 precipitation test work.

[0229] 2.2.3.2 TiO2 Precipitation

[0230] Titanium dioxide precipitation tests were performed in 0.5 L and 5 L reactors at 90 to 95 °C by hydrolyzing Ti ions in the secondary leach solution in hot water. Initially, the secondary leach solution was oxidized at room temperature with dilute H202 to obtain an ORP of ~ 150 to 200 mV. The required amount of DI water was heated to the test temperature in a reactor equipped with a thermometer and condenser. The oxidized solution was added slowly until the ratio of water to liquid became 1 : 1, and then the slurry was stirred to allow possible agglomeration of Ti02particles. The test was performed under a nitrogen blanket to minimize oxidation of Fe(II). At the end of the test, the slurry was filtered in a pressure filter and the liquid was stored for further processing. The solid was initially washed with 10% to 15% HC1 and then with DI water. The solid was dried at 60 °C overnight. Solid and liquid samples were submitted for analysis. The Ti02precipitation test liquor was combined to generate a bulk liquid for further processing.

[0231] 2.2.3.3 Acid neutralization and removal of iron from the Ti02precipitate liquor

[0232] The Ti02removal liquor reported a high free acid analysis (~ 70 g / L), which was neutralized by the addition of limestone. The acid neutralization liquor was used for iron removal. Iron removal tests were performed in a 5 L reactor using the neutralization liquor following the same procedure described in Section 2.2.2.5. The iron removal test liquor was homogenized for further downstream processing.

[0233] 2.2.3.4 Removal of magnesium / manganese from the Fe removal liquor

[0234] Initially, the Fe removal liquor was partially evaporated (~ 34 mass %) by heating the solution on a hot plate. The partially evaporated liquor was used for Mg / Mn removal following the same procedure described in Section 2.2.2.6.

[0235] 2.2.3.5 Regeneration of hydrochloric acid from the Mg / Mn removal liquor

[0236] The Mg / Mn removal final liquor was further evaporated to reach the required Ca concentration in the liquid so that ~ 20% w / w HC1 could be produced during the HC1 regeneration reaction. The HC1 test was performed in a 2 L glass reactor using the same procedure described in Section 2.2.2.7.

[0237] 3. Results

[0238] 3.1 Chemical and mineralogical analysis

[0239] The analysis of the Ti concentrate from Western Australia is given in Table 1. The elemental analysis was ~34% Fe, 0.34% V, 23.6% Ti, 2.2% Al, 0.8% Mg, 3.7% Si and <0.2% of the analysed Ca, Cr, Cd, Cu, Na, K and Zn. The mineralogy of the concentrate reported ilmenite, hematite, goethite and quartz phases and reasonable amounts of zirconolite and kaolinite minerals.

[0240] Table 1. Analysis of Ti concentrate from Western Australia.

[0241]

[0242] 3.2 Pre-leach test results

[0243] The purpose of the pre-leach was to check if it was possible to remove unwanted impurities prior to the primary leach test. The unwanted impurities were mainly monovalent cations such as Na and K, as these are chloride consuming elements that cannot be recovered. Three pre-leach tests were performed using 5% w / w HC1, 5% w / w H2SO4 and ~150 g / L of FeCl3solution at 65°C and ~20% w / w slurry density for 2 h. The leach solution analysis and the percentage of dissolved metals are given in Tables 2 and 3, respectively.

[0244] Table 2. Leach solution analysis of pre-leach tests using 5% w / w HC1, 5% w / w H2SO4 and 150 g / L FeCl3.

[0245]

[0246] Table 3. Dissolution of metals in pre-leach tests using HC1, H2SO4 and FeCl3solution.

[0247]

[0248] The liquid analysis data indicated that part of the Fe, Al, Mn, Ca, Mg, Na, K, Cu, Zn were dissolved in these tests, with very low Al and Mg dissolution for the FeCl3leach compared to the H2SO4 / HC1 leach. It was reported that about 23 to 40 mg / L V was dissolved for the HC1 and H2SO4leach, which was not ideal for the pre-leach tests. However, no V dissolution occurred in the FeCl3system. Due to the low dissolution of Na and K, it was not found that the pre-leach tests were essential for the Ti concentrate used for the flowsheet development studies. The Na and K analysis in the concentrate was also low (~0.1%), therefore, no further pre-leach tests were performed and the concentrate was used directly for the primary leach test work.

[0249] 3.3 Process flow-1:

[0250] 3.3.1 Primary leach

[0251] The objective of this primary leach was to dissolve as much of the impurities and V as possible, leaving ilmenite intact in the leach residue for secondary leaching. Initially, two primary leach tests were conducted at ~97°C and 20% w / w slurry density using ~21% w / w and ~17.5% w / w HC1 concentrations for 4h. Fe and Mg leach extraction data are shown in Figure Figure 5 Figure Figure 6 shows V and Al leach extraction data. It was found that Fe and Mg extractions were similar at 17.5% and 21% w / w HC1 concentrations. V and Al extractions were slightly higher at 21% w / w HC1 compared to 17.5% w / w HC1. Figure 5 and Figure 6 show that most of the Fe, Mg, V and Al extractions occurred within the first 2h of leaching; this indicates that a 2h leach time should be sufficient to perform a primary leach under the conditions used. It was found that Ti dissolution increased gradually during leaching at 17.5% and 21% w / w HC1 concentrations Figure 7 ). The rate of increase was relatively higher at 21% w / w HC1 compared to 17.5% w / w HC1. At 2h leach time, ~0.5g / L and ~1.0g / L Ti dissolution occurred for 17.5% and 21% w / w HC1 concentrations, respectively. Figure 7 It was clear that there would be some Ti dissolution (at least 0.5g / L Ti) during primary leaching at 17.5 to 21% w / w HC1 concentrations, and it would be difficult to minimise the Ti concentration to below 0.5g / L unless the acid concentration was further reduced. However, reducing the acid concentration would also reduce V extraction. Therefore, considering the higher V extraction in the primary leach, a 21% w / w HC1 concentration was chosen for further primary leach test work to generate a bulk leach solution.

[0252] Three pilot leach tests (PL-3, PL-4 and PL-5) were conducted at 21% w / w HCI for 2 h duration, keeping other conditions constant. The leach test results were found to be reproducible. The leach conditions of 21% w / w HCI, 97°C, 20% w / w slurry density and 1 to 2 h duration resulted in dissolution of 48% Fe, 69% V, ~51% Al, 98% Mg, ~16% Mn, 1.8% Ti and 0.4% Si. The mineralogy of the leach solids reported ilmenite, quartz and zirconolite phases and a small amount of rutile phase. The leach liquor analysis was 41.5 g / L Fe, 0.64 g / L V, 0.9 g / L Ti, 3 g / L Al, 2.1 g / L Mg and <0.1 g / L of Mn, Ca and Si, and free acid concentration of 140 g / L. The leach filter cake analysis reported 24 to 25% Fe, 27% Ti, ~1.5% Al, 4.5% Si, 0.15% V, 0.23% Mn and 0.02% Mg in the solids.

[0253] The leach liquor was treated in Stream-1 for downstream processing and the leach filter cake was used for secondary leaching in Stream-2.

[0254] 3.3.2 Neutralisation of primary leach liquor

[0255] The high free acid of the primary leach liquor was neutralised with Ti concentrate, resulting in a final free acid concentration of ~41 g / L. The metal concentrations in the neutralised liquor increased, resulting in an analysis of ~73 g / L Fe, 1.26 g / L V, 0.4 g / L Ti, 6 g / L Al, 4.1 g / L Mg and <0.2 g / L of Ca, Mg and Si in the liquor.

[0256] The Ti concentrate neutralised liquor was further neutralised with limestone to reduce the free acidity to <5 g / L prior to the Fe(III) reduction stage.

[0257] 3.3.3 Iron(III) reduction of neutralised leach liquor

[0258] The neutralised liquor was treated to reduce Fe(III) to Fe(II) by addition of Fe powder (Fe shot 120). The reduction tests were conducted at 70°C by adding Fe powder (Fe shot 120) in excess of the stoichiometric requirement under a nitrogen blanket to prevent air oxidation of Fe(II). For a typical test with 1.17 times stoichiometric Fe shot addition, Figure 8The pH and ORP profiles are presented in FIG. 2, which shows the increase in pH (to 1.47) and decrease in ORP (to -345 mV) with time. At ambient temperature, the final pH and ORP of the reduced liquor were ~1.9 and -400 mV, respectively. The oxidation-reduction potential (ORP) of the liquor / slurry decreased to a negative ORP with time, while the pH increased and remained below 2. Due to the increase in pH, ~5% V was lost from the solids. However, this V loss could be recovered by dissolving the precipitate in a HC1 solution. The final Fe concentration as Fe(II) in the reduced liquor was ~110 g / L.

[0259] The reduced liquor was treated at 70 °C under a blanket of nitrogen to remove V and Al by increasing the pH with the addition of limestone to precipitate V and Al together.

[0260] 3.3.4 Removal of vanadium and aluminum from the reduced liquor

[0261] V and Al were removed from the reduced liquor containing ~111 g / L Fe, 1.19 g / L V, and 6.1 g / L Al at 70 °C by increasing the pH of the liquor to ~4.0 with the addition of limestone and / or lime under a blanket of nitrogen to precipitate V / Al as hydroxides.

[0262] Initially, two V / Al precipitation trials (V / Al PN-1 and V / Al PN-2) were conducted by adding limestone and lime with / without the addition of H2O2 to understand the precipitation behavior of V and Al. H2O2 was added to increase the ORP of the reduced liquor to ~200 mV, as the initial ORP of the reduced liquor was about -300 mV. Partial oxidation was carried out with H2O2 before the addition of limestone. In the first two trials, the target pH was set to ~4.5. However, due to the partial oxidation / precipitation of Fe(II), the target pH could not be achieved with the addition of limestone alone. Therefore, after the addition of the calculated amount of limestone, an attempt was made to increase the pH by adding a small amount of lime slurry. However, even after the addition of lime, the pH of the reaction slurry remained below ~4. The samples collected after the addition of the calculated amount of limestone were analyzed, which resulted in >96% V and Al precipitation. This indicates that the addition of limestone is effective for V / Al precipitation and lime does not need to be added.

[0263] Based on the first two tests, a third test (V / Al PN-3) was performed by adding limestone only without any H2O2, where over 99% V and Al precipitation occurred from a feed solution having 1.19 g / L V and 6.1 g / L Al with ~111 g / L Fe, 0.45 g / L Mn, 18 g / L Ca, and 4 g / L Mg, resulting in a final solution V and Al analysis of <10 mg / L and 50 mg / L, respectively. A typical V / Al precipitation test solid analysis reported 2.6% V, 15.7% Al, 0.6% Fe, 0.01% Ti, 6.9% Ca, <0.001% Mn / Mg, and 0.3% Si.

[0264] 3.3.5 Removal of iron from V / Al removal solution

[0265] Fe removal was performed at 80°C using the V / Al removal solution (containing ~99 g / L Fe) by adding limestone (15% to 30% w / w slurry density) as a neutralizer and air as an oxidizer at a flow rate of ~5 L / min. The feed solution analysis is provided in Table 5 and the Fe precipitate analysis is provided in Table 6. The test required 3.5 h to completely remove Fe.

[0266] Table 5. Representative feed and final solution analysis for Fe removal test.

[0267]

[0268] Table 6. Representative precipitate solid analysis for Fe removal test.

[0269]

[0270] Effective complete removal of Fe was achieved by precipitating Fe as magnetite, leaving <1 mg / L Fe in the final solution. The iron removal solution analysis reported having <1 mg / L Fe, 184 mg / L Mn, ~53.2 g / L Ca, and ~2 g / L Mg.

[0271] 3.3.6 Removal of magnesium and manganese from iron removal solution

[0272] Mg and Mn removal was performed at 60°C using the Fe removal solution by raising the pH of the solution to ~9 with lime to precipitate Mg as Mg(OH)2, followed by oxidizing Mn(II) to Mn(III) / Mn(IV) with 7.5% H2O2 to precipitate Mn as Mn-oxides. The feed and final solution analysis for the Mg / Mn removal test at 60°C with lime and H2O2 addition is given in Table 7.

[0273] Table 7. Feed and final solution analysis for Mg / Mn removal test at 60°C with lime and H2O2 addition.

[0274]

[0275] Complete removal of Mg and Mn was achieved and the Mg / Mn oxide / hydroxide filter cake analysis was 0.02% Fe, 1.94% Mn, 22.5% Mg and 14.5% Ca.

[0276] 3.3.7 Evaporation of Mg / Mn removal liquor

[0277] The Mg / Mn depleted liquor was evaporated to obtain ~ 130 g / L Ca in the final liquor which was used for HCl regeneration test work.

[0278] 3.3.8 Regeneration of HCl from Mg / Mn removal evaporated liquor

[0279] HCl regeneration was carried out using the evaporated liquor at 80 to 85 °C with the addition of 92% to 97% of the stoichiometric requirement of H2SO4, where the addition of 97% stoichiometry produced 296 g / L HCl (26.6% w / w HCl) and 92% stoichiometry produced 260 to 270 g / L HCl (~ 23 to 24% w / w HCl) concentration. The precipitates produced during the HCl regeneration reaction process were mainly gypsum (CaSO4.2H2O) and some anhydrite (CaSO4) and small amounts of bryozite (CaSO4.05H2O as a hemihydrate). The filter cake washed with ~ 1.2 times gypsum saturated water reported ~ 114 to 126 g / L HCl (~ 11 to 12% w / w) and ~ 50 g / L HCl (~ 4.8% w / w) in the first and second washes, respectively. A third wash can be required for most of the residual entrained HCl based on chloride loss in the second wash filter cake.

[0280] The regenerated HCl was recycled for primary leaching of the Ti concentrate under the same leaching conditions where the leach extraction was ~ 44% Fe, 68% V, 55% Al, 99% Mg and 14% Mn as shown in Table 8. This leach data is very similar to the metal extraction data obtained with fresh 21% w / w HCl in the primary leach.

[0281] Table 8. Metal extraction of primary leach test with recycled regenerated HCl (21% w / w) at 96 to 98 °C at 20% pulp density for 2 h compared to fresh HCl (21% w / w) at 96 to 98 °C at 20% pulp density for 2 h.

[0282]

[0283] The test results for all stages of Process Flow-1 confirm that, Figure 1 Process Flow-1 of the proposed flow sheet shown is metallurgically viable.

[0284] 3.4 Treatment Stream-2:

[0285] 3.4.1 Secondary leaching of primary leach residue

[0286] Treatment Stream-2 begins with the secondary leaching of the primary leach residue with the aim of dissolving Ti minerals from the primary leach residue in the presence of a reductant in an HC1 + CaCl2solution. The composition of the primary leach residue used in the study is provided in Table 9:

[0287] Table 9. Analysis of primary leach residue used for secondary leach test work.

[0288]

[0289] The secondary leach test was conducted at 75 °C and 4.9% w / w pulp density in an HC1 + CaCl2solution with ~7 M HC1 (~21% w / w) and 300 g / L CaCl2for 5 h with ~0.59 g Fe-sand 120 added per gram of dry primary leach residue. The metal extractions and leach liquor analysis are given in Table 10.

[0290] Table 10. Metal extractions and leach liquor analysis for large sample secondary leach test with reductant addition. Conditions: 300 g / L CaCl2in HC1 + CaCl2solution with ~7 M HC1, 4.9% w / w pulp density and 75 °C.

[0291] The leach extraction was ~93% Fe, ~92% V, ~84% Ti and ~94% Mn with the resulting liquor analysis of ~40 g / L Fe, 0.04 g / L V, ~12 g / L Ti, 0.3 g / L Mn and 0.43 g / L Mg and a free acid concentration of ~133 g / L. XRD analysis of the secondary leach filter cake showed one small ilmenite peak and a significantly higher rutile peak, indicating that some of the dissolved Ti can have precipitated during the leach. Analysis of the secondary leach solids was ~6.5% Fe, 0.04% V, ~16.3% Ti, ~3.8% Al, ~16.5% Si and <0.1% Mn, Ca and Mg. Filter cake washing data indicated that two washing stages using 2 to 3 times the filter cake volume of wash solution can be sufficient to remove most of the entrained leach liquor with the second wash liquor resulting in an analysis of ~1 g / L Fe, ~0.25 g / L Ti, ~2 g / L Ca and 1 mg / L V. The secondary leach liquor was stored for downstream processing.

[0292] 3.4.2 Precipitation of Ti02from the secondary leach liquor

[0293] As the secondary leach liquor contains mainly titanium (III) chloride, it is oxidised to titanium (IV) chloride with H2O2 prior to TiO2 precipitation. The H2O2 consumption for the oxidation of Ti(III) chloride is calculated to be ~90 kg H2O2 (30%) per tonne of Ti concentrate, which would equate to 54,000 tonnes of 30% H2O2 per year for the processing of 600,000 tonnes of Ti concentrate sourced from Western Australia.

[0294] TiO2 precipitation is carried out by hydrolysis of titanium (IV) chloride in hot water at 95°C with a leach liquor to hot water ratio of ~1. Feed and final liquor analyses and Ti precipitation data are given in Table 11.

[0295] Table 11. Feed / final liquor analyses and Ti precipitation data for the pilot scale TiO2 precipitation test at ~95°C.

[0296]

[0297] Over 95% Ti precipitation occurred from the feed liquor containing ~12 g / L of Ti with a final liquor analysis of ~0.25 g / L Ti. The mineralogy of the precipitated TiO2 was found to be mainly rutile or a mixture of rutile and anatase. A typical TiO2 sample prepared during the test program was found to be very high purity (>99.5% purity) with a total impurity analysis of 0.24% including elemental analysis of Mg, Ca, Na, K, Al, V, Co, Cr, Cu, Fe, Mn, Mo, Nb, Ni, Pb, Y, Zn, Zr, P, As, Bi, S and Si. Another TiO2 sample prepared during the test work was also pure (>98% purity). However, the iron analysis reported higher (<1%); the total analysis of all impurities except Fe was only ~0.22%. This confirms that high purity TiO2 product can be produced from Ti concentrate using this flow sheet.

[0298] The final liquor was used for further downstream test work.

[0299] 3.4.3 Neutralisation of TiO2 precipitation final liquor

[0300] The TiO2 precipitation liquor was neutralised with limestone as the free acid analysis reported ~70 g / L to minimise the free acid concentration to below 10 g / L. The analysis of the neutralised liquor reported ~20 g / L Fe, ~30 to 40 ppm V / Al, ~0.43 g / L Ti, 0.2 g / L Mn / Mg and 81 g / L Ca. Ideally, in this acid neutralisation step, V and Ti are recovered from the TiO2 precipitation liquor by increasing the pH of the neutralised liquor, with V, Ti and Al precipitating together for further separation. This process step would be the same as the V / Al removal step described in Section 3.3.4 for stream-1.

[0301] There are two options for the treatment of the acid neutralised liquor: i) evaporation to concentrate the liquor prior to Fe removal; or ii) first remove the iron prior to evaporation. In this study, Fe removal was chosen first as evaporation can cause Fe(II) to be converted to Fe(III) which would require re-reduction prior to Fe removal. Therefore, the neutralised liquor was treated to remove Fe in the next step of the process.

[0302] 3.4.4 Removal of iron from the neutralised liquor

[0303] The neutralised liquor was used directly for Fe removal under similar conditions to Stream-1 Fe removal, where complete removal of Fe was achieved from a feed liquor with an analysis of ~20 g / L Fe.

[0304] Three Fe removal trials [FeR(S2)-1 to FeR(S2)-3] were performed at 80 °C and gas flow >5 L / min using limestone as the neutralising agent. The initial pH was increased by the addition of lime or limestone prior to the addition of air. In the first trial, Fe precipitation was performed using a ~20% w / w limestone slurry, but a 25% to 30% w / w limestone slurry was used in the later trials. A higher slurry density of limestone slurry was used in the later trials to reduce the amount of water from the limestone slurry as any additional water added would need to be evaporated in later stages of the process.

[0305] It was found that the Fe precipitation behaviour was almost linear with time Figure 9 ), resulting in complete removal of iron. Similar precipitation behaviour was also observed in the Stream-1 Fe removal stage. The Fe precipitation kinetics were slightly slower for the third trial [FeR(S2)-3], possibly due to the larger volume of feed liquor used (3 L) compared to the other two trials (feed volume ~2.5 L), where the air addition rate was the same. V, Ti and Al were almost completely precipitated during the initial pH increase to ~4 at 80 °C prior to the addition of air, resulting in an analysis of ~1 mg / L V / Al and ~5 mg / L Ti in the liquor. The feed and final liquor analysis and the precipitate solids analysis are given in Tables 12 and 13, respectively.

[0306] Table 12. Feed and final liquor analysis for Fe precipitation trials in Ti02precipitation neutralised liquor using limestone as the neutralising agent at 80 °C, gas flow >5 L / min.

[0307]

[0308] Table 13. Precipitate solids analysis for Fe precipitation trials in Ti02precipitation neutralised liquor.

[0309]

[0310] XRD analysis of Fe(S2)3test solids found goethite and magnetite formed during the Fe removal reaction.

[0311] 3.4.5 Removal of magnesium and manganese from the iron removal liquor and evaporation of the Mg / Mn removal liquor

[0312] The Fe removal liquor was partially evaporated prior to Mg / Mn removal. The Fe removal homogenate was partially evaporated (~34 mass%) by heating. The Ca concentration in the evaporated liquor increased from ~90 g / L to 146 g / L. This liquor was used for Mg / Mn removal with an analysis of ~0.2 g / L Fe, ~0.37 g / L Mg, 0.17 g / L Mn and 146.4 g / L Ca.

[0313] Complete removal of Mg / Mn was achieved from a feed liquor containing ~0.17 g / L Mn and 0.37 g / L Mg at 60 °C by raising the pH of the liquor to ~9 to 10 with lime and oxidising the Mn(II) by addition of dilute H2O2. Removal of Mg and Mn was >99% with 6 mg / L Mg and <1 mg / L Mn in the final liquor. The analysis of the feed and final liquor for the Mg / Mn removal test is given in Table 14.

[0314] Table 14. Feed and final liquor analysis for the Mg / Mn removal test with lime addition at 60 °C.

[0315]

[0316] The Ca analysis in the final liquor was ~150 g / L. The precipitate solids analysis reported ~5% Fe, 4.7% Mn, 10.3% Ca and 11.4% Mg.

[0317] The Mg / Mn removal liquor was further evaporated (~26.5 mass%) to obtain a ~231 g / L Ca concentration in the liquor for the HCl regeneration test work.

[0318] 3.4.6 Regeneration of HCl from the evaporated Mn / Mg removal liquor

[0319] HCl regeneration from the evaporated liquor was performed at ~85 °C using 53% stoichiometric requirement of H2SO4 which produced a 260 g / L HCl [20.4% w / w HCl (~7.1 M)] concentration in the HCl + CaCl2 final liquor while precipitating anhydrite (CaSO4) solids. Filtrate cake washing data using ~2 times the filtrate cake mass of anhydrite saturated water indicated that two washes should be sufficient to recover most of the entrained HCl from the filtrate cake. The first wash and second wash liquor analysis reported ~82 g / L HCl and ~20 g / L HCl respectively.

[0320] The recycled HCl + CaCh solution containing 7 M HC1 and ~245 g / L CaCh was used for secondary leaching, which resulted in low Ti extraction (67%) along with 91% Fe, 71% V, 7% Al, and 99% Mn extraction. Comparing the leaching data of Fe (~92%) and Ti (83% to 84%) obtained in SLR-8 / SLR-9 tests, the Ti extraction was very low in the HC1 + CaCh recycling leaching test. The inventors speculate that there could be two reasons for the low Ti extraction: i) the lower CaCh concentration in the liquor (less total chloride concentration); and ii) no effective reduction behavior in the presence of small mass of slurry (650 g) and small dosage (0.5 g / 10 min) of Fe shot, as the reaction with Fe powder is likely to occur on the surface of the slurry rather than with the bulk of the slurry. The inventors believe that this result can be improved with further process optimization to achieve >85% Ti extraction with the recycling of HC1 + CaCh solution.

[0321] The test results for processing stream-2 all stages confirmed that stream-2 of the proposed flow diagram Figure 1 ) is also metallurgically viable. Thus, the present study successfully demonstrated the operation of the proposed flow diagram Figure 1 ) for a titanium-bearing material, in particular this Ti concentrate sourced from Western Australia.

[0322] Example 2 - Recovery of titanium dioxide from a titanomagnetite concentrate

[0323] 1. Experimental methods

[0324] In this example, this two-step leaching technique was applied to a titanium-bearing magnetite titanium material to examine its leaching behaviour. The vanadium-titanium bearing magnetite concentrate used in the study was sourced from Australia and its composition is described below. The primary leaching of the concentrate was carried out in an HC1 solution, while the secondary leaching of the primary leaching residue was carried out in a mixed solution of HC1 and CaCh.

[0325] 2. Materials and methods

[0326] 2.1 Material

[0327] The analysis of the concentrate is given in Table 15, which shows 52% Fe, 9.5% Ti, 0.57% V, 1.3% Al, ~2% Si, 0.9% Mg, and ~0.3% each of Mn and Ca.

[0328] Table 15. Analysis of the titanium magnetite concentrate.

[0329]

[0330] 2.2 Method

[0331] 2.2.1 Primary leach

[0332] The primary leach of the titanium magnetite sample was conducted in a 2 L glass reactor using 20% w / w HC1 at 70 to 95 °C and 20% w / w slurry density for 2 to 4 h. The reactor was fitted with a glass lid connected to a condenser. The required amount of HC1 solution and concentrate was added to the reactor and placed in an oil (PEG 400) bath. The temperature of the oil bath was raised and once the reaction temperature was reached, the sample was collected and the reaction continued for 2 to 4 hours. The sample was collected and filtered every hour. The solids were initially top washed with ~15% HC1 solution and then re-slurried / washed with deionized (DI) water. At the end of the reaction, the reactor bulk slurry was filtered; the solids were thoroughly washed and dried in an oven at 60 °C.

[0333] 2.2.2 Secondary leach

[0334] The secondary leach trials were conducted in a 2 L glass reactor using the primary leach residue in HC1-CaCl2 mixed solution in the absence and presence of a reducing agent (Fe Grit 120) at 70 °C for 2 h to 4 h. The required amount of primary leach wet cake and HC1 CaCl2 solution was placed in the reactor, which was fitted with a condenser, thermometer, and ORP probe, and placed in a hot water bath. The reaction was continued for 0.5 h at the trial temperature, after which ~1 g of Fe grit 120 was added manually at fixed intervals of ~10 minutes (where appropriate) under a nitrogen blanket until the end of the reaction. The online ORP of the reaction was recorded during the leach with Fe grit addition. Samples were collected at intervals of 0.5 to 1 h and immediately filtered in a filter press. The solids were initially re-slurried / washed with 15% w / w HC1 followed by DI water. The final slurry was processed similarly to the collected samples. The liquid (filtrate) samples were immediately diluted for analysis to avoid any crystallization of the leach liquor upon storage at ambient temperature. The bulk filtrate was stored in an oven at ~60 °C to prevent ferrous chloride crystallization in the trials with Fe grit addition.

[0335] 3. Results and discussion

[0336] 3.1 Primary leach

[0337] First, three primary leach trials were conducted using 20% w / w slurry density in 20.1% w / w HC1 solution at temperatures of 70 °C, 85 °C, and 95 °C for 4 h to detect the dissolution behavior of the titanium magnetite concentrate and to produce leach residues for secondary leaching. Figure 10Fe and V extraction increased by 2h, after which there was no significant increase. Temperature had a higher effect on V extraction, with ~97% V extraction at 95°C at 1h, compared to ~90% V extraction at 70°C and 85°C. Figure 10 It was shown that Ti dissolution occurred mainly during the heating period (up to 1h at 70°C), after which the dissolved Ti precipitated and was reported in the leach residue.

[0338] Residue analysis of these initial leach tests is given in Table 16, which shows a decrease in Fe, V and Al and an increase in Ti and Si analysis as the leach temperature was increased. Secondary leach tests on these residues are given in Section 3.2, where the 70°C and 85°C test residues gave better Ti leach efficiency compared to the 95°C leach residue. The Ti analysis in the 1stleach liquor of the 70°C test was reported at ~0.12g / L, higher than the Ti analysis of the 85°C test leach liquor (0.01g / L). Based on the secondary leach performance and the 1stleach liquor Ti analysis, 85°C and a 2h duration was chosen as the optimal conditions for further 1stleach tests. The leach liquor analysis of the 85°C leach test was 111.5g / L Fe, 1.4g / L V, 0.01g / L Ti, 2g / L Al, 1.9g / L Mg, ~0.29g / L Mn, 0.26g / L Ca, 0.16g / L Si and ~2g / L free HCl.

[0339] Table 16. Leach residue analysis of temperature variation 1stleach tests.

[0340]

[0341] Another 1stleach test was performed at 85°C for 2 hours at a 20% w / w slurry density in 20.1% w / w HCl without sampling to produce enough leach residue for secondary leach tests. Figure 11 The leach extraction is given in, which shows ~79% Fe, 92% V, 59% Al, ~41% Mn, 82% Mg and ~6% Si extraction. The liquor analysis reported 115.6g / L Fe, ~1.6g / L V, 0.04g / L Ti, 2.2g / L Al, 0.3g / L Mn, 0.25g / L Ca, 1.95g / L Mg and 0.24g / L Si in the final leach liquor. The leach residue analysis was 30% Fe, 0.13% V, 25.8% Ti, 1.43% Al, 0.45% Mn, 0.41% Ca, 0.42% Mg and ~3.4% Si.

[0342] 3.2. Secondary leach

[0343] 3.2.1 Preliminary secondary leaching

[0344] Firstly, the tertiary secondary leaching trials used the residues from the 70°C, 85°C and 95°C primary leaching trials, and were conducted at 70°C in a HC1+CaCl2mixed solution with 17% to 18% w / w HC1 (~6 to 6.3M) and 230 to 240 g / L CaCl2at 2.2% w / w pulp density for 4h. No reducing agent was added in these trials. Due to the low pulp density used in the leaching (2.2% w / w), the HC1 and CaCl2concentrations in the HC1+CaCl2mixed solution were kept low in these trials (compared to 20% w / w HC1+300 g / L CaCl2solution). The Ti extraction in these trials is given in Table 17, which shows >98.5% Ti extraction from the 70°C and 85°C primary leaching residues and lower Ti extraction (~91%) from the 95°C primary leaching residue. This indicates that <85°C is a better temperature for primary leaching of titanium magnetite concentrate in order to achieve >98.5% Ti extraction in secondary leaching. The Fe extraction was almost similar (98.5% to 99.6%) for all three primary leaching residues. The Mg extraction for these leaching residues was ~89% to 92%, while the Al extraction decreased with increasing temperature of the primary leaching trials.

[0345] Table 17. Secondary leaching metal extraction under test conditions of 17% to 18% w / w HC1 (~6 to 6.3M) and 230 to 240 g / L CaCl2mixed solution, 2.2% w / w pulp density, 70°C and 4h.

[0346]

[0347] Figure 12 The secondary leaching liquor Ti analysis is given in Table 18, which shows that the Ti analysis was slightly higher in the trial with 85°C leaching residue, otherwise the Ti analysis was almost similar (6.7 to 7.7 g / L). The Ti analysis data indicates that most of the Ti leaching occurred within the first 1 h of reaction, with a slight further increase up to 2h, after which the Ti concentration remained similar. This indicates that a 2h leach should be sufficient to extract most of the Ti from the primary leaching residue. Therefore, further secondary leaching trials were conducted for a duration of 2h.

[0348] 3.2.1 Leaching at higher pulp density

[0349] Further secondary leaching experiments were conducted at higher pulp density of 6.2% w / w in the absence and presence of Fe grit 120 to examine the effect of reducing agent and leaching efficiency at higher pulp density. Leaching experiments were conducted using 85°C primary leach residue under experimental conditions of ~20% w / w HC1 and 300 g / L CaCl2mixed solution, 70°C and 2h. The leach extraction is given in Table 18. It was found that the metal extraction was slightly better in the presence of reducing agent compared to the absence of reducing agent except for Al and Mg. The Ti extraction was 89.8% and 91.2% in the absence and presence of Fe Grit 120, respectively. The leach liquor and leach residue analysis is given in Table 19 and Table 20, respectively.

[0350] Table 18. Metal extraction in secondary leaching at 6.2% w / w pulp density, ~20% w / w HC1 and 300 g / L CaCl2mixed solution, 70°C and 2h experimental conditions.

[0351]

[0352] Table 19. Leach liquor analysis in secondary leaching at 6.2% w / w pulp density, ~20% w / w HC1 and 300 g / L CaCl2mixed solution, 70°C and 2h experimental conditions.

[0353]

[0354] Table 20. Leach solid analysis in secondary leaching at 6.2% w / w pulp density, ~20% w / w HC1 and 300 g / L CaCl2mixed solution, 70°C and 2h experimental conditions.

[0355]

[0356] 4. Conclusions

[0357] It was found that the two-step leaching process is suitable for titanium magnetite concentrate to achieve high Ti extraction in secondary leaching in the presence and absence of Fe powder as reducing agent. The titanium extraction was 90% and 91% in the absence and presence of reducing agent under experimental conditions of 6.2% w / w pulp density, ~20% w / w HC1 and 300 g / L CaCl2mixed solution, 70°C and 2h with Ti analysis in the liquid of ~19.5 g / L. The addition of Fe powder during secondary leaching can be considered as a better option as Fe needs to be present as ferrous in the leach liquor before the leach liquor can be treated to precipitate Ti02.

[0358] The optimum parameters for the primary leaching of titanomagnetite were 85 °C and 2 h in 20% w / w HC1 solution at 20% slurry density, where ~79% Fe, 92% V, 59% Al, ~41% Mn, 82% Mg and ~6% Si extraction occurred.

[0359] Those skilled in the art will appreciate that the application described herein is susceptible to variations and modifications other than those specifically described. It is to be understood that the application includes all such variations and modifications which fall within the spirit and scope of the present application.

[0360] The terms "comprise", "comprises", "comprised", "comprising" as used in the specification (including the claims) are to be construed as specifying the presence of the stated features, integers, steps or components, but not to preclude the presence or addition of one or more other features, integers, steps, components or groups thereof.

Claims

1. A process for the recovery of titanium dioxide from a titanium-containing material, the process comprising the steps of: leaching the titanium-containing material in a first leaching step at atmospheric pressure and a temperature of 70°C to 97°C with a first leaching agent to produce a first leach solution comprising an undissolved first leach solid containing a titanium content and a first leach liquor, the first leaching agent comprising hydrochloric acid at a concentration of less than 23% w / w; separating the first leach liquor from the undissolved first leach solid; leaching the first leach solid in a second leaching step at atmospheric pressure and a temperature of 60°C to 80°C with a second leaching agent in the presence of a Fe powder reducing agent to produce a second leach solution comprising an undissolved second leach solid and a second leach liquor comprising a leached titanium content and an iron content, the second leaching agent comprising a mixed chloride solution comprising less than 23% w / w hydrochloric acid and one or more additional chlorides selected from alkali metal chlorides, magnesium chloride and calcium chloride; separating the second leach liquor from the undissolved second leach solid; precipitating titanium dioxide from the second leach liquor by adding heated or boiling water under an inert gas or nitrogen atmosphere to raise the temperature of the second leach liquor to 85°C to 100°C, thereby producing a treated second leach liquor and a titanium dioxide containing solid; separating the titanium dioxide containing solid from the treated second leach liquor; precipitating the iron content from the treated second leach liquor by adding a neutralising agent and an oxidising agent to the treated second leach liquor at a temperature of 70°C to 90°C to raise the pH of the second leach liquor to 4 to 8, thereby producing an iron removed slurry comprising an iron removed second leach liquor and an iron precipitate solid; separating the iron removed second leach liquor from the iron precipitate solid; and regenerating the second leaching agent for recycle to the second leaching step, thereby recovering titanium as titanium dioxide from the second leach solution.

2. The method of claim 1, wherein, The first leaching step is conducted using a first leaching agent comprising a 20% to 22% w / w HCI solution.

3. The method of claim 1, wherein, The titanium-containing material includes at least one valuable metal selected from iron, vanadium, manganese, magnesium or aluminium, and the first leach liquor is subjected to a step for the recovery of the at least one valuable metal therefrom.

4. The method of claim 3, wherein, The at least one valuable metal includes vanadium and / or aluminium, and the process includes a vanadium and / or aluminium removal step comprising: adding a neutralising agent to the first leach liquor under an inert gas or nitrogen atmosphere at a temperature of 50°C to 80°C to raise the pH of the liquor to 3 to 6, thereby precipitating vanadium and aluminium to produce a V / Al removed slurry, wherein the neutralising agent is selected from at least one of limestone, lime or MgO; and separating the V / Al removed slurry into a liquid portion comprising a V / Al removal liquor and a solid portion comprising a V / Al precipitate solid.

5. The method of claim 3, wherein, The at least one valuable metal includes iron, and the process includes an iron removal step comprising: adding a neutralizing agent and an oxidizing agent to the first leach liquor at a temperature of 70 °C to 90 °C to raise the pH of the liquor to 4 to 7, thereby precipitating iron to produce an iron-removed slurry; and separating the iron-removed slurry into a liquid portion comprising an iron-removed liquor and a solid portion comprising iron precipitate solids.

6. The method of claim 5, wherein, The oxidizing agent comprises at least one of an alkali metal peroxide, an alkali metal perchlorate, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, an alkali metal chlorate, chlorine gas, an alkali metal hypochlorite, hydrogen peroxide, perchloric acid, or an oxygen-containing gas.

7. The method of claim 5, wherein, The iron-removal step is performed after a vanadium and / or aluminum-removal step.

8. The method of any one of claims 5-7, wherein, The at least one valuable metal comprises manganese and / or magnesium, and the method comprises a manganese and / or magnesium-removal step comprising: adding a neutralizing agent and an oxidizing agent to the iron-removed liquor at a temperature of 60 °C to 90 °C to raise the pH of the liquor to 9 to 10, thereby precipitating Mg and / or Mn to produce a Mg / Mn-removed slurry; and separating the Mg / Mn-removed slurry into a liquid portion comprising a Mg / Mn-removed liquor and a solid portion comprising precipitated Mg and / or Mn solids.

9. The method of claim 8, wherein, The manganese and / or magnesium-removal step is performed after the iron-removal step.

10. The method of claim 8, comprising: regenerating the first leaching agent and recycling the first leaching agent to the first leaching step.

11. The method of claim 10, wherein, The first leaching agent is regenerated by: concentrating the chloride content of the Mg / Mn-removed liquor by removal of water to produce an evaporation liquor; reacting the evaporation liquor with at least 98% w / w sulfuric acid at a temperature of 30 °C to 90 °C under atmospheric conditions to produce 20% to 22% w / w hydrochloric acid and a solid precipitate, separating the precipitated solids from the hydrochloric acid liquor; and recycling the hydrochloric acid liquor to the first leaching step.

12. The method of claim 11, wherein, The chloride content comprises calcium chloride and the reaction between the evaporation liquor and concentrated sulfuric acid is performed at a temperature range of 80 °C to 85 °C, thereby precipitating only anhydrite.

13. The method of claim 4, comprising the following steps prior to precipitating vanadium and aluminum from the first leach liquor: neutralizing at least some of the free acid HCl in the first leach liquor by adding at least one of a feed titanium-containing material, limestone, lime, or MgO to the first leach liquor to produce a first liquor neutralization slurry comprising neutralized leach solids; and separating the first liquor neutralization slurry into a solid portion comprising neutralized leach solids and a liquid portion comprising a neutralized first leach liquor.

14. The method of claim 13, comprising the following steps after the neutralization step: reducing the neutralized first leach liquor by adding metallic iron at 45 °C to 75 °C to convert ferric chloride in the first leach liquor to ferrous chloride; and separating the reduced first leach liquor into a liquid portion comprising a reduction liquor and a solid portion comprising any unreacted solid iron powder.

15. The method of claim 14, wherein, The reduction is performed under an inert gas or nitrogen atmosphere and achieves an oxidation-reduction potential ORP of the liquor of less than 100 mV.

16. The method of any one of claims 1 to 7, wherein, The iron precipitate is magnetite. The method of claim 1, wherein the first leach liquor is separated into a solid portion comprising neutralized leach solids and a liquid portion comprising a neutralized first leach liquor.

17. The method of any one of claims 1 to 7, wherein, The second leaching step is performed using a second leachant comprising a mixed chloride solution of 20% to 22% w / w HCl and additional chloride with a total chloride concentration of 400 to 550 g / L.

18. The method of any one of claims 1 to 7, wherein, The second leaching step is performed for a duration of 2 to 6 h.

19. The method of any one of claims 1 to 7, wherein, The second leaching step comprises: a first leaching protocol performed in the mixed chloride solution, i.e. without addition of any iron powder; and a second leaching protocol performed in the mixed chloride solution with addition of iron powder.

20. The method of claim 19, wherein, The first and second leaching protocols of the second leaching step are performed as continuous leaching steps in the same leaching stage / vessel; or in separate leaching stages / vessels.

21. The method of claim 19, wherein, The first leaching protocol is performed for a duration of 1 to 2 h and the second leaching protocol is performed for a duration of 1 to 4 h.

22. The method of any one of claims 1 to 7, wherein, The neutralizing agent added to the treated second leach liquor to precipitate iron content therefrom comprises at least one of limestone, lime or MgO.

23. The method of any one of claims 1 to 7, wherein, The oxidizing agent added to the treated second leach liquor to precipitate iron content therefrom comprises one or more selected from alkali metal peroxides, alkali metal perchlorates, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, alkali metal chlorates, chlorine gas, alkali metal hypochlorites, hydrogen peroxide, perchloric acid and oxygen-containing gases.

24. The method according to any one of claims 1 to 7, comprising: introducing an oxidizing agent into the second leach liquor prior to the titanium dioxide precipitation step to oxidize any titanium III content to titanium IV by controlling the redox potential of the second leach liquor to be within 100 to 200 mV, wherein the oxidizing agent is selected from one or more of air, oxygen, alkali metal peroxides, alkali metal perchlorates, ammonium perchlorate, magnesium perchlorate, magnesium chlorate, alkali metal chlorates, chlorine gas, alkali metal hypochlorites, hydrogen peroxide and perchloric acid.

25. The method of claim 24, wherein, The oxidizing agent comprises hydrogen peroxide.

26. The method of any one of claims 1 to 7, wherein, The titanium dioxide precipitation step comprises hydrolyzing the titanium IV content of the second leach liquor to precipitate as titanium dioxide (TiO2) solids under an inert gas or nitrogen atmosphere.

27. The method according to any one of claims 1 to 7, comprising the steps of: adding a neutralizing agent to the treated second leach liquor under an inert gas or nitrogen atmosphere at a temperature of 50°C to 80°C to raise the pH of the liquid to 3 to 6, thereby precipitating vanadium and aluminum to produce a V / Al-removed slurry, the neutralizing agent being selected from at least one of limestone, lime or MgO; and separating the V / Al-removed slurry into a liquid portion comprising a treated second leach liquor from which V / Al has been removed and a solid portion comprising V / Al precipitate solids.

28. The method according to any one of claims 1 to 5, comprising the steps of: adding a neutralizing agent and an oxidizing agent to the second leach liquor from which iron has been removed at a temperature of 60°C to 90°C to raise the pH of the liquid to 9 to 10, thereby precipitating Mg and / or Mn to produce a Mg / Mn-removed slurry; and separating the Mg / Mn-removed slurry into a liquid portion comprising a Mg / Mn-removal liquor and a solid portion comprising precipitated Mg and / or Mn solids.

29. The method of any one of claims 1 to 7, wherein, regenerating the second leachant for recycling to the second leaching step includes: concentrating the chloride content of the treated second leachate by removal of water to produce a concentrated chloride solution; reacting the evaporation liquor with at least 98% w / w sulphuric acid at a temperature of 30°C to 90°C under atmospheric conditions to produce a mixed chloride solution having 20% to 22% w / w hydrochloric acid and additional chloride content in solution and a solid precipitate, separating the precipitated solid from the mixed chloride solution; and recycling the mixed chloride solution to the second leaching step.

30. The method of claim 29, wherein, the chloride content includes calcium chloride and the reaction between the evaporation liquor and concentrated sulphuric acid is conducted at a temperature range of 80°C to 85°C to precipitate only anhydrite.

31. The method of any one of claims 1 to 7, wherein, the treatment stages of the first leachate treatment step and the second leachate treatment step are combined for at least one of a vanadium and / or aluminium removal step, an iron removal step, or a manganese and / or magnesium removal step.

32. The method of any one of claims 1 to 7, wherein, the neutralising agent in the process includes MgO and the process includes a Mg removal step in which Mg(OH)2 is precipitated using lime and a MgO regeneration stage in which Mg(OH)2 is calcined at 300°C to 400°C to regenerate MgO for recycling as the neutralising agent in the process.

33. The method of any one of claims 1 to 7, wherein, the titanium-containing material includes an ore body containing a titanium mineral selected from ilmenite, rutile or leucoxene.

34. The method of any one of claims 1 to 7, wherein, the titanium-containing material includes vanadium associated with titanomagnetite, a vanadium-containing mineral.

35. The method of any one of claims 1 to 7, wherein, the titanium-containing material includes a titanium-containing leach residue.

36. The method of any one of claims 1 to 7, wherein, the titanium-containing material includes a mineral processing residue.

37. The method of any one of claims 1 to 7, wherein, the titanium-containing material is a titanium-containing ore.

38. The method of claim 37, wherein, the titanium-containing material is one or more of a concentrate of a titanium-containing ore, a modified ore of a titanium-containing ore and a tailings of a titanium-containing ore.

39. The method of any one of claims 1 to 7, wherein, the titanium-containing material includes at least one of ilmenite or titanomagnetite.

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