The present invention relates to a process of ferric leaching of metal ores and concentrates.

Uranium may be extracted from its ores by acid leaching, typically by sulphuric acid leach. The acid leaching process is made more effective by addition of a strong oxidising agent, particularly where the ore is refractory. Refractory ores contain acid consuming minerals, typically phosphate and carbonate minerals which neutralise the acid so increasing acid consumption. So, while a simple acid leach of black shales containing uranium (uraniferous shale) has been practised with 75% uranium recovery, the acid consumption was 130 kg phosphoric acid per tonne of uraniferous shale.

This problem is also seen with acidic ferric leaching of uranium ores. It is generally accepted that the use of ferric iron as a leaching medium is limited to ores which have a relatively low acid consumption, otherwise the quantity of acid needed to maintain an acidic environment becomes uneconomic.

The acid consumption of an ore is often dictated by the amount of carbonates or phosphate minerals present. Ores which are high in these types of acid consuming minerals are normally limited to leaching using a more expensive alkali medium such as sodium carbonate. Sometimes the use of alkali leaching methods results in poor extraction of elements in comparison to the use of acidic leaching. That is, elements such as radium, vanadium, nickel and molybdenum do not form readily soluble carbonates. Leach recoveries are therefore typically low.

Although the above discussion is focussed on uranium extraction, similar challenges of acid consumption are faced in the leaching of other metal ore types, for example laterites, clay like minerals which contain valuable metals nickel and cobalt. Ores containing other elements of economic interest may also be subject to high processing costs where a substantial proportion of the ore is comprised of acid consuming minerals. Such ores are more frequently considered for processing as metal grades fall. Grades for elements such as uranium and rare earth elements in deposits of potential economic interest may often be in the tens or hundreds of parts per million range.

A process which could be economically applied to the leaching of high acid consuming ores, especially complex ores, would be advantageous. However, such process would also be expected to be applicable to ores with lesser proportions of acid consuming minerals.

With this object in view, the present invention provides a process of leaching a metal containing ore comprising the steps of forming an acidic solution containing ferric iron; and contacting the acidic solution containing ferric iron with the metal containing ore to extract elements by oxidative leaching with ferric iron wherein ferric iron is precipitated to generate acid for the leaching step. In such process, ferric iron has a dual role acting 1 ) as a leaching reagent; and 2) providing acid by precipitation to supplement an external acid addition to the leaching step. Therefore, acid should not be applied in excess to avoid precipitation, rather precipitation is to be encouraged.

The chemistry of ferric iron precipitation proceeds as follows:

Fe ₂ (SO ₄ ) ₃ + 6H ₂ O <→ 2Fe(OH) ₃ + 3H ₂ SO ₄

Though a substantial proportion of the ferric hydroxide is soluble in mineral acid, especially sulphuric acid, allowing regeneration of ferric iron for use in ferric leaching, some ferric iron is precipitated as goethite or goethite type precipitates from which ferric iron regeneration may not be practicable.

Acid, being a mineral acid selected from the group consisting of sulphuric, hydrochloric and nitric acids, present in the acidic ferric iron containing solution for the metal leaching step may comprise a portion derived from any external addition to the process and a portion generated, internally, from ferric ion precipitation. Any external addition of mineral acid may be provided only to act as a carrier for the ferric iron which is the dominant reactant for leaching. The ratio of portion of acid contributed by external addition to portion of acid contributed by internal generation through ferric iron precipitation would be decided on a project specific basis according to the economics of iron and acid provision.

Metal ore should be comminuted before leaching in a tank or vat leach. The ore is conveniently, from economic perspective, in un-beneficiated state for leaching. However, beneficiation may be conducted prior to leaching. In such case, a beneficiation step comprising flotation or gravity separation may be used for mineral concentration. So, where the ore contains organic carbon, particularly graphite, with which valuable metal minerals are associated, flotation allows more effective recovery and leaching for metal recovery from these minerals. Alternative beneficiation steps could be implemented.

The process is particularly advantageous for application to ores containing a substantial proportion of acid consuming minerals selected from the group consisting of carbonate minerals and phosphate minerals. This is because acid consuming minerals have previously required higher acid consumption and higher acid reagent cost. Generation of acid by ferric iron precipitation reduces these acid reagent costs, through neutralisation of such minerals. The process has, in particular, been shown as effective in treating complex ores containing substantial carbonate and phosphate content. Phosphate minerals, such as apatite and fluorapatite, often contain uranium and other metals including vanadium and molybdenum as well as rare earth elements. However, ferric leaching of such minerals has previously been found less advantageous than acid or alkaline leaching processes. The process also allows recovery of phosphorus from acid consuming phosphates.

Where the ore contains both apatite and carbonate minerals, these minerals may be separable, for example by flotation, to enable more effective metal recovery in the leach step. In such case, the susceptibility - to flotation - of phosphate minerals is exploited.

The ore may contain uranium in the form of uraninite, coffinite, brannerite or other uranium minerals. Such uranium may or may not be in association with radium, an element which is difficult to remove by alkaline or acid leaching because of the insolubility of radium sulphate or radium carbonate in comparison with uranyl sulphate or uranyl carbonate.

The process may advantageously be applied to extraction of rare earths from ores whether or not such ores contain a substantial proportion of acid consuming minerals in the form of carbonates and phosphates, particularly phosphates in the form of apatite minerals. The Applicant has shown that rare earth elements (including lanthanum (La), cerium (Ce), neodymium (Nd), yttrium (Y), samarium (Sm), dysprosium (Dy), europium (Eu) and gadolinium (Gd)) may be extracted, depending on mineralogy using mildly acidic ferric medium at temperatures close to ambient, here meaning a temperature of less than 100°C, more preferably less than 80°C and most preferably less than 70°C. Higher temperature leaching is not precluded This offers significant advantage over prior processes where the leaching of rare earth elements normally requires the use of high strength acids or alkalis at high temperature. Rare earth minerals treatable by the process include REE fluoro-carbonate (bastnaesite), containing La, Ce and Nd; monazite containing Ce, La and Nd and Y-phosphate minerals.

The Applicant's process may also be applied to extraction of base metals, especially including nickel, zinc, cobalt and copper from ores containing these. Vanadium, rhenium and molybdenum may also be extracted using the process as may precious metals including silver.

Advantageously, and in order to optimise extraction of elements, the residue from the ferric leaching step is, following separation from pregnant solution, re-leached or washed in a mild acid solution, conveniently of 10% acid or less, to remobilize elements which have co-precipitated with ferric ion during the acidic ferric leach. Such a process may be described as a two step process. The acid used for the washing step may again be selected from the group consisting of sulphuric, hydrochloric and nitric acids. Conveniently, the same acid is used for washing as is present in the acidic ferric leaching solution as for washing as this reduces reagent inventory. Less acid may be required in the two step process than if only a single acidic ferric leaching step was to be used in which sufficient acid was added to inhibit precipitation of valuable elements in the form of jarosite, goethite and other refractory minerals.

The process of ferric leaching of a complex ore containing both phosphate and carbonate acid consuming minerals in accordance with embodiments of the present invention is now described with reference to one application and comparative examples and examples relating to that application. The following description is not intended to limit the scope of the invention. The process may be applied, for example, to treatment of metal bearing ores sourced from a mineralised zone ("the Berlin Deposit") containing both sandstone and carbonate rocks and material dominant in either rock type may be treated for metal extraction according to the above described process. The mineralised zone exposed in trenches typically has a tan-coloured clay-rich footwall containing several hundred ppm U ₃ O ₈ . The tan coloured layer is typically overlain by a mature, well-sorted fine to medium grained calcareous sandstone that contains variable amounts of interstitial bituminous hydrocarbons. This mature sandstone is consistently mineralised. The clean, well-sorted sandstone fines upward into a laminated mudstone-siltstone unit in which uranium grades typically drop to several tens of ppm U ₃ O ₈ . The sandstone may contain fracture fillings of green to pale blue to orange botryoidal coatings identified as variscite (AIPO ₄ .2H ₂ O) and childrenite ((Fe,Mn)AIPO ₄ (OH) ₂ H ₂ O). These are typically developed in the weathering environment from the primary phosphate minerals and occur within interstices in the sandstone. These interstices are also filled by apatite, Fe-AI-Ti-Cu-Ca-Cr-bearing phosphates, roscoelite (a V-Ba mica), Y- phosphates including churchite, monazite (REE-phosphate) and numerous U- bearing phosphates of the autunite (Ca(UO ₂ )(PO ₄ )2.10-12H ₂ O) and meta-autunite sub-groups.

Some sandstones consist of micro-mosaic brecciated quartz grains.

These are breccias, that if the matrix was removed, the fragments would fit together like a jig-saw puzzle to reconstitute the original grains. The matrix between these fragments consists of fine grained apatite, churchite, Fe-AI- phosphates +/-U, fine grained zircon, roscoelite and Fe-AI +/- Cu-Ti-V-Cr phosphates.

Other sandstone rock material for treatment in accordance with the process include coarse clastic domains dominated by quartz with interstitial domains of mica-apatite-chlorite-pyrite-rutile-zircon in contact with more apatite- rich domains which seal from general quartz, mica, iron-oxide, zircon, rutile and a host of metals. These apatite-rich domains host such metal minerals as silver poor tetrahedrite, Ni-S (millerite), pyrite (FeS ₂ ), sphalerite (ZnS) and fine grained U-Ti-bearing minerals such as brannerite. Sphalerite may be associated with organic carbon in the form of graphite.

Carbonate material for treatment according to the process is dominated by calcite (CaCOs) with lesser amounts of fluorapatite and quartz. Also present as minor phases detected by XRD are muscovite, dolomite, pyrite, chlorite and sphalerite. Table 1 below provides the relative mineral concentrations in the ore.

Nature of Mineralisation

Uranium present in materials to be treated from the Berlin deposit contains any or all of the three major uranium phases:

• Uraninite/pitchblende (UO2) which is likely to be most abundant of the uranium phases in the materials to be treated. It may occur as disseminated, very fine particles that typically range between 5μηη and 10μηη in size. Uraninite is closely associated with graphite though it also occurs at grain boundaries of calcite, apatite and quartz.

• Coffinite (U(SiO )i _-x (OH) _4x ) which may occur in material to be treated in small (<10μηη) particles, for example hosted by mica. Micaceous hosts for coffinite are often located in the interstices between calcite, apatite, quartz and sulphide grains. • Brannerite((U,Cu,Ca)(Ti,Fe)2O6) which may occur in multi-crystalline laths

Rare earth minerals treatable by the process may include REE fluoro- carbonate (bastnaesite), containing La, Ce and Nd; monazite containing Ca, La and Nd and Y-phosphate minerals.

In the material to be treated, bastnaesite may be present in quartz-mica- sulphide-bearing parts of carbonate rocks as above described; monazite particles may be hosted by quartz and calcite. The Y-phosphate may occur as inclusions within an alaskite matrix.

The metallurgical challenge was extracting a suite of the above identified economically valuable elements from a complex ore, a challenge magnified by the carbonate-rich nature of the majority of the host rock that leads to high acid consumption and potentially adverse economics despite good recoveries of uranium and other elements mentioned above. The following comparative examples demonstrate the challenge.

Example 1 (Comparative) Acid Leaching

Un-beneficiated carbonate rich ore, extracted from the mineralised zone described above, was crushed and subjected to acid leaching with baseline tests using sulphuric and hydrochloric acid. In all sulphuric acid leach tests, sodium chlorate was added to achieve a target oxidation reduction potential (ORP) of 500 mV when measured using a platinum electrode against a silver/silver chloride electrode.

8

Recoveries of uranium were moderate to good but with high to very high levels of acid consumption.

Alkaline leaching was tried as an alternative to acid leaching.

Example 2 (Comparative) Alkaline Leaching

The same ore as in Example 1 was subjected to aggressive alkaline leach using sodium carbonate. Pulp density of ore in alkaline leach solution was held at 2%.

Table 3

Recoveries of uranium were not statistically different (67% and 68% extractions) under the very different separation conditions (noting temperatures of 250°C and 90°C, temperatures requiring heating of the leach vessel). The recovery of vanadium, phosphate and nickel was poor. Molybdenum recovery was good especially under the 90°C conditions.

The comparative examples demonstrated high acid consumption for moderate metal recoveries (acid leaching) and suitability of alkaline leaching for only a small component of the suite of metals of economic interest present in the ore, an acidic ferric leaching process having flowsheet as shown in the Figure was adopted.

The flowsheet shows a leaching vessel 10 for leaching of unbeneficated crushed whole complex carbonate rich ore from the above described deposit; a solid/liquid separation stage 20; an acid washing stage 30; a further solid/liquid separation stage and metal recovery. An acidic leach solution of ferric sulphate (as source of ferric iron) was formed by dissolving ferric sulphate in water. The natural pH of the ferric sulphate solution is 1 .5 to 2.5 so addition of mineral acid to achieve required acidity may be avoided where ferric iron precipitation generates sufficient sulphuric acid. However, a portion of sulphuric acid could be contributed, by external addition, to act only as a carrier for ferric iron which is the dominant reactant for leaching. The ratio of the portion of acid contributed by any external addition to portion of acid contributed by internal generation through ferric iron precipitation would be decided on a project specific basis according to the economics of iron and acid provision.

The acidic ferric iron solution was introduced as stream 12 to leach stage 10, comprising one or more leach vessels, and crushed un-beneficiated carbonate-rich ore stream 13 from the deposit was leached in the leach vessel(s) 10 under agitation.

Ore residue was separated from the leach solution by filtration in solid/liquid separation stage 20.

Pregnant solution 14 from leach vessel 10 could be directed to metal recovery stage 40 as shown in the Figure. However, the ore residue still contains appreciable quantities of metals of economic interest after the first leach stage involving acidic ferric leach solution conducted in leach vessel 10.

Therefore, an acid washing or "re-leach" stage 30 is provided as a second leach stage. Acid washing involved contacting of the ore residue separated in solid/liquid separation stage 20 with a dilute or weak mineral acid solution, for example of sulphuric or hydrochloric acid.

Pregnant solution 36 is directed to metal recovery after a solid-liquid separation stage 33. Again, filtration is suitable for this solid/liquid separation stage 33. Leached ore residue 34 may then be disposed of.

The pregnant solution 36 from acid washing stage 30 was analysed for metal extractions from the two step ore leaching process. Solvent extraction would likely be adopted for metals recovery in metal recovery stage 40 as such process is very suitable for recovery of uranium and rare earth elements. Other processes could be used for recovery of vanadium, nickel, phosphorus and molybdenum (as well as other metals present in leach solution in economic quantities).

The following example provides more detail about one embodiment of the above described leach process.

Example 3 Ferric Leaching

The same complex ore containing uranium, as subjected to alkaline and acid leaching in the Comparative Examples, was subjected to mild acidic ferric leaching of ore with acidic ferric sulphate solution at atmospheric pressure and a temperature of 65°C for 48 hours. The natural acidity of a ferric sulphate solution (pH 1 .5 to 2.5) was relied on and no external addition of sulphuric acid was made in the ferric leaching stage. Ferric iron consumption of 84 kg/t ore was indicated through the testwork.

Residues from the acidic ferric leaching were then subjected to "re- leaching" or washing using a 10% solution of either hydrochloric acid or sulphuric acid to solubilise elements co-precipitated with ferric compounds (as ferric hydroxide or/and hydronium jarosite or/and goethite) in the acidic ferric leaching stage. Acid consumption, in the "re-leaching" or washing step, of 125 kg/t ore was indicated through the testwork.

The average extractions of elements from nine batch tests together with

the extractions obtained for a suite of metals are shown in Tables 4 and 5. Table 4 - Extraction of Major Elements from Berlin Ore using acidic ferric leaching

Silver extractions only for residues subjected to weak Hydrochloric acid re-leaching.

The extractions of lesser elements from the 7 of the 9 tests which still may

be of economic importance are summarized below in Table 5.

Table 5 - Extraction of Other Elements from Berlin Ore using acidic ferric leaching

The dominant role of the ferric iron as reactant and the simple secondary role of the acid as carrier is well illustrated by comparison of metal recoveries from the acid leach example with those of the acidic ferric leach provided above. The acidic ferric leach results show higher extractions for uranium as well as a broad suite of elements than the acid leach results (see Example 1 (Comparative)). In particular, the two step leach process provides good to excellent recoveries for uranium (93-99%, average 97%) and phosphate (92- 99%, average 97%) as well as excellent to acceptable recovery for a range of metals of potential economic interest including vanadium (56-82%, average 74%), yttrium (80-96%, average 92%), neodymium (49-95%, average 74%), zinc (64- 100%, average 94%), nickel (50-77%, average 63%), molybdenum (43-61 %, average 52%) and rhenium (1 1 -71 %, average 41 %). The leach parameters investigated were particle size; pulp density; temperature; and ferric iron concentration.

Particle size

Although three different particle sizes were investigated (106μηη; 75μηη; 38μηη) some of the superior results were obtained from those tests using the larger size range of 106um. Although more extensive tests are required to verify the influence of particle size these results are promising in suggesting that an acceptable grind size associated with normal milling operations may be sufficient to liberate elements and expose them sufficiently for leaching.

Temperature

The influence of temperature was noted from the test work, with superior extractions generally obtained at 65°C compared to 40°C. These temperatures are close to ambient and relatively easy to obtain cost effectively at commercial scale. All tests were conducted at atmospheric pressure and no noxious gases or fumes were generated.

Ferric Iron

Ferric iron concentrations of either 25 g/l or 50 g/l were tested with both achieving comparable metal recoveries. Ferric iron consumption may be reduced by using higher pulp densities.

Acid Wash Step

It would be expected that weak sulphuric acid, in concentration about 2% to 10%, would be used as the medium of choice for the acid wash step but hydrochloric acid was also investigated. The concentrations used were 10%. In general terms a greater variety of elements and marginally higher total extractions were obtained in the presence of chloride ions in comparison to sulphate. Although this is to be expected, given typically greater chloride solubility, it does not necessarily imply that a chloride route would be commercially superior. This is largely due to higher reagent costs and potential difficulties in selective recovery of elements from a chloride medium where chlorine gas handling may be required.

Less acid may be required in the two step process than if only a single acidic ferric leaching step was to be used in which sufficient acid was added to inhibit precipitation of valuable elements in the form of jarosite and other refractory minerals such as goethite, if this forms above 60°C.

Modifications and variations to the method of ferric leaching of an ore in accordance with the present invention may be apparent to the skilled reader of this disclosure. Such modifications and variations are deemed within the scope of the present invention.