PROCESS OF PROVIDING TITANIUM DIOXIDE AND/OR VANADIUM OXIDE

Abstract

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

Claims

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

2. The process of claim 1, wherein the alkaline material has a pH of less than 8.0.

3. The process of claim 1 or claim 2, wherein the alkaline material comprises borax.

4. The process of any one of the preceding claims, wherein the alkaline material is added to the feedstock in an amount of 3 to 20 wt %, based on the total weight of the feedstock, preferably wherein the alkaline material is added to the feedstock in an amount of 5 to 15 wt %, based on the total weight of the feedstock.

5. The process of any one of the preceding claims, wherein the roasting in step (c) is at a temperature in the range of 600 C. to 800 C.

6. The process of any one of the preceding claims, wherein the roasting in step (c) is for 0.5 hours to 4 hours.

7. The process of any one of the preceding claims, wherein the roasted calcine has a pH of less than 9.0, preferably wherein the roasted calcine has a pH in the range of 6.0 to 8.0.

8. The process of any one of the preceding claims, wherein the process comprises (d) leaching at least a portion of the roasted calcine under acidic conditions to provide a pre-cure leach solid and a pre-cure leach liquor; and (f) adding sulphuric acid and/or hydrochloric acid to at least a portion of the pre-cure leach solid to provide a first acidic solid.

9. The process of claim 8, wherein the leaching of the at least a portion of the roasted calcine in step (d) is at a pH of 1.0 to 3.0, preferably 1.0 to 2.0.

10. The process of claim 8 or claim 9, wherein the process comprises separating the pre-cure leach solid and the pre-cure leach liquor, preferably via filtration.

11. The process of any one of the preceding claims, wherein the process comprises (e) obtaining vanadium oxide from at least a portion of the pre-cure leach liquor via solvent extraction and/or precipitation.

12. The process of any one of the preceding claims, wherein the feedstock further comprises aluminium, and where the process further comprises obtaining aluminium oxide from at least a portion of the pre-cure leach liquor via precipitation.

13. The process of any one of the preceding claims, wherein the process further comprises pelletizing the first acidic solid before curing step (g).

14. The process of claim 13, wherein the first acidic solid is pelletized into pellets having an average diameter in the range of 5 to 80 mm

15. The process of any one of the preceding claims, wherein the curing in step (g) is at a temperature in the range of 100 C. to 200 C.

16. The process of any one of the preceding claims, further comprising pulverizing the at least a portion of the cured first acidic solid before leaching step (h).

17. The process of any one of the preceding claims, wherein the leaching of the at least a portion of the cured first acidic solid in step (h) is at a pH of less than 3.0.

18. The process of any one of the preceding claims, wherein the leaching of the at least a portion of the cured first acidic solid in step (h) is at a pH of less than 2.0.

19. The process of any one of the preceding claims, wherein the process comprises separating the post-cure leach solid and the post-cure leach liquor, preferably via filtration.

20. The process of any one of the preceding claims, wherein the hydrolysing at least a portion of the post-cure leach liquor in step (i) is at a temperature in the range of 50 C. to 80 C., preferably wherein the hydrolysing at least a portion of the post-cure leach liquor in step (i) is at a temperature in the range of 50 C. to 70 C.

21. The process of any one of the preceding claims, wherein the feedstock further comprises manganese, iron, chromium, and/or nickel, and wherein the process further comprises obtaining manganese, iron, chromium, and/or nickel from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation.

22. The process of claim 21, wherein the feedstock further comprises manganese, iron, chromium, and nickel, and wherein the process further comprises obtaining manganese, iron, chromium, and nickel from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation; and wherein the process further comprises melting the obtained manganese, iron, chromium, and nickel to provide an alloy comprising manganese, iron, chromium, and nickel.

23. The process of any one of the preceding claims, wherein the feedstock comprises a slag comprising titanium and/or vanadium, or wherein the feedstock comprises one or more of the group consisting of ilmenite, leucoxene, rutile, and complex calcium magnesium aluminium silicates.

24. The process of any one of the preceding claims, wherein the feedstock comprises 0.3 to 5.0 wt % vanadium and/or 5.0 to 85.0 wt % titanium dioxide, based on the total weight of the feedstock.

25. The process of any one of the preceding claims, wherein the process is a process of providing titanium dioxide and vanadium oxide, and wherein the process comprises (i) hydrolysing at least a portion of the post-cure leach liquor to provide titanium dioxide, and (j) obtaining vanadium oxide from at least a portion of the post-cure leach liquor via solvent extraction and/or precipitation.

26. The process of any one of the preceding claims, wherein step (f) further comprises adding water to at least a portion of the roasted calcine or at least a portion of the pre-cure leach solid.

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

Description

[0160] These and other aspects of the invention will now be described with reference to the accompanying Figures, in which:

[0161] FIGS. 1 and 2 are scanning electron microscope images of Material A under backscattered imaging.

[0162] FIG. 3 is a plot of percentage recovery for vanadium during the initial acid leach step, as a function of wt % borax.

[0163] FIG. 4 is a plot showing overall percentage recovery for vanadium for Example 1 as a function of wt % borax.

[0164] FIG. 5 is a plot showing overall recovery for vanadium for Example 2 as a function of wt % borax.

[0165] FIG. 6 is a plot showing the % recovery for TiO.sub.2 following the acid curing steps for Example 1.

[0166] FIG. 7 is a plot showing the % recovery for TiO.sub.2 following the acid curing steps for Example 2.

[0167] FIG. 8 is a plot of sulphuric acid consumption per ton of material against roasting temperature in degrees Celsius ( C.).

[0168] FIG. 9 is a plot of percentage recovery against acid consumption.

[0169] FIG. 10 is a scanning electron microscope image of precipitated TiO.sub.2 via hydrolysis at a lower temperature (less than 70 C.).

[0170] FIG. 11 is a scanning electron microscope image of precipitated TiO.sub.2 via hydrolysis at a temperature of 70-80 C.

[0171] FIG. 12 is a scanning electron microscope image of precipitated TiO.sub.2 via hydrolysis at a temperature of >80 C., i.e. 90-100 C.

[0172] FIG. 13 is a photo of precipitated TiO.sub.2 via hydrolysis at 70-80 C.

EXPERIMENTAL SECTION

Test Methods

[0173] pH is measured by a pH meter. Suitable pH meters are commercially available.

[0174] Diameters and average diameters of particles are measured by techniques known in the art, dependent on the particle size. For example, the average diameter of Titanium dioxide particles is measured by Scanning Electron Microscopy (SEM). Scanning electron microscope images were obtained using a commercially available scanning electron microscope using backscattered imaging.

[0175] Chemical analysis of leach liquors in parts per million (ppm) was performed using known ICP-OES techniques. Examples of analysers that may be used are well known in the art.

[0176] Chemical analysis of feedstocks can be performed using known XRF (X-ray fluorescence) and/or ICP-OES techniques.

Chemical Analysis and Mineralogy of Slag Feedstocks

[0177] A slag feedstock (Material A) was used for all of the Examples below. Chemical analysis (by XRF and ICP-OES) of this feedstock is shown in Table 2.

TABLE-US-00002 TABLE 2 Chemical analysis of Material A Sample TiO.sub.2 V.sub.2O.sub.5 Na.sub.2O CaO SiO.sub.2 MgO Al.sub.2O.sub.3 Fe S Mn Cr ID (wt %) (wt %) (wt %) (wt %) (wt %) (wt %) (wt %) (wt %) (wt %) (wt %) (wt %) Material 29.53 0.88 5.84 15.57 20.09 10.57 12.62 2.74 0.48 0.49 0.11 A

[0178] For mineralogical examination, a representative portion of Material A was cast in epoxy and re-cast in epoxy for transverse section exposure of the sample block for analysis using a scanning electron microscope. Mineralogical data of Material A is shown in Table 3 from which the following observations can be made: [0179] The main phase is the complex CaMgAlSiTiVO, which appears to contain most of the titanium and vanadium; [0180] Rutile, which is one of the titanium-bearing minerals, accounts for 15% of the total weight of the material. Based on this, about 50% of the titanium exists in the rutile form while the rest of the titanium is present in the complex calcium magnesium aluminium silicate.

TABLE-US-00003 TABLE 3 Mineralogical data for Material A Mineral Distribution List % Weight CaMgAlSiTiVO Phase 63.34 Rutile 15.2 AlCaMgOVTiCrFe Phases 6.73 Perovskite 4.18 Fe-oxides (Hematite) 2.24 CaSiMgAlO low Ti Phases 2.18 Rutile-CaAlMg Phases 1.48 Olivine 0.98 CaMgOTi Phases 0.62 Kassite-Mg Al Si Phases 0.50 Ilmenite 0.36 Calcite 0.20 Quartz 0.14 Pyrrhotite 0.08 Kassite 0.07 Dolomite 0.07 Chlorite 0.05 Olivine-CaTi Phases 0.04 Anhydrite 0.03 Braunite 0.03 Others 0.03 Biotite 0.03 Actinolite 0.02 Kaolinite 0.01 Muscovite 0.01 K-feldspar 0.01 [Unclassified] 1.33 OTHERS 0.03

[0181] Scanning electron microscope images of Material A under backscattered imaging are presented in FIGS. 1 and 2. FIGS. 1 and 2 show that rutile and perovskite are embedded in the complex CaMgAlSiTiVO matrix.

EXAMPLES

Example 1

[0182] A titanium and vanadium-containing slag feedstock (Material A) was mixed with 5-20 weight % borax and roasted (calcined) at different temperatures in an oxidising atmosphere, in the range of 600-790 degrees Celsius. The roast calcines were leached at room temperature (i.e. about 20 C.) in sulphuric acid media (pH of around 1.8). Up to 55 wt % of the vanadium present in Material A was recovered into solution during leaching. FIG. 3 is a plot of percentage recovery for vanadium during the initial acid leach step. The materials were roasted at 700, 750 and 790 C. with varying wt % amounts of borax in air (in the absence of coal) and leached in sulphuric acid media (pH of around 1.8). The recovery of vanadium was dependent on roasting temperature, as shown by FIG. 3.

Example 2

[0183] Example 1 was repeated but roasting was undertaken in a reductive atmosphere and hence coal was added during the roasting step. The roast calcine was leached in acidic environment and the recovery of vanadium into leach solution was similar to the recovery in example one. Therefore, it was concluded that coal has no effect on increasing the recovery of vanadium.

[0184] The acid leach residues from Examples 1 and 2 were admixed with water and 400-600 kg sulphuric acid per ton of acid leach residue and cured for 30 minutes to 3 hours at a temperature of between 160 C. to 250 C. The cured materials were highly acidic. They were leached out in water (without addition of acid). Most of the residual vanadium that remained after the previous steps was recovered into the leach solution. Titanium oxysulphate that had formed during curing was also recovered during this leaching step.

[0185] A summary of results showing the overall recoveries of vanadium and titanium oxide (V and TiO.sub.2) is shown in Table 4. Further information is provided in FIGS. 4 to 8.

TABLE-US-00004 TABLE 4 Results showing overall recovery of vanadium and titanium dioxide after acid curing the leach residue (i.e. the overall recovery from the first acid leaching step and the later acid curing and leaching steps). Type of experiment % Recovery of Vanadium % Recovery TiO.sub.2 Oxidative roasting 74-92 64-91.7 (Example 1) Reductive roasting 70-83 68-80 (Example 2)

[0186] FIG. 4 is a plot showing overall percentage recovery for vanadium for Example 1 as a function of wt % borax (the materials were roasted at 700, 750 and 790 C. with varying amounts of borax in the absence of coal and leached in acid media at a pH of around 1.8. Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached). A comparison of FIGS. 3 and 4 shows that mixing the leach residue with water and sulphuric acid followed by curing and leaching increases the overall recovery of vanadium.

[0187] FIG. 5 is a plot showing overall recovery for vanadium for Example 2 as a function of wt % borax (the materials were roasted at 700, 750, and 790 C. respectively, with varying amounts of borax in the presence of coal at a coal/borax ratio of 0.8. The roasted samples were leached in acid media at pH of around 1.8. Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached.

[0188] FIG. 6 is a plot showing the % recovery for TiO.sub.2 as a function of wt % borax following the acid curing steps for Example 1. The materials were roasted at 700, 750 and 790 C. with varying amounts of borax in air (in the absence of coal). The roasted samples were leached in acid media (pH of around 1.8). Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached.

[0189] FIG. 7 is a plot showing the % recovery for TiO.sub.2 as a function of wt % borax following the acid curing steps for Example 2. The materials were roasted at 700, 750 and 790 C. in the presence of coal at a coal/borax ratio of 0.8. The roasted samples were leached in acid media (pH of around 1.8). Sulphuric acid was added to the leach residue, and the first acidic solid was cured and then leached.

[0190] Results showing sulphuric acid consumption in the first acid leaching step (step (d)) against roasting temperature for the first acid leaching step, are shown in FIG. 8. FIG. 8 is a plot of sulphuric acid consumption per ton of feedstock against roasting temperature in degrees Celsius ( C.). FIG. 8 shows that acid consumption is sensitive to both quantity of borax and roasting temperature. In essence, consumption of sulphuric acid increases with increase in roasting temperature and amount of borax. It can be seen that consumption of sulphuric acid generally increases when samples are roasted above 800 C. Therefore, it is thought to be advantageous to roast samples below 800 C.

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

[0191] The effect of pH during leaching of the roasted calcine samples in sulphuric acid media was studied by leaching the samples at different pH and the results are shown in Table 6. It is evident from the results in Table 6 that acid consumption increases with decrease in leaching pH and this is thought to be due to decreased selectivity at lower pH or higher acidity. Nonetheless, the highest recovery of vanadium was achieved at a pH of 1.4 and the weight loss was higher than at any other pH.

TABLE-US-00005 TABLE 6 Effect of pH during leaching of the roasted calcine samples in sulphuric acid media % Recovery Acid consumption Leaching pH Vanadium % Weight loss (kg/ton ore) 0.8 26.78 16.00 664 1.2 22.30 13.40 233 1.4 39.73 17.90 220 1.6 21.40 14.00 182 1.8 23.38 15.10 171 2.2 19.13 12.10 159

Effect of Amount of Acid Addition

[0192] The effect of acid addition per ton of material was studied because acid, in this case sulphuric acid, is one of the major costs in TiO.sub.2 production. The results showing the dependency of recoveries of TiO.sub.2 and vanadium on acid consumption are shown in FIG. 9. FIG. 9 is a plot of percentage recovery against acid consumption. The roasted calcine samples were leached in sulphuric acid at a pH of 1.8, and sulphuric acid was added to the residue. This was subsequently cured at a temperature of 160 C. to 250 C.

[0193] FIG. 9 shows that recoveries for both TiO.sub.2 and vanadium are dependent on acid consumption. The recoveries increase as sulphuric acid addition is increased, thought to be due to increased sulphation of TiO.sub.2 and vanadium. FIG. 9 also shows that increasing acid consumption per ton to more than 700 kg per ton of material does not have a significant effect on the recovery of TiO.sub.2 and vanadium.

[0194] Based on the results shown in FIG. 9, it can be concluded that the optimum acid consumption is around 700 kg per ton of feedstock material. Recoveries of 68 and 74% for TiO.sub.2 and vanadium, respectively, were achieved. The acid consumption is lower than known processes, in which acid consumption is generally more than 1,000 kg per ton of ore (feedstock).

Effect of Size of Particles

[0195] Screen analysis was performed on samples taken from the cured leached residue (post-cure leach solid) to understand the distribution of vanadium and TiO.sub.2 according to particle size after leaching step (h). The results are shown in Table 7. The finest screen size fraction (less than 0.038 mm) had the lowest percentage of vanadium and TiO.sub.2. Table 7 shows that more vanadium and TiO.sub.2 was unobtainable in the coarser size fractions. Put another way, it was more difficult to liberate vanadium and TiO.sub.2 from the larger particles. It is therefore thought to be advantageous to pulverize or crush the cured first acidic solid before leaching.

TABLE-US-00006 TABLE 7 Screen analysis of the cured leached residue sample Particle size (diameter) (mm) % Mass % Vanadium % TiO.sub.2 More than 0.075 19.64 0.68 15.76 More than 0.063 to less than 20.68 0.63 16.18 0.075 More than 0.038 to less than 26.83 0.64 12.44 0.063 Less than 0.038 32.85 0.36 7.87

Effect of Pelletizinq Before Curing

[0196] Two experiments were carried out to investigate the effect of pelletizing before curing. Samples of first acidic solids were prepared by mixing Material A with borax, roasting, leaching, and adding sulphuric acid and water. The first acidic solid samples (using Material A and according to steps (a) to (f)) were pelletized, having a diameter of 5 to 80 mm. The pellets were cured for 30 minutes to 2 hours at 160-250 C. The cured pellets were then pulverized, taking care to avoid fines generation, and the resulting particles were leached out in an acidic environment. The results showing the recoveries of vanadium and TiO.sub.2 are shown in Table 8. These results can be compared to results shown in FIG. 9, for example. The recoveries of both vanadium and TiO.sub.2 were higher for similar levels of acid consumption, which is thought to be due to increased sulphation when the pellets are being baked. Without being bound by theory, it is thought that sulphur dioxide gas does not escape easily from the pellets, thereby leading to increased sulphation and consequently increased percentage recovery of vanadium and TiO.sub.2.

TABLE-US-00007 TABLE 8 Summary of results for samples that were calcined, leached in acid, and cured via pelletizing Acid consumption (kg sulphuric acid/ton % Recovery feedstock) Vanadium TiO.sub.2 500-750 70-78 85-88 600-800 80-85 90-92
Obtaining Vanadium from Pre-Cure Leach Liquor

[0197] Previously, vanadium has conventionally been extracted by leaching roasted calcines in water (generally in an alkaline environment due to the make-up of the calcines).

[0198] The present inventors roasted samples in air in the presence of borax for durations less than and more than an hour, respectively. The roasted calcine samples were then treated as follows: [0199] Leaching the roasted calcines sample in water (alkaline environment) and re-leaching the residue in acidic environment; or [0200] Directly leaching the roasted calcine samples in an acidic environment.

[0201] The summary of results for the samples is shown in Table 9 which shows that water leaching is not as effective at extracting vanadium from the roasted calcine samples. A majority of vanadium together with boron are extracted when the roasted calcine samples are directly leached out in acidic environment. Similarly, extraction of aluminium is more efficient when the roasted calcine samples are directly leached in an acidic environment. Based on the results in Table 9, it can be concluded that calcined samples should preferably be directly leached out in acidic environment to maximise vanadium/titanium recovery.

TABLE-US-00008 TABLE 9 Chemical analysis of the various leach liquors using ICP-OES technique in parts per million (ppm). The samples had been roasted (calcined) in air in the presence of borax Roasting Al Ca Co Cu Fe Mg Mn Sample Time (hr) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) Water (alkaline) leach <1 804 4.2 0.1 2.8 14.4 0.1 1.2 Water (alkaline) leach >1 787 3.2 0.1 5.8 14.7 0.1 0.8 Water (alkaline) leach followed <1 678 429 0.2 7.1 335 635 82 by Acid leach at pH = 1.8 Water (alkaline) leach followed >1 1206 415 0.4 9.8 291 581 61 by Acid leach at pH = 1.8 Direct Acid leach at pH = 1.8 <1 2528 500 0.08 9.3 361 436 61 Direct Acid leach at pH = 1.8 >1 1733 405 0.18 11.4 499 494 91 Roasting Ni Na K Cr V Ti B Sample Time (hr) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) (ppm) Water (alkaline) leach <1 0.1 3681 210 0.1 92 0.1 860 Water (alkaline) leach >1 0.1 3586 262 0.5 82 0.1 838 Water (alkaline) leach followed <1 0.4 712 17 5.2 47 128 320 by Acid leach at pH = 1.8 Water (alkaline) leach followed >1 0.5 989 18 11.7 78 122 152 by Acid leach at pH = 1.8 Direct Acid leach at pH = 1.8 <1 0.4 3191 139 11.7 269 192 1566 Direct Acid leach at pH = 1.8 >1 1.1 2879 120 7.2 223 343 1837

Example 3

[0202] Two samples of Material A were mixed with borax and then roasted in air at a temperature of 790 C. for (a) 1 hour and (b) 1.5 hours, respectively. The roasted calcine samples (a) and (b) were not subjected to any leaching steps. Instead, the roasted calcine samples (a) and (b) were mixed with sulphuric acid and water, pelletized, and then cured at a temperature of 160 C. to 250 C. for 0.5 to 1.5 hours. The cured samples were leached in water. Titanium dioxide and Vanadium and were recovered from the post-cure leach liquor via hydrolysis and solvent extraction/precipitation. The results are shown in Table 10 from which it can be observed that recoveries were 41-48% for vanadium and 39-44% for titanium.

[0203] The leach residues (the post cure leach solid) in Table 10 was again mixed with sulphuric acid and water and pelletized before re-curing at a temperature of 300 C. in order to increase recoveries of vanadium and titanium. The cured samples were leached in water. Titanium dioxide and V.sub.2O.sub.5 and were recovered from the post-cure leach liquor via hydrolysis and solvent extraction/precipitation. The results showing overall recoveries of vanadium and TiO.sub.2 are presented in Table 11. Table 11 shows that 85-90% of vanadium and 89-91% TiO.sub.2 were recovered. The weight losses were very high implying that most of the constituents were extracted into the leach liquor. Table 11 shows that that only about 30 weight % of the waste remains after extraction.

[0204] Furthermore, Tables 10 and 11 also indicate that roasting (calcination) should preferably be carried out for more than 1 hour, in order to achieve increased recoveries.

TABLE-US-00009 TABLE 10 Leach residue results for samples that were roasted in air, pelletized with sulphuric acid and cured, followed by leaching Roasting (calcination) % Weight % Recovery % Recovery Time loss Vanadium TiO.sub.2 1 hour 30.37 41.96 39.29 1.5 hours 34.00 47.92 44.15

TABLE-US-00010 TABLE 11 Leach residue results for the samples that were roasted in air at 790 C., pelletized and cured at 300 C., followed by leaching Roasting (calcination) % Weight % Recovery % Recovery Time loss Vanadium TiO.sub.2 1 hour 67.67 85.55 89.16 1.5 hours 70.33 90.09 91.48

Hydrolysis of Titanium Dioxide (TiO.SUB.2.)

[0205] Titanium dioxide was precipitated from the solutions (post-cure leach liquor) from Examples 1 to 3 via hydrolysis which involved heating the pregnant solution at a temperature of more than 50 C. The particle size of the hydrolysed titanium dioxide is of importance because fine particles can create problems during any subsequent chlorination steps.

[0206] It was found that growth of the titanium dioxide particles could be controlled by varying the hydrolysis temperature.

[0207] Scanning electron microscopy images for the TiO.sub.2 that was hydrolysed at varying temperature are shown in FIGS. 10-12. When the hydrolysis was performed at less than 70 C., it was found that the particle sizes (average particle diameter) were around 20-60 m, as shown by FIG. 10. When the hydrolysis was performed at higher than 70 C. to 80 C., it was found that the particle sizes (average particle diameter) were around 3 m to 20 m, as shown by FIG. 11. When the hydrolysis was performed at higher than 80 C., it was found that the particle sizes (average particle diameter) were less than 3 m, as shown by FIG. 12.

[0208] A digital image of the TiO.sub.2 precipitates which was produced via hydrolysis at more than 70-80 C. is shown in FIG. 13. As shown in FIG. 13, precipitated TiO.sub.2 had white colour, thought to be due to a low iron content. The digital image in FIG. 13 further shows that the precipitated TiO.sub.2 settled well.

[0209] FIG. 10 is a scanning electron microscope image of the precipitated TiO.sub.2 via hydrolysis at lower temperature (less than 70 C.).

[0210] FIG. 11 is a scanning electron microscope image of the precipitated TiO.sub.2 via hydrolysis at a temperature of 70-80 C.

[0211] FIG. 12 is a scanning electron microscope image of the precipitated TiO.sub.2 via hydrolysis at a temperature of >80, i.e. 90-100 C.

[0212] FIG. 13 is a photo of the precipitated TiO.sub.2 via hydrolysis at 70-80 degrees Celsius.

Solvent Extraction

[0213] Solvent extraction was carried out on Example 1 to extract vanadium and base metals (Cr, Ni, Co, Mn etc.) that dissolved in solution during leaching after acid curing. The procedure was as follows: [0214] The aqueous solution, after hydrolysis of titanium, was fed to the solvent extraction process. This liquor contained vanadium, nickel, iron, chromium, and aluminium. [0215] The initial pH of the liquor was determined and adjusted via addition of lime or sodium hydroxide or ammonia solution. [0216] An organic solution was prepared at a strength of 5%-40% v/v (extractant to diluent). [0217] The aqueous solution was contacted with the organic solution for 10-15 minutes in a ratio of 1:2 to aid the efficient separation of the loaded organic and raffinate. [0218] The loaded organic (solvent) is subjected to a two-stage scrubbing process using 2 gpl of H.sub.2SO.sub.4 for 10-15 minutes to extract the metal ions (Fe, Al, Ni and Cr) from the loaded organic. [0219] The raffinate contained the desired vanadium metal ions which were recovered and V.sub.2O.sub.5 obtained.

[0220] Chemical analysis of base metals extracted from the loaded organic is shown in Table 12. The main components were iron, chromium, nickel, and manganese.

TABLE-US-00011 TABLE 12 Chemical analysis of the extracted base metals % Mn % Fe.sub.2O.sub.3 % Cr.sub.2O.sub.3 % NiO 1.64 64.34 16.55 10.20

[0221] The base metals iron, chromium, nickel, and manganese were melted to obtain an alloy comprising manganese, iron, chromium, and nickel.