Background of the Invention In the processing of various mineral ores, acid leaching is frequently employed as a means of dissolving valuable species in the ore, so that the valuable species can be subsequently recovered from an aqueous phase, which is separated from the solid phase.

In such acid leaching, an oxidant is added to increase the rate at which oxidation takes place and/or to increase the amount of oxidisable species that is oxidised.

In the case of uranium ore, it is desirable to oxidise as much as possible of any uranium (IV) that is present in the ore to uranium (VI), because the uranium (VI) is more soluble in an aqueous medium than uranium (IV).

Oxidants such as sodium chlorate and pyrolusite that are used for this purpose are expensive and contribute significantly to the processing cost of such minerals. There is accordingly a need for reducing the consumption of such oxidants in mineral processing.

When ores are milled using mills such as rod mills and ball mills of which the rods or balls are made of steel, small amounts of elemental iron arising from wear of the iron components of such mills, are present in the ore. Such elemental iron has been shown by Wyllie, World Mining, February 1979, to have a strong reducing effect, and react with the oxidant in the acid leaching step.

In East German patent specification No. DD218780, a process is described in which reducing components in a uranium ore, including ferrous ions, are oxidised by intensively aerating a uranium ore slurry at a pH below 6, preferably at a pH between 4 and 5. In this process, any metallic iron present is probably converted to Fe (II) or Fe (III) because of the acidic conditions in which oxidation occurs. It is well known that Fe (II) and Fe (III) are insoluble at a pH above 6, and are still virtually insoluble at a pH of 5. The aeration of a uranium ore at a pH below about 5 accordingly results in the conversion of elemental iron to soluble species of iron. Although this process results in a reduction in oxidant

consumption, it has the disadvantages that it requires excessive aeration and that it consumes significant amounts of acid which is also costly.

Object of the Invention It is the object of the present invention to overcome or substantially ameliorate at least one of the above disadvantages.

Summary of the Invention According to a first aspect of the invention, there is provided a process for treating a slurry of a mineral ore, including the step of contacting the slurry with oxygen or an oxygen containing gas, at a pH of 6 or greater, whereby an oxidisable magnetic species in the slurry is at least partially oxidised to a substantially insoluble species.

According to a second aspect of the invention, there is provided a process for treating a slurry of a mineral ore, including the step of contacting the slurry with oxygen or an oxygen containing gas, at a pH of 6 or greater, whereby an oxidisable magnetic species in the slurry is at least partially oxidised to a species that consumes a lesser amount of oxidant added in a subsequent leaching step, than the oxidisable magnetic species.

The slurry is conveniently an aqueous slurry.

The oxygen containing gas may conveniently be air or air enriched with oxygen.

The mineral ore may be a uranium ore or any other type of mineral ore that needs to be oxidised in a subsequent leaching step.

The oxidisable magnetic species may be iron particles wherein iron is partially or substantially in its elemental form. The iron particles may originate from a milling process to which the mineral ore had been subjected prior to the treatment process according to the invention. Such milling process may include the passing of the ore through an ore mill, such as a ball mill or a rod mill, in which the ore is milled down to the required particle size distribution and in which the ore comes into contact with a steel surface such as a liner of the mill, or in which steel balls or rods are employed for the comminution of the ore, and such steel balls are worn away by the ore.

The substantially insoluble species is conveniently a species that does not consume oxidant or consumes only a small amount of oxidant or, alternatively, consumes a lesser amount of oxidant than the oxidisable species, in subsequent acid leaching of said mineral ore in the presence of said oxidant.

The process according to the invention may conveniently include a step, prior to the step in which the slurry is contacted with oxygen or a gas containing oxygen, in which at least a portion of the slurry is subjected to magnetic separation. In the magnetic separation step, a portion of the oxidisable magnetic species in such portion of the slurry, may be removed from the said portion of the slurry.

The invention includes in its scope a slurry of a mineral ore treated by the process of the invention.

In the event that a milled uranium ore contains more iron than is required in a subsequent acid leaching process in which the ore is leached under oxidising conditions, some of the iron may be removed from the ore by subjecting a portion of the ore to magnetic separation, whilst another portion of the ore may be subjected to a pre-leaching step in which it is contacted with air in accordance with the process of the invention. The two portions of the ore may be combined after air pre-leaching of the one portion and magnetic separation of iron from the other portion. In this way, just enough iron may be left in the ore for purposes of oxidising the uranium (IV) that is in the ore, to uranium (VI).

Advantageously, the pH may be towards the lower end of the range of 5 to 9.5, 5 to 9, 5. 1to9, 5. 1to8, 5. 1 to 7. 5, 5. 1 to 7, 5. 1to6. 5,5. 5to9, 6to9, 6to8, 6to7. 5, 6to7, 6 to 6.5, 6.01 to 9,6. 01 to 8.5, 6.01 to 7.5, 6.01 to 7, or 6.01 to 6.5 more usually 6.01 to 8.

A preferred pH range is between 5.01 and 7.5.

The temperature at which the process in accordance with the invention may be carried out, may range from about 0°C to about 100°C, usually from about 12°C to about 50°C and more usually from about 15°C to about 45°C with a preferred temperature being ambient temperature. Because of cost considerations, no heating or cooling of any of the process streams, in particular, neither the ore slurry nor the gas, would normally be done.

In one embodiment of the invention, the oxygen containing gas is air. In this embodiment of the invention, the mineral ore slurry may be contacted with the air for a period of from about 10 minutes to about 24 hours, preferably from about 1 hour to about 8 hours, more preferably from about 1 hour to about 5 hours.

According to a third aspect of the invention, there is provided a method of reducing the consumption of oxidant in an acid leaching step in which a valuable species is recovered from a mineral ore slurry, the method comprising the step of contacting the mineral ore slurry with oxygen or an oxygen containing gas, prior to the leaching step, at a pH of 5 or greater, whereby an oxidisable species in the mineral ore slurry is at least partially oxidised to an insoluble species.

According to a fourth aspect of the invention, there is provided a method of reducing the consumption of acid in an acid leaching step in which a valuable species is recovered from a mineral ore slurry, the method comprising the step of contacting the mineral ore slurry with oxygen or an oxygen containing gas, prior to the leaching step, at a pH of 5 or greater, whereby an oxidisable species in the mineral ore slurry is at least partially oxidised to an insoluble species.

According to a fifth aspect of the invention, there is provided a method of improving the recovery of a valuable species from a mineral ore slurry in an acid leaching step in which the valuable species is recovered from the mineral ore slurry, the method comprising the step of contacting the mineral ore slurry with oxygen or an oxygen containing gas, prior to the leaching step, at a pH of 5 or greater, whereby an oxidisable species in the mineral ore slurry is at least partially oxidised to an insoluble species.

In the processes according to the third, fourth and fifth aspects of the invention, the oxidisable species may, as an alternative, be oxidised to a species that consumes a lesser amount of oxidant added in a subsequent leaching step, than the oxidisable magnetic species.

According to a sixth aspect of the invention, there is provided a process for treating a slurry of a mineral ore, including the step of contacting the slurry with a first oxidant, whereby the slurry or a portion thereof, after said contacting step, consumes a lesser amount of a second oxidant added in a subsequent leaching step, than it would have consumed without said contacting step.

The first oxidant may be the same or different and may be selected from oxygen, oxygen enriched air, sulphur dioxide, pyrolusite and any other suitable oxidants.

Combinations of oxidants may be used. The first oxidant may be a cheaper oxidant in terms of its ability to oxidise metals than the second oxidant. The first oxidant preferably has a higher oxidation potential than an invaluable species forming part of the mineral ore slurry. The second oxidant preferably has a higher oxidation potential than a valuable species forming part of the mineral ore slurry, and which is to be recovered therefrom.

In the case of a mineral ore that contains uranium, the first oxidant may be oxygen or an oxygen containing gas, and the second oxidant may be pyrolusite.

The contacting step is preferably carried out under operating conditions, including pH, such that an invaluable species forming part of said mineral ore slurry, is at least partially converted by the first oxidant to a species which does not consume second oxidant or consumes less second oxidant in the subsequent leaching step than it would

have consumed without said contacting step. Where the mineral ore contains an iron species as well as uranium, the process is conveniently operated at a pH of 5 or greater.

The process is conveniently operated at a pH greater than pH 5 and at most pH 9 or 9.5.

According to a sixth aspect of the invention, there is provided a method of reducing the consumption of second oxidant in an acid leaching step forming part of a process for the recovery of a valuable mineral from a mineral ore slurry containing the valuable mineral, wherein, in the acid leaching step, the valuable mineral is solubilised by oxidising it with the second oxidant converting the valuable species from an insoluble to a soluble form, the method comprising the step of contacting the mineral ore slurry with a first oxidant, prior to the acid leaching step, causing a portion of the mineral ore slurry to be at least partially converted to a form which does not consume second oxidant in the acid leaching step or which consumes less second oxidant than the mineral ore slurry would have consumed without said contacting step.

The first oxidant may be oxygen or an oxygen containing gas, the mineral ore slurry is contacted with the oxygen or an oxygen containing gas at a pH of 5 or greater, and the slurry may comprise an iron species. The pH may be greater than pH 5 and at most pH 9 or 9. 5.

According to a seventh aspect of the invention, there is provided a method of reducing the consumption of acid in an acid leaching step forming part of a process for the recovery of a valuable mineral from a mineral ore slurry containing the valuable mineral, wherein, in an acid leaching step forming part of the process, the valuable mineral is solubilised by oxidising it with a second oxidant converting the valuable mineral from an insoluble to a soluble form, wherein the method comprises the step of contacting the mineral ore slurry with a first oxidant, prior to the acid leaching step, causing a portion of the mineral ore slurry or an invaluable species in the mineral ore slurry to be at least partially converted to a form which does not consume acid or which consumes less acid than the mineral ore slurry would have consumed in the acid leaching step if it had not been so contacted with said first oxidant.

The valuable mineral may be uranium, the invaluable species may be an iron species, the first oxidant may be oxygen or an oxygen containing gas, and the mineral ore slurry may be contacted with the oxygen or the oxygen containing gas at a pH of 5 or greater. The pH may be greater than pH 5 or pH 6 and at most pH 9 or pH 9.5.

According to an eighth aspect of the invention, there is provided a method of increasing the recovery of a valuable species in a process for the recovery of the valuable

species from a mineral ore slurry containing it, wherein the process includes a step of solubilising the valuable species in an acid leaching step in which the mineral ore slurry is oxidised to a soluble form by means of a second oxidant, the method including the step of contacting the mineral ore slurry with a first oxidant, prior to the acid leaching step, causing at least a portion of the mineral ore slurry or an invaluable species therein to be at least partially converted to a form which does not consume at least one of acid and oxidant in the acid leaching step, resulting in the increased recovery of the valuable species.

The valuable species may be uranium, the invaluable species may be an iron species, the first oxidant may be oxygen or an oxygen containing gas, and the mineral ore slurry may be contacted with the first oxidant such as oxygen or an oxygen containing gas at a pH of 5 or greater, optionally at a pH greater than pH 5. The pH may be 5.1, 5.2, 5.3, 5.4, 5.5, 5.6, 5.7, 5.8, 5.9, 6.0 or 6.1 and at most pH 9.5, 9.4, 9.3, 9.2, 9.1 or 9.0, for example.

According to a ninth aspect of the invention, there is provided a process for treating a slurry of a mineral ore comprising an oxidisable invaluable species and a valuable species which is recoverable by oxidising said valuable species in an acid leach, including the steps of (a) contacting the slurry with a first oxidant at a first pH; and (b) thereafter contacting the slurry, in the presence of a strong acid, with a second oxidant at a second pH which is lower than the first pH, wherein the first pH is such that the amount of an oxidised species derived from said invaluable species going into solution is such that the slurry or a portion thereof consumes a lesser amount of the second oxidant in step (b), than it would have consumed without said contacting step (a).

The first oxidant may be oxygen or an oxygen containing gas, the valuable species may be uranium and the invaluable species may be an iron species. The first pH may be greater than pH 5 and the second pH may be about 0.1 to about 4, preferably about 1 to 3, more preferably about 2. The second pH may be 0.5, 0.75, 1.0, 1.25, 1.5, 1.75, 2.0, 2.25, 2.5, 2.75, 3.0, 3.25, 3.5 or 4, for example.

According to a tenth aspect of the invention, there is provided a process for treating a slurry of a mineral ore, including the step of contacting the slurry with oxygen or an oxygen containing gas, whereby the slurry is at least partially converted to a slurry that

consumes a lesser amount of oxidant added in a subsequent leaching step, than the slurry would have consumed without said contacting step.

The insoluble species may be an iron species.

According to an eleventh aspect of the invention, there is provided a method of reducing the consumption of oxidant in an acid leaching step in which an insoluble valuable species is recovered from a mineral ore slurry by oxidising it with an oxidant converting the insoluble valuable species to a soluble form, wherein the method comprises the step of contacting the mineral ore slurry with oxygen or an oxygen containing gas, prior to the acid leaching step, causing a portion of the mineral ore slurry or an insoluble first species in the mineral ore slurry to be at least partially converted to a form which does not consume oxidant in the acid leaching step.

According to a twelfth aspect of the invention, there is provided a method of reducing the consumption of acid in an acid leaching step forming part of a process for the recovery of a valuable mineral from a mineral ore slurry containing the valuable mineral, wherein, in an acid leaching step forming part of the process, the valuable mineral is solubilised by oxidising it with an oxidant converting the valuable mineral from an insoluble to a soluble form, the method comprising the step of contacting the mineral ore slurry with oxygen or an oxygen containing gas, prior to the acid leaching step, causing at least a portion of the mineral ore slurry or an insoluble first species in the mineral ore slurry to be at least partially converted to a form which does not consume acid or which consumes less acid than the mineral ore slurry would have consumed in the acid leaching step if it had not been so contacted with oxygen or an oxygen containing gas.

According to a thirteenth aspect of the invention, there is provided a method of increasing the recovery of an insoluble valuable species in a process for the recovery of the insoluble valuable species from a mineral ore slurry containing it, wherein the process includes a step of solubilising the insoluble valuable species in an acid leaching step in which the mineral ore slurry is oxidised to a soluble form by means of an oxidant, the method including the step of contacting the mineral ore slurry with oxygen or an oxygen containing gas, prior to the acid leaching step, causing at least a portion of the mineral ore slurry or an insoluble first species therein to be at least partially converted to a form which does not consume at least one of acid and oxidant in the acid leaching step, resulting in the increased recovery of the insoluble valuable species.

The insoluble valuable species may be uranium or a compound thereof.

According to a fourteenth aspect of the invention, there is provided a process for treating a slurry of a mineral ore, including the step of contacting the slurry with oxygen or an oxygen containing gas, whereby the slurry or a portion thereof, after said contacting step, consumes a lesser amount of oxidant added in a subsequent leaching step, than it would have consumed without said contacting step.

The contacting step is preferably carried out under operating conditions, including pH and temperature, such that an insoluble species forming part of said mineral ore slurry, other than a valuable species to be recovered therefrom, is at least partially converted by the oxygen to a species which does not consume oxidant or consumes less oxidant in the subsequent leaching step than it would have consumed without said contacting step.

Where the insoluble species is an iron species, the process is preferably operated at a pH of 5 or greater such as up to pH 9.5 or 9.0.

According to a fifteenth aspect of the invention, there is provided a process for treating a slurry of a mineral ore comprising an oxidisable invaluable species and a valuable species which is recoverable by oxidising said valuable species in an acid leach, including the steps of (a) contacting the slurry with a first oxidant at a first pH; and (b) thereafter contacting the slurry, in the presence of a strong acid, with a second oxidant at a second pH which is lower than the first pH, wherein the first pH is such that the amount of an oxidised species derived from said invaluable species going into solution is such that the slurry or a portion thereof consumes a lesser amount of the second oxidant in step (b), than it would have consumed without said contacting step (a).

The first oxidant may be oxygen or an oxygen containing gas.

The valuable species may be uranium and the invaluable species may be an iron species and the first pH is preferably at least about pH 5. The amount of the invaluable species which is converted in step (a) to a species which is oxidisable by said second oxidant in step (b) may be such that the slurry or a portion thereof consumes a lesser amount of the second oxidant in step (b).

It is to be understood that the scope of the invention is not limited to the separation and beneficiation of only those minerals that have been mention above. The processes and methods of the invention may be applied in the separation and beneficiation of various other minerals such as those containing a valuable metal and in invaluable sulphide or

other compound which consumes oxidant in a leaching step, and which can be converted using a cheap source of a first oxidant to reduce the consumption of an second, more expensive oxidant.

Brief Description of the Drawings A preferred form of the present invention will now be described by way of example with reference to the accompanying drawings wherein: Figure 1 is a diagrammatic block flow diagram of one embodiment of a process in accordance with the invention; Figure 2 is a diagrammatic block flow diagram of another embodiment of a process in accordance with the invention; and Figure 3 is a diagrammatic block flow diagram of a further embodiment of a process in accordance with the invention.

Detailed Description of the Preferred Embodiments Referring to Figure 1, there is shown a diagrammatic block flow diagram of one embodiment of a process 10 in accordance with the invention. In the process 10, uranium ore slurry 12 is introduced into an aeration tank 14 in which the uranium ore slurry 12 is contacted with air 16 at atmospheric pressure, a pH of between 6 and 8 and ambient temperature of about 45°C.

The slurry 12 is contacted with the air 16 for a period of from about 2 to about 4 hours. The air 16 may optionally be enriched with oxygen.

After aeration in the tank 14, the slurry is transferred to a standard uranium leach tank 18, to which sulphuric acid 20 is added in order to reduce the pH to less than 2.5.

Oxidant in the form of pyrolusite 22 is also added to the leach tank 18, for conversion of Fe (II) to Fe (III) which in turn oxidises any uranium (IV) present in the ore, to uranium (VI).

By virtue of the pre-leaching step in which the ore slurry 12 is aerated, the amount of pyrolusite 22 used in the standard uranium leaching step is reduced, as is evident from Example 2. Apart from a reduction in the consumption of oxidant, the amount of the uranium recovered may also be increased. In some cases, a reduction in the consumption of acid was also observed.

Referring to Figure 2, there is shown a diagrammatic block flow diagram of another embodiment of a process 110 in accordance with the invention. In the process 110,

uranium ore slurry 112 is introduced into a magnetic separator 113 in which a magnetic fraction 112a is first separated from the ore, before it is fed to an aeration tank 114 in which the uranium ore slurry 112 is contacted with air 116 at atmospheric pressure, a pH of between 6 and 8 and ambient temperature of about 45°C. The magnetic fraction 112a is discarded.

The slurry 112 is contacted with the air 116 for a period of from about 2 to about 4 hours.

After aeration in the tank 114, the slurry is transferred to a standard uranium leach tank 118, to which sulphuric acid 120 is added in order to reduce the pH to less than 2.5.

Oxidant in the form of pyrolusite 122 is also added to the leach tank 118, for conversion of Fe (II) to Fe (III) which in turn oxidises any uranium (IV) present in the ore, to uranium (VI).

By virtue of the pre-leaching step in which the ore slurry 112 is aerated, the amount of pyrolusite 122 used in the 1 leaching tank 118 is reduced as in the case of the process shown in Figure 1. Apart from a reduction in the consumption of oxidant, the amount of the uranium recovered may also be increased. In some cases, a reduction in the consumption of acid was also observed.

Referring to Figure 3, there is shown a diagrammatic block flow diagram of a further embodiment of a process 210 in accordance with the invention. In the process 210, uranium ore slurry 212 is introduced into a magnetic separator 213 in which a magnetic fraction 212a is first separated from the ore slurry 212, before it is fed to an aeration tank 214 in which the uranium ore slurry 212 is contacted with air 216 at atmospheric pressure, a pH of between 6 and 8 and ambient temperature of about 45°C. The slurry 212 is contacted with the air 216 for a period of from about 2 to about 4 hours.

After aeration in the tank 214, the slurry is transferred to a standard uranium leach tank 218, to which sulphuric acid 220 is added in order to reduce the pH to less than 2.5.

Oxidant in the form of pyrolusite 222 is also added to the leach tank 218, for conversion of any uranium (IV) present in the ore, to uranium (VI).

A portion of the magnetic fraction 212a is discarded, whilst the remaining portion is intensively oxidised in a further tank 224 by the addition of sulphuric acid and by intensive aeration, preferably under pressure, for a time sufficient to dissolve all elemental iron particles. before the solution from that step, which contains ferric iron sulphate, is added to the uranium leach tank 218. Such ferric iron sulphate serves the purpose of oxidising the uranium (IV) present in the ore, to uranium (VI), whilst being

reduced to ferrous iron sulphate. In order to speed up the oxidation of elemental iron to first ferrous iron and thereafter to ferric iron, an oxidant may be added to the further tank 224. The oxidant may be selected from oxygen, oxygen enriched air, sulphur dioxide, pyrolusite and other suitable oxidants. Combinations of oxidants may be used.

By virtue of the pre-leaching step in which the ore slurry 212 is aerated, the amount of pyrolusite 222 used in the leaching tank 218 is reduced as in the case of the processes shown in Figures 1 and 2. Apart from a reduction in the consumption of oxidant, the amount of the uranium recovered may also be increased. In some cases, a reduction in the consumption of acid was also observed.

EXAMPLES Example 1: Effect on Oxidant Consumption of Drying of Ore, Before Leaching Laboratory tests were conducted in which uranium ore samples were dried at 30- 40°C. In subsequent leaching of the said ore samples, the consumption of oxidant in the form of pyrolusite was found to be approximately 1 kg per tonne. This compares favourably with the consumption of oxidant which is normally more than 3 kg per tonne.

The results of these tests are reflected in Table 1.

Table 1 : The Effect of Air Drying on Oxidant Consumption Leach Treatment Oxidant Reagent Consumption ORP U Residue No. Conditions (kgt-1) (mV) Extraction (ppm % U3O8 Oxidant Acid 433E Wet milled and Caro's 1.46 33.8 460/440 94.1 193 leached acid (as H202) 435G Caro's 1.26 30.2 460/436 94.2 191 acid (as H202) 4831 Pyrolusite 2.39 39.8 460/441 93. 7 206 478C Wet milled, Pyrolusite 1. 05 35.4 460/438 93.7 205 dried and leached 482C Pyrolusite 0.96 35.6 460/438 94. 1 193 498B Pyrolusite 1.32 43.5 460/430 93. 9 201

It was expected that air drying of the ore would have had only a marginal effect on the sulphide content of the ore, but that it caused some oxidisable species in the ore, probably elemental iron, to undergo changes that were responsible for the observed reduced oxidant demand.

As air drying of ore is not a practical or economic means of obtaining, in a commercial minerals processing plant, the reduction of oxidant consumption observed in the laboratory, two further tests were conducted. These were: oxidation by aeration of the ore slurry before leaching (Example 2), and removal of the species responsible for oxidant consumption by flotation (Example3) and magnetic separation (Example 4).

Example 2: Effect of Aeration Aeration was carried out on milled slurry having a neutral pH. After 3 hours aeration the oxidant consumption was comparable to that observed in the laboratory with drying of the ore-see Tables 1 and 2.

It is apparent that aeration dramatically reduces the subsequent oxidant consumption, possibly by the oxidation or passivation of some easily oxidised or passified component which would otherwise consume chemical oxidant in subsequent acid leaching.

Aeration appears to also reduce the concentration of uranium in the residue by up to 30 ppm U308, which is equivalent to an improvement in extraction of about 1%.

Table 2: The Effect of Aeration on Oxidant Consumption Leach Treatment Oxidant Reagent ORP U Residue No. Conditions Consumption (kg f') (mV) Extraction (ppm U308) (%) Oxidant Acid 478C Wet milled, Pyrolusite 1. 05 35.4 460/438 93.7 205 dried and leached 482C Wet milled, Pyrolusite 0.96 35.6 460/438 94. 1 193 dried and leached- repeat 498B Wet milled, Pyrolusite 1.32 43.5 460/430 93.9 201 dried and leached- new batch 4831 Wet milled Pyrolusite 2.39 39.8 460/441 93.7 206 and leached 498D Wet milled Pyrolusite & 2.63 61.3 460/457 94.4 183 and leached pH 1.8 497B Wet milled Air (pH 3) 2.86 33.6 280/460 93.5 214 and leached-then repeat pyrolusite 498C Wet milled Air (pH 3) 2.51 36.2 280/460 93.5 211 and leached-then repeat pyrolusite 497C Wet milled Air (pH 4) 2.52 33.1 240/460 93.3 218 and leached then pyrolusite 500F Wet milled Air (neutral) 1.64 37 94.3 185 and leached for Ih, then pyrolusite 500G Wet milled Air (neutral) 1.16 36.6 94.4 182 and leached for 3h, then pyrolusite 503D Wet milled Air (neutral) 2.06 33.8 461/426 94.1 194 and leached for 3h at 35°C, then pyrolusite 503E Wet milled Air (neutral) 1.54 47.7 462/437 95.1 160 and leached for 3h, then pyrolusite H 1. 8

Example 3: Effect of Removing the Magnetic Fraction In laboratory tests, a magnetic fraction, representing 1 wt% of the feed and containing 1.8% of the uranium in the feed, was removed. A laboratory Readings WHIMS separator was used (at 10A or 15500 gauss).

Leaching of the WHIMS bulk non-mags, after settling to the leach slurry density, required 1.21 kg t-'pyrolusite-see Table 3. Once again, this was comparable to that obtained by drying the ore before leaching.

The WHIMS separation was repeated for Leach 501A with aeration carried out before leaching as well. Uranium reporting to the magnetic fraction was 2.7% with a mass of 0.7%. Oxidant consumption is comparable to those using aeration or WHIMS alone.

The conclusion from this test is that the oxidising reagent consuming species is removed into a small magnetics fraction. This species is readily oxidised/passivated by air.

Table 3: The Effect of Pre-Processing Milled Slurry by WHIMS on Pyrolusite Oxidant Consumption Leach Treatment Conditions Reagent Consumption (kg t-') ORP U Residue No. (mV) Extraction (ppm % U308 Oxidant Acid 478C Wet milled, dried and 1.05 35.4 460/438 93.7 205 leached 4831 Wet milled and leached 2. 39 39. 8 460/441 93. 7 206 500E Wet milled, WHIMS 1. 21 41 460/437 93.6 208 and leached 91. 9 * 501A Wet milled, WHIMS 1.26 36.6 460/441 93.9 192 aerated and leached (91. 4) * * after accounting for loss to the mags product Example 4: Effect of Flotation Flotation is both an aeration process and a sulphide removal process. It was expected that, with flotation of the ore, it would be difficult to differentiate between the effects of aeration and sulphide removal with regard to their influence on subsequent reduction in the consumption of oxidant in leaching. For this reason, aeration was

compared with flotation. A preliminary set of flotation tests was carried out on primary an ore composite containing 0.744 % U308 and 0.66% sulphur as S, so that the likely response of the ore to flotation could be assessed. These flotation results are given in Table 4. Although the flotation concentrates were observed to contain sulphide minerals, they also contained a considerable quantity of ore slimes. On the basis of the analyses ie lowest U loss and comparable Fe and S removal, the conditions used in test Example 2 were chosen for the flotation of the ore before leaching.

Table 4: Flotation Test Results Test Settled Flotation Conditions Float Product Loss to No. Sludge floated Volume fraction (mL) (%) pH Collecto Conditioning/Weight % U30s % Fe % S U S r (g t') Flotation Time (g) (mins) I A 250 9. 0 80 5/5 7. 6 0.801 12.54 8.83 3.3 1B 220 9.1 20 10/5 6. 7 0.658 12.32 8.63 2.4 1 C 250 7. 3 20 10/5 10. 6 0.691 10. 90 7.12 3.9 1D 250 9.0 20 20/20 12.3 0.676 9. 89 6.28 4. 5 Results of the leach carried out after flotation, Leach 501B, are compared with other results in Table 5. The composition of the flotation concentrate compared with the magnetic concentrate suggests that other iron rich, low S, minerals (Fe, Mg and aluminosilicates) are recovered in a magnetic fraction whereas flotation is removing the sulphide minerals.

Table 5: The Effect of Flotation on Pyrolusite Oxidant Consumption Leach No. Treatment Conditions Reagent Consumption (kg t') ORP Residue (mV) Extraction (ppm O, ou Oxidant Acid 478C Wet milled, dried and 1. 05 35.4 460/438 93.7 205 leached 4831 Wet milled and leached 2. 39 39. 8 460/441 93.7 206 500E Wet milled, WHIMS 1. 21 41 460/437 93.6 208 and leached (91. 9) * 501A Wet milled, WHIMS 1.26 36.6 460/441 93.9 192 aerated and leached 91.4) * 501B Wet milled, floated and 3.73 45.4 460/436 94.1 190 leached (92. 3) * * after accounting for loss to the WHIMS magnetics or flotation products As shown in Table 5, flotation failed to achieve the same reduction in oxidant consumption observed with air drying of the ore. This failure may have been attributable, at least partially, by flotation reagents. The overall result was an increase in both oxidant and acid consumption.

Table 6: Summary of Oxidant Consumption with Respect to Treatment After Milling Leach No. 4831 495C 497B 497C 497D 498C 498D 500E 50OF SOOG 478C 498B Base case Aeration no Yes yes yes no yes no no yes yes Aeration pH 7/3 3 4 3 no neutral neutral Aeration 1/2 2 2-2 1 3 time (h) Comments Aerated pyro rpt pH 1. 8 WHIMS dried Dried after Neutral 2 h at 497B leach solids Solids pH3 Leach (24h) results Acid (kg/t) 39.8 53. 1 33.6 33. 1 33.3 36.2 60.9 41 37 36. 6 35. 4 43.5 Extraction 93.7 94.4 93.5 93.3 93.1 93.5 94.4 93.6 94.3 94.4 93.7 93.9 % Residue 206 182 214 218 225 211 183 208 185 182 205 201 (ppm U308) Pyrolusite (kg/t) Leach Time zu 2 1. 55 0.66 1. 99 1. 88 2.09 2.06 2.39 0.56 1. 00 0. 52 0. 65 4 2. 10 2. 49 0.72 1. 17 0.64 0.50 0.99 6 2.14 1.03 2.50 2.20 2.45 2. 47 8 2. 39 2. 63 1.00 1.37 0.91 0.68 1. 31 10 1. 25 2.85 2.53 2.45 2. 51 12 1. 21 1.64 1.16 1.05 1.32

Example 5: Effect of Ore Type In this Example, ores from other sources were tested. Results are compared in Table 7. The effect varies considerably with ore type, which might relate to ore hardness which in turn has an effect on grinding iron consumption.

Example 6: Effect of a Ceramic Grinding Mill In order to determine whether the observed reduction of oxidant consumption is associated with the type of milling or, more specifically, with metallic iron in the ore arising from milling, grinding tests were carried out using a porcelain mill and grinding media. After doing an assessment to determine the milling conditions necessary to obtain a standard grind, leaches were carried out on slurries prepared from such material, with and without aeration. Results are compared with those using a steel laboratory rod mill.

The results are summarised in Table 8.

The use of ceramic grinding media has dramatically reduced the oxidant demand.

This indicates that the metallic iron picked up during milling can be a significant factor in oxidant consumption during leaching. Oxidant consumption is still higher than that observed after aeration of a slurry, which has been milled in a steel laboratory mill-see Table 6.

Table 7: Application to Other Ore Types-3 h aeration Leach No. Ore Type Head Reagent ORP U Residue and Assay (%) Consumption (mV) Extraction (ppm Treatment (kg f') (%) UJO8) U308 S Oxidant Acid 4831 S5 0.327 0.82 2.39 39.8 460/441 93.7 206 standard 500G S5 1. 16 36.6 460/439 94.4 182 aerated 501C Sil 0.689 1. 5 6.95 42.6 458/426 95.9 282 standard 501 D Sil 2.88 43. 1 460/422 96.2 264 Aerated 501E S15 0.393 ? 5.12 28.5 461/433 95. 3 183 standard 501F S15 1.85 24 457/431 95.4 181 aerated 503F S 16 0. 482 3. 77 48.4 460/454 93. 3 321 503 G S16 3.21 47.2 463/454 93.0 339 aerated Table 8: Effect of Milling in a Ceramic Ball Mill Leach Ore Type Head Reagent pH ORP U Residue No. and Assay (%) Consumption (mV) Extraction (ppm Treatment k t''% U O U308 S Oxida Acid nt 503 A Sll 0.689 3.46 40.3 1. 95 460/421 95.2 334 Standard grind in ceramic mill 503 B Sil 0.689 3. 18 Probe 1. 95/460/379 96.9 213 Standard failed 0.44 grind in ceramic mill + aeration 503 C Sil 0.689 5.38 43. 5 1. 95 460/430 95.0 344 Standard grind in steel mill + 3h agitation 501C 0.689 6. 95 42.6 1.95 460/426 95.9 282 Standard grind in steel mill

Example 7: Milled Ore Sample From Site To further investigate the effect of laboratory milling compared with that observed in practice, a milled ore sample from a uranium mine's neutral thickener underflow was aerated. This slurry contained 53 wt% solids. Its density was adjusted to 50 wt% solids for leach tests. The ore was assayed to contain 0.257% U308 by the DNA method and 0.365% S (total) and 0.19% C by the Leco method.

Aeration of this milled ore resulted in a reduction in oxidant consumption from 2.5 to 2.1 kg t', or 16%, after leaching at standard pH/ORP, as shown in Table 9. When the severity of leaching was increased to pH 1.7 and ORP 500 mV, the same effect on oxidant consumption after aeration was observed, namely from 4.1 to 3.4 kg t-, or 16%.

Uranium concentrations in the solute after aeration and leaching were slightly lower, as was uranium extraction.

Drying the ore slurry reduced the oxidant consumption from 2.52 (test 508A) to 1.60 (test 509A) kg t'1, or by 37%, whereas treatment of the ore by wet magnetic separation (test 509B, lwt% to the magnetic fraction containing 0.578% U308 and 10.88% Fe203) reduced the oxidant consumption from 2.52 (test 508A) to 1.94 (test 509B) kg t-', or by 24%. This confirmed earlier results, reported in Table 3 ie that the treatment of ore to remove highly magnetic mineral phases is not as effective as drying the ore, in reducing the oxidant demand.

Increasing the aeration temperature from 45°C (508B/510B) to 65°C (510C) reduced the oxidant consumption from 2.5 to 1.78 kg t-l, or by 29%. Aerating at a higher pH of 8 reduced the oxidant consumption from 2.12 to 1.86 kg t', or by 12%. Increasing drying temperature from 40°C (509A) to 80°C (510A) reduced the oxidant consumption from 1.6 to 1.25 kg t-, or by 22%. All reductions in oxidant requirement were accompanied by a reduction in acid consumption.

Table 9: Application to Neutral Thickener Underflow 12/98 Leach Ore Type Head Reagent pH ORP U Residue No. and Assay (%) Consumption (mV) Extraction (ppm Treatment (kg tn) (%) U308) U308 S Oxidant Acid 508 A none 0.257 0.365 2.52 28.3 1. 95 460/43 93.5 167 0 510 B Repeat 0.257 0.365 2.48 26.6 1. 95 460/42 93.0 179 9 508 B Aerated, 45°C 0.257 0.365 2. 12 30.5 1. 95 460/43 93.3 173 3 5tOC Aerated, 65°C 0.257 0.365 1. 78 22.8 1. 95 460/43 92. 1 203 6 510 D Aerated, 0.257 0.365 1.86 25.6 1. 95 460/43 92.9 183 45°C, H 8 6 508 C none 0.257 0.365 4.06 51. 1 1. 70 500/44 94.8 134 9 508 D aerated 0.257 0.365 3.40 50.3 1.70 500/45 94.6 138 1 509 A Air dried, 0.257 0.365 1.60 29.7 1.95 460/43 93.8 159 40°C 4 510 A Air dried, 0.257 0.365 1.25 22.3 1. 95 460/43 92.9 183 80°C2 509 B WHIMS 0.254 ? 1.94 32.1 1.95 460 93.3 171

Example 8: Aeration of Uranium Slurry from pH 5.0-6. 5 The effect of pH on the aeration pre-leach was investigated. Three runs (respectively referred to as runs 541A, 541B and 541C) in each of which an aeration pre-leach step was incorporated, were carried out at nominal pH values of respectively 5.0, 5.5 and neutral (meaning no deliberate acid addition). Although a single figure is used to indicate the nominal pH, the actual pH varied during the contacting step. Variations of as much as about 1 pH unit were observed over a period of three hours. In the case of Run 541B, the initial pH was measured as 6.65, whereafter it dropped to 4.94 after 1 hour, 4.71 after two hours. Lime was then added to prevent the pH from dropping too far, so that it was 4.96 after 3 hours.

In each case, the aeration step was followed by a standard uranium leach. A leach without aeration (Run 541D) was also carried out, for comparison purposes. These results are shown in Table 10, and are compared with a leach on the same ore sample (F4) from which the magnetic fraction had been removed (Run 531B).

The magnetic fraction was removed using a strong rare earth magnet and was analysed by electron microscopy. It was found to consist mostly of particles of mild steel. However, there was also evidence of chlorite, quartz, iron oxide, apatite, ilmenite and pyrite in smaller amounts.

Table 10 Aeration pre-leach at different pHs (summary) Leach Treatment pH of ORP of Oxidant Solution assays after 3 hours of aeration No. Conditions aeration aeration addition (mV) (kg/t) Fe (mg/L) U S Cu (m) () (m 541A Neutral 6. 4-6.5 45-90 1.29 <0.1 14.9 6.3 <0. 1 aeration for 3 hours 541C Aeration for 5. 3-5. 5 70-125 1.34 4.7 (25-73 16.4 6.6 0.8 3hatpH5. 5.... 3 h at pH 5. 5 during aeration) 541B Aeration for 4. 7-5. 0 120-155 1.38 93 (44-180 48.9 6.4 2.2 3 hours at during aeration) pH 5 514D no aeration - - 2.65 531B Magnetic 1. 31 fraction removed

In each of Runs 541C, 541B and 541A, which incorporated an aeration pre-leach step, at nominal pH values of 5 to 6.5 respectively, a reduction was observed in the consumption of oxidant in the subsequent leaching step, though the iron concentration in solution during aeration was less than 200 mg/L.

From Table 10, it can be seen that the higher the aeration pH, the lower the final iron concentration in solution after 3 hours of aeration.

Example 9: Solution assays after aeration pre-leach For additional comparison, assays were done on the solutions of various runs and are compared with the results obtained from aeration pre-leaches at nominal pH values of 3 and 4. These results are shown in Table 11. It was noted that the aeration pre-leaches at nominal pH values of 3 and 4 did not result in a reduction of oxidant consumption in the subsequent leach. The iron concentration during the pre-leach was 1440 mg/L for pre- leaching at the nominal pH value of 3 and 2190 mg/L for pre-leaching at the nominal pH value of 4. It is believed that the dissolved iron in these cases would have been present in the form of Fe2+, since Fe is not very soluble above pH 3.

Table 11 Aeration pre-leach at different pHs (summary) Leach Treatment Acid pH ORP Extraction Solution Assays No. Conditions (kgt') (mV) (% U) Fe U S Cu (mg/L) (mg/L) (g/L) (mg/ L) 497B Aeration for 2 10. 2 2.99 170-40.6 2190 1310 10.5 101 hours at pH 3 277 497C Aeration for 2 4. 4 3. 88 50-16.5 1440 642 9.11 27.2 hours at pH 4 239 503B Neutral aeration 0 6.53 165-<0. 2 1.5 1.22 <0. 3 for 3 h at 45°C in tap water 503D Neutral aeration 0 7.34 85-<0. 2 1.1 0. 35 <0. 3 for 3 h at 35°C in tap water

Example 10: Results of various runs compared Various process parameters determined for a number of different runs, with aeration at varying nominal pH levels, were compared with three runs in which no aeration was applied. The results are set out in Table 12.

Table 12 Comparison of various process parameters (All Runs subjected to acid leaching for 12h) RUN Ore Aeration Aeration Fe in Leach Acid Oxidant U Fe in pH time (h) sol. pH Cons. Cons. recovery leachate mg/L kg/tonne kg/tonne % mg/L 497B S5 3 2 2190 1. 94 26.5 2.85 90.1 2610 497C S5 4 2 1440 1.95 25.6 2.53 90.9 2340 498C S5 3 2 2180 1.95 28.3 2. 51 91.0 2570 500F S5 Neutral I ND*2 1. 95 28. 8 1. 64 92.8 2790 500G S5 Neutral 3 Nid*2 1 95 28.9 1. 16 92.2 2640 501A S5 6.55 3 <0.2 1. 95 26.1 1. 26*' 91. 1 1670 503D S5 7.34 3 <0. 2 1.94 26.0 2. 06*3 92.9 2070 503E S5 7.02 3 <0.2 1. 8 32.2 1. 54 93.8 2790 541A F4 6.4-6. 5 3 0-1. 2 1.95 22.0 1. 29 97.3 2390 541B F4 4. 7-5.0 3 44-180 1.96 22.6 1. 38 96.7 2020 541 C F4 5. 3-5.5 3 47-73 1.95 21. 4 1. 34 96.5 2530 541D F4 - 0 ND*2 1. 96 24.1 2.65 95.9 2210 531A F4 - 0 ND*2 1. 98 22.8 2.54 95.5 3030 531B F4 - 0 ND*2 1.98 22.6 1.31*1 95.3 2050 NOTES: * Magnetic fraction removed *2 Not determined *3 Aeration was carried out at 35°