AU2014250661B2 - Process for the Recovery of Rare Earth Elements - Google Patents

Abstract

A process for recovering rare earth elements from a composite ore or concentrate that includes at least apatite group minerals, as well as at least one other rare earth containing mineral, the process including: a. subjecting the ore or concentrate to a sulphuric acid pre-leach to form a pre leach liquid phase and a pre-leach solid phase, the pre-leach liquid phase containing dissolved phosphates and dissolved rare earth elements, and the pre leach solid phase containing sulphuric acid-insoluble minerals and calcium sulphate precipitate; b. mixing the pre-leach solid phase with sulphuric acid and baking the mixture at a temperature in the range of 200 to 3000C to produce an acid-baked solid phase; and c. subjecting the acid-baked solid phase to a water leach, in water and/or in the pre-leach liquid phase, to dissolve rare earth elements in the acid-baked solid phase to form a water-leach liquid phase and a water-leach solid phase; wherein the water-leach liquid phase is at least partly recycled to the pre-leach and rare earth elements are recovered from the pre-leach liquid phase; and/or the water-leach liquid phase is not recycled to the pre-leach and rare earth elements are recovered from the water-leach liquid phase. I - on;> Ufr.~ I I 1My V1,

Description

PROCESS FOR THE RECOVERY OF RARE EARTH ELEMENTS

TECHNICAL FIELD [0001] The present invention relates generally to the recovery of rare earth elements from rare earth ores or concentrates containing those elements. More particularly, the invention relates to the recovery of rare earth elements from a rare earth composite ore or concentrate of rare earth-containing minerals that includes at least apatite group minerals, as well as other rare earth-containing minerals such as monazite and/or allanite group minerals. Such a composite ore or concentrate may also contain other rare earth phosphate and carbonate minerals, and/or rare earth silicate minerals.

BACKGROUND OF INVENTION [0002] Rare earth elements include all the lanthanide elements (lanthanum, cerium, praseodymium, neodymium, promethium, samarium, europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium, ytterbium, and lutetium), as well as the rare metals scandium and yttrium. For ease of discussion and because of their abundance and similar properties, the Total Rare Earth Elements (TREE) are often divided into three groups; the Light Rare Earths (LRE) which are lanthanum, cerium, praseodymium, and neodymium; the Middle Rare Earths (MRE) which are samarium, europium and gadolinium (promethium does not exist as a stable element); and the Heavy Rare Earths (HRE) which are terbium, dysprosium, holmium, erbium, thulium, ytterbium, and lutetium. Yttrium and Scandium are often added to this list although they are not strictly heavy rare earth elements.

[0003] During the past twenty years there has been an explosion in demand for many items that require rare earth elements, which now include many items that people use every day, such as computer memory, DVD's, rechargeable batteries, mobile phones, catalytic converters, small electric motors, magnets, and fluorescent lighting.

[0004] Rare earth elements also play an essential role in electricity generation from wind power, new generation electric vehicles, and military applications. In this

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2014250661 16 Oct 2014 respect, the military uses include night-vision goggles, precision-guided weapons, communications equipment, GPS equipment and batteries.

[0005] The increase in use of rare earth elements in new technology devices has lead to an increase in demand, and a need for diversification in a supply chain that has been dominated by China since the early 1990s. Indeed, the development of non-Chinese resources for the mining and processing of rare earths has expanded in recent years, particularly since China announced in 2010 that it will severely restrict its export of rare earth elements to ensure supply for domestic manufacturing.

[0006] Rare earths are relatively abundant in the Earth's crust, but discovered mineable concentrations are less common than for most other ores. The world’s resources are contained primarily in bastnasite and monazite. China and the United States constitute the largest percentage of the world's known rare-earth economic resources, while monazite deposits in Australia, Brazil, China, India, Malaysia, South Africa, Sri Lanka, Thailand, and the United States constitute the second largest segment.

[0007] Conventional methods for the extraction of rare earth elements from their ores or concentrates are described in the book “Extractive Metallurgy of Rare Earths by C.K. Gupta and N. Krishnamurthy, CRC Press, 2005.

[0008] Typically, rare earth elements have been extracted from monazite ores, or concentrates thereof, by processes of acid baking or caustic fusion. In “acid baking” (sometimes called “sulphation”), the ore or concentrate is mixed with concentrated sulphuric acid and baked at high temperatures (e.g. from 200 to 500°C) to break down the mineral matrix and convert the rare earth elements into sulphate salts that can then be brought into solution by dissolution in a water leach of the baked solids. Once the rare earth elements are in solution, they can be recovered by a number of different techniques including (a) stepwise neutralisation to separate a rare earth hydroxide from impurities, (b) double sulphate precipitation, or (c) rare earth oxalate precipitation. Processes also include methods of recovery and separation of uranium and thorium, which also are present in these ores, from the rare earth elements themselves.

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2014250661 16 Oct 2014 [0009] In “caustic fusion”, the ore or concentrate is mixed and heated to near the boiling point (about 150°C) with a concentrated sodium hydroxide (60-70% caustic soda) solution to break down the rare earth element containing minerals. The phosphate content of the ore can then be solubilised in a caustic water leach and separated from the residue that contains the rare earth elements, uranium and thorium. The sodium phosphate solution is then evaporated to produce a sodium phosphate by-product. The residue can be solubilised in any mineral acid and treated by a range of processes including (a) selective precipitation, (b) solvent extraction or (c) double sulphate precipitation, to produce a mixed rare earth product and separate uranium and thorium by-products.

[0010] Gupta and Krishnamurthy (2005) also describe processes for the extraction of rare earth elements from the mineral bastnasite (a rare earth fluorocarbonate). In China, bastnasite concentrates are processed by heating with concentrated sulphuric acid to 500°C in a rotary kiln. The residue is then treated with water to dissolve the soluble rare earth sulphates, which are then precipitated as double sulphate salts using sodium chloride as a reagent. Such rare earth double sulphate salts are treated with sodium hydroxide to convert them into a rare earth hydroxide which is then dissolved in hydrochloric acid prior to separation of the rare earths using solvent extraction.

[0011] The increasing demand for rare earth elements has led to interest in tapping non-traditional ore sources, containing multiple mineral components. For example, rare earth elements can be found in significant concentrations in apatite (a calcium phosphate mineral) and allanite (a calcium aluminium (and/or iron) silicate mineral) deposits. Such deposits may also contain monazite group minerals. These composite mineral ores are typically not amenable to processing by the conventional techniques mentioned above and have required new processes for extraction.

[0012] An example of such a new process is that described in the applicant’s own United States patent 7,993,612 relating to composite rare earths. In the ‘612 patent, a pre-leach process for the dissolution of rare earth phosphates is adopted, followed by the precipitation (by the addition of ammonia) of the rare earth phosphates, prior to an acid bake (of the type described above) of both the precipitate and the pre-leach residue. The acid bake is followed by a water leach to produce a water leach liquor

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2014250661 16 Oct 2014 rich in rare earth elements, and the water leach liquor is subjected to various further treatments to remove uranium (solvent extraction) and thorium, iron and aluminium (neutralisation), leaving a post-neutralisation liquor ready for the precipitation of a final product of a rare earth carbonate.

[0013] In one form, the ‘612 patent suggests the use of a nitric acid pre-leach. An advantage of a nitric acid process is that the post-precipitation pre-leach liquor produced contains both nitrate and phosphate and, in the absence of undesirably high amounts of deleterious elements such as uranium or thorium, the pre-leach liquor can later be used for fertiliser production. In this form, the levels of nitric acid utilised in the pre-leach are recommended to be 110% stoichiometric with respect to calcium, and at concentrations in the range of 30 to 60 wt%.

[0014] In another form, the ‘612 patent suggests the adoption of a hydrochloric acid pre-leach. Indeed, in this form, the levels of hydrochloric acid utilised in the preleach are also recommended to be as high as 110% stoichiometric with respect to calcium, and at concentrations in the range of 10 to 40 wt%.

[0015] However, both nitric acid and hydrochloric acid are expensive reagents, and the by-products produced in the pre-leach (calcium nitrate or calcium chloride) thus need to be recovered and re-used. For example, the calcium nitrate is typically used in fertiliser production, if the impurity levels of uranium and/or thorium permit, while the calcium chloride is typically used to regenerate hydrochloric acid for re-use. However, a significant investment in equipment is required to build and operate these by-product processes, further increasing the overall cost to the process.

[0016] It is an aim of the present invention to provide a simplified and more economic process, in terms of both capital expenditure and operating costs, for the recovery of rare earth elements from a composite ore or concentrate of rare earth minerals.

[0017] The above discussion of background is included to explain the context of the present invention. It is not to be taken as an admission that any of the material referred to was published, known, or part of the common general knowledge (in any country) at the priority date of any one of the claims.

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2014250661 16 Oct 2014

SUMMARY OF INVENTION [0018] The present invention provides a process for recovering rare earth elements from a composite ore or concentrate that includes at least apatite group minerals, as well as at least one other rare earth containing mineral, the process including:

a. subjecting the ore or concentrate to a sulphuric acid pre-leach to form a pre-leach liquid phase and a pre-leach solid phase, the pre-leach liquid phase containing dissolved phosphates and dissolved rare earth elements, and the pre-leach solid phase containing sulphuric acid-insoluble minerals and calcium sulphate precipitate;

b. mixing the pre-leach solid phase with sulphuric acid and baking the mixture at a temperature in the range of 200 to 300°C to produce an acid-baked solid phase; and

c. subjecting the acid-baked solid phase to a water leach, in water and/or in the preleach liquid phase, to dissolve rare earth elements in the acid-baked solid phase to form a water-leach liquid phase and a water-leach solid phase; wherein the water-leach liquid phase is at least partly recycled to the pre-leach and rare earth elements are recovered from the pre-leach liquid phase; and/or the water-leach liquid phase is not recycled to the pre-leach and rare earth elements are recovered from the water-leach liquid phase.

[0019] It is to be understood that in the above paragraph, and in equivalent descriptions throughout this specification, the word “water” is to include “wash water” as well as “process water”.

[0020] In a first form of the invention, the process includes treating the ore or concentrate with weak sulphuric acid in a sulphuric acid pre-leach to substantially dissolve most of the apatite and some of the contained rare earth elements, as well as other impurities, into a pre-leach liquid phase, leaving the sulphuric acid-insoluble minerals in a pre-leach solid phase along with a calcium sulphate precipitate. The pre-leach solid phase is subsequently mixed with a concentrated sulphuric acid and the mixture is baked to produce an acid-baked solid phase.

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2014250661 16 Oct 2014 [0021] In this first form of the invention, the acid-baked solid phase is leached with water or the pre-leach liquid phase, or a mixture of both, in a water leach to dissolve the rare earth elements in the acid-baked solid phase into a water-leach liquid phase, from which the rare earth elements can be recovered.

[0022] In a second form of the invention, the water-leach liquid phase is at least partly recycled back to the pre-leach so that excess sulphuric acid available in this phase can be used to economically leach incoming rare earth concentrate. If required, and if the recycled water-leach liquid phase contains insufficient acid to dissolve the apatite in the incoming concentrate and retain sufficient acidity to keep the dissolved rare earths in solution, extra sulphuric acid may be added to the preleach to ensure maximum dissolution of the apatite and rare earth minerals.

[0023] However, a more effective use of acid would be to add all the required acid to the acid bake of stage (b) above to maximise the overall rare earth extraction, with surplus acid being recycled to the pre-leach of stage (a). Conversely, if the recycled water-leach liquid phase contains an excess of acid to what is required for the preleach, then only part of the water-leach liquor may be recycled, with the excess waterleach liquor being sent to the rare earth recovery stage.

[0024] In this second form of the invention, it is the pre-leach liquid phase, that includes soluble rare earths from both the water leach of the acid-baked solid phase and from the pre-leach itself, that can then be used for the recovery of the rare earth elements.

[0025] In both forms of the invention, the rare earth recovery techniques include, but are not limited to, direct solvent extraction, precipitation as a rare earth carbonate, rare earth hydroxide or rare earth oxalate salt, or precipitation as a sodium rare earth double sulphate salt. A preferred form of the invention, utilising precipitation as a sodium rare earth double sulphate salt, will be further described below.

[0026] Turning now to a general description of the overall process of the present invention, it will be appreciated that the process may be applied to the direct treatment of a, for example, composite apatite-monazite ore. Preferably, however, such an ore would be first beneficiated by known processes of crushing and grinding followed by flotation and/or magnetic separation or other known beneficiation

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2014250661 16 Oct 2014 processes to produce a rare earth concentrate from the ore, rejecting barren material as gangue. The process may then be applied to a rare earth concentrate. Indeed, while the above general summary of the invention refers to the use of the process on both ore and concentrate, the following description of the invention will tend to refer only to a concentrate, albeit merely for ease of reference.

[0027] The first stage in the process summarised above is the sulphuric acid leach, sometimes referred to as a “pre-leach” as it precedes another substantive operational stage. In a preferred form, the majority of the apatite in the concentrate, shown below as a fluorapatite mineral, is dissolved in the leach and converted to soluble phosphoric acid and hydrofluoric acid together with insoluble calcium sulphate, according to the following Equation (1):

Ca5(PO4)3F(s) + 5H2SO4 (aq) + xH₂O = 5CaSO4.xH₂O (s) + 3H3PO4 (aq) + HF (aq)

Apatite Sulphuric Acid Calcium Sulphate Phosphoric Hydrofluoric

Acid Acid [0028] With this in mind, the liquid phase formed from this pre-leach can thus be said to contain both dissolved phosphates and dissolved rare earth elements from the dissolution of acid-soluble rare earth minerals. In this respect, it will be understood by a skilled addressee that a reference to the majority of the apatite in the concentrate dissolving is a reference to most but not all of the apatite. The aim is to reduce the amount of acid-soluble calcium and phosphate in the pre-leach solid phase as these may interfere with the acid bake leading to less than economic extractions of the rare earths in the subsequent water leach.

[0029] It is difficult to determine how much acid-soluble calcium and phosphate remain in the pre-leach solid phase due to the presence of acid-insoluble calcium and phosphate minerals. In this regard, it will be appreciated that the addition of acid in the pre-leach should be sufficient to ensure that the majority of the apatite, an acid soluble mineral, in the concentrate has been dissolved.

[0030] The amount of sulphuric acid required for the pre-leach will depend on the mineralogy of the concentrate and in particular the amount of apatite in the concentrate. In this respect, apatite and rare earth elements associated with the apatite will dissolve during this pre-leach, as will other acid soluble minerals in the concentrate that will include iron and aluminium containing minerals. The acid

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2014250661 16 Oct 2014 addition will thus preferably be such as to maximise the dissolution of the apatite but minimise the dissolution of the other acid soluble minerals that contain impurity elements such as iron and aluminium.

[0031] An advantage of using sulphuric acid, or a sulphuric acid containing solution, in the pre-leach rather than other mineral acids, such as hydrochloric acid or nitric acid, is that sulphuric acid is much more economic to use than other mineral acids. It is also the acid used in the subsequent acid bake stage, meaning that only a single acid storage facility is required. It is advantageous, from an economic point of view as well as from a processing point of view, to use the same acid in the pre-leach. Also, because the sulphate from the acid is captured in the solid phase as insoluble calcium sulphate, and is eventually deposited in a tailings storage facility along with other process residues, there is no requirement for treatment of any soluble calcium salt that might be the by-product of the use of other mineral acids, such as the nitric and hydrochloric acids of the prior art.

[0032] In a preferred form, the concentrate is mixed in the pre-leach with a weak sulphuric acid of from 5% to 40% w/w, and preferably about 20% to 25% w/w. The concentration of the sulphuric acid is chosen so that the solids concentration in the acid-concentrate mixture in the pre-leach is from 10% to 40% w/w, and preferably around 25% w/w. However, the optimum solids concentration will depend on the components of the concentrate, and will ideally be adjusted to obtain efficient leaching of the apatite at the maximum solids concentration.

[0033] The amount of sulphuric acid required for the pre-leach should be at least sufficient to dissolve most of the apatite in the ore or concentrate. Preferably, the amount of sulphuric acid added, or present in the recycled water leach liquid phase, will be 100% to 150% of the stoichiometric requirement, and more preferably from 100% to 120%, for the dissolution of apatite according to the Equation (1).

[0034] Preferably, added sulphuric acid will be pre-mixed with water to the required concentration and cooled to ambient temperature prior to adding to the concentrate for the pre-leach. This assists to ensure that the temperature of the leach is as low as possible without introducing external means of cooling the concentrateacid mixture.

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2014250661 16 Oct 2014 [0035] The heat of reaction produced by the reaction of the sulphuric acid with the concentrate will produce a temperature rise in the reaction vessel. Under optimal conditions, the temperature of the reaction should not increase to a temperature greater than 80°C and more preferably to a temperature less than 60°C. This is to minimise the surplus sulphuric acid required to keep rare earths from precipitating as phosphates or sulphates. Another reason is that the amount of water associated with the calcium sulphate precipitate will depend on the reaction temperature, with the hemihydrate (CaSO₄.¹/2H2O) favoured at elevated temperature and gypsum (CaSO4.2H2O) favoured at low temperature. At higher temperatures, the calcium sulphate crystals become finer and more difficult to filter. In addition, significant coprecipitation of rare earth elements in calcium sulphate can occur during the formation of hemihydrate (CaSO4.¹/2H2O), while rare earth incorporation into gypsum (CaSO₄.2H₂O) is less.

[0036] In a preferred form, the pre-leach reaction between the added acid, or the water-leach liquid phase, and the apatite minerals in the concentrate will be quite fast, with the reaction time preferably being between 15 minutes and 60 minutes, and more preferably being about 30 minutes. In this respect, at longer reaction times, more slowly reacting minerals also dissolve increasing the quantities of impurities in the solution. It is preferred, therefore, to limit the reaction time to minimise the dissolution of the impurity minerals.

[0037] The concentration of the rare earth sulphate salts generated by the preleach (in the pre-leach liquid phase) will depend on the concentration of the rare earths in the concentrate, the relative abundance of dilute-acid soluble and refractory rare earth containing minerals, the sulphuric acid concentration in the pre-leach, the acid-concentrate mass ratio, and the temperature. Some of the rare earths dissolved from the apatite may be re-precipitated along with the calcium sulphate. However, this does not appear to have a significant effect on their later re-dissolution after the acid bake in the water leach stage.

[0038] Following the sulphuric acid pre-leach reaction, the pre-leach liquid phase containing the dissolved minerals is separated from the pre-leach solid phase, the solid phase containing leach residue and calcium sulphate precipitate, by any standard method of solid-liquid separation (with or without washing of the solid

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2014250661 16 Oct 2014 phase). Suitable methods, that will include the washing of the pre-leach solid phase, may be counter-current decantation or filtration including a wash cycle. In this respect, to assist in the water balance, wash water from this separation stage may optionally be recycled to the pre-leach and/or water leach stages.

[0039] Following this separation stage, the pre-leach solid phase is passed to the acid bake of step (b) above, while the pre-leach liquid phase, in the first form of the invention, is passed to the water leach of step (c) above, and in the second form of the invention, to the recovery of the dissolved rare earth elements.

[0040] Turning now to a general description of the acid bake of step (b) above, it will be understood that the purpose of the acid bake is to “crack” or break down the rare earth containing minerals that were not solubilised in the relatively weak sulphuric acid pre-leach. These minerals may include monazite group minerals and/or allanite, as well as other rare earth containing minerals the exact nature of which will depend on the mineralogy of the rare earth concentrate. During the acid bake, these rare earth minerals are converted to water-soluble sulphate salts as shown in the following Equation (2) for the dissolution of monazite, where the term “RE” refers to the rare earth elements:

2RE(PO₄) + 3H₂SO₄ = (RE)₂(SO₄)₃ + 2H₃PO₄

Monazite Sulphuric Acid RE Sulphate Phosphoric Acid [0041] The pre-leach solid phase is preferably mixed with a concentrated sulphuric acid solution, preferably greater than 90% w/w sulphuric acid, and more preferably 94% to 98% sulphuric acid, prior to the acid bake. The amount of sulphuric acid added to the pre-leach solid phase depends on the rare earth concentration in the residue that must be sulphated, and also the nature of the other minerals in the residue that can also react with the concentrated sulphuric acid. The acid addition is preferably at least the stoichiometric requirement for sulphation of the rare earth minerals, and more preferably well in excess of this amount, to drive the sulphation reaction to near completion. Preferably, the acid addition is from 400 to 4,000 kilograms per tonne of residue, and more preferably from 600 to 1,500 kilograms per tonne of residue. The optimal acid addition should be determined by experiment to be the most economical amount to maximise the sulphation of the rare earth minerals at the minimal addition of sulphuric acid.

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2014250661 16 Oct 2014 [0042] The pre-leach solid phase and sulphuric acid mixture is then preferably baked in a baking machine, which in one form will be a rotary kiln but may also be any other suitable baking device, for a time in the range of from 30 minutes to 4 hours at a temperature in the range of from 200 to 300°C. Preferably, the baking temperature is in the range from 220 to 250°C and the baking time is in the range of from 1 to 2 hours.

[0043] Acidic fumes emitted during the acid bake of step (b) above are preferably collected and condensed. This acid may be recycled to the acid bake process.

[0044] Following the acid bake, the acid-baked solid phase is preferably cooled prior to the next stage of the process, being the water leach of step (c) above. In this respect, the acid-baked solid phase is mixed in the water leach stage of step (c), in the first form of the invention, with the pre-leach liquid phase separated from the products of the pre-leach, or, in the second form of the invention, with water. Extra water may be added to the water leach if required, and/or wash water from the subsequent solid-liquid separation stage may also be used. The amount of liquor added to the water leach will ideally be sufficient to ensure the majority of the rare earth sulphate salts are solubilised in this stage. This will depend on the temperature of the water leach, which preferably will be as low as possible without the use of external cooling methods, although external cooling methods may be used if required.

[0045] In this respect, the temperature of the water leach will depend on the temperature of the acid-baked solid phase, the temperature of the pre-leach liquid phase, the ratio of water or pre-leach liquor to baked solids feeding into the water leach stage, and the heat of reaction of the water leach itself, which is mainly due to the hydration of residual sulphuric acid from the acid bake. Preferably, the temperature of the water leach should be below 80°C, but more preferably below 60°C.

[0046] The maximum concentration of the dissolved rare earths that can be achieved in the water-leach liquid phase will depend on the temperature of the water leach liquid phase, as the solubility of the rare earth sulphate salts decrease with increasing temperature, and also the concentration of the dissolved phosphoric and residual sulphuric acid from the acid bake and, in the first form of the invention, from

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2014250661 16 Oct 2014 the sulphuric acid pre-leach. Preferably, the total concentration of rare earth elements in the water-leach liquid phase will be as high as possible, although generally the concentration will vary from 5 g/L to 50 g/L, or more specifically from 10 g/L to 20 g/L.

[0047] The water-leach liquid phase will generally be acidic, due to the excess sulphuric acid from the baking process and the presence of dissolved phosphoric acid. Depending on the mineralogy of the rare earth concentrate being processed, the water-leach liquid phase will also contain significant concentrations of impurity elements such as iron and aluminium, as well as possible by-product elements such as thorium and uranium.

[0048] Following the water leach process of step (c) above, the water-leach liquid phase may be separated from the water-leach solid phase by any known suitable technique. In this respect, it will be appreciated that the water-leach solid phase (the residue from the water leach) is generally then required to be neutralised and stored in a tailing storage facility.

[0049] It is believed that sulphuric acid, or a process solution containing sulphuric acid, has not previously been contemplated for use in a pre-leach prior to an acid bake and water leach due to the understanding that it would react with acid-soluble calcium containing minerals to form insoluble calcium sulphate species that would interfere with any further processing of the rare earths, particularly the more valuable middle and heavy rare earths (MRE and HRE), specifically incorporating them into the calcium sulphate solid phase and making them less available for dissolution and later recovery from solution. However, preliminary experimental work that compared direct acid baking of beneficiation concentrate with sulphuric acid (the current art) to the acid bake of a sulphuric acid pre-leach residue (as in the present invention), the recovery of MRE and HRE during the subsequent water leach stage reduced from 8590% (MRE) and 80-85% (HRE) with pre-leach to only 70% (MRE) and 59% (HRE) with direct acid bake for the same overall acid addition.

[0050] However, the present inventors have surprisingly found that the use of sulphuric acid in such a pre-leach process, to dissolve apatite minerals in an ore or concentrate, can result in the associated rare earths being partially precipitated with

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2014250661 16 Oct 2014 the calcium sulphate into the pre-leach solid phase, and that this calcium sulphate does not interfere with the recovery of the co-precipitated rare earth elements during the following acid bake and water leach stages of the process.

[0051] It will therefore be appreciated that, because sulphuric acid then may be used in both the pre-leach and the acid bake, the processing of the combined preleach and water leach liquors in either form of the invention is significantly simplified, such as there being no need for a pre-acid bake precipitation step as described in the applicant’s own US patent 7,993,612 mentioned above. Additionally, significant economic advantages also arise due to the use of a single acid for the entire rare earth recovery process rather than two different acids. Also, because the sulphate from the acid is captured in the pre-leach solid phase as insoluble calcium sulphate, and is eventually deposited in a tailings storage facility along with other process residues, there is no requirement for the treatment of any soluble calcium salt, resulting from the use of other mineral acids, such as the nitric and hydrochloric acid of the prior art.

[0052] The subsequent recovery of the rare earth elements from the water-leach liquid phase or the pre-leach liquid phase (depending upon which form of the invention is adopted) may also be accomplished by any known suitable technique. In a preferred form, the rare earth elements will be recovered from these aqueous phases by precipitation as a sodium rare earth double sulphate salt, a general outline of which will now be provided, before turning to several examples showing the working of the process of the present invention. In this respect, for the sake of convenience, reference in the following paragraphs (that describe the rare earth element recovery processes) will be made to the “pregnant liquor” rather than to the “water-leach liquid phase” or the “pre-leach liquid phase”.

[0053] To precipitate the rare earth elements, the pregnant liquor may be treated with a concentrated solution, or a solid salt, of sodium sulphate or sodium chloride, to produce a solid phase containing most of the rare earth elements as a sodium rare earth double sulphate precipitate, plus a barren liquor containing uranium, thorium, phosphate and other impurities. This precipitation stage may be carried out at a temperature from ambient temperature to the boiling point of the mixture, preferably at a temperature between 80°C and the boiling point of the mixture, and more preferably

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2014250661 16 Oct 2014 at a temperature between 95°C and the boiling point of the mixture. This is because, in general, rare earth double sulphate salts have a decreasing solubility at higher temperatures, meaning that more of the rare earths are precipitated at higher temperatures compared to lower temperatures. The precipitation reaction will preferably be allowed to proceed for a time period in the range of from 0.5 to 8 hours, or preferably from 1 to 4 hours, or more preferably from 1 to 2 hours.

[0054] The amount of salt added is preferably equivalent to at least the stoichiometric requirement for the precipitation of the rare earths (designated again as RE) as double sulphate salts according to the following Equation (3):

RE₂(SO₄)₃ + Na₂SO₄ = 2NaRE(SO₄)₂ RE Sulphate Sodium Sulphate Sodium RE Sulphate [0055] Preferably, the amount of added salt is between 2 and 12 times the stoichiometric requirement, and more preferably between 4 and 8 times the stoichiometric requirement. Because the pregnant liquor also contains excess sulphuric acid, sodium chloride may be used as the precipitating agent, the sulphuric acid providing the sulphate ions for the precipitation reaction. In this respect, although sodium chloride is in general a cheaper reagent, and is only slightly less effective than sodium sulphate as a precipitating reagent, it is preferred to use sodium sulphate as this reagent can, if required, be recycled and re-used. In addition, the corrosive effects of chloride ions in acidic solutions are avoided.

[0056] Following the precipitation reaction, the aqueous barren liquor phase may be separated from the solid double sulphate precipitate. The barren liquor phase, which will contain minor amounts of the more soluble rare earth elements, especially the heavy rare earth elements, as well as most of the impurity elements, is sent for neutralisation and disposal in the normal manner.

[0057] Following solid-liquid separation, the solid double sulphate precipitate may be treated with a concentrated sodium hydroxide solution to convert the solid double sulphate precipitate to a solid rare earth hydroxide solid phase and a solution phase containing excess sodium hydroxide and sodium sulphate. This reaction can be described by the following Equation (4):

NaRE(SO₄)₂ (s) + 3NaOH = RE(OH)₃ (s) + 2Na₂SO₄

Sodium RE Sulphate Sodium Hydroxide RE Hydroxide Sodium Sulphate

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2014250661 16 Oct 2014 [0058] Preferably, the quantity of sodium hydroxide added is at least sufficient to convert all the double sulphate salt to the rare earth hydroxide solid, more preferably the quantity of sodium hydroxide added is from 100% to 120% of the stoichiometric requirement, and more preferably between 100% and 110% of the stoichiometric requirement. The reaction tends to be faster at elevated temperature. By way of example, the reaction may be carried out at about 70°C to minimise the reaction time.

[0059] The reaction may also be carried out in the presence of an oxidant such as air or another suitable oxidising agent to assist in the oxidation of cerium(lll) hydroxide to cerium(IV) hydroxide. Alternatively, the rare earth hydroxide solid may be treated as described below in paragraph [0062], [0060] Following this reaction, the sodium sulphate containing liquor phase may be separated from the rare earth hydroxide precipitate by any known suitable procedure. This liquor phase, which will contain virtually no rare earth elements, may be sent for recovery of the sodium sulphate or sodium chloride reagent. It can be treated by conventional techniques in a crystalliser or a concentrator to recover a concentrated solution of the salt, or a solid salt that can be re-used in the double sulphate precipitation stage of the process.

[0061] The solid rare earth hydroxide phase may then be re-dissolved in an acidic solution, preferably a hydrochloric acid solution, for removal of impurities if required, and separation of the rare earth elements, by conventional techniques of solvent extraction and/or ion exchange, into separate rare earth products that may be mixed, or individual rare earth oxides, rare earth carbonates, rare earth oxalates or other rare earth salts.

[0062] In another form of the invention, the solid rare earth hydroxide phase may be heated by standard procedures to a temperature between 120 and 160°C for a period of time in the range of from 2 to 12 hours in an oxidising atmosphere (such as air) to convert the cerium (III) hydroxide in the precipitate to cerium (IV) hydroxide. This oxidation reaction is faster at higher temperatures so that a temperature between 150°C and 160°C is preferred to lessen the reaction time required to between 2 and 4 hours. All other rare earths contained in this hydroxide solid phase remain in the +(lll) oxidation state, allowing the separation of cerium, which is in general the most

SPCN-995859

2014250661 16 Oct 2014 abundant and lowest value rare earth, from the remaining rare earths, prior to the subsequent separation stage.

[0063] The rare earth hydroxide then may be dissolved in dilute hydrochloric acid at ambient temperature, maintaining the pH between 1.5 and 4, to dissolve the remaining rare earths away from the cerium, which remains in the solid phase as a cerium(IV) hydroxide, as per the following Equation (5):

Ce(OH)₄ (s) + RE(OH)₃ (s) + 3HCI = RECI₃ (aq) + 3H₂O + Ce(OH)₄ (s)

Cerium Hydroxide RE Hydroxide Hydrochloric Acid RE Chloride Cerium Hydroxide [0064] Most of the impurity thorium and uranium, as well as minor quantities of other rare earths, will also remain with the cerium(IV) hydroxide solids and thus may need to be removed by known standard techniques to form a purified cerium product. The rare earth chloride solution, containing only minor amounts of the cerium, can then be treated for impurity removal and separation of the rare earths by known techniques of solvent extraction and/or ion exchange into separate rare earth solid products such as carbonates, oxalates or oxides.

BRIEF DESCRIPTION OF DRAWINGS [0065] Figure 1 is a schematic representation of a flow diagram in accordance with a first preferred embodiment of the process of the present invention, being a cocurrent process.

[0066] Figure 2 is a schematic representation of a flow diagram in accordance with a second preferred embodiment of the process of the present invention, being a counter-current process.

[0067] Figure 3 is a schematic representation of a flow diagram of a test procedure used in the recycle aspect of the counter-current process in Figure 2.

DETAILED DESCRIPTION OF FLOW DIAGRAMS [0068] Before providing a more detailed description of the preferred embodiments of the present invention, with reference to experimental data, it will be useful to provide some explanation of the flow diagrams of Figure 1 and Figure 2. The flow

SPCN-995859

2014250661 16 Oct 2014 diagram of Figure 3 will be further described below when discussing the experimental data.

[0069] Figure 1 generally illustrates a co-current process for extracting one or more rare earth elements from an ore or concentrate consisting of the rare earthcontaining apatite group minerals, and at least one other rare earth containing mineral. The process includes treating the ore or concentrate with sulphuric acid in a sulphuric acid leach (hereafter called a “pre-leach”), preferably a weak acid leach, to substantially dissolve most of the apatite and some of the contained rare earth elements into a pre-leach liquid phase and leaving the sulphuric acid-insoluble minerals in a pre-leach solid phase along with a calcium sulphate precipitate.

[0070] The pre-leach solid phase (2) is then separated from the pre-leach liquid phase (1) by a standard process such as filtration or counter-current decantation (CCD), and subsequently the separated pre-leach solid phase (2) is mixed with concentrated sulphuric acid in a mass ratio of acid to solids of between 0.4 and 4.0, preferably from 0.6 to 1.5, depending on the nature of the solids.

[0071] The process then includes baking the mixture in a kiln at a temperature of from 200 to 300°C, for a period of time in the range of 30min to 4 hours, in an acid bake stage to produce an acid-baked solid phase (3). The acid-baked solid phase (3) is leached with the pre-leach liquid phase (1), or part thereof with added water as required, for between 15 minutes and 120 minutes, in a water leach stage to dissolve the rare earth elements in the acid-baked solid phase (3) into a water-leach liquid phase (4). Optionally, part of the pre-leach liquid phase may be sent directly to double sulphate precipitation.

[0072] This second aqueous phase (4) is separated from the water-leach solid phase, being an acid-insoluble residue (5), the latter of which can be treated by neutralisation prior to disposal in a suitable tailings storage facility. It is to be understood that in this description of the process, the word “water” includes wash water as well as process water.

[0073] Once the rare earth elements have been extracted from the ore or concentrate into the water-leach liquid phase (4), they can be recovered from the aqueous solution by a range of techniques. These techniques include, but are not

SPCN-995859

2014250661 16 Oct 2014 limited to, direct solvent extraction, precipitation as a rare earth carbonate, rare earth hydroxide or rare earth oxalate salt, or precipitation as a sodium rare earth double sulphate salt.

[0074] According to a preferred aspect of the present invention, the rare earth elements are recovered from the water-leach liquid phase (4) by precipitation as a sodium rare earth double sulphate salt. This is shown in Figure 1 by way of the treatment of the water-leach liquid phase (4) with sodium sulphate or sodium chloride, preferably sodium sulphate, either as a solid or in a concentrated solution, in a double sulphate precipitation stage, to produce a solid phase (7) containing most of the rare earth elements as a sodium rare earth double sulphate precipitate and a barren liquor (6) containing uranium, thorium, phosphate and other impurities from the ore or concentrate. This precipitation stage may be carried out at a temperature from ambient temperature to the boiling point of the mixture, preferably at a temperature between 80°C and the boiling point of the mixture.

[0075] Following that, in the process of Figure 1, the aqueous barren liquor phase (6) is separated from the solid double sulphate precipitate (7) prior to neutralisation and disposal. The solid double sulphate precipitate (7) is then treated with sodium hydroxide solution, preferably with aeration to convert the double sulphate precipitate to a rare earth hydroxide solid phase (8) containing cerium (IV) hydroxide and other rare earth(lll) hydroxides, and a solution phase (9) containing excess sodium hydroxide, if added, and sodium sulphate. Preferably the quantity of sodium hydroxide added is at least sufficient to convert all the double sulphate salt to the rare earth hydroxide solid.

[0076] The aqueous sodium sulphate solution phase (9) is then separated from the solid rare earth hydroxide phase (8). The sodium sulphate or sodium chloride solution phase (9) is treated by conventional techniques in a concentrator or a crystalliser to recover a concentrated solution of the salt, or a solid salt, that can be re-used in the double sulphate precipitation stage of the process [0077] Optionally but not shown in Figure 1 or Figure 2, if the solid rare earth hydroxide phase (8) has not been treated by aeration to convert Ce(lll) to Ce(IV) hydroxide, it may be heated and dried in an oxidising atmosphere then redissolved in

SPCN-995859

2014250661 16 Oct 2014 a mildly acidic solution, preferably a hydrochloric acid solution, for removal of impurities (including cerium) if required, and separation of the rare earth elements, by the conventional techniques of solvent extraction and/or ion exchange, into separate rare earth products that may be mixed or individual rare earth oxides, rare earth carbonates, rare earth oxalates or other rare earth salts.

[0078] Having described the co-current process of Figure 1, an alternative process, a counter-current process that relies upon the recycle of the water-leach liquid phase, or part thereof, to the pre-leach stage, will now be described.

[0079] The principal difference between the co-current process of Figure 1 and the counter-current process of Figure 2 is that the water-leach liquid phase in the counter-current process is recycled back to the pre-leach stage, in part or in total, so that the excess sulphuric acid available in the water leach stage can be used to economically leach incoming rare earth concentrate. If required, and if the water leach liquor contains insufficient acid to dissolve the apatite in the incoming concentrate and retain sufficient acidity to keep the dissolved rare earths in solution, the counter-current embodiment may still be adopted, but with extra sulphuric acid added to the pre-leach stage to ensure maximum dissolution of the apatite mineral. Conversely, if too much acid is present in the water leach liquor, then only part of the water leach liquor can be recycled, with added water as required, with the remaining water leach liquor proceeding directly to rare earth recovery.

[0080] Subsequently, the pre-leach liquid phase, that in this second embodiment will be a pregnant liquor that includes soluble rare earths from both the water leach of the acid-baked solid phase and from the pre-leach itself, can be used for the recovery of the rare earths from solution by any of the known methods described above. Preferably, the recovery of the rare earths is undertaken by direct solvent extraction or more preferably by double sulphate precipitation.

[0081] An advantage of this counter-current embodiment is that there is a more economical use of sulphuric acid than in the co-current process, due to the use of water leach liquor containing residual sulphuric acid in the initial pre-leach of the concentrate. Having said that, a disadvantage of the counter-current embodiment is that the acid addition to the acid bake stage, and the pre-leach stage if required, will

SPCN-995859

2014250661 16 Oct 2014 ideally be carefully controlled to ensure the maximum solubility of the dissolved rare earth sulphate minerals is not exceeded. This can occur with the addition of excessive use of sulphuric acid in either process. However, under controlled conditions, the counter-current process can be more economical in the use of sulphuric acid than the co-current process.

[0082] With particular reference to Figure 2, the process includes treating the ore or concentrate with a sulphuric acid containing water leach liquor in a sulphuric acid leach (hereafter called a “pre-leach”), again preferably a weak sulphuric acid, this leach liquor including part or all of a water leach liquid phase recycled from later in the process plus added sulphuric acid or water as required. As with the first embodiment of Figure 1, the pre-leach aims to substantially dissolve most of the apatite and some of the contained rare earth elements into a pre-leach liquid phase, leaving the sulphuric acid-insoluble minerals in a pre-leach solid phase along with a calcium sulphate precipitate.

[0083] The pre-leach solid phase (12) is then separated from the pre-leach liquid phase (11) by a standard process such as filtration or counter-current decantation (CCD), and subsequently the separated pre-leach solid phase (12) is mixed with concentrated sulphuric acid in a mass ratio of acid to solids of between 0.4 and 4.0, preferably between 0.6 and 1.5, depending on the nature of the solids.

[0084] The process includes baking the mixture in a kiln at a temperature of from 200 to 300°C, for a period of time in the range of 30 minutes to 4 hours, in an acid bake stage to produce an acid-baked solid phase (13). The acid-baked solid phase (13) is leached with added water, for between 15 minutes and 120 minutes, in a water leach stage to dissolve rare earth elements still in in the acid-baked solid phase (13) into a water-leach liquid phase (14). This water-leach liquid phase (14) is separated from the water-leach solid phase (15), being an acid-insoluble residue, the latter of which can be treated by neutralisation prior to disposal in a suitable tailings storage facility.

[0085] The water leach liquid phase (14), that contains excess sulphuric acid from the acid bake stage, together with some dissolved rare earth elements, is then recycled, in total or in part, to the pre-leach to assist with economically dissolving the

SPCN-995859

2014250661 16 Oct 2014 apatite mineral (with added sulphuric acid as desired) in the incoming rare earth ore or concentrate (16). Optionally, if only part of the water leach liquid phase is recycled, the remainder is directed to double sulphate precipitation.

[0086] Once the rare earth elements have been extracted from the ore or concentrate into the pre-leach liquid phase (11), they can be recovered from the preleach liquid phase by a range of techniques. These techniques include, but are not limited to, direct solvent extraction, precipitation as a rare earth carbonate, rare earth hydroxide or rare earth oxalate salt, or precipitation as a sodium rare earth double sulphate salt.

[0087] According to a preferred aspect of the present invention, the rare earth elements are recovered from the pre-leach liquid phase (11) by precipitation as a sodium rare earth double sulphate salt as described for the co-current process in Figure 1.

DESCRIPTION OF EXPERIMENTAL DATA [0088] Attention will now be directed to a description of experimental data developed to illustrate the preferred embodiments of the present invention.

Co-Current Sulphuric Acid Pre-Leach (Tests 1 to 3) [0089] Three batches of beneficiation concentrate were used in Tests 1-3 respectively. The head analyses of the three batches are listed in Table 1. Mineralogical analysis of the concentrate showed that the main rare earth containing minerals were apatite, allanite and monazite. The monazite was much finer than the apatite and was present as inclusions in the apatite matrix. Other minor rare earth containing minerals included crandallite, goyazite and bastnasite. Thorium and uranium were associated mainly with the monazite, but were also contained in the apatite matrix.

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2014250661 16 Oct 2014

Table 1 - Head Assay of Beneficiation Concentrate

Batch #	LRE1 %	MRE2 %	HRE3 %	Y %	TREE4 %	Al %	Ca %	Fe %	P %	S %	Th	U
1	5.66	0.26	0.05	0.09	6.07	4.87	17.9	2.47	8.93	0.13	4710	427
2	5.62	0.24	0.05	0.09	6.00	4.86	17.9	2.50	9.01	0.15	5430	439
3	5.25	0.23	0.04	0.08	5.60	3.68	18.9	2.43	8.88	0.21	4660	412

¹ Light Rare Earths (Sum of La, Ce, Pr, Nd) ² Middle rare Earths (Sum of Sm, Eu, Gd) ³ Heavy rare Earths (Sum of Tb, Dy, Ho, Er, Tm, Yb, Lu ⁴Total Rare earths (Sum of LRE, MRE & HRE+Y) [0090] The sulphuric acid leach tests were conducted at 50°C, in a sealed baffled reactor vessel on a hot plate and agitated by an overhead stirrer at approximately 400 rpm. A condenser was fitted to the vessel to minimise evaporation. The required volume of leach liquor was placed in the reactor vessel and was pre-heated to about 35°C and then a calculated mass of beneficiation concentrate was added. The exothermic reaction took the temperature up to ~50°C and was maintained at this temperature for the duration of the reaction. After 30 minutes, a slurry sub-sample was taken, vacuum filtered into a filtrate and solid sample and the solid washed thoroughly, dried at 105°C to a constant mass and assayed. These were the samples used for the mass balance calculations. Filtrate samples for assay were diluted 1:20 with 1% (v/v) nitric acid in deionised water to prevent precipitation. The bulk residue recovered from vacuum filtration was washed twice with deionised water at a wash ratio of 1:1 solid:water. The wash solutions were also analysed to perform an overall elemental mass balance. The washed bulk solids were dried overnight at 105°C.

[0091] Three preliminary pre-leach tests were conducted as a function of acid dose. The acid dose was calculated according to the calcium content of the sample and assuming one mole of calcium in the concentrate would react with approximately one mole of sulphuric acid. The calcium to sulphuric acid molar ratios tested were 90%, 100% and 110% of the stoichiometric requirement, calculated to be 404, 448 and 491 kg/t of 98% w/w sulphuric acid addition, respectively. The sample mass of about 800 g was used with -20% w/w sulphuric acid that kept the initial percentage solids at about 30% w/w. This appeared to be the upper limit of percent solids that could be used in these tests, which would allow sufficient mixing of the concentrate with the acid.

SPCN-995859

2014250661 16 Oct 2014 [0092] The calculated percentage metal dissolution data obtained are summarised in Table 2 and the composition of the pre-leach liquors are shown in Table 3.

Table 2 - Percentage Dissolution Data for Preliminary Pre-Leach Tests

Test No.	Molar Ratio % HzSO^Ca	98% H2SO4 kg/t	Dissolution (% w/w)
LRE	MRE	HRE	Y	TREE	Al	Ca	Fe	P	Th	U
1	90	404	6.5	10.7	18.0	25.7	7.1	19.5	1.7	28.7	61.3	12.8	26.7
2	100	448	9.7	13.9	20.3	28.6	10.2	14.2	1.7	24.2	65.2	20.5	26.0
3	110	491	10.7	16.3	24.2	28.6	11.3	15.3	1.8	23.9	64.1	21.7	26.1

Table 3 - Composition of Pre-Leach Liquors

Test No.	Pre-Leach Liquor Composition
LRE g/L	MRE mg/L	HRE mg/L	Y mg/L	TREE g/L	Al g/L	Ca g/L	Fe g/L	P g/L	S g/L	Th mg/L	U mg/L
1	2.02	153	52	123	2.35	5.34	1.69	4.20	26.6	7.58	356	60
2	2.94	190	57	130	3.31	3.66	1.55	3.26	28.4	10.7	566	58
3	3.02	215	65	126	3.42	3.76	1.60	3.04	25.6	11.4	554	56

[0093] The main purpose of the sulphuric acid leach (referred to as the pre-leach) is the dissolution of the apatite mineral. As calcium is re-precipitated as a calcium sulphate, this is best indicated by the dissolution of phosphate. The complete dissolution of apatite required at least a H₂SO₄:Ca molar ratio of 1.0 (or 100%), or an acid dose of 448 kg of 98% sulphuric acid per tonne of concentrate. However, the phosphate was not totally dissolved as other weak acid insoluble phosphate minerals, predominantly monazite type minerals, are still present in the residue of this leach.

[0094] It is presumed that the rare earths that dissolve are predominantly those associated with the apatite mineral and other minor rare earth containing minerals that are solubilised in the pre-leach reaction. This was confirmed by mineralogical analyses. Some of the dissolved rare earths may also be co-precipitated with gypsum that is formed during the reaction. This, however, appears to not interfere with their later dissolution during the acid bake and water leach stages.

[0095] The results indicated that the percentage dissolution of rare earth elements for each acid dose was in the order Y > HRE > MRE > LRE. Some iron, aluminium,

SPCN-995859

2014250661 16 Oct 2014 uranium and thorium also dissolved in this reaction. A slight mass gain of about 5% solids was observed. Based on these results, and to ensure maximum dissolution of the apatite minerals, an acid dose of 491 kg/t (acid:Ca molar ratio of 110%) was chosen as the optimum for further testing.

Acid Bake and Water Leach (Tests 4 and 5) [0096] Two Acid Bake I Water Leach tests (Tests 4 and 5) were conducted on the residue from Pre-Leach Test 3 at concentrated sulphuric acid (98% w/w) additions of 863 and 1,078 kg/t of concentrate. The dry residue from the sulphuric acid pre-leach test was placed in a stainless steel tray and the required mass of concentrated sulphuric acid (98% w/w H2SO4) was added directly to the feed solids, while mixing thoroughly to ensure an even distribution. The mixture was then transferred to a preheated furnace at 250°C for baking. A temperature probe was placed directly into the sample to ensure the bake temperature of 250°C was achieved in the sample. The baking duration was two hours at temperature.

[0097] The entire baked product was cooled, weighed, and re-pulped with the required mass of deionised water. The resultant slurry was agitated for an hour at sufficient intensity to keep the solids suspended. The temperature rose to about 50°C from ambient temperature due to the heat of dissolution of the sulphuric acid retained in the solids after baking. After 60 minutes, a slurry sub-sample was taken, weighed and assayed as described in Example 1. The bulk slurry was also similarly treated.

[0098] The calculated metal extraction data are shown in Table 4 and the compositions of the water leach liquor are shown in Table 5. The overall summary of Pre-Leach I Acid Bake extraction data obtained for Pre-leach Test 3 and Acid Bake I Water Leach Test 5 are shown in Table 6.

Table 4 - Percentage Dissolution Data for Preliminary Acid Bake / Water Leach Tests

Test No.	98% H2SO4 kg/t1	Extraction (% w/w)
LRE	MRE	HRE	Y	TREE	Al	Ca	Fe	P	Th	U
4	863	95.5	84.7	66.5	65.5	94.5	25.6	10.6	34.1	88.2	99.3	97.9
5	1,078	96.5	87.0	71.2	70.6	95.6	37.3	19.3	47.0	91.2	99.4	97.9

¹ Adjusted per t of concentrate as 98% H₂SO₄

SPCN-995859

2014250661 16 Oct 2014 [0099] The extraction data showed that slightly higher percent RE extraction was obtained at the higher acid dose of 1,078 kg/t of 98% sulphuric acid. The extraction was in the order LRE > MRE > HRE > Y. The extraction of calcium, aluminium and iron was also greater at the higher acid dose. Thorium and uranium dissolutions were nearly 100% in both cases. The discharge residue contained 0.24% TREE for Test 5.

Table 5 - Water Leach Liquor Compositions for Preliminary Acid Bake Tests

Test No.	Water Leach Liquor Composition (mg/L or g/L)
LRE g/L	MRE mg/L	HRE mg/L	Y mg/L	TREE g/L	Al g/L	Ca g/L	Fe g/L	P g/L	S g/L	Th g/L	u mg/L
4	11.9	410	53	76	12.4	1.60	1.15	1.02	6.70	49.6	1.14	90
5	11.8	392	53	74	12.3	2.40	1.90	1.52	6.90	71.4	1.14	86

[0100] The water leach liquors contained about 12 g/L total rare earths (TREE) and about 50 and 71 g/l sulphur (as sulphate) respectively. The TREE concentrations may be increased by lowering the amount of water addition to the water leach.

Table 6 - Overall Pre-Leach and Acid Bake Extraction Data¹

Sample / Test	98% H2SO4 kg/t	Feed and Residue Grades (% w/w)	Overall Extraction (%)2
LRE	MRE	HRE	Y	TREE	LRE	MRE	HRE	Y	TREE
Feed		5.66	0.26	0.05	0.09	6.07
Pre-Leach Test 3	491	4.62	0.20	0.04	0.06	4.92	10.7	16.3	24.2	28.6	11.3
Acid Bake/WL Test 5	1,0783	0.18	0.03	0.01	0.02	0.24	84.9	72.6	54.1	48.7	83.6
Overall	1,569	0.18	0.03	0.01	0.02	0.24	96.9	89.1	79.2	79.7	96.2

¹ Extractions based on feed and residue solids assays.

² Acid Bake extraction adjusted relative to concentrate.

³ Adjusted per t of concentrate as 98% H₂SO₄.

[0101] With the total acid dose of 1,569 kg/t to Pre-Leach and Acid Bake, 96.2% TREE was extracted, including 96.9% LRE, 89.1% MRE, 79.2% HRE and 79.7% yttrium. Given that the main aim of the acid bake step is the extraction of TREE, and especially the more valuable Nd/Pr, MRE, and HRE, into the water leach liquor, an acid dose of 1,000 kg/t of 98% sulphuric acid was chosen as the optimum for further testing.

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2014250661 16 Oct 2014

Combined Pre-Leach, Acid Bake and Water Leach (Tests 6 and 7) [0102] In the first embodiment of the invention, two combined Pre-Leach, Acid Bake and Water Leach (SAPL/AB/WL) tests were conducted in which the Pre-Leach Liquor was forwarded to the Water Leach of the Acid Baked Pre-Leach residue. The tests were done on beneficiation concentrate that originated from three different batches and therefore had slightly different head assays, as shown in Table 1. The feed acid solution to Test 6 was pre-heated to 50 °C while in Test 7 it was pre-heated to 35 °C prior to concentrate addition.

[0103] Test 6 pre-leached solids were washed with deionised water at a ratio of 1:1 feed solids to water, while Test 7 pre-leached solids were washed with 1% w/w H₂SO₄ in deionised water, to prevent possible re-precipitation of the dissolved rare earths, at a ratio of 0.6 to 1.0 water to leach slurry by mass. For both tests, two displacement washes were performed and the resulting wash waters were combined and used as required together with primary filtrate for the subsequent water leach.

[0104] In the water leach for Test 6, a solids to liquor ratio of 1.4 was used to target a TREE concentration of ~15 g/L in the final water leach liquor. The feed to water leach comprised all of the primary filtrate from the sulphuric acid pre-leach plus some of the combined wash water to make up the balance. In Test 7, the feed liquor to water leach was prepared from the pre-leach primary filtrate and combined wash water at a ratio to target around 12 g/L TREE concentration in the final water leach liquor. The test results are summarised in Tables 7 and 8 below.

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Table 7 - Overall Pre-Leach and Acid Bake Extraction Data for Tests 6 and 7¹

Sample / Test	98% H2SO4 kg/t	Feed and Residue Grades (% w/w)	Overall Extraction (%)2
LRE	MRE	HRE	Y	TREE	LRE	MRE	HRE	Y	TREE
Feed		5.66	0.26	0.05	0.09	6.07
Pre-Leach Test 6	491	4.36	0.19	0.04	0.06	4.65	15.6	20.2	28.0	29.5	16.1
Acid Bake/WL Test 6	1,0953	0.16	0.03	0.01	0.02	0.22	82.1	70.2	54.0	53.5	80.9
Overall Test 6	1,586	0.16	0.03	0.01	0.02	0.22	97.7	90.4	82.0	83.0	97.1
Feed		5.62	0.24	0.05	0.09	6.00
Pre-Leach Test 7	413	4.50	0.20	0.04	0.06	4.80	5.7	3.5	10.5	23.0	6.0
Acid Bake/WL Test 7	12373	0.21	0.03	0.01	0.02	0.27	90.5	82.7	62.8	54.9	89.4
Overall Test 7	1,650	0.21	0.03	0.01	0.02	0.27	96.2	86.3	73.3	77.9	95.4

¹ Extractions based on feed and residue solids assays.

² Acid Bake extraction adjusted relative to concentrate.

³ Adjusted per t of concentrate as 98% H₂SO₄

Table 8 - Combined SAPL/AB/WL Primary Filtrate Liquor and Wash 1 Compositions

Test	Stage	Sample	Free Acid (g/L)	Liquor Composition (mg/L or g/L)
LRE g/L	MRE	HRE	Y	TREE g/L	Al g/L	Ca g/L	Fe g/L	P g/L	S g/L
6	PL	Primary Filtrate	115	3.19	198	57	115	3.56	3.90	1.60	3.46	36.0	24.0
		Wash 1		1.76	109	31	66	1.97	2.21	0.980	1.81	19.9	12.8
	AB	Primary Filtrate	307	14.7	570	95	171	15.6	9.92	0.514	6.46	29.2	102
		Wash 1		7.93	312	49	84	8.38	4.92	0.867	3.05	14.0	50.8
7	PL	Primary Filtrate	54	1.14	92	31	68	1.33	1.16	1.77	1.40	20.6	3.70
		Wash 1		0.256	25	9	18	0.308	0.350	0.744	0.371	5.36	1.95
	AB	Primary Filtrate	251	11.7	493	82	131	12.4	5.10	1.02	4.54	21.6	49.6
		Wash 1		6.63	221	38	77	6.96	3.19	1.16	2.79	12.8	31.0

[0105] The data in Table 7 shows that that by using a lower acid dose in the preleach in Test 7 versus Test 6, the extraction of TREE was lower in this stage. This was despite the phosphorus content of the concentrate being reduced from about 9% to 3.2% in Test 6 and to 2.8% in Test 7 (data not shown in Table). The TREE concentrations of the pre-leach residues were 4.65 and 4.80% for Tests 6 and 7 respectively.

SPCN-995859

2014250661 16 Oct 2014 [0106] Despite the differences in pre-leach performance, the subsequent acid bake tests generated residues containing similar residual TREE values. The Test 7 water leach discharge contained 0.27% TREE in comparison to 0.22% TREE in the Test 6 residue. In other words, the poorer recovery of rare earths in the pre-leach was made up for in the acid bake and water leach for Test 7 relative to Test 6.

[0107] The pre-leach primary filtrate in Test 6 contained 3.6 g/L TREE and 115 g/L free acid (24 g/L total sulphur) compared to 1.3 g/L TREE with 54 g/L free acid (total sulphur 3.7 g/L) in Test 7. The management of total sulphur transferred to water leach liquor in Test 7 was reflected where the data showed only 50 g/L sulphur in WLL in comparison to 102 g/L in Test 6. The Test 6 water leach solution was unstable and on aging, the TREE concentration in solution fell from 15.6 g/L TREE to 9.8 g/L within a few days. The Test 7 water leach liquor contained 12.4 g/L TREE and did not form any residues on standing. From this data, it appears a solution concentration of about 15 g/L TREE in water leach liquor at 50°C is achievable at a maximum of about 50 g/L total sulphur. Above this temperature or sulphur content there is a risk of the rare earth sulphates exceeding their solubility in the water leach solution.

[0108] The data from Test 7 indicated 95% TREE extraction comprised of 96% LRE, 86% MRE, 73% HRE and 78% yttrium with a total sulphuric acid (98% w/w) addition at 1650 kg/t based on concentrate. However, Test 6 showed much higher extractions up to 97% TREE with 1586 kg/t acid addition based on concentrate, but as discussed above the final solution was not stable. For continuous processing, in which the water leach solution is not aged, the optimal total acid addition to the concentrate samples used, to maximise RE extraction is probably between 1586 and 1650 kg/t of 98% sulphuric acid. Alternatively, further dilution of the water leach liquor could be used to ensure the rare earths remain in solution.

Additional Acid to Pre-Leach and Partial Recycle of Water Leach Liquor (Test 8) [0109] In a variation on both forms of the invention, a counter-current recycle test consisting of a sulphuric acid pre-leach, acid bake of residue, and water leach of baked solids was conducted using acid additions consistent with Tests 6 and 7 described above. A schematic diagram of the test involved is shown in Figure 3, outlining how the solids, filtrates and wash waters were circulated. In addition to the

SPCN-995859

2014250661 16 Oct 2014 acid contained in water leach solution fed to pre-leach (measured by free acid titration), additional concentrated sulphuric acid was added to the water leach solution to ensure that the total acid additions employed were 110% H₂SO₄:Ca molar ratio for the sulphuric acid pre-leach tests.

[0110] For simplicity, 1000 kg of 98% sulphuric acid per tonne of pre-leach residue was added to the acid bake. Only the primary water leach filtrate was used in the pre-leach with no wash water added. Since a significant portion of primary water leach solution was held up in the unwashed filter cake, this primary filtrate represented only a fraction of the total water leach solution generated. In these tests, the associated quantity of concentrate used in pre-leach was not scaled down proportionally with the fraction of recovered water leach solution. Therefore, the execution of this test is equivalent to a process where only part of the water leach solution is recycled to the pre-leach, with the unused water leach solution (or wash solutions) and the whole of pre-leach solution forwarded to rare earth recovery.

[0111] The same sample of concentrate (Batch 1) used in the previous co-current tests was used in these tests, and the experimental methods used were also as described for the co-current process.

[0112] The recycle test was initiated by an acid bake of the combined residues obtained from the three preliminary sulphuric acid pre-leach tests (Tests 1-3). These were used in the initial acid bake and water leach test (named Test 8 AB1). The primary water leach liquor (filtrate) from this test was used to pre-leach a sample of rare earth concentrate (Batch 1 concentrate.) in Test 8 PL1. An additional 101 kg/t of 98% sulphuric acid was added to meet the 110% stoichiometric addition target for the pre-leach. The primary filtrate from this pre-leach test was retained for later double sulphate precipitation tests. The residue from this pre-leach test was forwarded to a second acid bake test (Test 8 AB2). Sulphuric acid (1000 kg/t of sulphuric acid per tonne of pre-leach residue) was added to the acid bake, and all of the available bulk displacement wash water from AB1 and PL1 was used as required in the water leach component of this test.

[0113] The primary water leach liquor from AB2 was then used in a second preleach test (PL2) of Batch 1 concentrate along with added 98% sulphuric acid (330

SPCN-995859

2014250661 16 Oct 2014 kg/t). It is believed that this extra acid addition was too high, pushing the solubility of the rare earths in solution beyond their saturation point and resulting in precipitation of rare earths from the pre-leach liquor. This conclusion is based on the near 50% reduction in TREE concentration in solution versus the 15% precipitation of rare earths based on solids assays. The filtrate was stored for future testing.

[0114] The residue proceeded to a final acid bake and water leach test (Test 8 AB3), with 1000 kg/t of 98% sulphuric acid per tonne of pre-leach residue added to the acid bake. All of the available wash water from the former pre-acid bake and preleach tests was used in the water leach. This ensured no re-dissolved rare earths were lost and also was intended to approximate the expected water balance for the continuous process. This completed the tests.

[0115] The solids feed and residue compositions data from the tests are summarised in Table 9 and the composition of the pre-leach and water leach primary liquors are shown in Table 10.

Table 9: Feed and residue solids compositions - counter-current recycle tests

Recycle	Test 8	Acid Dose kg/t	Sample	Solids Composition (% w/w)
LRE	MRE	HRE	Y	HRE + Y	TREE	P	S
1	AB1	1000	Feed	4.51	0.17	0.03	0.06	0.09	4.77	2.84	9.94
			Residue	0.20	0.04	0.01	0.02	0.04	0.27	0.17	14.3
	PL1	101	Feed	5.66	0.26	0.05	0.09	0.14	6.07	8.93	0.13
			Residue	4.71	0.18	0.03	0.06	0.09	4.98	3.31	10.1
2	AB2	1000	Feed	4.71	0.18	0.03	0.06	0.09	4.98	3.31	10.1
			Residue	0.19	0.03	0.01	0.02	0.03	0.25	0.22	16.1
	PL2	330	Feed	5.66	0.26	0.05	0.09	0.14	6.07	8.93	0.13
			Residue	5.11	0.21	0.04	0.06	0.10	5.42	4.77	9.32
3	AB3	1000	Feed	5.11	0.21	0.04	0.06	0.10	5.42	4.77	9.32
			Residue	0.23	0.04	0.01	0.02	0.04	0.31	0.22	16.2

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2014250661 16 Oct 2014

Table 10: Liquor compositions, primary filtrates - counter-current recycle tests

Test 4 - Recycle	Test 8	Liquor Composition (g/L)
TREE	S	P
1	AB1	17.0	93.6	11.1
-	PL1	16.8	69.8	47.0
2	AB2	20.4	116	27.4
-	PL2	10.3	141	50.0
3	AB3	11.5	140	39.4

[0116] The overall extractions for each recycle of these tests are summarised in

Table 11. The acid doses adjusted per tonne of concentrate are also included.

Table 11: Overall RE extractions of the counter-current recycle tests¹

Recycle	Sample/ Test	98% H2SO4 kg/t	Overall Extraction (% )4
LRE	MRE	HRE	Y	TREE
	Feed
1	Tests 1-32	432	14.5	29.7	38.4	28.0	15.6
-	Test 8 AB1	1,073	82.9	60.1	42.4	52.4	81.1
Overall	SAPL + AB	1,506	97.4	89.8	80.8	80.4	96.6
2	Test 8 PL1	101	10.0	7.2	16.4	14.7	8.70
-	Test 8 AB2	1,323	106.54	80.4	60.5	62.4	104.34
Overall	SAPL + AB	1,434	96.5	87.6	76.9	77.1	95.7
3	Test 8 PL2	330	16.2	3.30	11.8	8.0	15.0
-	Test 8 AB3	1,287	112.64	91.0	66.1	68.5	110.64
Overall	SAPL + AB	1,617	96.4	87.7	77.9	76.5	95.5

'Extractions based on feed and residue solid assays.

² The feed solids to AB1 were a blend of material from Test 1 (-45%), Test 2 (-45%) and Test 3 (-10%). The values presented are the weighted averages for these SAPL tests.

³Acid Bake extractions were adjusted to be based on concentrate rather than pre-leach residue.

⁴ Indicates some LRE precipitation in PL1 and PL2 carrying through to acid bake..

[0117] Although the acid additions have not been optimised in these tests, the results show that if the pre-leach is operated such that there is a net precipitation of rare earths, then effectively all of these rare earths are recovered in the subsequent acid bake and water leach. This test also demonstrated that the pre-leach could be conducted without recycling back the full flow of water leach liquor. Therefore, if necessary, the recirculating loads between water leach and pre-leach can be controlled through a bleed of water leach solution. At high acid additions there was some precipitation of rare earths from the process liquors into the solid phases,

SPCN-995859

2014250661 16 Oct 2014 especially in the pre-leach phase. These precipitated rare earths were, however, recovered during the washing of the solids.

Counter-Current Sulphuric Acid Pre-Leach (Tests 9 to 17) [0118] In the second embodiment of the invention, a series of counter-current tests were conducted in which water leach liquor generated from an acid bake and water leach of a pre-leach residue was forwarded to a sulphuric acid pre-leach. The pre-leach residue used in these tests was generated from pre-leaches using sulphuric acid and water. With the pre-leach solution discarded, the water leach tests were conducted using water consistent with the counter-current flowsheet. The details of the initiating pre-leaches are presented here as they will have some impact on the counter-current test that follows. It is worth noting that the initiating pre-leach followed by acid bake and water leach embodies the first form of the invention, while the acid bake and water leach followed by a pre-leach embodies the second form. Although additional detail can be extracted from these results about the first embodiment of the invention, discussion in this section will focus on the second embodiment of the invention exclusively.

[0119] The tests were done on beneficiation concentrate that originated from two significantly different batches (Batch 2 and Batch 3) and therefore different head assays, as shown in Table 1. The pre-leach Tests 9 and 10 were conducted as per paragraph [0090] with the following differences. Tests 9 and 10 pre-leached solids were displacement washed twice with DI water at a wash ratio of 1:1 feed solid to DI water, while the water leach solids were washed thoroughly. A bulk pre-leach was conducted at ambient temperature to generate initial pre-leach residue for Tests 11 to 17, and the filter cake was washed thoroughly. Pre-leach tests 11 to 17 were conducted at ambient temperature (feed solutions around 20 °C) in sealed flasks with magnetic stirrers. After 60 minutes the contents were treated as per the subsample (paragraph [0090] above) except no solid sample was withdrawn for assay. All solids were forwarded to the following acid bake. No additional sulphuric acid was added to water leach solutions feeding into pre-leach tests. Acid bake Tests 9 and 10 were conducted as per paragraphs [0096] and [0097] above. The water leach tests following acid bake Tests 11 to 17 were conducted at ambient temperature (feed solutions around 20 °C) in sealed flasks with magnetic stirrers. After 120 minutes the

SPCN-995859

2014250661 16 Oct 2014 contents were treated as per the subsample (paragraph [0090] above). The results are presented in Tables 12 to 14.

Table 12 - SAPL and AB/WL Residue Solid Compositions for Tests 9 to 17

Test	Stage	98% H2SO4 kg/t	Sample	Mass Ratio1 w/w	Solids Composition (% w/w)
LRE	MRE	HRE	Y	TREE	P
9	Feed		Batch 2	1	5.62	0.24	0.05	0.09	6.00	9.01
	PL 1	482	PLR 1	1.10	4.33	0.18	0.03	0.06	4.60	2.99
	AB	923	WLR	1.10	0.15	0.03	0.01	0.02	0.20	0.47
	PL 2	0	PLR 2	1.12	6.14	0.20	0.03	0.06	6.43	2.69
10	PL 1	629	PLR 1	1.21	3.71	0.15	0.03	0.05	3.94	2.76
	AB	1511	WLR	1.08	0.17	0.02	0.01	0.02	0.22	0.27
	PL 2	0	PLR 2	1.11	4.73	0.18	0.03	0.05	5.00	2.91
11	Feed		Batch 3	1	5.25	0.23	0.04	0.08	5.60	8.88
	PL 1	510	PLR 12	1.02	4.38	0.17	0.03	0.05	4.63	1.68
	AB 1	1530	WLR 1	0.89	0.39	0.009	0.002	0.01	0.41	0.17
	AB 2	1515	WLR 2	0.88	0.26	0.014	0.007	0.01	0.29	0.19
12	AB 1	1378	WLR 1	0.89	0.39	0.009	0.003	0.01	0.41	0.19
	AB 2	1362	WLR 2	0.95	0.20	0.013	0.006	0.01	0.23	0.19
13	AB 1	1225	WLR 1	0.91	0.33	0.010	0.003	0.01	0.35	0.21
	AB 2	1214	WLR 2	0.95	0.25	0.012	0.005	0.01	0.28	0.22
14	AB 1	1122	WLR 1	0.92	0.21	0.012	0.006	0.01	0.23	0.22
	AB 2	1122	WLR 2	0.95	0.15	0.015	0.008	0.02	0.19	0.23
15	AB 1	1020	WLR 1	0.91	0.15	0.013	0.006	0.01	0.17	0.23
	AB 2	1016	WLR 2	0.98	0.17	0.015	0.007	0.01	0.20	0.28
16	AB 1	816	WLR 1	0.95	0.15	0.013	0.007	0.01	0.17	0.27
	AB 2	813	WLR 2	1.03	0.20	0.024	0.009	0.01	0.24	0.42
17	AB 1	612	WLR 1	0.91	0.17	0.017	0.009	0.02	0.21	0.33
	AB 2	613	WLR 2	0.88	0.46	0.052	0.013	0.02	0.55	0.53

¹ Ratio of residue (Pre-Leach or Water Leach) mass to that of concentrate (mass change) ² Common Sulphuric Acid Pre-Leach residue used as feed into the first Acid Bake Water Leach for Tests 11 through to 17 inclusive.

Table 13 - Water Leach and Pre-Leach Solution Compositions for Tests 8 to 16

Test	Stage	Liquor Composition (mg/L or g/L)
FA1 g/L	LRE g/L	MRE	HRE	Y	TREE g/L	P g/L	S g/L	Al g/L	Fe g/L	Th g/L	U
9	AB	135	15.5	557	73	117	16.3	10.9	54.4	1.78	1.36	1.41	112
	PL 2	11	9.07	447	91	161	9.77	36.6	18.5	3.46	3.22	1.04	140
10	AB	198	11.4	313	38	60	11.9	8.50	87.0	3.90	3.08	0.95	74
	PL 2	79	9.89	421	70	120	10.5	21.8	43.4	4.68	4.30	1.33	101

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2014250661 16 Oct 2014

11	AB 1	259	8.06	282	40	62	8.44	4.24	93.4	1.88	2.28	0.77	66
	PL 2	192	7.90	351	65	113	8.43	19.9	72.4	2.16	2.62	1.14	93
12	AB 1	236	8.64	303	44	69	9.05	4.28	85.4	1.52	2.08	0.90	68
	PL 2	152	8.33	360	67	119	8.88	21.8	58.6	1.78	2.32	1.17	92
13	AB 1	201	8.73	310	44	69	9.16	4.28	74.2	1.48	1.98	0.88	70
	PL 2	121	8.95	385	70	121	9.52	22.4	49.2	1.76	2.36	1.19	91
14	AB 1	170	9.79	348	46	74	10.3	4.36	64.2	1.32	1.72	0.90	74
	PL 2	90	9.99	422	72	131	10.6	23.6	39.0	1.60	2.04	1.15	98
15	AB 1	151	9.82	340	42	71	10.3	4.26	57.6	1.06	1.46	0.89	73
	PL 2	77	9.85	409	67	117	10.4	20.6	33.8	1.28	1.70	1.11	93
16	AB 1	112	10.7	362	45	71	11.1	4.24	44.6	0.70	1.00	0.92	76
	PL 2	46	10.6	426	64	114	11.2	18.8	23.2	0.90	1.22	1.15	95
17	AB 1	71	10.2	383	46	73	10.7	3.58	28.8	1.46	1.60	0.88	78
	PL 2	6.5	10.2	430	63	110	10.8	13.5	12.6	1.68	1.82	0.96	96

¹ Calculated free H₂SO₄. Calculated by assuming that protons make up the difference in normality between cations and anions from solution assay. Protons are first assigned to P to make H₃PO₄, with the surplus assigned as free H₂SO₄.

Table 14 - Summary of Overall RE Extraction for Combined SAPL/AB/WL Tests 9 to 17¹

Test	Stage	98% H2SO4 kg/t2	Overall Extraction (%)
LRE	MRE	HRE	Y	TREE	P	Al	Fe	Th	U
9	PL 1	482	14.9	17.2	24.1	27.7	15.3	63.4	14.8	26.2	29.7	23.8
	AB	923	97.0	87.5	76.2	81.5	96.3	94.3	10.5	27.2	98.6	98.4
	PL 2	0	-22.2	4.6	24.5	33.4	-19.9	66.6	-7.0	8.7	-12.2	19.7
10	PL 1	629	20.3	25.1	30.8	33.3	20.8	63.1	13.4	21.4	34.2	12.9
	AB	1511	96.7	90.9	82.1	76.5	96.0	96.8	14.3	32.0	98.2	98.2
	PL 2	0	6.2	14.1	26.2	39.2	7.2	64.0	8.3	18.9	20.6	18.8
11	PL 2	0	-1.3	13.8	25.1	29.4	0.0	81.3	2.5	6.5	36.7	29.5
	AB 2	1515	95.6	94.6	87.0	88.6	95.4	98.2	27.3	44.3	99.5	98.5
12	PL 2	0	-2.7	11.1	23.5	28.2	-1.5	89.9	3.2	4.5	26.0	25.6
	AB 2	1362	96.4	94.7	87.3	87.8	96.1	98.0	23.6	39.6	99.5	98.8
13	PL 2	0	1.8	14.6	26.1	29.6	2.9	91.9	3.4	7.0	29.6	23.1
	AB 2	1214	95.5	94.9	89.0	87.8	95.3	97.7	20.5	33.1	99.5	98.5
14	PL 2	0	1.6	14.4	26.0	31.7	2.8	97.2	3.4	5.9	24.1	25.6
	AB 2	1122	97.4	93.9	83.8	75.6	96.8	97.5	26.1	41.5	99.5	98.5
15	PL 2	0	0.7	13.7	24.7	26.0	1.8	82.6	2.8	4.6	21.9	21.2
	AB 2	1016	96.9	93.6	83.6	87.4	96.6	97.0	21.3	31.1	99.5	98.4
16	PL 2	0	-0.1	12.3	19.0	23.7	0.9	72.5	2.4	4.0	21.9	19.6
	AB 2	813	96.2	89.3	78.6	86.7	95.6	95.1	16.1	22.4	99.4	97.7
17	PL 2	0	0.7	9.7	17.5	21.7	1.5	48.7	2.8	4.2	8.1	19.9
	AB 2	613	92.3	80.1	73.4	77.4	91.5	94.8	37.5	37.5	99.5	95.1

¹ Extractions based on concentrate and residue solids assays with the exception of Pre-Leach Tests 11 to 17 which are based on Leach Solution (change) and concentrate feed.

² Based on concentrate.

SPCN-995859

2014250661 16 Oct 2014 [0120] Batch 2 concentrate was used leached in Tests 9 and 10 under similar temperature conditions. The only material difference in conditions between these tests was the acid addition. Following on from an acid bake with 1511 kg/t of acid, the counter-current pre-leach (PL2) in Test 10 resulted in 7% overall TREE extraction, while in Test 8 pre-leach gave -20% overall TREE extraction (indicative of REE precipitation from the pre-leach solution) after an acid bake with 923 kg/t acid. Thorium also precipitated in Test 9 suggesting that there was insufficient surplus sulphuric acid to stabilise the solution resulting in metal phosphate precipitation.

[0121] It can also be seen from Table 12 that Test 17 had less residual sulphuric acid than Test 9 without the significant rare earth (and thorium) precipitation. However, Test 17 pre-leach was conducted from an ambient initial temperature (about 20 °C prior to the heat of reaction causing the temperature to rise) compared to a 35 °C initiation temperature in Test 9. Therefore, it appears that lower temperatures in the pre-leach allow the rare earths to persist in solution with low residual acid.

[0122] It can be seen from Tests 11 to 17 that high overall extractions of total rare earths (95-97%) can be achieved in the counter-current configuration over a wide range of total acid additions from 613 to 1515 kg/tonne of concentrate. There is some variation in phosphate extraction over this range, with a significant drop of in extraction at the lower overall acid additions, while aluminium, iron, thorium and uranium extractions are relatively consistent over the full range of acid additions. The optimum condition would be an economic trade-off between revenue and costs which may vary with time.

Double Sulphate Salt Precipitation (DSP) - Reagent Dosage (Tests 18 to 20) [0123] The double sulphate precipitation process is known to be best for the precipitation of LRE and MRE and less preferred for precipitation of the HRE. It is also known to give a good separation of the rare earths from most of the impurity elements such as aluminium, iron, phosphorus and uranium. Calcium and thorium, however, tend to follow the rare earths into the precipitate.

[0124] Double sulphate precipitation tests were conducted at greater than 95°C, in a sealed and baffled glass reactor vessel. The vessel was agitated using a Rushton

SPCN-995859

2014250661 16 Oct 2014 turbine impeller driven by an overhead stirrer at approximately 400 rpm. A condenser was fitted to the vessel to minimise evaporation. A sensor inserted into the slurry recorded the temperature of the pulp, and the vessel was heated by a temperaturecontrolled hot plate. Test feed solutions varied from 0.6 to 1.2 L. After the test solution reached the desired temperature, a calculated mass of anhydrous sodium sulphate dry salt was added to the reactor within 1 -5 minutes. The sodium sulphate addition was calculated based on the stoichiometric relationship given in Equation (3) above and the quantity of TREE in solution. The precipitate formed immediately upon the addition of sodium sulphate.

[0125] After two hours, the slurry was weighed and filtered. The solution was collected for analysis and the solids were washed, firstly by repulping the residue using a 2.5-5% (w/w) solution of sodium sulphate followed by repulping twice with deionised water. The wash solutions were also analysed to perform an overall elemental mass balance. The washed solid was dried to a constant mass in an oven at 105°C, and a sub-sample was submitted, together with the filtrate, for analysis.

[0126] Three DSP tests were carried out using water leach liquor produced from tests similar to that described above, but varying the dose of sodium sulphate from a Na:TREE molar ratio of 6 to 8 to 12. All three tests were conducted at >95°C using dry sodium sulphate powder, to minimise solution dilution, for a two hour duration. The feed liquor composition is given in Table 15 and the percentage precipitations of the rare earths and impurity elements are presented in Table 16.

Table 15 - Water Leach Liquor Composition for feed to DSP tests

Free Acid g/L	Solution Composition (g/L or mg/L)
LRE g/L	MRE mg/L	HRE mg/L	Y mg/L	TREE g/L	Al g/L	Ca g/L	Fe g/L	P g/L	S g/L
251	11.4	313	38	60	11.9	3.90	1.31	3.08	8.50	87.0

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2014250661 16 Oct 2014

Table 16 - Summary of Percentage Precipitation Data for DSP tests

Test	Na: TREE Molar Ratio	% Precipitation
LRE	MRE	HRE	Y	HRE+Y	TREE	Al	Ca	Fe	P	s	Th	U
18	12	100	95.5	63.1	33.2	44.5	98.6	0.0	60.7	0.0	0.0	9.9	61.1	0.0
19	8	99.4	93.9	58.5	20.9	35.3	98.2	0.0	52.3	0.0	0.0	9.4	54.2	0.0
20	6	99.3	92.2	54.2	15.5	30.5	98.0	0.0	48.5	0.0	0.0	8.3	49.8	0.0

[0127] The results show that the LRE recovery was virtually complete at the lowest Na:TREE molar ratio of 6. However, the MRE and HRE recoveries at this dose were only 92% and 54%, respectively. Doubling the reagent dose increased the respective recoveries to 96% and 63%. Yttrium precipitation rose from 16 to 33%. The preferred reagent dose will therefore depend on the value of the rare earths precipitated, especially the HRE and the cost of the reagent used. The precipitation of the impurities calcium and thorium also increased with reagent dose. However, the extent of precipitation of the impurities aluminium, iron, phosphorus and uranium was virtually zero.

[0128] A person skilled in the art will understand that there may be variations and modifications other than those specifically described. It is to be understood that the invention includes all such variations and modifications. The invention also includes all steps, features, compositions and compounds referred to, or indicated in this specification, individually or collectively, and any and all combinations of any two or more of the steps or features.

Claims (17)

1. The claims defining the invention are as follows:

   1. A process for recovering rare earth elements from a composite ore or concentrate that includes at least apatite group minerals, as well as at least one other rare earth containing mineral, the process including:

   a. subjecting the ore or concentrate to a sulphuric acid pre-leach to form a pre-leach liquid phase and a pre-leach solid phase, the pre-leach liquid phase containing dissolved phosphates and dissolved rare earth elements, and the pre-leach solid phase containing sulphuric acid-insoluble minerals and calcium sulphate precipitate;

   b. mixing the pre-leach solid phase with sulphuric acid and baking the mixture at a temperature in the range of 200 to 300°C to produce an acid-baked solid phase; and

   c. subjecting the acid-baked solid phase to a water leach, in water and/or in the pre-leach liquid phase, to dissolve rare earth elements in the acidbaked solid phase to form a water-leach liquid phase and a water-leach solid phase; wherein the water-leach liquid phase is at least partly recycled to the pre-leach and rare earth elements are recovered from the pre-leach liquid phase; and/or the water-leach liquid phase is not recycled to the pre-leach and rare earth elements are recovered from the water-leach liquid phase.
2. 2. A process according to claim 1, wherein the concentration of sulphuric acid used in the pre-leach is such as to maximise the dissolution of the apatite but minimise the dissolution of other acid soluble minerals that contain impurity elements such as iron and aluminium.
3. 3. A process according to claim 1 or claim 2, wherein the sulphuric acid used in the pre-leach has a concentration in the range of from 5% to 40% w/w.
4. 4. A process according to any one of claims 1 to 3, wherein the concentration of the sulphuric acid in the pre-leach is such that the solids concentration in the leach is from 20% to 40% w/w.

   SPCN-995859
5. 5. A process according to any one of claims 1 to 4, wherein the temperature in the pre-leach does not rise above 60°C.
6. 6. A process according to any one of claims 1 to 5, wherein the pre-leach is limited to being between 15 minutes and 60 minutes.
7. 7. A process according to any one of claims 1 to 6, wherein the pre-leach solid phase is mixed with a sulphuric acid solution of greater than 90% w/w sulphuric acid, prior to the acid bake.
8. 8. A process according to any one of claims 1 to 7, wherein the amount of sulphuric acid mixed with the pre-leach solid phase prior to the acid bake is at least the stoichiometric requirement for sulphation of the rare earth minerals.
9. 9. A process according to claim 8, wherein the amount of sulphuric acid mixed with the pre-leach solid phase prior to the acid bake is from 500 to 4,000 kilograms of sulphuric acid per tonne of residue.
10. 10. A process according to any one of claims 1 to 9, wherein the acid bake is limited to being between 30 minutes and 4 hours.
11. 11. A process according to any one of claims 1 to 10, wherein the temperature in the water leach does not rise above 60°C.
12. 12. A process according to any one of claims 1 to 11, wherein the concentration of rare earth sulphates in the water-leach liquid phase is from 5 to 30 g/L.
13. 13. A process according to any one of claims 1 to 12, wherein the rare earth elements recovered are recovered by precipitation as a sodium rare earth double sulphate salt.
14. 14. A process according to any one of claims 1 to 13, wherein the water-leach liquid phase is recycled to the pre-leach and rare earth elements are recovered from the pre-leach liquid phase.
15. 15. A process according to any one of claims 1 to 13, wherein the water-leach liquid phase is not recycled to the pre-leach and rare earth elements are recovered from the water-leach liquid phase.

    SPCN-995859

    2014250661 16 Oct 2014
16. 16. A process according to claim 15, wherein water is added as required for the water leach.
17. 17. A process according to any one of claims 1 to 16, wherein the water-leach solid phase is neutralised prior to disposal.