WO2011116426A1

WO2011116426A1 - Process for leaching refractory uraniferous minerals

Info

Publication number: WO2011116426A1
Application number: PCT/AU2011/000338
Authority: WO
Inventors: Gary Vernon Rorke; Robert John Ring; Anthony Robert Gee; Neil Garrard; Daniel Arthur Kittelty; Vanessa Gaye Liebezeit; David Bojcevski
Original assignee: Bhp Billiton Olympic Dam Corporation Pty Ltd; Australian Nuclear Science And Technology Organisation
Priority date: 2010-03-24
Filing date: 2011-03-24
Publication date: 2011-09-29
Also published as: AU2011232311A1

Abstract

A process for leaching uranium from a refractory uraniferous mineral including the step of treating a slurry containing said uraniferous mineral and ferrous ions in a reactor with an acid and an oxidant to increase the ORP in the reactor to a value of at least 550m V vs Ag/Ag Cl reference electrode in order to oxidise the ferrous ions to ferric ions, which are used to oxidise said refractory uraniferous mineral and contribute to leaching of uranium from said uraniferous mineral,wherein the slurry has an elevated total dissolved iron content.

Description

PROCESS FOR LEACHING REFRACTORY URANIFEROUS MINERALS

This invention relates to a high redox process for

leaching uranium from a refractory uraniferous mineral .

Recovery of uranium from its ores is commonly carried out by a process which includes leaching the ore or a

concentrate thereof, typically by using a leachant

including sulfuric acid and an oxidising agent (oxidant) . The dissolved uranium is usually separated from the leach solutions by an ion exchange or solvent extraction

process. This produces a purified uranium-containing solution from which uranium is recovered by precipitation, commonly by treating with ammonia to precipitate uranium yellowcake product.

Uranium ore deposits may contain one or more different uranium containing minerals. The most commonly occurring uranium ore mineral is uraninite (UO₂ or U₃O₈) . However, uranium is also present in other less commonly occurring ore minerals, such as brannerite ( (U, Ca, Ce) (Ti, Fe) 2Ο6) and coffinite (U (Si0₄) i-_x (OH) _4x) . These uranium ore

minerals are regarded as refractory, in that they are resistant to acid attack and it can be difficult to recover significant amounts of uranium from these minerals using conventional leaching processes typically employed to treat uraninite. The applicant's Olympic Dam copper uranium and gold deposit has been operating for many years . The current mine operations include an underground mine as well as an integrated metallurgical processing plant. Uranium is recovered from copper flotation tailings by sulfuric acid leaching followed by solvent extraction. In recent years, the applicant has been focussing on test programs to develop new processes and flowsheets to improve the efficiency of uranium recovery at the Olympic Dam mine and operations .

In the current leach process, feed grades are around 500 to 600 ppm uranium expressed as equivalent U₃O₈ with the leaching process resulting in a residue grade in the order of 120 ppm (equivalent U₃O₈) . The overall recovery of 75 to 80 % uranium is attributed primarily to the uranium mineralogy where a significant proportion of uranium is contained within refractory brannerite and coffinite minerals . There are areas of the deposit that contain higher levels of brannerite and coffinite and hence it would be

desirable to develop an improved extraction process for such ores, thus producing a more efficient overall uranium extraction and recovery.

The redox potential of the uranium leaching process currently conducted at the Olympic Dam operations is around 430 mV (vs Ag/AgCl reference electrode) or lower. The present invention is based upon the realisation of the present inventors that by significantly increasing the redox conditions (ORP) during at least part of the acid leaching process, the recovery from refractory uraniferous minerals is greatly enhanced as compared with the

conventional acid leaching process. In addition, in a preferred embodiment the leaching process of the invention can be conducted with significantly lower acid consumption as compared with the conventional acid leaching process.

Accordingly, in a first aspect, there is provided a process for leaching uranium from a refractory uraniferous mineral including the step of treating a slurry containing said uraniferous mineral and ferrous ions in a reactor with an acid and an oxidant to increase the ORP in the reactor to a value of at least 550mV vs Ag/AgCl reference electrode in order to oxidise the ferrous ions to ferric ions , which are used to oxidise said refractory

uraniferous mineral and contribute to leaching of uranium from said uraniferous mineral .

In an embodiment, the ORP is 560mV or higher. In another embodiment the redox potential is 600mV or higher.

In an embodiment, the refractory mineral is one or more of brannerite and coffinite.

In an embodiment, the oxidant is one or more of SO₂/O₂, sodium chlorate, manganese dioxide, peroxide, or Caro' s acid. Preferably, the oxidant is SO₂/O₂. An advantage of using SO₂/O₂ is the relatively lower cost and higher availability as compared with more expensive oxidants , such as sodium chlorate.

In an embodiment, the slurry also contains ferrous ions and the oxidant oxidises the ferrous ions to ferric ions , which are then used to oxidise said uraniferous mineral . In an embodiment, the slurry has an elevated total

dissolved iron content. The total dissolved iron content may be greater than 5 g/L, preferably greater than 10 g/L, more preferably greater than 15 g/L. The total dissolved iron content may be less than 100 g/L, preferably less than 70 g/L, more preferably less than 60 g/L. It has been found by the inventors that when the leach solution contain high dissolved iron, the effect of ORP on uranium solubility is significantly greater than when the leach solution contains relatively low dissolved iron. The maximum total dissolved iron content may correspond to the maximum solubility of iron. Preferably, the oxidation of ferrous ions to ferric ions is effected by treating the slurry with one or more inlet gases containing sulfur dioxide and oxygen, as shown in Eqn 1:

6Fe²⁺ + 3S0₂ + 30₂ ~> 6Fe³⁺ + 3S0₄ ^2~ (1) wherein the delivery rate of the sulfur dioxide gas is oxidation rate limiting, and controlling the concentration of dissolved oxygen in said slurry at an optimum value at which there is an excess of oxygen dissolved in said slurry. This ferrous oxidation process is the subject of applicant ' s copending patent application No

PCT/AU2009/001528 the entire disclosure of which is incorporated herein by reference. The sulfur dioxide and oxygen gases may be introduced into the solution in the form of a gaseous mixture. However, in one embodiment the sulfur dioxide and oxygen are introduced in separate gas streams.

The redox potential of the leaching process of the

invention is dominated by iron species . The redox

potential is therefore a function of the log of the ferric to ferrous ion ratio and is typically expressed in mV (vs Ag/AgCl reference electrode throughout this document) .

In an embodiment, the high redox step of the leaching process of the invention is conducted for a period of time of one hour or greater. The high redox step may be

conducted for a period of time of greater than 5 hours , such as from 10 to 12 hours.

The refractory uraniferous mineral may be included in an ore, an ore concentrate, or a waste material such as ore tailings, slag or dust. In an embodiment, the refractory uraniferous mineral is included in an ore or an ore concentrate which optionally further includes one or more non-refractory uranium mineral/s, such as uraninite, and/or gangue mineral/s, such as siderite and/or chlorite. The ferrous ions may be derived from dissolution of one or more of iron containing gangue mineral/s and

hydrometallurgical process solutions. The ferrous ions may be derived from a ferrous containing liquor derived from the hydrometallurgical processing of ores,

concentrates, waste materials such as tailings, or

combinations thereof. The liquor may comprise or be part of a barren leach solution, a pregnant leach solution, a slurry, a raffinate, a thickener overflow, a solution resulting from reduction of ferric ions , or any other suitable liquor derived from a hydrometallurgical plant. Ferrous ions may also be derived from the dissolution of ferrous containing gangue minerals such as siderite or chlorite. Ferrous also arises from the reaction of ferric ions with base metal sulfide minerals, such as chalcocite, bornite, covellite, chalcopyrite , pentlandite, and

sphalerite .

The pH of the high redox leaching step is ore body

dependant but is typically less than 2.5. In a preferred embodiment, the pH is 1.5 or less. More preferably, the pH is less than 1.3.

In one preferred embodiment, the process is low acid consuming. In that embodiment, the slurry is preferably treated to precipitate jarosite which thereby releases additional sulfuric acid for use in further leaching of the refractory uraniferous mineral . The precipitation of potassium jarosite and release of sulfuric acid proceeds according to the following chemical reaction (Eqn 2) :

3Fe₂(S0₄)₃ + K₂S0₄ + 12H₂0 -> 2KFe₃ (S0₄) ₂ (OH) ₆ + 6H₂S0₄ (2) The precipitation of potassium jarosite in particular also results in a meaningful reduction in downstream

maintenance costs as the potassium is removed under this process almost completely. The high ORP leach step at a temperature of 70°C results in <100 ppm potassium in solutions exiting the leach process, thus significantly reducing the slow precipitation of jarosite in downstream processing such as in Continuous Counter Current

Decantation (CCD) , Solvent Extraction (SX) or Resin-in- Pulp (RIP) circuits.

The temperature of the high redox leaching step is

typically elevated having a minimum temperature of 40°C. The preferred temperature is between 70 and 80°C and may be operated up to 90°C.

The high redox leaching step is preferably conducted under atmospheric pressure conditions. In an embodiment, the high redox leaching step comprises one stage in a multiple stage leaching operation. In a preferred embodiment, the high redox leaching step

comprises one stage of a two stage leaching process .

During the high redox leaching stage, the solids density in the slurry is controlled at a first value which is sufficient to facilitate dispersion and hence

solubilisation of the oxidant (principally oxygen)

throughout the slurry. Subsequently, the slurry is treated in a second leaching stage at a second value of solids density which is higher than the first value.

Typically, the high redox leaching stage is conducted in one or more first reactors and the oxidised slurry is then passed to a thickener before proceeding to the second leaching stage. The first value of solids density is typically less than 55wt% , preferably less than 45 wt% , more preferably less than 40wt%, more preferably less than 30 wt%, more preferably less than 25 wt% , such as about 20 wt% . The second value of solids density is typically greater than 20 wt%, preferably greater than 30 wt% , more preferably greater than 45 wt%, such as about 55 wt%. The purpose of the two stage leaching operation is to allow optimisation of both the oxidation process and the time available for leaching the refractory uranium

minerals. The oxidation process relies upon the transfer of oxygen into solution and the rate of oxygen

solubilisation increases in the process liquor as the pulp density is reduced. The oxygen transfer requirement combined with the requirement to continue the oxidation for the time required to leach and oxidise the bulk of ferrous iron from gangue or sulfide minerals determines the operating time for the first stage of the leach that will be required in order to ensure ORP is maintained for the remainder of the leach process. The slurry is thickened prior to the second leaching stage in order to provide a more economic leach as the total volume of the slurry can then be reduced by a factor of 3 or more, reducing tankage volume required and hence the capital requirements proportionately .

The leaching process of the present invention may be used in the treatment of refractory uraniferous mineral- containing ore or a tailings or concentrate .

Figure 1 is a schematic diagram of pilot plant equipment for carrying out the process of the invention.

Figure 2 exemplifies the control parameters of tank 112.

Figure 3 is a graph showing the Redox profiles through the leach process.

Figure 4 shows the residual unleached uranium profile across the process stages. Figure 5 is a graph plotting the uranium dissolution against ORP in the SIL reactors of the pilot plant. Figure 6 is a graph showing the uranium tenor in leached residues for Runs 6 to 11.

Figure 7 is a comparison of total sulfur added to the pilot unit in the form of both acid and sulfur dioxide gas compared to the sulfur added to the conventional uranium leaching process as acid only at Olympic Dam.

Figure 8 illustrates U deportment in (a) uraninite (b) coffinite and (c) brannerite when leached using a

conventional acid leach and using a process of the

invention .

Figure 9 illustrates U deportment in (a) uraninite (b) coffinite and (c) brannerite when leached using a process of the invention under different values of ORP and pH.

Figure 10 plots the % change in uranium extraction from 440 mV to 550 mV vs iron concentration (ppm) for samples leached under low iron (squares) and high iron (diamonds) concentrations.

Example 1 In Example 1, the process of the invention was conducted in a slurry of a uranium ore sourced from tailings from applicant's Olympic Dam Uranium mine. Figure 1 schematically shows the pilot plant equipment indicated generally at 110 for carrying out the process of the invention in Example 1. The pilot plant 110 included two 0.2 m³ reactor tanks, 112, 113 in parallel into which was pumped a slurry feed at a rate of from about 50 to 100 L/h each from slurry supply tanks 111a and 111b via a slurry stock tank 114, a stream of ferric containing liquor from solution feed tank 115, and a stream of sulfuric acid from acid supply tank 117. Each reactor tank 112, 113 also included a dissolved oxygen sensor 138, a redox sensor 146 and an agitator 118 comprising an impeller 120 driven by a motor 122. The percentage of solids in each reactor is typically around 20% (W/W) . A relatively low solids density is required in the leaching reactors in order to enable efficient dispersion of oxygen throughout the slurry.

An oxidant comprising a mixture of sulfur dioxide and oxygen gas was fed into the mixture of slurry, acid and ferric liquor in each reactor tank 112, 113 via a

respective conduit 132 and gas diffuser 150 which diffused the gaseous mixture into solution below the impellers 120, by which the gas could be effectively dispersed throughout the slurry mixture. The temperature within each reactor 112, 113 was controlled by means of respective heat exchangers 152 and measured by respective thermometers 154.

The source of iron in the oxidised liquor was largely from gangue minerals such as residual sulfides, siderite and chlorite, that occurred with the uranium ore.

After oxygen was adequately dispersed throughout the slurry, the solids density could be increased to more conventional levels. Accordingly, the overflow from each reactor 112, 113 reported to a thickener 160. The

thickener underflow was operated at a solids density appropriate for the size of the thickener outlet 162 which was 45%, for the pilot plant although it is expected that the solids density would be higher (such as around 55%) for a commercial plant.

The thickener underflow containing the at least partially leached solids reported to a train of additional leaching tanks, in this case, three 0.2 m³ stirred leaching tanks 164a, 164b and 164c, for completion of leaching. Each leaching tank included an agitator 166a, 166b, and 166c and was fed independently with sulfuric acid from acid supply tank 117, and had independent temperature control.

The thickener overflow reported back to the solution feed tank 115. The level of oxidised liquor in the solution feed tank was kept substantially constant by the addition of other plant liquors, such as raffinate derived from a uranium solvent exchange circuit. The oxidised liquor is recycled back for use in the reactor tanks 112, 113 in order to maintain an optimal reactor solids percentage (such as 20 wt%) .

The conditions and measured operating parameters under which each pilot plant Run was conducted are presented in Tables 1 and 2, respectively. In Tables 1 and 2 and in Figure 5 the acronym "SIL" means "sulfur dioxide and oxygen in leach" .

Each Run was conducted on a different ore sample

unavoidably having a slightly different mineralogy. Table 1 : Conditions of Pilot Plant Runs and Purpose

In Table 1, Run 9 was a comparative example against which the process of the invention may be compared. The redox conditions under which Run 9 was conducted were

essentially the same as those conditions in the

conventional uranium leaching process which is currently used at Olympic Dam. Figure 2 exemplifies the process control parameters namely sulfur dioxide and oxygen gas flows, dissolved oxygen and the ORP of tank 112. The data shows the ramp up of the process to steady state from start up. Table 2: SIL Pilot Plant Runs and Operating Parameters

Run

Operating Parameters (Avg.)

No.

Solids

SIL Reactor Acidity REDOX DO

Feed pH

Re en . hrs (g/L) (mV) (ppm)

(kg/h)

Run 5 64 2 1.2 12.6 562 6.7

Run 6 55 2 0.9 17.3 567 8.1

Run 7 65 2 1.0 13.9 601 6.4

Run 8 61 2 0.9 19.2 603 6.5

Run 9 64 2 1.1 15.8 435 4.5

Run 10 64 1 1.3 13.1 556 6.6

Run 11 57 2 1.2 12.0 583 6.2 In Table 2, the process parameters for the SO₂/O₂ sparged reactor tanks 112,113 are shown. As is evident, the process was able to be controlled to oxidise the ferrous to reach the ORP (REDOX) values required. Run 9 showed a REDOX value of 435mV in the absence of the process of the invention compared to the controlled potentials for the other Runs . The Run 9 value would have ordinarily been lower, but there was a considerable level of ferric ions in the raffinate used as dilution liquor.

Figure 3 shows the typical REDOX profiles across the various stages in the leach process. In Figures 3 and 4, "SIL" refers to SIL reactors 112,113, "Thickener U/F" refers to underflow from thickener 160, and "LR1 to LR3" refer to leach reactors 164a, b and c, respectively. The standard SIL experiments all exited in the +500mV range as hoped. The Run (10) where the SIL retention was reduced to an hour discharged at only 480mV. This is in

accordance with what would be expected in terms of the rate of ferrous production by the gangue minerals . The shorter the residence time in SIL the less of the total ferrous that is likely to leach is leached. Thus with the 1 hour residence time less ferrous is oxidised.

Figure 4 shows the un-leached uranium profile in the solid residue in various stages across the leach plant,

expressed as equivalent U₃O₈ (ppm) in solid residue. One point of interest is that the uranium in residue in the SIL reactors 112,113 is already lower than the discharge residue in the comparative Run (Run 9) . The profiles also indicate that although the majority of leaching has happened in the SIL reactors 112,113 the rest of the leach train (reactors 164a,b,c) contributes up to an additional 20ppm of uranium dissolution. Figure 5 shows the correlation between the SIL ORP and % uranium extraction in the SIL stage of the process. The graph clearly shows the strong relationship between the ORP setpoint in the SIL process and the initial extraction in spite of changes in relative uranium and gangue

mineralogy between runs .

Figure 6 shows only the residue data of the trends

illustrated in Figure 4. For each Run, overall uranium containing residue amounts (expressed as equivalent U₃O₈

(ppm) in solid) are shown for a particular ore concentrate which is subjected to the leaching process of the

invention ("SIL leach") and a conventional leaching process at the current Olympic Dam operations ("Tails Leach (OD) ") . As noted previously, each Run was conducted under varying acid leach conditions on a different ore sample, having a slightly different mineralogy.

Accordingly, in Figure 6, the uranium tenor of residues from the comparative conventional OD tails leach vary slightly despite the conditions of the comparative leaches being the same. Nevertheless, the effect of the leaching process of the invention may be observed by comparing the uranium tenor in residues of the inventive process within a particular run and of the conventional leaching process, as well as by comparing uranium tenors in residues between Runs . There is significant improvement in uranium

dissolution when using the process of the invention as compared with the conventional leach. Runs 6 and 8 show the greatest improvement in uranium dissolution. This is attributed to a higher acid amount for Run 6 and to higher REDOX conditions for Run 8 which exhibited an ORP value above 600 mV. The SIL leach average increase in uranium dissolved was 46 ppm (average) U₃O₈ in comparison with the conventional ODC tails leach plant performance.

It is noted that, although Run 7 was also conducted at a REDOX value above 600mV, the uranium tenor in the residue was higher than in Run 8. This is attributed to a

difference in mineralogy between the two samples.

Figure 7 compares the sulfur addition to the process of the invention conducted in the Pilot unit and the

conventional leach process currently conducted at Olympic Dam. Run 6 was a high acid Run and so has an

understandably greater sulfur demand (of acid and SO₂) because of greater gangue mineral dissolution. On average the total addition rate of sulfur to the process of the invention is equal to or less than that represented by the conventional tails leach process. This data demonstrates that the acid consumption by the process of the invention is equal to or less than the acid requirement of the conventional leaching process currently used at Olympic Dam.

Example 2 : The deportment of uranium in various uraniferous minerals after using the process of the invention was compared with that of a standard acid leach.

Figures 8(a), (b) and (c) show the respective amounts of uranium in feed material and in leach residues when uraninite, coffinite and brannerite are leached under the indicated pH and ORP conditions . The leach conducted at a pH of 1.35 and an ORP of 420 mV is typical of standard or conventional leach conditions . Two leaches were conducted using the inventive process at an ORP of 550 mV and at pH values of 0.85 and 1.15, respectively.

As is evident from Figure 8 (a) , there is little difference in the amount of uranium in leach residues resulting from leaching uraninite using the conventional process or using a process according to the invention. Further, at the elevated ORP of 550 mV, the uranium deportment is little influenced by the pH of solution.

Figure 8 (b) illustrates deportment of uranium in coffinite residues and it is clear that the amount of uranium in leach residue is significantly lower using the inventive process as compared with using the conventional leach. In addition, at the higher ORP of 550 mV, there is a reduced amount of uranium in the leach residue when the pH of the process is lower.

Figure 8 (c) illustrates deportment of uranium in

brannerite residues. Again, there is a significant reduction in the amount of uranium in leach residues when using the process of the invention. Again, the amount of uranium in the leach residue is further reduced at a lower pH value.

Example 3

Figure 9(a) illustrates uranium deportment in uraninite when leached using the process of the invention which is conducted at an ORP of either 550 or 700 mV and at a pH of either 0.65 or 1.15. It is clear from this Figure that the increase in ORP at a given pH has little effect on the amount of uranium in the residue. At the ORP of 700 mV the amount of uranium in the leach residue was slightly lower at the higher pH value .

Figure 9 (b) illustrates uranium deportment in coffinite when leached using the process of the invention. The amount of uranium leached is similar under all ORP and pH values. It is evident that the process of the invention is very effective in leaching uranium from coffinite. Figure 9(c) also indicates the higher solubility of uranium from brannerite when leached using the process of the invention. The amount of uranium remaining in leach residue is similar under all ORP and pH values .

Example 4

The effect of iron concentration in the leach solution on uranium extraction at ORP conditions of either 440 mV or 550 mV was investigated. The results are shown in Figure 10 which plots the % change in uranium extraction from 440 mV to 550 mV vs iron concentration (ppm) . As is apparent from the graph, when the leach solution contain high dissolved iron, the effect of ORP on uranium solubility is significantly greater than when the leach solution

contains relatively low dissolved iron. Under the

particular conditions exemplified in Example 4, the effect of ORP on uranium solubility appears to decline as iron concentration exceeds about 50,000 ppm.

Accordingly, the process of the invention enables

significantly higher uranium recovery from uranium ores and concentrates, particularly those containing refractory uraniferous ores , without increasing acid requirement and in many cases , resulting in an actual reduction in acid consumption . References to prior art in this specification are provided for illustrative purposes only and are not to be taken as an admission that such prior art is part of the common general knowledge in Australia or elsewhere. In the claims which follow and in the preceding

description of the invention, except where the context requires otherwise due to express language or necessary implication, the word "comprise" or variations such as "comprises" or "comprising" is used in an inclusive sense, i.e. to specify the presence of the stated features but not to preclude the presence or addition of further features in various embodiments of the invention. The invention described herein is susceptible to variations, modifications and/or additions other than those specifically described and it is to be understood that the invention includes all such variations , modifications and/or additions which fall within the spirit and scope of the above description.

Claims

1. A process for leaching uranium from a refractory uraniferous mineral including the step of treating a slurry containing said uraniferous mineral and ferrous ions in a reactor with an acid and an oxidant to increase the ORP in the reactor to a value of at least 550mV vs Ag/AgCl reference electrode in order to oxidise the ferrous ions to ferric ions , which are used to oxidise said refractory uraniferous mineral and contribute to leaching of uranium from said uraniferous mineral ,wherein the slurry has an elevated total dissolved iron content.

2. A process for leaching uranium from a refractory uraniferous mineral according to claim 1, wherein the ORP is 560mV or higher, preferably 600mV or higher.

3. A process for leaching uranium from a refractory uraniferous mineral according to claim 1 , wherein the refractory mineral is one or more of brannerite and coffinite .

4. A process for leaching uranium from a refractory uraniferous mineral according to claim 1 , wherein the oxidant is SO₂/O₂, sodium chlorate, manganese dioxide, peroxide, or Caro' s acid.

5. A process for leaching uranium from a refractory uraniferous mineral according to claim 1 , wherein the oxidant is S0₂/0₂.

6. The process for leaching uranium from a

refractory uraniferous mineral according to claim 6 wherein the total dissolved iron content is greater than 5 g/L, preferably greater than 10 g/L, more preferably greater than 15 g/L.

7. The process for leaching uranium from a

refractory uraniferous mineral according to claim 7 , wherein the total dissolved iron content is less than 100 g/L, preferably less than 70 g/L more preferably less than 60 g/L.

8. A process for leaching uranium from a refractory uraniferous mineral according to claim 1 , wherein

oxidation of ferrous ions to ferric ions is effected by treating the slurry with one or more inlet gases

containing sulfur dioxide and oxygen in order to oxidise said ferrous ions to ferric ions , wherein the delivery rate of the sulfur dioxide gas is oxidation rate limiting; and

controlling the concentration of dissolved oxygen in said slurry at an optimum at which there is an excess of oxygen dissolved in said slurry.

9. A process for leaching uranium from a refractory uraniferous mineral according to claim 1, wherein the refractory uraniferous mineral is included in an ore or an ore concentrate or flotation tailings , which optionally further includes one or more non-refractory uranium mineral/s and/or gangue mineral/s.

10. A process for leaching uranium from a refractory uraniferous mineral according to claim 10, wherein the non refractory mineral is uraninite.

11. A process for leaching uranium from a refractory uraniferous mineral according to claim 1 , wherein the ferrous ions are derived from one or more of iron

containing gangue mineral/s and hydrometallurgical process solutions .

12. A process for leaching uranium from a refractory uraniferous mineral according to claim 1 , wherein the process is low acid consuming.

13. A process for leaching uranium from a refractory uraniferous mineral according to claim 13, wherein the slurry pH and temperature are adjusted such as to promote the precipitation of jarosite and thereby release

additional sulfuric acid for use in further leaching of the refractory uraniferous mineral .

14. A process for leaching uranium from a refractory uraniferous mineral according to claim 5, wherein the solids density in the slurry is controlled at a first value which is sufficient to facilitate dispersion and solubilisation of said oxidant throughout the slurry.

15. A process for leaching uranium from a refractory uraniferous mineral according to claim 14, wherein said first value is less than 55wt% , preferably less than 45 wt% , more preferably less than 40wt%, more preferably less than 30 wt% , more preferably less than 25 wt% , such as about 20 wt%.

16. A process for leaching uranium from a refractory uraniferous mineral according to claim 14, further

including subsequently treating the slurry in a second leaching stage at a second value of solids density which is higher than said first value.