WO1996010096A1

WO1996010096A1 - Process and solution for extracting metal

Info

Publication number: WO1996010096A1
Application number: PCT/AU1995/000643
Authority: WO
Inventors: Christopher John Browne; Peter Robert Hutchison
Original assignee: Tioxide Australia Pty. Ltd.
Priority date: 1994-09-29
Filing date: 1995-09-28
Publication date: 1996-04-04
Also published as: NZ293523A; PE18296A1; CO4440543A1

Abstract

A process for extracting one or more metals from a metal containing material, as herein defined, including the step of treating said metal containing material with an aqueous, acidic solution containing effective concentrations of iron-containing species, sulphuric acid and one or more leach rate accelerants present in the co-product solution from the manufacture of TiO2 by the Sulphate Process, whether said leach rate accelerants are present as a result of using a co-product solution or are separately compounded.

Description

PROCESS AND SOLUTION FOR EXTRACTING METAL

FIELD OF THE INVENTION

This invention relates to a process for extracting metal from a metal containing material and a solution for use in such a process. The process is particularly useful for extracting metals by dissolution of a metal mixture, metal alloy or metal compound in an acidic solution, thereby releasing metal or metals for recovery in such forms as metallic particles, as ions in solution or by precipitation as metal compounds. The process finds particular application in the economic utilisation of waste co-products of mineral processing operations. The process can further extend to economic utilisation of metal containing waste, such as mine tailings, in order to extract metals and/or metal compounds of economic significance.

BACKGROUND QF THE INVENTION

With increasing world wide concern about pollution of the environment, governments have introduced stricter controls on the disposal of waste by¬ products and effluent of industrial and mining processes. One such process is the production of titania, Tiθ2, from titanium-bearing ore or concentrate. Titania is an important compound and principally finds use as a pigment in the manufacture of paint. A significant portion of Tiθ2 is produced by the so called "Sulphate Process". As used throughout the specification, the term "Sulphate Process" means a process in which a titanium bearing compound, such as ilmenite (FeTiOβ), beneficiated ilmenite, titanium slag, or similar material, after being dried and milled, is digested with concentrated sulphuric acid to ultimately give Tiθ2 and the co-products FeS04 and H2SO4 by the following chemical reactions ( in the case of ilmenite):

FeTiθ3 + 2 H2SO4 → TiOS04 + FeS04 + 2H₂0 (1)

(ilmenite) OH^"

TiOSθ4 → Tiθ2nH2θ + H2SO4 (2)

Heat TiOnH2θ → Tiθ2 + nH₂0 (3) The Tiθ2 can be precipitated from solution using either the Mec lenberg or Blumenfeld process. The Ti0₂ nuclei or seed, prepared from, for example, TiCI₄, may be introduced at this stage if the Mecklenberg process is used. The precipitate may comprise a solid hydrate of Ti0₂. The precipitate undergoes a first filtration and the acidic filtrate including H2SO4, FeSθ4 and other metal ions and compounds is recovered. The Ti0₂ precipitate is then washed with water prior to being leached under reducing conditions to remove remaining trace contaminants. The washings have a similar composition to the acidic filtrate, only more dilute. A second filtration follows the leaching step and a second filtrate, even more dilute than the pre leach washings is recovered. The filtered TIO2 precipitate is preferably again washed with water to further improve purity. The final washings again include some H2SO4, FeSθ4 and other metal ions and compounds but are even more dilute than the second filtrate.

Until the recent introduction of tighter controls on effluent disposal, the recovered filtrate solutions and both pre and post leach washings have been individually and/or collectively disposed of as waste effluent into natural waterways, leading to unsightly discolouration of waterways, or neutralised at high cost and increased environmental burden.

Clearly, there is a need for an effective, environmentally acceptable and economic means of utilising these waste or co-product acid streams from the titania manufacturing process.

One partial solution to the problem has been to use feedstock with lower Fe content. One such feedstock which has found considerable use is titanium slag, which is produced by the electric arc smelting of ilmenite. Titanium slag has a much lower Fe content (around 8% Fe) than ilmenite and is rich in Ti (approximately 70 to 80% T1O2). By using titanium slag as a feedstock, the amount of iron in the waste effluent is significantly reduced. However, the problem of disposal of H2SO4 in the effluent still remains. Although H2SO4 could be concentrated and recycled to an earlier stage of the process, it is not always economical to do this.

In their search for an environmentally acceptable, yet practically and economically useful utilisation for the waste, or co-product acid streams from the T1O2 manufacturing process, the inventors have discovered that treatment of metal containing material, such as metal compounds, metal mixtures, metal alloys, mineral processing products and by-products, including mineral concentrates, tailings, and by-products of metal refining with the waste or co- product acid streams from the TIO2 manufacturing process is surprisingly effective in extracting the metal or metals of interest. These acid streams are particularly useful for extracting metal or metals from mineral processing products and by-products and other metal containing wastes. The waste or co-product acid streams can extract metals from tailings produced from processing extremely fine grained ore, from which it is normally uneconomical to extract metals. The solution is effective in leaching minerals such as metal sulphides and oxides. The solution is particularly effective in leaching transition metals and noble metals, such as one or more of Zn, Pb, Cu, Ag, Au, Ni and Mn. The rate and degree of leaching using the waste or co-product solution can be significantly and surprisingly higher than that using a solution having an equivalent concentration of H2SO4 alone or that using a solution having equivalent concentrations of H2SO4 and Fe ions. It is therefore apparent that the waste acid or co-product solution includes additional component/s which assist in accelerating the rate of leaching of the metal containing material relative to H₂S0₄ and H₂S0₄/Fe solutions.

SUMMARY OF THE INVENTION

According to the present invention, there is provided a process for extracting one or more metals from a metal containing material, as herein defined, including the step of treating said metal containing material with an aqueous, acidic solution containing effective concentrations of iron-containing species, sulphuric acid and one or more leach rate accelerants present in the co-product solution from the manufacture of Ti0₂ by the Sulphate Process, whether said leach rate accelerants are present as a result of using a co-product solution or are separately compounded.

The present invention also provides a process for extracting one or more metals from a metal containing material, as herein defined, including the step of treating said metal containing material with an aqueous acidic solution comprising the co-product solution from the manufacture of TiO₂ by the Sulphate Process.

As used throughout the specification, the term "metal containing material" is intended to mean any material containing one or more metals and includes metal or metal oxide mixtures, such as by-products of metal refining, metal alloys, or metal compounds such as ore minerals, mineral concentrates and tailings.

Preferably, the aqueous acidic solution is the filtrate recovered from the first filtration of the titanium containing precipitate. However, the aqueous acidic solution may comprise the filtrate or washings from any step of the Sulphate Process, either singly or in combination with one or more filtrate or washings from other steps of the process.

For most applications of the process of the invention, it is preferred that prior to and/or during treatment of the metal containing material with the acidic solution, an oxidising agent, such as an oxidising solution or an oxidising gas, is introduced into the acidic solution. The oxidising agent is preferably an oxidising gas, such as oxygen gas or an oxygen containing gas such as air. If an oxidising gas is used, it is preferably introduced by bubbling the gas through the acidic solution. The oxidising gas is preferably introduced into the acidic solution during treatment of the metal containing material with the acidic solution. A preferred means of introducing the oxidising gas into the solution is by using an aeration tube or a glass frit type aerating tube.

Agitation of the resultant solution and slurry is preferred to adequately suspend the solid particles to be leached and to disperse the oxidising agent.

The present invention also provides an aqueous acidic solution for extracting one or more metals from a metal containing material, wherein said solution is produced as a waste or a co-product from the manufacture of Tiθ2 by the Sulphate Process.

In another form of the invention, a metal is extracted from a metal containing material, using an aqueous acidic solution including one or more components of the co-product solution from the manufacture of Tiθ2 by the

Sulphate Process, whether present as a result of using a co-product solution or separately compounded. The ratio of the metal containing material to the aqueous acidic solution required for the leaching process will depend on many factors, such as the concentration of active components in the aqueous acidic solution, the concentration and form of metal in the material to be leached and the form of treatment to which the process is applied. However, for most applications, the ratio of the metal containing material to solution will be between 0.001 and 300% w/v. Preferably the range is from 10 to 50% w/v.

As previously stated, the preferred aqueous acidic solution is the filtrate recovered from the first filtration of the titanium containing precipitate of the Sulphate Process.

Preferably, the acidic solution is treated with an oxidising agent. The oxidising agent may be an oxygen containing gas such as 0₂ or air. Alternatively, the oxidising agent may be a peroxide, nitric acid or a hypochlorite. The oxidising agent is preferably an oxidising gas such as air, oxygen, ozone or mixtures thereof. More preferably, the oxidising agent is 0₂(g). The oxidising gas may be bubbled through the acidic solution, such as by using an aerating tube.

The amount of oxidising agent introduced to the aqueous acidic solution is preferably 1 to 2 times the stoichiometric amount needed to achieve the desired reaction rate. Where oxygen gas is the oxidising agent, it is preferably introduced into the solution at a rate of from 0.001 to 2.0 grams of oxygen per litre of leaching solution per minute. For example, in the case of material containing 5% Zn and 5% Pb, it may be necessary to feed oxygen at a rate of between 0.05 and 0.10 grams of oxygen/litre/minute into a leaching solution containing 20% w/v of metal containing material in order to achieve 90% or above reaction in 60 minutes.

The leaching process may be conducted over a wide pressure range. Preferably, the process is conducted at atmospheric pressure.

Chemical analyses of the aqueous acidic solutions formed as a co-product of Tiθ2 manufacture indicate that these solutions include sulphuric acid, iron and typically at least one of the following: titanium containing species (dissolved and/or particulate), manganese, chromium, magnesium, aluminium, vanadium and chloride. Other constituents which may be present include silicon, zirconium, 96/10096 PC-7AU95/00643

6 zinc, arsenic, barium, cadmium, copper, lead, mercury, nickel, niobium, thorium, uranium and selenium.

The pH of the aqueous acidic solution is of course acidic. Preferably, the pH of the solution is no higher than 6.5. More preferably, solution pH is in the range of 0 to 5.

Unless otherwise specifically indicated, the following discussion relates to the preferred aqueous acidic co-product solution, being the filtrate recovered from the first filtration of the titanium-containing precipitate formed using the Sulphate Process. However, it is to be understood that the invention is not limited to using that preferred solution. In order to facilitate the following discussion, that preferred aqueous acidic co-product solution will be hereafter referred to as "Tioleach" solution.

The "Tioleach" solution may have up to 250 g/l iron. Typically, however, the maximum iron concentration is around 80 g/l, such as in Tioleach solutions derived from the Sulphate Process using ilmenite feedstock. Preferably those solutions have up to 60 g/l iron. For those solutions, the minimum iron concentration may be around 20 g/l. However, where the solution is derived from the Sulphate Process using titanium slag feedstock, the maximum iron content can be around 20 g/l. Some solutions have iron concentrations as low as 0.1 g/l, whereas in other embodiments the iron concentration may be 5 g/l or higher. Preferably, the minimum iron concentration is 1 g/l.

In an embodiment of the invention, the process includes the further step of treating the Tioleach solution in order to effect partial removal of iron from solution. This step may be included where the Tioleach solution is derived from the Sulphate Process using ilmenite feedstock.

The Tioleach solution may have free H2SO4 present in solution at a concentration up to that for pure sulphuric acid, such as around 1800 g/l. However, typically the concentration of H2SO4 is at most 500 g/l, such as up to 350 g/l. In some embodiments of the invention, the concentration of H₂S0₄ can be up to 250 g/l. In other embodiments, H2SO4 has a minimum concentration of 1 g/l, such as 10 g/l or higher. In a preferred embodiment, the concentration of H2SO4 is in the range of from 150 g/l to 350 g/l. Depending on solution chemistry, titanium may be present in the Tioleach solution as either dissolved titanium compounds or particulate titanium compounds, or as both dissolved and particulate compounds. The total concentration of titanium in the Tioleach solution will be hereinafter expressed as the equivalent concentration of Ti0₂. The total concentration of Tiθ2 in the Tioleach solution, may be up to 200 g/l. Usually, the total titanium concentration is no higher than 20 g/l. The total titanium concentration expressed as equivalent concentration of Tiθ2, may have a minimum of 0.1 g/l. However, typically the minimum concentration is 2 g/l. The maximum dissolved Tiθ2 may be around 150 g/l, with a preferred upper limit for dissolved Tiθ2 being around 15 g/l. Lower limits of dissolved Tiθ2 concentration may be around 0.1 g/l although typically the lower limit is around 1 g/l. In a preferred embodiment, the concentration of dissolved Ti0₂ is between 1 and 15 g/l.

Particulate titanium containing compounds may be present in the Tioleach solution in one or more forms, including trivalent or tetravalent titanium containing oxides, hydroxides, hydrous oxides, sulphates, hydroxysulphates, chlorides or oxychlorides. Examples of titanium containing compounds include Tiθ2, TiO(OH)₂, Ti0₂.2H₂0, Ti(OH)₃, TiCI₄, TiOCl2 and Ti(OH)_xCI₍₄._x). As stated above, concentrations of titanium containing compounds will be expressed throughout this specification as the equivalent concentration of Tiθ2- The size of titanium-containing particulate solids may range from approximately 10 μm down to colloidal size particles. Typically, the particle size may range from 0.1 to 1 μm, such as from 0.2 to 0.5 μm.

The Tioleach solution preferably also includes chloride species. Chloride may be present at a concentration of up to 20 g/l. However, in some embodiments, it is present at a concentration of 10 g/l or less. Typically, the minimum chloride concentration is around 0.5 g/l.

Manganese may be present in the Tioleach solution at a concentration of up to 20 g/l. However, in some embodiments, the manganese concentration is no higher than 2 g/l. Other embodiments of the solution have lower manganese concentration, such as a minimum of 0.01 g/l. Typically, however, the minimum concentration of manganese is 0.15 g/l. The Tioleach solution may further have up to 4 g/l chromium Typical chromium concentrations are up to 0.4 g/l. The minimum chromium concentration may be 0.001 g/l, such as 0.03 g/l or higher.

Magnesium may also be present in the Tioleach solution up to 60 g/l. In some embodiments, the maximum magnesium concentration is around 5 to 6 g/l. Other embodiments have a minimum magnesium concentration of around 0.01 g/l, however a concentration of at least 0.15 g/l is typical.

Aluminium may also be present in the Tioleach solution at a concentration up to 30 g/l. Typical concentrations are up to 5 g/l, with concentrations up to 3 g/l being preferred. The minimum concentration of aluminium may be around 0.01 g/l, with a minimum of 0.25 g/l being typical.

The Tioleach solution may further have up to 10 g/l vanadium, such as up to 0.75 g/l. The minimum vanadium concentration may be around 0.001 g/l, although typically it is 0.05 g/l or higher.

The Tioleach solution may have up to 100 g/l total suspended solids. Preferably, the solution has up to 10 g/l total suspended solids.

Typical concentration ranges for other components of the Tioleach solution are given in, but not limited by, Table I.

TABLE I

Element Concentration (mg/l)

Arsenic 0.04 - 1.4

Barium 0.03 - 30

Cadmium 0.002 - 1.0

Copper 0.40 - 4.0

Lead 0.20 - 9.0

Mercury 0.001 - 0.90

Nickel 3.0 - 12.0

Niobium 0.03 - 0.3

Silicon 7.0 - 35.0

Thorium 0.03 - 18.0 Uranium 0.03 - 6.0

Zinc 2.0 - 300

Zirconium 4.0 - 400

The leaching process may be conducted over a wide temperature range from, for example, 0 to 300°C, with the higher end of the temperature range covering embodiments undergoing pressure leaching. However, it is preferred that the Tioleach solution is reacted with the metal containing material at an elevated solution temperature. For most process conditions, leaching rate increases with increasing temperature. Preferably, the solution temperature is at least ambient temperature. More preferably, the solution temperature is at least 60°C. For some process conditions, leaching rate increases sharply at 70°C and above. Leaching rates of some ore tailings are optimised between 70°C and 90°C. However, for other embodiments, the solution temperature is 90°C or higher.

Leaching rate is also dependent on the chemical and physical form of metal in the metal containing material, such as particle size, chemistry and percentage of constituent particles and overall metal content. Where the metal containing material is mine tailings containing extremely fine grained and intermixed ore minerals, or where the average particle size of the metal containing material is unacceptably high, leaching rate can be relatively slow. Grinding of such material prior to leaching can assist in releasing ore minerals, resulting in an improvement of leaching rate.

The process of the invention can also be used in heap leaching. Leaching rate can be improved, where appropriate, by agitating the

Tioleach solution. Agitation is effected to ensure the metal containing material is adequately suspended in the solution and the oxidising species is adequately dispersed in the Tioleach solution. However, agitation is not always appropriate, for example in the case of heap leaching. Agitation during the leaching process can have the disadvantage of causing foaming or frothing of the solution. Foaming can entrain solids and physically separate them from the leaching solution, thereby making the handling of the solution difficult. An effective amount of the foam control agent may therefore be advantageously added to the leaching solution. One such foam control agent is calcium lignosulphonate. It may be present at a concentration of up to 1 % w/v. However, for most applications, the calcium lignosulphonate has a maximum concentration of 0.05% w/v, such as around 0.025% w/v. The minimum concentration of calcium lignosulphonate is typically around 0.0001% w/v.

Where the metal containing material contains metal sulphides, the leaching process can result in formation of free sulphur and/or sulphur compounds which may coat unreacted metal containing particles. This phenomenon can prevent or reduce reaction of the coated particles with the leaching solution, thereby adversely affecting the leaching rate. Coating by sulphur-containing material is particularly problematic where the material being leached contains chalcopyrite. This problem can be alleviated by including the step of attrition of the metal containing material. This may be effected by addition of an attriting agent to the leaching solution during agitation thereof. The attriting agent assists to physically remove the sulphur containing coating by attrition, thereby exposing the surface of the unreacted particles to the leaching solution. A suitable attriting agent is particulate Si0₂, such as sand. Where an attriting agent is used, it is preferably present in an amount which is approximately equal to the amount of metal containing material. Thus the ratio of sand to metal containing material is preferably 0.2:1 to 1.5:1.

Attriting can also be effected by increasing the ratio of solid metal containing material to aqueous solution in the reaction mixture. Calcium ligonsulphonate, in addition to its defoaming properties, also acts as a dispersent of free sulphur and/or sulphur compounds. Thus, the addition of both sand and calcium lignosulphonate to the Tioleach solution further enhances leaching rate.

The Tioleach solution may additionally contain one or more leaching promoters. Such promoters include copper ions and/or ions derived from carboxyllic acids, such as acetic acid. The copper may be added to solution such as by adding copper sulphate, CuS0₄.5H₂0. Alternatively, the copper may be already present in the Tioleach solution, such as where copper is incorporated into the waste or co-product solution during the Sulphate Process, or where copper has been released into the solution as a result of leaching copper containing materials, eg. tailings, mineral concentrates etc. Acetate ions may be added as acetic acid. The preferred concentration of copper ions is about 0.6 g/l. Acetate ions, if present, are preferably present at a concentration of about 1.25 g/l, expressed as equivalent amount of acetic acid.

DESCRIPTION OF THE DRAWINGS The invention will become more readily apparent from the following exemplary description in connection with the accompanying drawings and Examples:

FIGURE 1 is a schematic diagram of one embodiment of apparatus which can be used in the leaching process of the present invention. FIGURE 2 is a graph plotting the amount of copper leached (percent) versus time for Example 8 (diamonds) and Comparative Examples 1 and 2 (triangles and squares, respectively).

FIGURE 3 is a graph plotting the amount of zinc leached (percent) versus time for Example 8 (diamonds) and Comparative Examples 1 and 2 (triangles and squares, respectively).

FIGURE 4 is a graph showing the amount of copper and zinc leached (percent) versus time for Example 9. The squares represent zinc and diamonds represent copper.

FIGURE 5 is a graph plotting the amount of copper leached (percent) over time for Example 10 (diamonds) and Comparative Examples 3 (triangles) and 4 (squares).

FIGURE 6 is a graph plotting concentration (ppm) of copper (diamonds) and reacted iron (squares) in solution for Example 10.

FIGURE 7 is a graph plotting the amount of zinc leached versus time for Example 11 (squares) and Comparative Examples 5 (triangles) and 6 (diamonds). FIGURE 8 is a graph plotting the amount of zinc leached versus time for Examples 12 (open squares), 13 (diamonds), 14 (triangles) and 15 (closed squares).

FIGURE 9 is a graph plotting the amount of zinc leached versus time for Examples 16 (squares), 17 (triangles) and 18 (diamonds).

FIGURE 10 is a graph plotting the amount of zinc leached versus time for Examples 19 (squares), 20 (triangles) and 21 (diamonds).

FIGURE 11 is a graph plotting the amount of zinc leached versus time for Examples 22 (triangles), 23 (diamonds) and 24 (squares). FIGURE 12 is a graph plotting the amount of zinc leached versus time for

Examples 25 (diamonds) and 26 (squares).

FIGURE 13 is a graph plotting the amount of zinc leached versus time for Examples 27 (open squares), 28 (triangles), 29 (diamonds) and 30 (closed squares). FIGURE 14 is a graph plotting the amount of copper leached versus time for Example 31 (triangles) and Comparative Examples 7 (squares), 8 (diamonds) and 9 (open squares).

FIGURE 15 is a graph plotting the amount of copper leached versus time for Examples 32 (open diamonds), 33 (triangles), 34 (open squares) and 35 (closed diamonds) and Comparative Example 10 (closed squares).

FIGURE 16 is a graph plotting the amount of Zn leached versus time for Examples 36 (circles), 37 (squares) and 38 (triangles).

FIGURE 17 is a graph plotting the amount of Zn recovered versus time for Examples 39 (diamonds) and 40 (squares). FIGURE 18 is a graph plotting the amount of Cu recovered versus time for

Examples 41 (diamonds) and 42 (squares).

FIGURE 19 is a graph plotting the amount of Cu leached versus time for Example 43 (squares) and Comparative Example 11 (diamonds).

FIGURE 20 is a graph plotting the amount of Zn recovered versus time for Examples 44 (triangles) and 46 (diamonds).

FIGURE 21 is a graph plotting the amount of Cu recovered versus time for Examples 45 (diamonds) and 47 (squares). DETAILED DESCRIPTION OF THE PREFERRED EMBODIMENTS

The following non-limiting Examples illustrate, in detail, embodiments of the invention. Although the Examples relate principally to extraction of metals from ore minerals, and particularly from waste mine tailings, it is to be understood that the invention is not limited to that application. It has been found that the process of the invention can extract, inter alia, greater than 90% of zinc, lead and copper, up to about 80% silver and more than 40% gold from ore minerals. EXAMPLES 1 to 3

Compositions of Tioleach solutions produced as co-product effluent from the manufacture of Tiθ2 from titanium slag feedstock are provided as Examples 1 and 2 in Table II. In Table II, Example 3 provides the composition of a Tioleach solution produced as co-product effluent from the manufacture of Tiθ2 from ilmenite feedstock. In Table II, "NA" means not analysed.

It is evident from comparison of Examples 1 and 2 with Example 3 that the Ti slag derived solutions of Examples 1 and 2 are lower in such elements as Fe, Mn, Zn, As, Ba, Cu, Pb, Th and U and higher in such components as H2SO4 Mg, Al, Cr, V and Zr, than the ilmenite derived solution of Example 3.

TABLE II

Element Example 1 Example 2

Example 3

(mg/l) (mg/l)

(ma/\)

Aluminium 1980 2060 363

Arsenic <1 0.07 1

Barium <10 0.05 <30

Cadmium <1 0.002 0.6

Chromium 168 297 33

Copper <1 0.76 3

Lead <1 0.27 6

Magnesium 4400 5270 210

Mercury <1 <0.001 0.62 1 i 4

Nickel 8 5.6 8.4

Niobium <1 0.03 0.228

Silicon 20 14 23.1

Thorium NA 0.04 12.3

Uranium NA 0.07 3.6

Vanadium 496 600 99

Zinc 3 4.4 247.2

Zirconium 33 23 7.5

Iron 17000 16000 60,000

Manganese NA 240 1620

H₂S0₄ 215000 247000 150000

Tiθ2(acid soluble) 3870 3980 3600

Tiθ2(dissolved) 3240 3960 3000

Total Suspended Solids 470 4150 1500

Total Dissolved Solids NA 322500 370000

EXAMPLES 4 to 7

Mine Tailings including fine grained lead, zinc, silver and gold-bearing ore were made into a 20% slurry and treated with a Tioleach solution from the manufacture of Tiθ2 from titanium slag. The mine tailings contained 6 to 8% lead, in the form of oxides and sulphides, 6 to 8% zinc, in the form of a sulphide, combined with some silver as sulphide and a high quantity of gold bearing pyrite. The acidic solution acted as a leachant and was effective in dissolving the minerals to allow extraction of the metals. An oxidising gas, such as air, was bubbled through the acidic solution as it reacted with the minerals.

It is to be noted that zinc and lead sulphides can be dissolved without oxidation of the sulphur present in the sulphides. This is advantageous by not increasing the sulphate concentration in solution.

SUBSTITUTE SHEET (RULE 26.) Table III presents the leach times required to achieve 98% recovery of Zn and Pb from the tailings as a function of solution temperature.

TABLE

Example Metal Solution T(°C) Leach Time ^mins-

4 Zn 60 100

5 Pb 60 60

6 Zn 90 75

7 Pb 90 40

It is believed that at a solution temperature of 90°C, zinc is dissolved in solution as zinc sulphate and lead precipitates out of solution as lead sulphate.

EXAMPLES 8 and 9 and COMPARATIVE EXAMPLES 1 and 2

The Tioleach solution produced from the manufacture of Tiθ2 from titanium slag using the Sulphate Process was used to leach metals from samples of copper and silver ore concentrate. The Tioleach solution contained 270 grams per litre of sulphuric acid and 16 grams per litre of ferrous iron. Leach rates of the acidic co-product solution was compared with those of a H2SO4 solution and a H2SO4 + Fe2⁺ solution made from a sulphuric acid solution doped with FeSO4. The comparative solutions had concentrations of H₂S0₄ and Fe²⁺ similar to those of the Tioleach solution. Figure 1 shows one embodiment of apparatus used in the leaching process. Beaker 10 sits in water bath 20 having a constant temperature. To beaker 10 is added a solution of 30 grams of ore concentration and 300 ml of leachant. The solution is agitated by glass stirrer 30. Air from an aquarium air pump (not shown) is bubbled through the solution via glass frit type aerating tube 40 at a flow rate of 1 litre per minute.

In Examples 8 and 9 and Comparative Examples 1 and 2 a copper/silver ore concentrate was mixed with an acidic solution at a ratio of 10% solids. Leaching was conducted at a solution temperature of 70°C, while stirring and air was bubbled through the solution at a rate of 1 litre/minute. Copper and zinc levels in solution were measured against time and are presented in Figures 2 and 3. Figures 2 and 3 show copper and zinc leaching rates from a copper ore concentrate having 8% Cu, 7% Zn and 8% Pb and at a solution temperature of 70°C. In Figures 2 and 3, the diamonds represent Example 8 in which 30 grams of ore concentrate was reacted with 300 ml of Tioleach solution, including 270 g/l sulphuric acid and 16 g/l ferrous iron, the triangles represent Comparative Example 1 in which 30 grams of ore concentrate was reacted with 300 ml of a solution comprising 270 g/l H₂S0₄ and 16 g/l Fe²⁺ and the squares represent Comparative Example 2 in which 30 grams of ore concentrate was reacted with 300 ml of a solution having 270 g/l H2SO4.

It is evident from Figures 2 and 3, that the leach rate of copper and zinc by the Tioleach solution in Example 8 is considerably higher than that for the solution containing equivalent concentrations of H2SO4 and Fe²⁺ only (Comparative Example 1), or H2SO4 only (Comparative Example 2). Clearly then, one or more of the additional constituent/s in the Tioleach solution accelerate the leaching rate of the ore concentrates. For Example 9, Figure 4 shows leaching rates of copper and zinc from the copper ore concentrate by the Tioleach solution at a solution temperature of 90°C.

All of Figures 2 to 4 show that for any given set of conditions, the leach rate for zinc is higher than that for copper.

EXAMPLE 10 and COMPARATIVE EXAMPLES 3 and 4

The Tioleach solution from the manufacture of T1O2 from titanium slag using the Sulphate Process was used as a leachant for a copper and iron ore concentrate derived principally from chalcopyrite and pyrite ore. As for Examples 8 and 9, the acidic co-product solution contained 270 g/l H2SO4 and 16 g/l ferrous iron. Also similarly to Examples 8 and 9, leach rate of the Tioleach solution of Example 10 was compared with that of a H2SO4 solution and a H2SO4 + Fe²⁺ solution made from sulphuric acid with ferrous sulphate. The apparatus shown in Figure 1 was again used for Example 10 and Comparative Examples 3 and 4.

The copper and iron ore concentrate was mixed with an acidic solution in the ratio of 30 grams concentrate and 300 ml solution, at a solution temperature of 70°C. Throughout the leaching process, the solution was stirred and air was bubbled through the solution at a rate of 1 litre/minute.

Figure 5 shows the leach rate of copper, for the three acidic solutions. The diamonds represent results for Example 10 in which 30 grams of concentrate was leached with 300 ml of Tioleach solution; the triangles represent results for Comparative Example 3 in which 30 grams of ore concentrate was leached with 300 ml of a solution comprising 270 g/l and 16 g/l Fe²⁺; and the squares represent results for Comparative Example 4 in which 30 grams of ore concentrate was reacted with 300 ml of a solution having 270 g/l H2SO4.

It is evident from Figure 5 that the leach rate for copper is again highest using the acidic co-product solution (Example 10) and is considerably higher than that for either of Comparative Examples 3 or 4.

Figure 6 is a plot of concentration of ions in solution (ppm) vs time (minutes) for Example 10. The diamonds represent concentration of Cu²⁺ in solution and the squares represent the difference between the actual concentration of Fe²⁺ in solution and the original concentration of Fe²⁺ in solution, prior to commencement of the leaching process. The increase in Fe²⁺ ions in solution is probably due to leaching of iron containing compounds in the concentrate.

It is noted again that copper sulphides can be dissolved without oxidation of the sulphur present in the sulphides to sulphate.

The following Examples 11 to 30 and Comparative Examples 5 and 6 describe the results of leaching ore tailings having a particle size of less than 38 microns and an average composition of 6 to 9% Zn, 7-8% Pb, 0.5% Cu, 250-300 ppm Ag and 2 to 4 ppm Au. The tailings are produced as a 60% pulp density slurry having the following mineralogy: principally pyrite (FeS₂), with some sphalerite (ZnS) and Galena (PbS) and minor amounts of Tetrahedrite (4Cu₂S. Sb₂S₃), Chalcopyrite (CuFeS₂), Arsenopyrite (FeAsS₂), Barytes (BaS0₄), Pyrrhotite (Fe^S), Argentite (Ag₂S) and gold. These minerals are very fin grained and intermixed, making metal recovery by conventional means ver difficult.

Results of the leaching tests on the ore tailings are presented in terms of % Zn leached vs time. However, it should be noted that leaching rates of othe metals, such as Cu and Pb, follow approximately the same path as for Zn.

All leaching tests were conducted at atmospheric pressure.

EXAMPLE 11 AND COMPARATIVE EXAMPLES 5 AND 6 320g of tailings were treated with a leaching solution comprising : 1.6 I o

"Tioleach", 1 g Cu ions; 2ml acetic acid; 320g sand; and 0.5g calcium lignosulphate. The solution was agitated and heated to a temperature of 90°C. Oxygen gas was bubbled through the solution at 60 l/hour.

Figure 7 shows % Zn leached over time of Example 11 (squares compared with zinc recovery from Comparative Example 5 (triangles), in whic 250 g/l of H₂S0₄ was used instead of Tioleach, and from Comparative Example (diamonds), in which 250 g/l H₂S0₄ and 16 g/l Fe²⁺ replaced Tioleach. Th results indicate that, at least for the conditions of Figure 7, the leaching rate fo Tioleach solution is between approximately 150 and 250% higher than thos rates for H₂S0₄ or H₂S0₄ / iron solutions. Comparison of Comparative Example 5 and 6 shows that the leaching rate of the H₂S0₄ / iron solution is generall higher than that of H₂S0₄ per se, with the difference between respective leaching rates increasing over time.

FXAMPLES 12 o 15

The effect on leaching rate of grinding the tailings prior to leaching i illustrated in Figure 8. Each of Examples 12 to 15 involved reacting 320 grams o dry tailings equivalent with 1600 ml of Tioleach containing 1 gram Cu ions an 2ml acetic acid with agitation at a temperature of 90°C, and a feed rate of oxyge of 50 l/hr. In Example 12 (open squares) the mine tailings were unground an had an average particle size of less than 38 microns. The tailings of Example 1 (diamonds) were ground. As illustrated in Figure 8, grinding results in a marke increase in the leaching rate of Zn, particularly during the first 30 minutes of the leaching process when there is more than 100% increase in leaching rate.

Examples 14 (triangles) and 15 (closed squares), additionally included an attriting agent, sand, in the leach mixture. The mine tailings of Example 15 were ground, whereas in Example 14, the mine tailings remained unground. Comparison of Examples 14 and 15 again illustrate the increase in leaching rate of Zn that grinding effects. Comparison of Example 12 with Example 14 and Example 13 with Example 15, respectively, indicate that addition of sand to the reaction mixture improves leaching rate and that the difference in leaching rate between reaction mixtures including sand and those not including sand appears to increase with time.

EXAMPLES 16 to 21

Figure 9 illustrates leaching rates for three different tailings compositions. In each case, 320 grams of unground dry tailings equivalent were reacted with 1600ml Tioleach at 90°C with agitation and at a feed rate of oxygen of 50 litre/hour. In Example 16, the tailings contained 7.4% Zn, Example 17 contained 8.2% Zn and Example 18 contained 6.9% Zn. It is evident that each different tailings composition has a different leach rate. The leaching rate of Example 16 was highest, especially in the first 15 to 20 minutes. The leaching rates of Examples 17 and 18 were lower than that of Example 16, with the rate of Example 17 increasing relative to Example 18 after about 15 minutes and exceeding the rate of Example 18 after about 40 minutes. For each Example, the initial leaching rate was rapid until approximately 50% Zn dissolution and thereafter proceeded at a slower, constant rate. It is thought that the initial leach rate may be determined by the particle size of the relevant ore minerals.

The leaching conditions of Examples 19 to 21 were similar to those for Examples 16 to 18, respectively, with the exception that sand was added to the reaction mixture, at a ratio of 1:1 sand to tailings. Figure 10 shows the results of those Examples. Comparison of Figure 10 with Figure 9 shows that at short leaching times, such as up to 15 to 20 minutes, the presence of sand makes little difference to leaching rate. However at longer leaching times, Examples 19 to 21 each show increased leaching rates as compared with their respective counterparts in Examples 16 to 18. After 90 minutes of reaction time, the leach rate of the tailings composition 7.4% Zn had increased by more than 50%, the rate of tailings composition 8.2% Zn had increased by approximately 45% and that of tailings composition 6.9% Zn had increased by approximately 60%. It is noted that each different tailings composition still has a different leach rate.

EXAMPLES 22 to 24

The effect of temperature on leaching rate is depicted in Figure 11. In each of Examples 22, 23 and 24, 320 grams of unground dry tailings equivalent were reacted with 1600 ml Tioleach solution containing sand in a ratio of sand: tailings of 1 :1 and having a feed rate of oxygen gas of 50 litre/hour. The temperature of leaching of Examples 22, 23 and 24 were 70°C, 80°C and 90°C, respectively. Figure 1 1 clearly shows that, at least for the particular conditions of Examples 22 to 24, increasing temperature resulted in an increase in leaching rate. For example at a reaction time of 75 minutes, an increase of temperature from 70°C to 90°C leads to an increase of more than 125% of leaching rate.

EXAMPLES 25 and 26 The effect on leaching rate of the addition of Cu ions and acetate ions to the leaching solution is investigated in Examples 25 and 26. In both Examples, 320gm of dry mine tailings equivalent were reacted with 1600ml of Tioleach solution at a temperature of 90°C and a feed rate of oxygen gas of 50 litre/hour. Sand was added to the reaction mixture at a ratio of sand: tailings of approximately 1 :1 and the mixture was agitated. Figure 12 shows that the leaching rate of Example 26, having 4g CuSO₄.5H₂O and 2ml acetic acid added to the reaction mixture, is higher than that of Example 25, having no additions.After a reaction time of 90 minutes, the leaching rate of Example 26 was almost 40% higher than that of Example 25. EXAMPLES 27 to 30

In Examples 27 to 30, 320 grams of unground mine tailings were reacted with 1600ml of Tioleach solution with agitation at a temperature of 90°C and a feed rate of oxygen gas of 50 litre/hour. The results are presented in Figure 13. Example 28 illustrates the effect on leaching rate of adding 0.025% w/v calcium lignosulphonate to the reaction mixture, compared with Example 27 having no calcium lignosulphonate. It is evident that addition of calcium lignosulphonate increases leaching rate. For example, at approximately 80 minutes reaction time, there is almost a 20% increase in leaching rate between Examples 27 and 28.

Comparison of Examples 29 and 30 also indicates the improved leaching rate that can result from addition of calcium lignosulphonate. Those Examples further illustrate the improvement in leaching rate that addition of sand brings (Example 29) as well as the further improvement in leaching rate that results from the combined addition of sand and calcium lignosulphonate (Example 30). Comparison of Examples 27 and 30 indicates that at, for example, 80 minutes of reaction time, the leaching solution including lignosulphonate and sand has a leaching rate approximately 80% higher than a solution not having those additives.

EXAMPLE 31 and COMPARATIVE EXAMPLES 7 to 9

The following Example 31 and Comparative Examples 7 to 9 relate to leaching of the ore mineral chalcopyrite (CuFeS₂) and are presented in Figure 14. Example 31 (triangles) shows leaching rate of Cu from 320 gram chalcopyrite using 1600 ml Tioleach solution with agitation, including sand in a ratio of 1 :1 sand to chalcopyrite, 0.4g lignosulphonate, a feed rate of oxygen of 50 litre/hour, and a temperature of 90°C. Comparative Example 7 (squares) shows the results of leaching chalcopyrite using a leaching solution comprising a mixture of 0.25 M Fe₂(S0₄)₃ and 0.5 M H₂S0₄. The concentration of chalcopyrite in the leaching solution is 0.2 g/l.

Comparative Example 8 shows the best reported results of leaching chalcopyrite using a typical pressure leaching process. Comparative Example 9 shows the leaching rate of chalcopyrite predicted theoretically by thermodynamics.

The leaching rates achievable by using a Tioleach solution are significantly higher than those achieved by a H₂S0₄/Fe solution or by pressure leaching. Further, after 60 minutes of leaching, the slope of the leaching curve for Example 31 more closely approximates the slope of Comparative Example 9 than either of Comparative Examples 7 or 8. For example, at a reaction time of approximately 180 minutes, the amount of copper leached is approximately 50% of the theoretically achievable amount, compared with approximately 10%, for Comparative Example 7, and approximately 27%, for Comparative Example 8.

EXAMPLES 32 to 35 and COMPARATIVE EXAMPLE 10

Figure 15 shows the leaching rate over time of chalcopyrite treated with various solutions. In each case, 320g of chalcopyrite was reacted with 1600ml of solution containing 320g sand, 1g Cu ions, 2ml acetate ions (as acetic acid) and 0.4g calcium lignosulphonate. The solution temperature was 90°C and the solution was agitated while oxygen was fed into solution at 50 litre/hour.

Example 32 (open diamonds) represents leaching of chalcopyrite using a Tioleach solution. Comparative Example 10 (closed squares) represents the results of reacting chalcopyrite with an acidic solution including 250 g/l H₂S0₄ and 16 g/l Fe as FeS0₄. As previously discussed in relation to earlier Examples, the Tioleach solution is considerably more effective in leaching metals than is a solution containing H₂S0₄ and Fe ions only.

Example 33 (triangles) represents leaching of chalcopyrite using the H₂S0₄/Fe solution of Comparative Example 10 to which has been added 10ml concentrated HCI. The results show that the addition of chloride ions leads to an increase of leaching rate, at least in the early stages of the reaction. After approximately 180 minutes, the leaching rate begins to decrease.

Example 34 (open squares) represents leaching of chalcopyrite with the H₂S0₄/Fe solution to which has been added Ti0₂, as unfiltered titanyl sulphate (20ml). The leaching rate of this solution was only slightly higher than the H₂S0₄/Fe solution of Comparative Example 10 up to approximately 180 minutes of reaction time. However, after 180 minutes, the leaching rate of Example 34 rapidly increased relative to Comparative Example 10.

Example 35 (closed diamonds) relates to leaching of chalcopyrite using the H₂S0₄/Fe solution of Comparative Example 10 to which was added 0.26g Ti ions and 0.78g Cl ions at 90 minutes reaction time. It is evident from Figure 15 that prior to addition of Ti ions and Cl ions, the leaching rate of Example 35 was, as would be expected, approximately the same as for Comparative Example 10. After addition of Ti and Cl ions at 90 minutes, the leaching rate starts to increase and is equal to the leaching rate for Example 33 (HCI addition) at approximately 120 minutes. Subsequently, leaching rate further increased and at 240 minutes was more than twice the leaching rate of Comparative Example 10. It is to be noted in Figure 15 that from approximately 120 minutes onwards, the slope of the curve for Example 35 is approximating the slope of the curve for Example 32.

The results for Examples 32 to 35 and Comparative Example 10 indicate that, at least under the particular conditions of those reactions, the level of chloride species in solution increases leaching rate, particularly at short reaction times. Titanium containing species also affect leaching rate, with an increase in leaching rate becoming more pronounced at longer reaction times. The addition of both Ti and Cl to the leaching solution, leads to an even greater increase in leaching rate.

EXAMPLES 36 to 38

Figure 16 shows leaching rate over time of tailings treated with Tioleach solution at varying ratios of solids: solution. In each case, tailings were reacted with Tioleach solution containing calcium lignosulphonate whilst agitated at an oxygen feed rate of 2 l/min and a temperature of 70°C. Example 36 (circles) represents a concentration of tailings in solution of 100 g/l, Example 37 (squares) represents a concentration of 200 g/l and Example 38 (triangles) represents 300 g/l. By increasing the ratio of solids: solution, at least under the conditions of Examples 36 to 38, there is an increase in leaching rate, particulariy at relatively low reaction times. However, comparison of Examples 37 and 38 indicate that for those concentrations of solids, the difference in respective leaching rates decreases at longer reaction times. It is believed that the increase in solids: solution ratio results in greater attrition of the sulphur containing material coating the surfaces of solid particles, thereby increasing the exposure of unreacted solid particles to the leaching solution.

EXAMPLES 39 and 40

Figure 17 displays the results of leaching tailings with Tioleach solutions from two different sources. In each case, 320g of tailings were reacted with 1.6 litre of Tioleach solution containing 320g sand, 3.5g CuS0₄.5H₂0, 0.5g calcium lignosulphonate and 2ml acetic acid. The solution temperature was 90°C and the solution was agitated while 0₂ was fed into solution at 60 l/hour.

The Tioleach solution used in Example 39 (diamonds) was derived from the Sulphate Process using titanium slag as feedstock. In Example 40 (squares), the Tioleach solution was derived from ilmenite feedstock. The ilmenite derived solution had undergone partial removal of iron from solution. Figure 17 shows that at least under the conditions of Examples 39 and 40, the Tioleach solution derived from ilmenite generally gives only slightly higher zinc recovery than the Tioleach solution derived from titanium slag, with the difference in zinc recovery between the respective Examples increasing over time.

EXAMPLES 41 and 42

Figure 18 displays the results of leaching chalcopyrite with Tioleach solution derived from feedstocks comprising titanium slag (Example 41 , diamonds) and ilmenite (Example 42, squares) respectively. The ilmenite derived solution had undergone partial removal of iron from solution. In each case, 100g chalcopyrite was reacted with 1.6 litre Tioleach solution containing 320g sand, 4g CuS0₄.5H₂0, 0.4g calcium lignosulponate and 2ml acetic acid. The solution temperature was 95°C and the solution was agitated while 0₂ was fed into the solution at 70 l/hour. Figure 18 shows that, at least under the conditions of Examples 41 and 42, the Tioleach solution derived from ilmenite feedstock in the Sulphate Process generally gives slightly higher copper recovery than the Tioleach solution derived from titanium slag.

EXAMPLE 43 and COMPARATIVE FXAMPI F 1 Figure 19 shows the results of leaching a dross containing 58% lead as sulphide and/or oxide and 23% copper as cuprous sulphide (matte). Diamonds represent the results of Comparative Example 11 , in which 400g of the dross was leached with 1600ml of 250 g/l H₂S0₄ at temperature of 90°C with agitation of the solution and with an oxygen feed rate of 1.5 l/min. Squares represent the results of Example 43, in which 400g of the dross was leached with 1600ml Tioleach at 85°C with agitation and an oxygen feed rate of 1.5 l/minute.

Figure 19 illustrates the considerably higher leaching rate of copper from the dross using the Tioleach solution compared with the results using the H₂S0₄ solution. In fact, for a reaction time of 120 mins, essentially all the copper in the dross had been leached by the Tioleach solution, whereas only approximately

40% of the dross had been leached by the H₂S0₄ solution.

EXAMPLES 44 to 47 In Examples 44 and 45, a Tioleach solution from the manufacture of Ti0₂ from titanium slag using the Sulphate Process was used to leach metals from samples of mineral processing tailings, and copper concentrates respectively.

The results were compared to those using an acidic leach solution having concentrations of H₂S0₄ and Fe ions and Ti and Cl ions which were equivalent to those of the Tioleach solution (Examples 46 and 47).

In each of Examples 44 to 47, 320g of tailings were treated with 1 ,600 mis of leach solution and agitated in an open container at 90°C. Oxygen was introduced using treatment at a rate of 50 litre/hour.

In Figure 20, triangles represent Example 44 and Diamonds represent Example 46. The graph shows that a similar leach rate is achieved for each of the two solutions for recovery of Zinc from a mineral tailings sample having the composition as described for Examples 11 to 30.

SϋBSTITl-TE SHEET (RULE 26) Similarly, in Figure 21, Example 45 (diamonds) and Example 47 (squares) exhibit similar leaching rates for recovery of copper from a copper concentrate sample having the same composition as described for Examples 31 to 35.

Finally, it is to be understood that various alterations, modifications, and/or additions may be introduced into the constructions and arrangements of parts and/or steps previously described without departing from the spirit or ambit of the invention.

Claims

1. A process for extracting one or more metals from a metal containing material, as herein defined, including the step of treating said metal containing material with an aqueous, acidic solution containing effective concentrations of iron-containing species, sulphuric acid and one or more leach rate accelerants present in the co-product solution from the manufacture of Ti0₂ by the Sulphate Process, whether said leach rate accelerants are present as a result of using a co-product solution or are separately compounded.

2. The process of claim 1 , further including the step of oxidising a sufficient amount of iron in said iron-containing species to ferric iron prior to and/or during treatment of said metal containing material with said aqueous, acidic solution.

3. The process of claim 1 or claim 2, wherein said one or more leach rate accelerants include a titanium-containing compound which is particulate and/or dissolved in solution.

4. The process of any one of claims 1 to 3, wherein said one or more leach rate accelerants include a chloride containing species.

5. The process of claim 3 or claim 4, wherein the titanium-containing compound is selected from or derived from the group comprising oxides, hydroxides, hydrous oxides, sulphates, hydroxysulphates, chlorides or oxychlorides of titanium.

6. The process of claim 5, wherein the titanium-containing compound is selected from the group comprising Ti0₂, TiO(OH)₂, Ti0₂.2H₂0, Ti(OH)₃, TiCI₄, TiOCI₂and Ti(OH)_xCI₍₄__x).

7. The process of claim 2, or any one of claims 3 to 6 when appended to claim 2, wherein said step of oxidising includes introducing an oxidising gas into said solution.

8. The process of claim 7, wherein said oxidising gas is air or oxygen.

9. The process of claim 8 wherein said oxidising gas is oxygen.

10. The process of claim 9, wherein said oxygen is introduced into said solution at a rate of from 0.001 to 5.0 grams of oxygen per litre of solution per minute.

11. The process of any one of claims 1 to 10, wherein the pH of said aqueous acidic solution is 6.5 or less.

12. The process of claim 11 , wherein the pH is from 0 to 5.

13. The process of any one of claims 1 to 12, wherein the concentration of iron in said solution is less than 80 g/i.

14. The process of any one of claims 1 to 12, wherein the concentration of iron is less than 20 g/l.

15. The process of any one of claims 1 to 14, wherein the concentration of H₂S0₄ is less than 350 g/l.

16. The process of any one of claims 1 to 14, wherein the concentration of H₂S0₄ is less than 150 g/l.

17. The process of any one of claims 1 to 16, wherein said metal containing material is an ore mineral or a mineral concentrate.

18. The process of any one of claims 1 to 17, wherein said metal containing material comprises mine tailings or a by-product or waste product of mineral processing including metal refining.

19. The process of claim 18, wherein said mine tailings are ground prior to treatment with said aqueous acidic solution.

20. The process of any one of claims 1 to 19, wherein said one or more metals are selected from the Transition metals and Noble metals.

21. The process of any one of claims 1 to 20, wherein said one or more metals are selected from Zn, Pb, Cu, Ag, Au, Ni and Mn.

22. The process of any one of claims 1 to 21, wherein the temperature of said aqueous acidic solution is between ambient and 90°C.

23. The process of any one of claims 1 to 21 , wherein the temperature of said aqueous acidic solution is between 90°C and 300°C.

24 The process of any one of claims 1 to 22, wherein the temperature of said aqueous acidic solution is between 70°C and 90°C.

25. The process of any one of claims 1 to 24, wherein said solution is agitated during said treatment.

26. The process of any one of claims 1 to 25, wherein said aqueous acidic solution further includes an effective amount of a foam control agent.

27. The process of claim 26, wherein said foam control agent is calcium lignosulphonate.

28. The process of any one of claims 1 to 27, wherein said process furthe includes the step of attriting the metal containing material.

29. The process of claim 28, wherein said step of attriting comprises adding a effective amount of sand to said solution.

30. The process of claim 28, wherein.said step of attriting comprise increasing the ratio of said metal containing material to said aqueous acidi solution.

31. The process of any one of claims 1 to 30, wherein said aqueous acidi solution further includes one or more leaching promoters.

32. The process of claim 31 , wherein said one or more leaching promoters ar selected from copper ions and acetate ions.

33. The process of any one of claims 1 to 32, further including the step o partially removing iron from said aqueous acidic solution.

34. A process for extracting one or more metals from a metal containin material, as herein defined, including the step of treating said metal containin material with an aqueous acidic solution comprising the waste or co-produc solution from the manufacture of Ti0₂ by the Sulphate Process.

35. A process for extracting one or more metals from a metal containin material, substantially as herein defined with reference to any one of Examples to 47.