This invention relates to the extraction of uranium, particularly from its ores, by an acidic or alkaline leaching process route. Background of the Invention.

Two leaching processes are well known for the extraction of uranium from its ores. The first such process is the alkaline or carbonate leaching process and the second such process is the acid extraction process.

The alkaline leaching process has been described by A.R. Burkin in "Extractive Metallurgy of Uranium" and proceeds according to the following route:

U0 ₃ + 3Na ₂ C0 ₃ + H ₂ 0 → Na ₄ U0 ₂ (C0 ₃ ) ₃ + 2NaOH (2)

2Na ₄ U0 ₂ (C0 ₃ ) ₃ + 6NaOH → Na ₂ U ₂ 0 ₇ + 6Na ₂ C0 ₃ + 3H ₂ 0 (3) This process is very selective and is therefore desirably adopted in cases where the uranium ore has high acid consumption. Uranium is recovered from the process by precipitation of the sodium diuranate which is filtered and dried and exported for further processing.

However, notwithstanding the advantages of the alkaline leach process from the point of view of selectivity and corrosion minimisation, it generally has a major cost disadvantage relative to the acidic leach process in that fine grinding of the ore is required. Furthermore, carbonate reagents are expensive.

For this reason, the dominant processing route in the world today is the acidic processing route which proceeds as follows: Fe + H ₂ S0 ₄ → Fe≥+ + S0 2- + H ₂ (4)

Mn0 ₂ + 2H ₂ S0 ₄ + 2Fe≥+ → 2Fe3+ + Mn2+ + 2H ₂ 0 + 2S0 2- (5)

U0 ₂ + 2Fe3+ → U0 ₂ 2+ + Fe2+ (6)

U0 ₃ + 2H+ → U0 ₂ 2+ + H ₂ 0 (7)

In this route, the uranium oxide, in the case of oxidic ores such as uraninite and pitchblende, is attacked by ferric ions which are generated by

oxidation of ferrous ions generated by dissolution of iron present in the uranium ore by the acid, usually sulphuric acid though other mineral acids such as nitric and hydrochloric acids may be suitable for the purpose. A further source of iron is the grinding media which grind the uranium ore to a suitable size for leaching. Such attack converts U(IV) ions present in the ore to U(VI) ions in soluble uranyl cation (U0 ₂ 2+) form which is amenable to recovery by solvent extraction or other suitable processing steps.

At the present time, the oxidation of ferrous ions to ferric ions is undertaken by the addition of pyrolusite (Mn0 ₂ ), sodium chlorate or other metallic compound oxidants to the leach circuit. The introduction of such oxidants presents real difficulties. Firstly, pyrolusite may be expensive or require to be fine ground to ensure process efficiency. Therefore uranium processing plants may be required to pay high costs for pyrolusite or provide capital intensive grinding plant for the purpose of grinding. At some uranium mines, grinding to 80% passing 0.074 mm is required. Such grinding is undoubtedly expensive and if it could be avoided a great benefit to uranium producers could be achieved. Secondly, pyrolusite introduces manganese or a non-payable element to the leach process. In addition to being non-payable, manganese is subject to undesirable side reactions which reduce process efficiency and which may affect the efficiency of a solvent extraction process for recovery of uranium. In the case of sodium chlorate, expense due to the occurrence of undesirable side reactions are reason enough to avoid use of this compound. In particular, the chlorate ion tends to degrade to chloride ions which may attack electrodes used for the electrowinning of copper or other metals downstream of the leach circuit.

In spite of the adverse nature of oxidants presently used by the industry, there is a persistence with them because other oxidants either do not promote the required conversion of tetravalent uranium to hexavalent uranium or necessitate very expensive equipment to enable their use.

Summary of the Invention

It is therefore the object of the present invention to provide a process for extraction of uranium which enables the use of economic oxidants heretofore not known. With this object in view, the present invention provides, in a first aspect, a process for the extraction of uranium from a uranium containing material by a leaching process comprising dissolution of said material, said dissolution being catalysed by a catalytic agent.

The leaching process may involve, for example, acid leaching by an acid such as sulphuric acid or Caro's acid; or alkaline leaching, by an alkaline agent, for example, alkaline carbonate.

In a second aspect, the present invention provides a process for the extraction of uranium from a uranium containing material by an acid leaching process comprising dissolution of a uranium mineral present in said uranium containing material by ferric ions wherein said ferric ions are generated by oxidation of ferrous ions at least a portion of which are formed by reaction of acid with iron or an iron containing mineral present in said uranium containing material, said oxidation being catalysed by a catalytic agent to thereby enable achievement of an effective ratio of said ferric to ferrous ions to provide substantial dissolution of said uranium mineral.

Adsorbent catalytic agents such as those based on carbon.e.g activated carbon, and ion exchange resins are especially convenient, readily obtainable catalytic agents for use in the process . Other catalytic agents may also be employed. It is to be observed that, whereas contaminant metal ions are difficult and/or expensive to remove from solutions, extraction of solid catalytic agents, such as adsorbent activated carbon, is as straightforward as physical separation; for example by screening or filtration.

In a preferred embodiment, gaseous oxidants which are more economic and less detrimental to process efficiency than chlorate and pyrolusite, but presently not employed by the uranium industry, such as air, oxygen, ozone or an elemental oxygen containing gas having oxidising properties may be utilised under ambient pressure conditions in the presence of an adsorbent catalytic

agent such as activated carbon. These gases are advantageous in that conventional gas delivery equipment may be used to supply them to a leaching vessel.

Such equipment has a substantial cost advantage over grinding equipment and operating expenses compare favourably with expenses for conventionally used oxidants. Importantly, there is little or no addition of contaminant metal ions, such as manganese, chlorate or chloride to the leaching process and acid consumption, in the case of acid leaching processes, may be reduced by the avoidance of undesirable side dissolution reactions and absorption caused by addition of conventional metallic oxidants.

In addition, where gaseous oxidants are used, conventional gas supply equipment which requires less maintenance than specialised grinding or chemical storage facilities may be employed, reducing capital costs.

Other oxidants that could be used are solid oxidants selected, for example, from manganese dioxide, permanganates, peroxides, chlorates, chlorites, hypochlorites, chromates, dichromates and persulphates. Alkali metal, such as sodium or potassium, salts may be especially preferred.

It should be noted that the present invention is most effective within a certain range of oxidation-reduction potential (ORP). Use of solid oxidants may not allow the required ORP range to be rapidly attained so gaseous oxidants, which typically allow such rapid attainment, are to be preferred. A further advantage that may accrue is better control over the leaching process as lag in attaining desired ORP, which is likely to be encountered with solid oxidants, may be avoided using gaseous oxidants. Both types of oxidant may be employed together with some advantage, if desired.

If a solid oxidant is employed, this may be introduced directly to the leaching stage via the leachant or feed, optionally in the form of an aqueous solution.

Other sources of ferrous ions may be used than iron minerals present within the uranium ore. For example, iron or iron compounds may be added to the leach solution though this may be found uneconomic and it is therefore

preferred that sufficient iron to practice the process is available within the ore to be treated.

Detailed Description of the Invention

The invention will be better understood from the following detailed description of a preferred embodiment thereof made with reference to . the drawing in which:

Figure 1 illustrates a flowsheet for an acid leaching process for the extraction of uranium;

Referring now to Figure 1 , a uranium ore comprising, for example, uraninite (having a theoretical formula U0 ₂ ) is transported from the pit to a primary gyratory crusher from which ore to be treated is recovered after screening as undersize and is stored in a fine ore bin. The oversize ore is recycled to the crushing step.

Further comminution involves grinding of the fine ore in a rod mill with admixture of water. Cyclone separation is then undertaken to separate an overflow which is thickened in a thickener, the underflow of which is sent to the leaching step. Cyclone underflow is fed to a ball mill for further size reduction.

Leaching takes place in tanks where the ore is agitated with sulphuric acid and an oxidant in the presence of an adsorbent catalytic agent which catalyses the oxidation of ferrous ions formed by acid attack on iron mineral(s) present in the ore, iron compounds available in the leach solution or iron introduced by the grinding media to ferric ions which oxidise the uranium (IV) oxide to the hexavalent uranyl cation (U0 ₂ ²⁺ ) which is soluble in acid solution.

In a conventional acid leaching plant, the oxidant is manganese dioxide in pure or mineral (pyrolusite) form. Otherwise the oxidant generally employed is sodium chlorate. Air or oxygen have not been employed as oxidants because their slow rates of dissolution in acid solution prevent a substantial degree of oxidation of ferrous to ferric ions and hence oxidation of uranium (IV) to uranium (VI). This situation may be changed in accordance with the inventive process by addition of a catalytic adsorbent agent, conveniently activated carbon, though

any resin or adsorbent having sufficient reactivity may also be employed. The addition of the adsorbent catalytic agent ,in an appropriate quantum, catalyses oxidation of ferrous ions to ferric ions by enabling a sufficient rate of availability in acid solution of oxygen sourced from oxidants as above discussed and particularly oxidising gases such as air, oxygen, ozone, elemental oxygen containing gases or mixtures of these gases that an oxidation reduction potential of at least +300 mV is attained in the acid leaching process and economically viable uranium extraction rates are obtained.

The most desirable quantum of activated carbon addition appears to be 10 to 200 g C/kg ore, preferably 14 to 70 g/kg ore, but other additions may be suitable depending on ore type and plant operating conditions. Other kinds of adsorbents may require to be added in different amounts.

However, conditions of pH and temperature may affect the efficiency of the activated carbon or other agent in catalysing the leaching process. With respect to pH, a range of 0.7 to 2 is preferable. Below pH 0.7 carbon may be degraded. Above pH 2, base metals such as copper, lead and zinc; and silica dissolve, possibly with adverse effects.

With respect to temperature, the temperature may be maintained, with acceptable extraction, at 30°C or even lower. Heating to higher temperatures, especially say 60°C or above, in the leach tanks by addition of live steam or use of other heating arrangements may be employed to achieve even more favourable kinetics.

In accompanying Table 1 , it can be clearly seen that the addition of 25 g/|_ carbon to an acidic solution containing 2 9/ total iron enables attainment of a superior oxidising environment measured in terms of Eh and ferrous/ferric ion conversion over time when oxygen is introduced at a rate of 6 l/min to the unsatisfactory situation when oxygen is introduced at the same rate to the leach solution in the presence of a small quantity or substantial absence of a catalyst.

TABLE 1 Comparison of Ferrous to Ferric Ion Conversion at 1.25 g/L and 25g/L Additions of Activated Carbon and 6 I /min Oxygen as Oxidant

1.25g/L Activated Carbon 25g/L Activated Carbon Time (Hrs) [Fe(ll)] % conversion Eh fFe(ll)] % conversion En

(mg/L) tFeπn→FeflllY ) (mg/L) 'Feπn→Feπim

0 2000 0 563 2000 0 563

1 2000 0 578 1714 14.3 628

2 2000 0 598 1350 32.5 663 3 2000 0 599 1036 48.2 668

4 2000 0 608 821 58.95 679

20.5 1930 3.5 649 50 97.5 789

It is to be observed, in this regard, that the conversion of uranium (IV) to uranium (VI) cannot occur at an economic rate if the required rate and extent of conversion of ferrous to ferric ions does not occur. This is dependent upon rate of dissolution and availability of oxidant.

A further advantage that may occur using the present process is that, using pyrolusite, manganese dioxide or sodium chlorate, there may occur a period during leaching in which there is no value in introducing these reagents. This period corresponds with the time when ferrous ions commence coming into solution as a result of acid attack of iron, pyrite or other iron containing minerals in the uranium ore. In this period, the effect of these metallic compound oxidants on oxidation of uranium (IV) is negligible and possibly counter-productive as they may react with components such as hydrogen sulphide and hydrogen evolved during the acid leach. Clearly, such reactions would cause excessive reagent consumption. This should not be a significant problem if gaseous oxidants, as particularly preferred in the present invention, are employed. Therefore, these gaseous oxidants can, in the presence of a catalyst, be introduced from commencement of the leach process thereby achieving a small degree of oxidation of uranium (IV) to uranium (VI) in a period where conventionally used metallic compound oxidants are ineffective. To this extent

the catalytic oxidation reaction described above may proceed in parallel with direct oxidation of uranium (IV) but independently thereof.

Following the leaching stage, the barren ore and adsorbent catalytic agent, e.g activated carbon, may be separated from the pregnant liquor. Conveniently, such separation is achieved by a multiple-stage counter-current decantation process employing several thickeners. The underflow constituents barren ore and activated carbon are separated with the barren ore usually being neutralised and disposed of or recycled. The activated carbon is recycled to the leaching stage though it may require regeneration or treatment to remove adsorbed species before recycle. The overflow from the thickeners is clarified by sand filtration to ensure that suspended solids are prevented from entering the uranium recovery circuit. in another embodiment, the activated carbon may be retained in the leach tanks by screens with occasional regeneration as required. If appropriate, additional stripping stages to recover species adsorbed onto the activated carbon may be conducted. Of course, activated carbon may be substituted by other catalytic agents such as ion exchange resins, if desired.

The clarification of pregnant liquor is especially relevant in the case of uranium recovery by solvent extraction or resin ion exchange. In the case of resin ion exchange, strong base anion exchange resins are used to adsorb anionic uranium complexes which exclude metal cations.

Solvent extraction is also used to treat clarified acid liquors. Typically, the pregnant liquor is passed through a series of mixture/settler units in which the pregnant liquor is contacted with an organic solvent, as for example referred to in A.R. Burkin, "Extractive Metallurgy of Uranium", such as an amine being 5% Alamine 336 and 2% isodecanol. The process there described involved four stages and enables recovery of a uranium strip liquor grading 3-4 g/|_ U ₃ 0s. This uranium is recovered from the strip liquor by precipitation with ammonia to form ammonium diuranate. The precipitation reaction proceeds as follows:

2UO ₂ S0 ₄ + 6NH ₄ OH → (NH ₄ ) ₂ U ₂ 0 ₇ i + 2 (NH ₄ ) ₂ S0 ₄ + 3H ₂ 0(8).

This precipitate is then thickened, washed and dewatered by calcination to obtain a uranium product grading at least 90% by weight U ₃ 0 ₈ .

In another embodiment of the present invention, a uranium-rich ore or concentrate which contains precious metals values such as gold or silver may also be treated. In such a case an acid leaching stage would be preferred as the first stage of the process in which uranium is extracted. In the second stage, a leaching e.g cyanidation process is employed to enable extraction or precious metals from the pulp obtained from acid leaching the ore or concentrate. In cases where such an ore or concentrate contains cobalt such a process route is preferred to avoid formation of cobalticyanide ions which interfere with ion exchange recovery processes.

The invention will be better understood from the following description of an example of an embodiment thereof. Example Copper/uranium flotation tailings-grading 0.8-0.9 kg/t U ₃ 0 ₈ in the form of

55% uraninite (uranium oxide), 25-30% uraninite disseminated in haematite and sulphide minerals and 15 to 20 % coffinite and brannerite, the latter minerals being regarded as generally refractory to leaching-was treated. In this regard, the complex brannerite [(U,Ca,Fe,Th,Y)(Ti _I Fe) ₂ θ ₆ ] is regarded as unleachable and silicate coffinite [U(Si0 ₄ )ι- _x (OH) _x ] dissolves only slowly with an estimated 33% uranium in this mineral estimated to be leachable. The occurrence of these minerals reduces the probable extractable uranium in the tailings to 65-70% of total uranium.

A sufficient quantity of tailings was slurried, by agitation, in water to produce a slurry grading 50% by weight flotation tailings. Sufficient sulphuric acid solution was then added to the pulp to achieve a pH of 1.5. Leaching took place at 30°C and 60°C, the temperature being maintained constant for the duration of the leaching process.

Comparative leaching tests were conducted for the following instances:

(a) Sufficient, approximately 2 kg/t tailings, industrial grade sodium chlorate (NaCI0 ₃ ) was added to the pulp to establish an oxidation-

reduction potential (ORP) at leach commencement of greater than +300 mV.

(b) Carbon, in the form of activated carbon, and oxygen oxidant (to 5 substitute chlorate) were introduced to the pulp at additions 14, 35 and 70 g C/kg ore and 1.5 l/min respectively.

Oxygen was introduced to the leaching vessel through a single nozzle located just below the impeller. Ideally, oxygen or other gases are to be 10 introduced in fine dispersion through sparging or similar operation. Rate of addition may be controlled to achieve a desired ORP.

The leaching test was conducted for 24 hours with samples being taken for analysis at conclusion of that period. The results are provided in Table 2 below:

15 TABLE 2 Leaching of Copper/Uranium Flotation Tailings

Uranium Recovery (% recoverable U by weight)

20 24 hr Leach Recovery Acid Cons iumption (kg/t)

(as 100% H ₂ S0 ₄ )

Conditions Temp (°C) Temp (°C)

30°C 60°C 30°C 60°C

NaCIOs (2kg/t) 81.2 75.5 7.71 8.04

25 0 ₂ with C (g/kg)

0 ₂ /14 T=35°C T = 35°C

63.3 78.7 6.22 13.54

0 ₂ /35 T=35°C T = 35°C

68.8 77.6 7.28 1 1.94

30 O ₂ 70 80.3 89.4 9.20 14.08

The uranium recovery with chlorate is comparable with that at carbon addition of 14 g C/kg ore at oxygen input 1.5 l/min. However, this effect is achieved without the disadvantages encountered in using chlorate or pyrolusite as described above. Better control over ORP may also be achieved. The present invention is not limited in its application to a uranium extraction process as described above and modifications may be developed by those skilled in the art. For example, the invention is equally applicable to alkaline leaching uranium extraction schemes. Such modifications fall within the scope of the present invention.