CONTAINING ORES

BACKGROUND TO THE INVENTION his invention relates to a process for beneficiation of uranium from uranium containing ores, and in particular low grade Witwatersrand ores, by means of a flotation process.

Uranium flotation has been evaluated historically in a number of different applications. However, none of these provide a viable solution for the economic recovery of uranium from low grade uranium containing ores. So, for example, one prior art process suggests a two stage uranium flotation process designed to float apatite, as a consequence of which uranium, which is associated with the apatite, is recovered. However, in ores where uranium is not associated with apatite, such as in the Witwatersrand ores, this process is not applicable.

Test work done in Canada indicated that reagents such as Cupferron™, Kelex-100™ and TOPO™ could be used as flotation collectors to recover uranium, with Cupferron™ showing the most promising results. However, Cupferron™ has since been identified to be a highly likely carcinogen and hence would not be suitable in commercial applications. In addition, the particle size of the ore tested was 85% passing 44μι·η, which is smaller than the particles that are typically found in Witwatersrand milled products, which is 85% passing 75pm. In addition, the ore floated in the prior art test work contained approximately 985ppm of U ₃ 0 ₈ , which is significantly higher that the 50ppm to 100ppm of U ₃ 0 ₈ that are found in the targeted Witwatersrand ores. When this prior art process was applied to low grade Witwatersrand ores, the consumption rate of reagents was up to 4 kg per tonne of ore floated, which is very high. Although reasonable recoveries were achieved, the grade of uranium in the concentrate produced was still too low to recover the uranium economically, making the process uneconomical for low grade ore flotation.

In the Witwatersrand ores uranium occurs naturally with gold, ranging in grade from less than 50ppm up to approximately 500ppm. Depending on the uranium sale price, typically only sources containing more than about 100ppm of uranium (as U ₃ 0 ₈ ) can be processed economically with current process technologies. However, the largest majority of underground sources and surface tailings contain uranium concentrations of less that 100ppm, which make these sources uneconomic to process for uranium recovery.

In order to make recovery of uranium from such low grade sources economic, a low operating cost beneficiation step is required, such as flotation. Sulphide flotation has been utilised in the past and is currently still in use as a means of upgrading uranium. However, these process methodologies provide a very low yield of uranium into the concentrate. Typically a maximum of about 40% is achievable using sulphide flotation methodology.

It is accordingly an object of the present invention to provide a process for the beneficiation of uranium from low grade ores and particularly to provide a low operating cost process giving a high yield of uranium into the concentrate at an economic grade. SUMMARY OF THE INVENTION

According to the present invention, there is provided a process for beneficiating uranium, for example from uranium containing ore particles or tailings, the process including the step of obtaining an aqueous feed slurry containing uranium, with a density of 20% or greater, by weight, particles, the particles typically having a size in the range of between 80% and 90% passing 75pm, typically 85% passing 75pm; and passing the slurry through an oxide flotation step where the slurry is subjected to flotation in the presence of a collector for oxide minerals and the flotation slurry density is regulated at 20-45%, preferably 25-45%, more preferably 30-45%, typically 35-40% by weight solids.

The oxide flotation step should be carried out at a pH of less than 9, preferably less than 7, typically 3 to 5, most preferably 4.0 to 4.5.

Preferably, the collector for oxide minerals is added incrementally in multiple steps in a cumulative amount of less than 400g/t, typically greater than 100 but less than 200 g/t, typically in a least three steps and up to six steps:

Step 1 : add 50-110g/t, preferably 70-90g/t, collector and float for 10 to 20 minutes, typically 15 minutes with a mass recovery (mass pull) into the concentrate of 7 to 11%, preferably 9%;

Step 2: add 20-90g/t, preferably 50-70g/t, collector and float for 10 to 20 minutes, typically 15 minutes with a mass recovery (mass pull) into the concentrate of 5 to 9%, preferably 7% mass pull;

Step 3: add 20-50g/t, preferably 30-40g/t, collector and float for 10 to 20 minutes, typically 15 minutes with a mass recovery (mass pull) into the concentrate of 5 to 9%, preferably 7% mass pull. The collector for oxide minerals may be a fatty acid collector, an acidic phosphoric acid mono- and di-ester mixture or a combination of the two collectors, for example an alkyl hydroxamate anionic collector (fatty acid) such as AM2™ and/or Flotinor SM15™.

Prior to the oxide flotation step, the slurry may be passed through a sulphide flotation step where the slurry is subjected to flotation -in the presence of a collector for sulphide minerals and the flotation slurry density is at 20-45%, preferably 25-45%, more preferably 30-45%, typically 35-40% by weight solids.

The sulphide flotation step may be carried out at a pH of 3 to 8, preferably a pH of 6 or lower depending on the sulphide collector used.

The sulphide flotation step may be carried out in the presence of a sulphide activator which preferably also destroys cyanide, such as lead nitrate or copper sulphate, for example75-250g/t, preferably 75-100g/t copper sulphate.

The sulphide flotation step may be carried out in the presence of a frother, for example 40-200g/t, preferably 40-55g/t Dowfroth™.

Preferably, the density and pH of the slurry is adjusted and sulphide activator added in a conditioning step prior to the sulphide flotation step.

The collector for sulphide minerals is preferably a xanthate such as Sodium Isobutyl xanthate (SIBX) or Potassium amyl xanthate, preferably Potassium amyl xanthate. Other collectors that may be used are Senkol™ (a sulphur based collector available from Senmin (Pty) Ltd, or AeroPromotor™ available from Cytec Company).

The collector for sulphide minerals may be added in an amount of 50- 200g/t, preferably 90-120g/t. The sulphide flotation step may be conducted for 10 to 20 minutes, preferably 13 to 15 minutes, with a mass recovery (mass pull) into the concentrate of 5 to 7%, preferably 6%.

BRIEF DESCRIPTION OF THE DRAWINGS

Figure 1 is a basic flowchart of the process of the present invention; is a graph showing the recoveries of uranium and gold that are achieved by the process of the invention, using AM2™ as an oxide collector; is a graph showing the recoveries of uranium and gold that are achieved by the process of the invention, using Flotinor SM15™ as an oxide collector;

Figure 4 is a graph showing the cumulative recovery of uranium achieved by the process of the invention, using Flotinor SM15™ as an oxide collector;

Figure 5 is a graph showing the uranium grades in the concentrate used in Figure 4;

Figure 6 is a graph showing the gold and uranium grades obtained with a feed slurry of 38% solids, sulphide flotation performed at pH6, and oxide flotation at pH 4 using Flotinor SM15™; and

Figure 7 is a graph which demonstrates the process selectivity for uranium and gold even when applied to high grade Witwatersrand ores. DETAILED DESCRIPTION OF THE INVENTION

The basic concept of the invention is a froth flotation process to beneficiate uranium economically from both low grade tailings dump ores or high grade ores; by upgrading into a concentrate. The total flotation time required is approximately 74 minutes (depending on feed grade, flotation kinetics and cut-off grade required) with multiple stages of reagent addition for satisfactory uranium recovery, with the first stage being the sulphide flotation for pyrite recovery. Potassium amyl xanthate was selected as the sulphide collector for the process of this invention because xanthates increase gold recovery. The selectivity of the process for recovery of uranium in the concentrate is enhanced by maintaining the slurry medium at a density similar to the feed. The slurry is maintained at an optimal % solids in the slurry of between 25% and 40% solids, preferably 35-40%. The percentage solids impacts on both the flotation kinetics and selectivity of the valuable minerals into the concentrate. The uranium collector (a collector of oxide minerals) used for the process is either a fatty acid collector, an acidic phosphoric acid mono- and di-ester mixture or a combination of the two collectors. Examples of these include AM2™ (available from a South African company - Axis House (Pty) Ltd) or Flotinor SM15™ (available from Clariant Corporation). Flotinor SM15™ flotation reagent is an acidic phosphoric acid mono- and di-ester mixture and AM2™ is an alkyl hydroxamate anionic collector (fatty acid). Full scale tests and pilot plant testing with sulphur paydam material have shown that it is possible to achieve in the combined concentrate of the sulphide rougher and 3-stage uranium rougher, a uranium recovery of at most 71.5% in 37% of the mass. This can be achieved, together with control of the feed within the specified feed solids % range and pH range; by a more even distribution of collector down the flotation bank by using up to 50% of the collector at the top of the bank, and subsequent smaller additions in the downstream cells. Thereby, a reduction of reagent consumption is also achieved. With reference to Figure 1 , the present invention relates to a process for beneficiating uranium from uranium containing ore particles or tailings by means of flotation, by forming an aqueous feed slurry 10 with a density of 20% or greater, typically 35%-45% by weight ore particles (the ore particles have a size in the range of between 80% and 90% passing 75pm, typically 85% passing 75pm), and then passing the slurry 10 through the following stages: . .. -

STAGE 1 : Conditioning

1. Maintaining the feed aqueous slurry 10 at a constant density of 25- 40% solids or 1.28t m ³ .

2. The addition of acid 12 (sulphuric acid) to the feed slurry to achieve a pH of 6;

3. Conditioning with agitation for a residence time of 25 minutes by adding a sulphide activator 14 which also destroys cyanide (75- 10Og/t copper sulphate) with agitation and for a residence time of 25 minutes (only required if the ore contains sulphide minerals).

STAGE 2: Sulphide Flotation

4. Addition of a collector for sulphide minerals 16 (90-120g/t Potassium amyl xanthate) and after 2 minutes a frother 18 (40-55g/t Dowfroth™ available from The Dow Chemical Company) is added.

5. Cell aeration self induced or forced air is started and sulphide flotation is done for 14 minutes with a mass recovery (mass pull) into the concentrate of 6% to form a first froth concentrate carrying the uranium material.

6. removal of a first froth concentrate 19. STAGE 3: Oxide Flotation

7. Step 1 : Adjust slurry pH to within the range of 4-4.5 with acid 20 (sulphuric acid), and condition with a collector of oxide minerals 22 (70-90g/t Flotinor SM1 ⁵ ™ and/or AM ² ™) for 15 seconds.

8. Cell aeration self induced or forced air is started and float for 15. minutes with a mass recovery (mass pull) into the concentrate of 9% (depending on feed grade, flotation kinetics and cut-off grade required).

9. Step 2: addition of collector of oxide minerals 22 (50-70g/t Flotinor SM15™ and/or AM2™) and flotation for 15 minutes for 7% mass pull (depending on feed grade, flotation kinetics and cut-off grade required). The flotation slurry density must be regulated at 35-40% solids by the addition of water 24. pH must be maintained at 4.0 - 4.5 with the addition of acid 26 (sulphuric acid).

10. Step 3 : Addition of a collector of oxide minerals 22 (30-40g/t Flotinor SM15™ and/or AM2™) and flotation for 15 minutes for 7% mass pull (depending on feed grade, flotation kinetics and cut-off grade required). Adjust density by addition of water 24 to maintain 35-40% solids.

Note: Additional steps may be added depending on feed grade, flotation kinetics and cut-off grade required.

11. Removal of a second froth concentrate 28. STAGE 4: Extended gold flotation

12. Gold scavenger flotation of the tails from the uranium flotation by addition of a collector of oxide minerals 30 (20-30g/t SM15™ and/or AM2™) and obtaining 7% mass pull in 15 minutes. The residence time of the oxide flotation varies depending on the ore, as it is determined by a number of factors including the flotation kinetics, the mineralogy and the uranium cut off grade or breakeven grade of the ore. The extent of the gold flotation similarly, is influenced by the gold cut off grade of the ore and the flotation can be continued until the cut-off grade is reached. Stages 1 and 2 part of the process may not be applied if there is a negligible sulphur content of the feed ore and similarly the extended gold flotation Stage 4 may be omitted if the feed ore has negligible gold value.

The first and second froth concentrates 19 and 28 obtained from the process may be treated in a number of ways to obtain a uranium product, for example the froth may be pre-treated by an oxidation process or aggressive leach process such as pressure oxidation, followed by a uranium leach and then a gold leach.

The invention will now be further described an illustrated with reference to the following non-limiting examples and the Drawings.

Example 1 - West Wits tailings laboratory flotation with SM15™

The first example is for using SM-15™ as a collector. The sample was conditioned for 30minutes at a pH of 4.5 to 5 with 200g/t of CuS0 ₄ . After conditioning 100g/t of Dowfroth™ 250 was added and 100 g/t of (Potassium amyl-xanthate) KAX sulphide collector and conditioned for an additional 2 minutes. A sulphide float was then conducted for 7 minutes.

After the sulphide float, the pH was adjusted to 4 and conditioned for 20 minutes using 800 g/t of SM15™. Two 5 minute rougher float concentrates was recovered from the uranium float.

Conditions for this example are provided in Table 1 below. Table 1

The results are depicted in Figure 2. An overall U308 recovery of 83% was achieved using SM15™ as a uranium collector. Using only a sulphide float the recovery is generally limited to approximately 40%.

Example 2 - West Wits tailings laboratory flotation with AM2™

The second example is for using AM2™ as a collector. The sample was conditioned for 25 minutes at a pH of 4.3 to 5 with 200g/t of CuS0 ₄ . After conditioning 10Og/t of Dowfroth™ 250 was added and 100 g/t of of (Potassium amyl-xanthate) KAX sulphide collector and conditioned for an additional 2 minutes. A sulphide float was then conducted for 7 minutes.

After the sulphide float, the pH was adjusted to 4 and conditioned for 7 minutes using 800 g/t of AM2™. Three 5 minute rougher float concentrates was recovered from the uranium float.

Conditions for this example are provided in Table 2 below.

Table 2

Dosage (g/t) Time (mln)

Stage

H ₂ SO„ CuSO, KAX Dow250 AM2 Conditioning Float pH 4.3 200 - - 25

Rougher 1 7

- - 100 100 2

Rougher 2 pH 4 - - - 800 7 5

Rougher 3 pH 3.54-4.4 - - - 800 3 5

Rougher 4 pH 5.3 - - - 40 3 5 The results are depicted in Figure 3. An overall U308 recovery of 89% was achieved using AM2™ as a uranium collector in a mass pull of 36%. This provided a concentrate with a grade higher than that of the uranium processing cost cut-off grade.

Example 3 - Sulphur paydam pilot plant scale flotation at pH 4 with SM15™

The following test was performed on a pilot plant to test the amenability of the sulphur paydam material on a continuous process. The pilot plant feed rate was 13L/min and each flotation cell was 180L, representing either a sulphide rougher or uranium rougher. The configuration of this test was one sulphide rougher and 3 uranium roughers. The feed was conditioned with copper sulphate for 30-40 minutes. S 15™ was added directly to each of the uranium rougher cells. Dilution water was added in uranium roughers 2 and 3. The total SM15™ dosage was 200g/t in stage additions with 80g/t addition in the first uranium rougher followed by smaller additions in the subsequent stages. The sulphide flotation was performed at pH 4 with a feed solids % of 40% and copper sulphate consumption 70g/t and PAX consumption of 71g/t.

Figure 4 depicts the cumulative uranium recovery with the mass pull or mass % of the combined concentrate (pyrite and uranium); for two different runs with SM15™ obtained from pilot plant runs with Sulphur paydam tailings material. The sulphide flotation for both runs were at a pH of 4 with a feed slurry of 38.5% and 40% solids. 71% uranium recovery is obtained with a 37% mass pull. The run with the lower %mass pull of 31.7% achieved a uranium recovery of 65.3%

Figure 5 depicts the corresponding uranium grades in the concentrate, for the mass pulls depicted in Figure 4, for runs with 0.11 and 0.12kg U308/t feed grade. Example 4 - Sulphur paydam pilot plant scale flotation at pH6 and uranium flotation at pH 4 with SM15™

The configuration for this test included a gold scavenger stage and the pH adjustment for the first uranium rougher was done in the cell with dilute acid to a pH of 4. Dilution water was added in uranium roughers 2 and 3. The sulphide rougher was controlled at a pH of 6 with the addition of dilute acid. The reagent consumptions were: 119g/t of PAX, 75g/t of copper sulphate, with 4 stage additions of SM15™.

Results of this test are shown in Figure 6, which shows gold and uranium concentrate grades obtained with feed grade 0.141kg U308/t and 0. 9g Au/t. The gold and uranium grades obtained in this run was obtained with a feed slurry of 38% solids and the sulphide flotation was performed at pH 6 with the uranium flotation at pH 4 with SM15™.

Example 5 - Reef pilot plant scale flotation with SM15™

The same pilot plant cells and feed rate as in Examples 3-4 was used for this test. The pH was 6 in the sulphide rougher and pH control to 4.5 was effected in the uranium rougher 1. The configuration was a sulphide rougher, three uranium roughers and a gold scavenger stage. Sufficient dilution water was added to Uranium rougher 3 only. The feed % solids for this test was 40% and the sulphide flotation was run at a pH of 6 and the pH in the first uranium rougher was adjusted to a pH of 3.9. The reagent consumptions were: 80g/t of copper sulphate, 73g/t of PAX, with 4 stage additions of SM15™.

Figure 7, which shows gold and uranium concentrate grades obtained with feed grade 0.373kg U ₃ 0 ₈ /t and 7.706g Au/t, demonstrates the process selectivity for uranium and gold even when applied to high grade

Witwatersrand ores. The mass pulls for this particular run was low and indicates the particularly high gold recovery obtainable in a small mass pull.