WO2021072534A1

WO2021072534A1 - Process for the recovery of titanium dioxide, vanadium and iron compounds from various materials

Info

Publication number: WO2021072534A1
Application number: PCT/CA2020/051376
Authority: WO
Inventors: Bryn Harris; Carl White
Original assignee: 9203-5468 Quebec Inc. Dba Nmr360
Priority date: 2019-10-15
Filing date: 2020-10-15
Publication date: 2021-04-22

Abstract

A process for the recovery of titanium dioxide from various feedstocks comprising titanium, vanadium and iron compounds is disclosed. The process comprises: leaching the feed material with a first HCl solution (e.g. 20% by weight or more), to obtain a first leachate comprising iron compounds and a magnetic concentrate comprising titanium dioxide; leaching the magnetic concentrate comprising titanium dioxide with a second HCl solution (e.g. 20% by weigh or more), in the presence of a reductant, to obtain a second leachate comprising soluble titanium and a solid residue; removing the solid residue; and hydrolysing the second leachate in presence of an oxidant to obtain pure titanium dioxide and HCl solution. The process also allows recovering the vanadium that is almost always associated with titanium, while ensuring reducing conditions without the necessity to add large quantities of reductant to account for the reduction of ferric iron to ferrous.

Description

PROCESS FOR THE RECOVERY OF TITANIUM DIOXIDE, VANADIUM AND IRON COMPOUNDS FROM VARIOUS MATERIALS

CROSS-REFERENCE TO RELATED APPLICATIONS

[0001] The present patent application claims the benefits of priority of U.S. Provisional Patent Application No. 62/915,093 entitled “PROCESS FOR THE RECOVERY OF TITANIUM DIOXIDE, VANADIUM AND IRON COMPOUNDS FROM REFRACTORY MATERIALS”, and filed at the United States Patent and Trademark Office on October 15, 2019, the content of which is incorporated herein by reference.

FIELD OF THE INVENTION

[0002] The present invention relates generally to a process for the recovery of titanium dioxide, vanadium and iron compounds from various feedstocks, in particular, but not limited to, where the titanium is present in a weathered and highly oxidised form making it refractory to conventional recovery techniques, and more particularly from such ores originating in Western Australia and Eastern Africa.

BACKGROUND OF THE INVENTION

[0003] Most of the world’s titanium and vanadium is recovered from ores which are readily amenable to either direct acid leaching or to a smelting or roasting process of some kind. In the smelting processes or those processes designed to generate synthetic rutile, any vanadium values present in the ore are lost.

[0004] Lower-grade titanium-bearing feeds, such as ilmenite, whilst generally amenable to direct acid leaching for either the upgrading or solution recovery of titanium, are very often associated with low-level radioactive materials, which become concentrated into one fraction or another, generally the solid phase, making the processing of these materials problematic and requiring precautions to deal with said radioactivity.

[0005] Refractory titanium-bearing titaniferous magnetites, particularly those found in Western Australia and Eastern Africa, as well as non-refractory magnetites as found in Quebec and Eastern Canada, and semi-refractory magnetites in Southern Africa, are generally free from associated radioactive components. The refractory magnetites are highly resistant to conventional acid leaching processes. However, such materials represent a large and untapped source of titanium, and also contain significant vanadium values, making recovery of the latter attractive.

[0006] Processes for the recovery of titanium dioxide from ores are known. Ilmenite (nominally FeTiCb) and rutile (nominally TiCte) are the two most commercially -important feedstocks for titanium metal and titanium dioxide pigment production, with the latter being preferred, since it has a higher concentration of titanium. However, there is now insufficient naturally-occurring, good-quality rutile to supply the demands of the market, and therefore processes have been developed to produce “synthetic rutile,” principally from ilmenite, but also other titanium-bearing materials. In Western Australia, the source of 40% of the world’s ilmenite and 25% of its rutile, mineral sands containing ilmenite and rutile are separated by physical means to produce a titanium concentrate containing between 45-65% TiCh. There are also substantial deposits of weathered titaniferous magnetite in this area, but they have not been exploited because either the titanium component of the mineral is difficult to leach, and/or the magnesium content is too high for smelting.

[0007] The most common method that has been developed to upgrade these concentrates to synthetic rutile (>90% T1O2) is the Becher Process, originally commercialized in the early 1960s. In the so-called modified Becher Process, as described by H.R. Harris and I.E. Grey in U.S. Patent 5,900,040, entitled “Roasting of Titaniferous Materials”, issued on May 4, 1999, the ilmenite is broken down through a high-temperature (1200°C) reduction of the iron fraction to metallic iron using coal and sulphur. This is followed by re-oxidation to make very fine iron oxide, which then separates readily by physical means from the coarser-grained TiCh. This complicated thermal treatment of the ilmenite ore also renders the titanium component inert to acid leaching. The synthetic rutile so-formed then normally undergoes a dilute acid wash, before it is sold to pigment manufacturers and the low-grade iron oxide is returned to the mine, where it is deposited as waste. The disadvantages of this approach are that it is energy intensive, does not recover a useful iron product, and nor is it possible to recover any of the associated values commonly found with titanium, such as vanadium.

[0008] A similar process has been developed by Austpac Resources, as described by E.A. Walpole and J.D. Winter, entitled “The Austpac ERMS and EARS Processes for the Manufacture of High-Grade Synthetic Rutile by the Hydrochloric Acid Leaching of Ilmenite”, published in Chloride Metallurgy 2002, Volume II, 32^nd Annual Hydrometallurgy Meeting, Edited by E. Peek and G. Van Weert, and published by CIM. This process aims to treat and upgrade ilmenite-bearing sands from Australia’s Murray -Darling Basin. Normal physical separation methods for these ilmenite-bearing sands are not effective, due to the presence of garnet, which has a similar specific gravity to that of ilmenite. Austpac has developed a two- step roasting technique (800-1000°C) using low-grade coal, which renders the ilmenite component magnetic, while at the same time rendering the iron leachable, and converting the titanium into an insoluble form.

[0009] After roasting, the treated ilmenite is recovered by way of magnetic separation and is subsequently leached with hydrochloric acid. This places substantially all of the iron and other soluble mineral components into solution, leaving behind a high-grade T1O2 that then undergoes a final calcining step, followed by magnetic separation to remove residual iron that could not be leached (>97% T1O2) in order to make a feed material suitable for pigment manufacture.

[0010] Where electricity costs are low, for example Canada (Quebec), ilmenite ores with a low magnesium content can be smelted to produce pig iron and a titanium-rich slag. The smelting process is competitive, because the pig iron is sold as a value-added co-product. Rio Tinto Iron & Titanium (RTIT) operates an ilmenite smelting furnace at its Sorel-Tracy complex. The furnace and RTIT’s business model are predicated on the treatment of such a relatively low- grade T1O2 feedstock, as this provides the necessary balance between slag and pig iron.

[0011] The pyrometallurgical, i.e. the modified Becher and RTIT processes aside, the hydrometallurgical approaches generally have one of two objectives, namely to remove the iron only, leaving behind a titanium phase which can be calcined to synthetic rutile, or to dissolve both the iron and titanium and effect a subsequent separation either by hydrolysis or solvent extraction combined with hydrolysis. The former, i.e. the removal of iron to leave behind synthetic rutile is by far the currently -preferred process. This rutile can be further processed by either of the so-called chloride or sulphate processes to generate pigment-grade titanium dioxide. However, such processes can also result in a concentration of any radioactive components into the synthetic rutile phase.

[0012] A process for the extraction of iron from iron-containing titaniferous ores is described in U.S. Patent 2,406,577, entitled “Extraction of Iron from Titaniferous Ores”, issued on August 27, 1946, by H.V. Alessandroni and B.V. Sterk. It is stated that it is known that concentrated hydrochloric acid will leach both iron and titanium into solution from an iron- bearing titaniferous ore if the temperature is below about 70°C, but that if the temperature is maintained above 70°C any titanium that may be dissolved is re-precipitated. Improved results are obtained, based on the observation that when chloride salts are added to hydrochloric acid, the acid will more effectively remove iron from iron-containing titaniferous ores. The patent discloses the leaching of a titaniferous ore with a solution of hydrochloric acid of a specific gravity of approximately 1.10 and at least 0.5 mole of a soluble chloride e.g. alkali metal chlorides, alkaline earth metal chlorides and aluminum chloride, at a temperature between 70°C and the boiling point of the solution. The specific gravity of 1.10 is stated to correspond to a concentration of hydrochloric acid of about 230 g/L, i.e. about 21% hydrochloric acid.

[0013] A process for leaching ilmenite is described in U.S. Patent 3,903,239, entitled “Recovery of Titanium Dioxide from ores”, issued on September 2, 1975, by S.A. Berkovich. The process comprises contacting ilmenite, or a concentrate thereof, with a concentrated hydrochloric acid lixiviant solution at a temperature of about 15-30°C to solubilize and leach from the ore at least 80% and preferably at least 95% of the iron and titanium values. The ore may be pre-treated prior to contact with the concentrated hydrochloric acid to increase the rate of dissolution of titanium and iron values during leaching. The pre-treatment is a smelting step that may include oxidation at elevated temperature e.g. 600-1000°C in the presence of air or oxygen, followed by a reduction of at least part of the iron oxide in the ore with carbon or carbon monoxide.

[0014] U.S. Patent 6,375,923 (W.P.C. Duyvesteyn et al.), entitled “Processing Titaniferous Ore to Titanium Dioxide Pigment”, issued on April 23, 2002, describes a chloride-based leaching process for the recovery of titanium dioxide. A complementary publication entitled “The Altair T1O2 Pigment Process and its Extension into the Field of Nanomaterials” by D. Verhulst, B Sabacky, T. Spitler and W. Duyvesteyn, pages 417-432, Chloride Metallurgy 2002 Volume II, 32^nd Annual Hydrometallurgy Meeting, Edited by E. Peek and G. Van Weert, published by CIM, further describes ahydrometallurgical process for producing pigment-grade titanium dioxide from a titaniferous ore. The process comprises leaching the ore with a solution of hydrochloric acid at a temperature of at least 50°C to provide a leachate of titanium chloride, ferrous chloride, ferric chloride and impurity chlorides, a residue of undissolved solids and sufficient excess hydrochloric acid to prevent precipitation of titanium dioxide. No pre treatment of the ore is described, hence the process is applicable only to those ores wherein the titanium component is readily soluble. [0015] U.S. Patent 6,500,396 V.I. Lakshmanan et al.) entitled “Separation of Titanium Halides from Aqueous Solutions”, issued on December 31, 2002, describes a method for the production of titanium materials from titanium-bearing ore. In embodiments, the ore or concentrate is leached with an aqueous solution of a hydrogen halide, especially hydrochloric acid, at a temperature of at least 90°C, followed by solid/liquid separation and extraction with an immiscible organic phase. In other embodiments, the ore is leached with the hydrogen halide in the presence of an oxidizing agent. A variety of oxidizing agents are disclosed, including air, hydrogen or other peroxides, or sodium or other perchlorates. In the leach solution, iron is solubilized and titanium is converted into titanium dioxide. No pre-treatment process for the ore is described, and the process therefore appears to be applicable only to ores wherein the titanium component is readily soluble. However, the addition of an oxidising agent is surprising, in that such a compound would promote the formation of titanium dioxide, which is insoluble, rather than effecting the dissolution of the titanium mineral.

[0016] A further patent, U.S. 7,803,336 (V.I. Lakshmanan et al.) entitled “Process for the Recovery of Titanium in Mixed Chloride Media”, issued on September 28, 2010, describes a process for leaching a value metal from a titanium-bearing ore material. In this process, the ore material is leached at atmospheric pressure with a lixiviant comprising a chloride and hydrochloric acid. The leaching conditions are such that the titanium is leached and remains in solution. The temperature is maintained at less than 85°C, and the concentration of hydrochloric acid is preferably less than 20% (mass ratio). The preferred chloride is magnesium chloride, and the lixiviant may contain an oxidant e.g. sodium chlorate or chlorine. No pre-treatment process for the ore is described, and the process therefore appears to be applicable only to ores wherein the titanium component is readily soluble.

[0017] B.T. Judd, in European Patent Application EP 0 186 370, entitled “Titanium Dioxide Pigment Production from Ilmenite”, issued December 10, 1985, describes a process for the production of titanium dioxide by two-stage hydrochloric acid leaching, with particular reference to feed materials prevalent in New Zealand. In the first stage, using dilute acid, impurities such as phosphorus are removed, and in the second stage using concentrated acid, the titanium is dissolved along with the iron. No other pre-treatment of the ore is described, so that this process also appears to be applicable wherein the titanium component is readily soluble. [0018] T. Ogasawara, and R. Veras Veloso de Araujo in an article entitled “Hydrochloric Acid Leaching of a Pre-reduced Brazilian Ilmenite in an Autoclave”, published in Hydrometallurgy 56 (2000), pages 203-216, describe a process for recovering titanium from an oxidised ore, wherein it was first necessary to thermally pre-reduce the ore in order to effect titanium leaching. In this process, the objective is to leach both titanium and iron, and selectively recover titanium dioxide from the leach solution. The pre-reduction step is necessary to render the titanium soluble, and then only in an autoclave.

[0019] N. El-Hazek, T.A. Lasheen, R. El-Sheikh and S.A. Zaki, in an article entitled “Hydrometallurgical Criteria for T1O2 Leaching from Rosetta Ilmenite by Hydrochloric Acid”, published in Hydrometallurgy 87 (2007), pages 47-50, describe a process opposite to selective iron dissolution and insolubilization of T1O2 from a titanium-bearing ore, wherein the titanium is dissolved along with the iron. In this process, the parameters for leaching both titanium and iron were established, showing that >90% of both metals could be dissolved in strong hydrochloric acid. Subsequently, approximately 10% by weight of the ore of iron powder was added in a single addition to ensure reducing leaching conditions and effect reduction of all of the ferric iron to ferrous (the objective for the subsequent solvent extraction step). It was noted, however, that only a portion of the iron was effective, with appreciable amounts being dissolved simply in the acid. This would render such a process ineffective due to the large quantities of metallic iron that would be consumed.

[0020] Further, it was noticed that some of the titanium was reduced to its trivalent state, which is very difficult to hydrolyse to titanium dioxide. It was also noted that addition of iron actually reduced the efficiency of the leaching process under comparable conditions to when it was not present, and that increasing the amount of reductant was counterproductive to optimum titanium extraction.

[0021] M.H.H. Mahmoud, A.A.I. Afifi and I.A. Ibrahim, in an article entitled “Reductive Leaching of Ilmenite Ore in Hydrochloric Acid for Preparation of Synthetic Rutile”, published in Hydrometallurgy 73 (2004), pages 99-109, describe a similar process wherein the addition of metallic iron appreciably enhances the leaching of titanium from an oxidised ilmenite ore by hydrochloric acid. The amount of iron added (110 kg/tonne of ore) was calculated to be in excess of that needed to reduce all of the ferric iron present to ferrous. The objective of the process was to produce a synthetic rutile, which was achieved by first allowing the metallic iron to reduce the ferric iron to ferrous at the boiling point of the solution (110°C), together with a portion of the titanium to its trivalent state. Thereafter, it is claimed that the reduced titanium reacted with fresh ferric iron to generate ferrous iron and soluble titanium, which subsequently hydrolyses to form synthetic rutile.

[0022] The aim of most of the above processes is to produce a synthetic rutile, which is subsequently used as feed to a pigment process. The exceptions to this are where a solvent extraction step is proposed, where the intent is to generate pigment quality titanium dioxide from the solvent strip solution. The recovery of vanadium from such ores has not been considered in any of the foregoing processes.

[0023] It is also clear that in the majority of cases, direct leaching of titanium into solution is not possible without some form of pre-treatment step. Applicant has found in the course of testing many samples from around the world that there are a large number of titanium-bearing ores which are not readily-amenable to direct acid leaching. For example, with ores from Quebec, the titanium component readily dissolves in hydrochloric acid along with the iron, whereas with ores from Western Australia, it does not dissolve at all. In light of all of the foregoing, it would therefore be advantageous to be able to define a process which could recover titanium from all types of ores, but especially those wherein the titanium is not amenable to direct acid leaching without recourse to either a pyrometallurgical pre-reduction step, or the necessity to add large quantities of metallic iron, the large quantities being necessary because any iron addition to ensure a reducing leach preferentially reacts to reduce the large amounts of ferric iron present (magnetite, for example, being two thirds ferric iron, one third ferrous). This is because subsequent processing steps require that any dissolved iron be in the ferrous state.

[0024] In particular, such a process would not only also recover the vanadium that is almost always associated with the titanium, which none of the above processes are able to do, but also ensure reducing conditions wherein the necessity to add large quantities of reductant to account for the reduction of ferric iron to ferrous is eliminated, and also wherein rutile is the predominant phase of the titanium dioxide recovered.

SUMMARY OF THE INVENTION

[0025] In accordance with a broad aspect of the present invention, there is disclosed a process for recovering titanium dioxide, vanadium and iron compounds from various titanium ores, but in particular those ores wherein the titanium is refractory and does not dissolve directly in acid, and more particularly, such weathered ores as are found, for example, in Western Australia.

[0026] According to a first aspect, the invention is directed to a process for the recovery of titanium dioxide from a feed material comprising titanium, vanadium and iron compounds, the process comprising: a) leaching the feed material with a first HC1 solution having a hydrochloric acid content of 20% by weight or more, to obtain a first leachate comprising iron compounds and a magnetic concentrate comprising containing titanium dioxide minerals; b) leaching the magnetic concentrate comprising titanium dioxide with a second HC1 solution having a hydrochloric acid content of 20% by weight or more, in the presence of a reductant, to obtain a second leachate comprising a soluble titanium salt and a solid residue, said solid residue containing preferably insoluble gangue minerals; c) removing the solid residue; and d) hydrolysing the second leachate in presence of an oxidant to obtain pure titanium dioxide and HC1 solution.

[0027] According to a preferred embodiment, wherein when a certain amount of vanadium compounds is present in the feed material, the process further comprises before step a): leaching the feed material with a third HC1 solution having a concentration of between 1 and 5 wt.%, to produce a leachate comprising vanadium compounds dissolved in the HC1 solution and a magnetic concentrate comprising iron compounds and titanium dioxide; and separating the magnetic concentrate from the leachate comprising vanadium compounds before processing the magnetic concentrate comprising iron compounds and titanium dioxide as the feed material in step a) of the process. More preferably, the third HC1 solution has a hydrochloric acid concentration of 1 to 2 % by weight.

[0028] According to a preferred embodiment, in step a), the first HC1 solution has a hydrochloric acid content of 30-35% by weight. [0029] According to a preferred embodiment, in step b), the second HC1 solution has a hydrochloric acid content of 30-35% by weight.

[0030] According to a preferred embodiment, step a) is performed at a temperature from ambient to boiling point, and is preferably about 85-90°C.

[0031] According to a preferred embodiment, step b) is performed at a temperature of from 80 to 100°C, preferably of from 85 to 90°C.

[0032] According to a preferred embodiment, the feed material is a refractory vanadium bearing titaniferous magnetite feed material.

[0033] According to a preferred embodiment, in step b), the reductant comprises iron metal, aluminium metal, zinc metal, magnesium metal, or a mixture thereof. More preferably, the reductant is iron metal.

[0034] According to a preferred embodiment, the reductant is an organic compound, such as but not limited to, formic acid or hydrazine.

[0035] According to a preferred embodiment, step d) further comprises the steps of diluting the second leachate of titanium chloride with water and adding titanium dioxide seed.

[0036] According to a preferred embodiment, step d) further comprises the steps of diluting the second leachate of titanium chloride with hydrochloric acid with no seed added to precipitate titanium dioxide as the brookite phase. More preferably, the added hydrochloric acid is at a concentration of at least 10% HC1.

[0037] According to a preferred embodiment, the HC1 solution obtained in step d) is recycled at least in part to be used to leachate the feed material in step a).

[0038] According to a preferred embodiment, in step d) is the oxidation is carried out by hydrogen peroxide, chlorine or electrolytically.

[0039] According to a preferred embodiment, the feed material before step a) is first crushed and/or ground to form particles with an optimum particle size. Preferably, the optimum particle size is dependent upon the nature of the feed material. For instance, the optimum particle size is in the range of 10 to 325 mesh, preferably 50 to 200 mesh, and more preferably about 100 mesh or 150 microns. [0040] Advantageously, the process as disclosed herein would not only also recover the vanadium that is almost always associated with the titanium, which none of the above processes are able to do, but also ensure reducing conditions wherein the necessity to add large quantities of reductant to account for the reduction of ferric iron to ferrous is eliminated, and also wherein rutile is the predominant phase of the titanium dioxide recovered.

BRIEF DESCRIPTION OF THE DRAWINGS

[0041] The above and other aspects, features and advantages of the invention will become more readily apparent from the following description, reference being made to the accompanying drawings in which:

[0042] Figure 1 is a flowchart illustrating the process according to a preferred embodiment of the invention.

DETAILED DESCRIPTION OF EMBODIMENTS OF THE INVENTION

[0043] A novel process for recovery of titanium dioxide, vanadium and iron compounds from various feedstocks is disclosed. Although the invention is described in terms of specific illustrative embodiments, it is to be understood that the embodiments described herein are by way of example only and that the scope of the invention is not intended to be limited thereby.

[0044] The terminology used herein is in accordance with definitions set out below.

[0045] As used herein % or wt.% means weight % unless otherwise indicated. When used herein % refers to weight % as compared to the total weight percent of the phase or composition that is being discussed.

[0046] By "about", it is meant that the value of weight %, time, pH, volume or temperature can vary within a certain range depending on the margin of error of the method or device used to evaluate such weight %, time, pH, volume or temperature. A margin of error of 10% is generally accepted.

[0047] The description which follows, and the embodiments described therein are provided by way of illustration of an example of particular embodiments of principles and aspects of the present invention. These examples are provided for the purposes of explanation and not of limitation, of those principles of the invention. In the description that follows, like parts and/or steps are marked throughout the specification and the drawing with the same respective reference numerals. The process as disclosed herein allows the recovery of titanium, vanadium and iron oxides from various materials, but especially those materials wherein the titanium is difficult to leach (refractory), and with particular reference to those emanating from Western Australia.

[0048] The process may comprise an initial high-acid concentration leach of the magnetic fraction of the primary ore at 90°C which dissolves a substantial portion of the iron, particularly that in the ferric oxidation state, leaving behind an upgraded titanium dioxide residue suitable for further processing. The iron leach solution can be treated by conventional means to recover hematite and vanadium, if present, and recycle the acid within the process.

[0049] The iron leach residue may be then further leached with concentrated hydrochloric acid, preferably at 60°C, with the addition of a reductant, which may be iron, aluminium or magnesium metal, or a suitable organic reductant such as formic acid or hydrazine. These materials generate a reactive form of hydrogen (sometimes known as nascent hydrogen) under these conditions, thereby causing the titanium to dissolve under appropriate redox potential of -200 to -400 mV (versus Pt-Ag/AgCl electrodes).

[0050] Oxidation and hydrolysis of the reducing leach liquor by adding water and heating the solution to 90°C promotes the formation a chemically pure titanium dioxide, with predominantly the rutile crystal structure. Alternatively, if the reducing leach is diluted with hydrochloric acid, then the chemically pure titanium dioxide is formed with the brookite crystal structure.

[0051] Referring to Figure 1, there is shown a schematic representation of a process wherein titanium dioxide, a vanadium product and an iron product are recovered from a refractory vanadium-bearing titaniferous magnetite feed material, which may be, but is not limited to, an ore or a magnetic concentrate.

[0052] Further referring to Figure 1, the feed material 10, containing titanium, vanadium and iron values is first crushed and ground to give a suitable particle size. The optimum particle size will be dependent upon the nature of the feed, but will typically be in the range of 10 mesh to 325 mesh, more typically 50 mesh to 200 mesh, and optimally, 100 mesh (150 microns). The material is fed to a primary acid leach 11 employing for instance a combination of recycled hydrochloric acid 12 and recycled spent titanium recovery liquor 27. The concentration of the acid will vary depending on the nature of the ore, but is not normally less than 20% (azeotropic), and more normally about 30-35%, and preferably 33%.

[0053] Optionally, and not shown, a dilute acid leach, at a preferred concentration of 1-5%, more preferably of 1-2%, may be used to selectively dissolve the vanadium where it is amenable to such an action. Such acid concentration may vary as it depends on how much vanadium compounds are in the feed material.

[0054] The primary leach 11 dissolves as much iron as possible and all of the vanadium from the feed material, with minimal co-dissolution of titanium. The temperature of the leaching reaction can be from ambient to boiling, and is preferably 85-90°C, such temperature ensuring that any titanium that does dissolve will re-precipitate.

[0055] Solid-liquid separation 14 of the primary leach slurry 13 may be effected by any convenient means, such as, but not limited to, flocculation and thickening, fdter press or vacuum belt fdter. A portion of the fdtrate 29 may be recycled in order to increase the iron and vanadium concentrations in order to facilitate their subsequent recoveries, and to consume as much acid as possible.

[0056] Under the conditions of the primary (iron) leach, predominantly ferric iron dissolves, leaving a significantly upgraded, but highly refractive titanium dioxide residue 15. This residue is an ilmenite or pseudo-ilmenite phase.

[0057] This primary (iron) leaching stage achieves two important results. By leaching with concentrated acid, but without the addition of any reductant, a substantial portion, typically 50- 60%, of the iron phase, and in particular the ferric iron component, is dissolved. Unlike the processes described above by El Hazel et al. and Mahmoud et al, this permits the subsequent leaching of titanium 16 to be carried out with a minimum addition of reductant, because there is very little ferric iron remaining present, which necessarily consumes reductant and for which the above processes are designed, as would be the case if the iron were not removed.

[0058] The principal chemical reaction taking place is the dissolution of ferric iron as shown in reaction (1), with a secondary reaction dissolving some magnetite as shown in reaction (2).

Fe₂0₃ + 6HC1 2FeCl₃ + 3H2O (1) Fe₂03*Fe0 + 8HC1 2FeCls + FeCh + 4H₂0 (2)

[0059] The primary leach reaction may be carried out in any suitable reactor, preferably using a cascade of CSTRs (Continuous Stirred Tank Reactor).

[0060] The impure titanium dioxide mineral solids 15 remain refractory. Others skilled in the art have found it necessary to either pre-reduce the solids in a separate calcining step in order to make them dissolve in acid. The solids 15 are leached 16 with recycled hydrochloric acid 12 and the addition of a reductant 17.

[0061] It has been discovered that an ORP (oxidation-reduction potential) of at least -150 mV, and preferably -50 mV relative to the Pt-Ag/AgCl Electrode, preferably causes the titanium to dissolve in acidic solutions, such as hydrochloric acid solutions. Metallic iron is capable of achieving this, but also adds significant amounts of dissolved iron, which can co-hydrolyse during the subsequent titanium hydrolysis step to recover pure titanium dioxide. Metallic aluminium, metallic zinc and metallic magnesium may also be used to achieve the desired ORP value, and are resistant to hydrolysis under the conditions wherein titanium is hydrolysed, and may be preferred over metallic iron. However, all three metals result in impurities in solution which have to be accounted for at some point in the flowsheet, and each is significantly more costly than iron.

[0062] Alternative reductants which do not add impurities, such as organic reductants, e.g. formic acid, or hydrazine may also be considered.

[0063] The reaction is carried out at a temperature of ambient to 100°C, preferably at 80-90°C. Normally, where titanium is readily soluble, leaching reactions such as this are carried out at <50°C in order to prevent premature titanium hydrolysis. However, with the reductive leach, the form of titanium in solution is in the trivalent state, as opposed to the more normal tetravalent state, and a higher temperature is preferred to facilitate the reaction.

[0064] In addition to maintaining close control on ORP and temperature, a high acidity in the solution may be preferably maintained, and the concentration of titanium in the solution should be preferably limited. Both of these are necessary in order to maximise titanium extraction. Preferably, a free hydrochloric acid concentration above 20%, more preferably at 25% or higher, is maintained. It has further been found that there is an upper limit of titanium concentration, above which the titanium will spontaneously start to hydrolyse, even though it is in its trivalent form, which is much more difficult to hydrolyse than its tetravalent analogue. The maximum level of titanium in solution is preferably about 40 g/L, with the optimum being in the range of 30-35g/L.

[0065] The mechanism of the reducing leach 16 is that it is necessary to generate a reactive form of hydrogen, sometimes referred to as nascent (or monatomic) hydrogen, which all of iron, aluminium, zinc and magnesium will do in strong hydrochloric acid. The simplified chemical reaction sequence taking place is shown in reactions (3, a, b c and d) for the metals, with the reaction for formic acid (3e) being shown as a generalisation for organic reductants. Reaction (4) shows the second stage wherein the titanium is reduced, where the reactive hydrogen, irrespective of its source, is represented as H· rather than the more common gaseous molecular form of hydrogen, H2.

Fe +2HC1 FeCk + 2H· (3 a) A1 + 3HC1 Aids + 3H· (3b) Zn + 2HC1 ZnCk + 2H· (3c) Mg + 2HC1 MgCk + 2H· (3d) HCOOH CO2 + 2H· (3e)

TiCk + H· + HC1 TiOCl + H2O (4)

[0066] Because the reaction is extremely complicated, and several parameters have to be monitored simultaneously, namely temperature, ORP (and hence reductant addition rate), titanium concentration and free acid concentration, it has been found that contrary to normal practice of using a series of CSTRs, the reaction is also readily carried out in batch mode.

[0067] Solid-liquid separation 19 of the leach slurry 18 may be effected by any convenient means, such as, but not limited to, flocculation and thickening, fdter press or vacuum belt fdter. The solids 20 from the leach will be predominantly unreacted gangue minerals which may be disposed of. [0068] The titanium chloride fdtrate solution 21 proceeds to titanium hydrolysis 22. Here, the solution is first diluted 24 to give a titanium concentration in the range of 5-15 g/L, and preferably in the range 7-10 g/L. This ensures that there is sufficient water present to effect the hydrolysis to form titanium dioxide. The temperature is raised to 80-100°C, preferably 85- 90°C to effect the precipitation of titanium dioxide. In this step, part of the initial wet solids are used as seed. The seeding allows forming the rutile crystal phase of titanium dioxide which is the preferred structure, and this is achieved by initially creating the rutile crystal and using it as seed. In the absence of any seed, the anatase phase of titanium dioxide will also precipitate, which requires subsequent calcining to convert it to rutile.

[0069] The diluted trivalent titanium is then oxidised 23 back to its tetravalent state. Oxidation is preferably controlled in the range +50 to +200 mV (relative to the Pt-Ag/AgCl), preferably +100 to +150 mV in order to oxidise the trivalent titanium to its tetravalent state, but to leave the ferrous iron untouched. Any suitable strong oxidant, such as hydrogen peroxide, chlorine or ozone may be used. Weaker oxidants, such as oxygen, are not effective, since the control of the ORP is very difficult. Oxidation via electrolysis of the hydrochloric acid already present to generate chlorine in-situ may also be preferentially employed. In order to ensure that no ferrous iron is oxidised, a heel of trivalent titanium is ideally left in solution, typically 1-5 g/L, and preferentially about 1 g/L. By this methodology, substantially 100% of the titanium in solution 35 is hydrolysed and precipitated. The reaction employing chlorine is shown in equation (5) and with hydrogen peroxide in equation (6).

2TiOCl +Ch 2TiOCh (5)

2TiOCl + H2O2 + 2HC1 2TiOCb + 2H₂0 (6)

[0070] Solid-liquid separation 26 may be effected by any convenient means, such as, but not limited to, flocculation and thickening, fdter press or vacuum belt fdter. The solids 28 are highly chemically pure titanium dioxide, mostly in the form of rutile suitable for pigment manufacture.

[0071] In a separate embodiment, instead of diluting with water, it has been discovered that by diluting with hydrochloric acid, preferably at a concentration of 10%, the brookite form of titanium dioxide is precipitated instead of rutile.

[0072] The fdtrate 27 is preferentially concentrated by a suitable technique such as, but not limited to, reverse osmosis, and recycled to the primary (iron) leach. [0073] The filtrate 29 from the primary (iron) leach is treated 30 for vanadium precipitation. In this stage, a suitable precipitant 31 such as hydrogen sulphide gas or NaSH (sodium hydrosulphide) is added to selectively and quantitatively precipitate the vanadium. The reaction is optimally carried out at 85-90°C. Such precipitants are also reductants, so it is important to control the ORP, again in the range of +230 to +400 mV (Pt-Ag/AgCl) in order that only the vanadium reacts, and that no ferric iron is reduced. The precipitate is a vanadium poly sulphide, being a mixture of various vanadium sulphides. The reaction for pentavalent vanadium with hydrogen sulphide gas is shown in equation (7).

2VOCI3 +IOH2S 2V2S5 + 2H2O + 6HC1 (7)

[0074] Solid-liquid separation 33 of the precipitation slurry 32 may be effected by any convenient means, such as, but not limited to, flocculation and thickening, filter press or vacuum belt filter. The solids 34 are highly chemically pure vanadium polysulphide, which may be calcined under suitable conditions to generate a pure form of vanadium pentoxide 35.

[0075] The filtrate 36 is treated by any suitable means 37 to recover the hydrochloric acid 12 for recycle in the flowsheet, and iron as a pure hematite product 40, after solid-liquid separation 39 and slurry 38 may be effected by any convenient means, such as, but not limited to, flocculation and thickening, filter press or vacuum belt filter. The filtrate 41 may be recycled, or a bleed 42 taken to remove any impurities such as calcium or magnesium.

[0076] Iron precipitation and acid recovery 37 from the iron leach filtrate may be effected by any suitable means. One option would be to use magnesia to precipitate the iron and spray roast the subsequent magnesium chloride solution, with the magnesia being recycled within the unit operation, but with a relatively dilute (17-18%) hydrochloric acid being recovered.

[0077] Figure 1 depicts another option using the methodology disclosed in the current inventors’ patent: U.S. 9,889,421 B2, entitled “Process For the Recovery of Metals and Hydrochloric Acid”, issued February 13, 2018, herein incorporated by reference. This allows for concentrated hydrochloric acid to be recovered and recycled to the leach together with a pure hematite product.

[0078] Another option would be to use the process disclosed in Applicant’s application WO 2019/006545, herein incorporated by reference. [0079] The process will now be illustrated by way of examples. These examples are provided for illustration of certain embodiments of the invention, and is not intended as limitations thereof.

Examnle 1

[0080] A sample of titaniferous magnetite ore, analysing 23.2% Fe, 13.2% Ti and 0.25% V was ground to minus 100 mesh (150 microns) and leached with hydrochloric acid additions of 2-5 tonnes of 100% HC1 per tonne of ore, at 95°C for a period of 6 hours. The maximum vanadium extraction achievable was 70%, which resulted in an average vanadium solution concentration of 160 mg/L. Thus, only 70% of the vanadium contained in the ore was acid- soluble. Under these conditions, 55-60% of the iron contained in the ore was also dissolved. Titanium dissolution was <1%.

Example 2

[0081] A magnetic concentrate, derived from a Western Australian ore and analysing 34.8% Fe, 32.3% T1O2, and 0.77% V2O5 with a mean particle size of 150 microns was leached at 30% solids loading at 90°C in 32% hydrochloric acid for 6 hours. The solution generated was characteristic of the colour of concentrated ferric chloride, and analysed 56 g/L ferric iron, 0.5 g/L ferrous iron, 200 mg/L Ti and 1.11 g/L V. The solids, 61.7% of the original mass, after the leach analysed 23% Fe and 50.7% T1O2, indicating that very little titanium (3.2%) was dissolved in this step, but that 59.2% of the iron was removed, leaving an enriched titanium bearing concentrate. The analysis of the leach solution indicates that it was predominantly ferric iron which was dissolved in this step. The leach solution analysis also indicates that 95% of the V was recovered.

[0082] Example 2 demonstrates that it is possible to remove a significant fraction of the iron from the ore, and that this fraction is predominantly ferric iron, with very little corresponding titanium extraction. This demonstrates and means that the requirement for the addition of reductant in the subsequent titanium leaching step is significantly reduced, since predominantly all of the ferric iron, which would consume reductant, is removed. The example also demonstrates that substantially all of the vanadium was recovered in this step. Example 3

[0083] The residue from the above leach was reacted with hydrochloric acid and metallic iron at 85°C for six hours at 7% solids. Sufficient iron powder was added slowly throughout the leach in order to maintain the ORP at -150 mV (relative to the Pt-Ag/AgCl electrode), and the free acidity was maintained at 25%. The residue fall from the leach was 24.5%, and the final solids analysed 28.5% T1O2, indicating extraction of 80% of the Ti. An XRD analysis of the final leach solids showed them to 55% kaolinite, which had been present as only a trace amount in the original feed.

[0084] This example demonstrates the ability of the reducing leach to dissolve titanium even from difficult-to-leach feed materials.

Example 4

[0085] A feed solution, analysing 36 g/L Ti and 40 g/L iron, 100% in the ferrous state, was initially oxidised with hydrogen peroxide at an ORP of 10-150 mV (versus Pt-Ag/AgCl), and then diluted by 2.5 times with water prior to being fed into a 3-reactor cascade with a residence time of 6 hours at 85 °C. The titanium concentration of the filtrate averaged 500 mg/L, indicating that 97.5% of the titanium in the feed had been precipitated. White solids, averaging 99.5% T1O2 were obtained. XRD analysis of these solids indicated the rutile structure, with minor anatase and amorphous material (<10%).

[0086] Example 4 demonstrates that substantially all of the titanium in the feed to hydrolysis can be recovered in a state of high purity, even from a solution containing a high concentration of iron, and that rutile is the predominant phase.

Example 5

[0087] The test of Example 4 was repeated, except that the dilution was carried out with 10% HC1. The results showed the product to be >96% brookite, with the balance being anatase and pseudorutile.

[0088] Example 5 demonstrates that adjusting the precipitation conditions can permit the formation of either the rutile or the brookite phase of titanium dioxide. Example 6

[0089] 42g of a sample of titanium-vanadium alloy baghouse dust were leached with hydrochloric acid at an acid loading of 800 kg/tonne of dust over a period of 60 minutes. The total volume of the recovered solution was 1.2 L. The temperature during leaching was kept below 70 °C in order to prevent any premature titanium hydrolysis, and the ORP of the solution was -340 mV versus the Pt-Ag/AgCl electrode. The final solution analysed 34.0 g/L Ti and 2.1 g/L V, and the amount of solids left after the leach was 0.3 g dry basis.

[0090] The leach filtrate was divided into two parts. In the first part, the filtrate ORP was first adjusted to +165 mV versus Pt-Ag/AgCl, and was then pumped into 1 L of water at 90-95°C at a rate of 5 mL/minute. It took 30 minutes before any sign of solids were seen. The solids formed after one hour’s reaction time were white, but the slurry could not be filtered. In the second part, the same experiment was repeated, except that this time, the ORP was first adjusted to +100 mV, and instead of water, 10 g/L HC1 was used, and the temperature was reduced to 80°C. This time, white solids were seen instantaneously, and large, snowflake-like crystals were seen, characteristic of brookite. The solids formed after one hour filtered rapidly.

[0091] Example 6 demonstrates the conditions necessary to hydrolyse titanium solutions and form solids that can be readily filtered.

[0092] While illustrative and presently preferred embodiments of the invention have been described in detail hereinabove, it is to be understood that the inventive concepts may be otherwise variously embodied and employed and that the appended claims are intended to be construed to include such variations except insofar as limited by the prior art.

Claims

CLAIMS What is claimed is:

1. A process for the recovery of titanium dioxide from a feed material comprising titanium, vanadium and iron compounds, the process comprising: a) leaching the feed material with a first HC1 solution having a hydrochloric acid content of 20% by weight or more, to obtain a first leachate comprising iron compounds and a magnetic concentrate comprising titanium dioxide minerals; b) leaching the magnetic concentrate comprising titanium dioxide with a second HC1 solution having a hydrochloric acid content of 20% by weight or more, in the presence of a reductant, to obtain a second leachate comprising a soluble titanium salt and a solid residue; c) removing the solid residue; and d) hydrolysing the second leachate in presence of an oxidant to obtain pure titanium dioxide and HC1 solution.

2. The process of claim 1, wherein when a certain amount of vanadium compounds is present in the feed material, the process further comprises before step a): leaching the feed material with a third HC1 solution having a concentration of between 1 and 5 wt.%, to produce a leachate comprising vanadium compounds dissolved in the HC1 solution and a magnetic concentrate comprising iron compounds and the titanium dioxide minerals; and separating the magnetic concentrate from the leachate comprising vanadium compounds before processing the magnetic concentrate comprising iron compounds and titanium dioxide minerals as the feed material in step a) of the process.

3. The process of claim 2, wherein the third HC1 solution has a hydrochloric acid concentration of 1 to 2 % by weight.

4. The process of any one of claims 1 to 3, wherein in step a), the first HC1 solution has a hydrochloric acid content of 30-35% by weight.

5. The process of any one of claims 1 to 4, wherein in step b), the second HC1 solution has a hydrochloric acid content of 30-35% by weight.

6. The process of any one of claims 1 to 5, wherein step a) is performed at a temperature from ambient to boiling point.

7. The process of claim 6, wherein the temperature is about 85-90°C .

8. The process of any one of claims 1 to 7, wherein step b) is performed at a temperature of from 80 to 100 °C.

9. The process of claim 8, wherein the temperature in step b) is from about 85 to 90°C.

10. The process of any one of claims 1 to 9, wherein the feed material is a refractory vanadium-bearing titaniferous magnetite feed material.

11. The process of any one of claims 1 to 9, wherein the feed material is a dust generated from the production of titanium alloys.

12. The process of any one of claims 1 to 11, wherein in step b), the reductant comprises iron metal, aluminium metal, zinc metal, magnesium metal, or a mixture thereof.

13. The process of claim 12, wherein the reductant is iron metal.

14. The process of any one of claims 1 to 11, where the reductant is an organic compound.

15. The process of claim 14, wherein the organic compound is formic acid or hydrazine chloride.

16. The process of any one of claims 1 to 15, wherein step d) further comprises the steps of: diluting the second leachate of titanium chloride with water; and adding titanium dioxide seed.

17. The process of any one of claims 1 to 15, wherein step d) further comprises the steps of: diluting the second leachate of titanium chloride with hydrochloric acid, with no seed added to precipitate titanium dioxide as the brookite phase.

18. The process of claim 17, wherein the added hydrochloric acid is at a concentration of at least 10 g/L HC1.

19. The process of any one of claims 1 to 18, wherein the HC1 solution obtained in step d) is recycled at least in part to be used to leachate the feed material in step a).

20. The process of any one of claims 1 to 19, wherein in step d) is the oxidation is carried by hydrogen peroxide, chlorine or electrolytically.

21. The process of any one of claims 1 to 20, wherein the feed material before step a) is first crushed and/or ground to form particles with an optimum particle size.

22. The process of claim 21 , wherein the optimum particle size is dependent upon the nature of the feed material.

23. The process of claim 21 or 22, wherein the optimum particle size is in the range of about 10 to 325 mesh.

24. The process of claim 23, wherein the optimum particle size is in the range of about 50 to 200 mesh.

25. The process of claim 23 or 24, wherein the optimum particle size is about 100 mesh or about 150 microns.