Recovery of Rare Earth Elements

Technical Field

The present disclosure relates to processes for treating ores containing rare earth elements and to processes for recovering rare earth elements from refractory minerals.

Background

It has been predicted that there will be an under supply of the critical rare earth elements ("REE"s) by 2015. This shortage will have a dramatic effect on emerging energy efficiency applications, as well as for technology and military applications. The shortage is being driven by reduced supply from China, which has 95% of global REE reserves.

Although alternative sources of supply are being discovered, one of the problems of bringing new mines into production is the combined difficulty of beneficiation of the ore and the separation of high purity fractions of the REEs. This problem is particularly relevant to laterite-hosted mineralization, where the REEs have been efficiently concentrated to ore grades by weathering processes and movement of groundwater.

However, extraction of the REEs has proven to be technically challenging from both laterite ores and unweathered sources, such as alkaline igneous intrusions that have not been subjected to subtropical weathering. Current processes require the use of large volumes of expensive and/or hazardous acid such as HF, H2SO4 and HCIO4. This problem of beneficiation and extraction of the REEs is primarily caused by the refractory (difficult to dissolve) nature of many REE-bearing minerals. This problem can be exacerbated by the mineralogical complexity of the ores and by the very fine- grained character of REE minerals in laterite profiles.

Similar difficulties exist for the recovery of other so-called specialty metals such as niobium, tantalum, zirconium and hafnium from ores. Summary of the Disclosure

It is to be understood that throughout this specification, the following terms have the below stated meanings, unless specified otherwise:

"lanthanides" means the elements having atomic numbers 57 to 71 (lanthanum to lutetium)

"rare earth elements" means one or more of the lanthanides, yttrium (atomic number 39), scandium (atomic number 21 ), zirconium (atomic number 40), hafnium (atomic number 72), niobium (atomic number 41 ) and tantalum (atomic number 73). Although some of these elements are not lanthanides they display similar geochemical properties and/or are commonly concentrated in lanthanide containing ores. Therefore, for the purposes of this specification they are considered to be rare earth elements.

"light rare earth elements" means the elements having atomic numbers 57 to 62, that is lanthanum (La), cerium (Ce), praseodymium (Pr), neodymium (Nd), promethium (Pm) and samarium (Sm)

"heavy rare earth elements" means the elements having atomic numbers 63 to 71 , that is europium (Eu), gadolinium (Gd), terbium (Tb), dysprosium (Dy), holmium (Ho), erbium (Er), thulium (Tm), ytterbium (Yb) and lutetium (Lu). According to a first aspect, the present invention provides a process for treating an ore containing at least one refractory mineral comprising a rare earth element, the process comprising the steps:

(i) combining the ore with a solid flux to form a mixture;

(ii) heating the mixture formed in step (i) to form a fused mixture;

(iii) cooling the fused mixture to form a fused solid; and

(iv) treating the fused solid with an acidic solution to dissolve the rare earth element.

The ore may comprise primary and/or primary secondary refractory minerals. In one embodiment, the ore comprises a laterite ore. The laterite ore may have secondary rare earth element bearing minerals together with residual primary minerals (such as pyrochlore) that that have been concentrated during the weathering process. In another embodiment, the ore comprises rare earth element bearing minerals hosted in alkaline igneous rocks. The processes of the present disclosure are particularly applicable to the recovery of rare earth elements from laterite ores, which has previously proven to be difficult. The ore may comprise any one or more of aeschynite, allanite (orthite), anatase, ancylite, apatite, bastnasite, brannerite, britholite, cerianite, cheralite, churchite, eudialyte, euxenite, fergunsonite, florencite, gadolinite, huanghoite, hydroxylbastnasite, kainosite, loparite, monazite, mosandrite, parasite, samarskite, synchisite, thalenite, xenotime and yttrotantalite. When the ore is a niobium or tantalum ore it may comprise any one or more of coltan, columbite, euxenite, fergusonite, polycrase, pyrochlore, samarskite -(Y), simpsonite, tantalite, zirkelite and wodginite.

When the ore is a zirconium or hafnium ore it may comprise any one or more of zircon, zirconolite and baddeleyite. The ore may comprise tailings from a metallurgical process, such as a hydrometallurgical process.

Preferably the ore has a particle size less than 200pm, 150pm, 125pm, 100pm, 75pm, 50pm, 30pm or 20pm, more preferably less than 10pm, even more preferably between 5pm and 10pm. In this regard, the particle size of the ore is preferably equivalent or approximately equivalent to the liberation size of the rare earth element so that it is available to contact with the solid flux.

For the purposes of the present invention, reference to a particular "particle size" means that at least 80% of the particles are of the recited size or smaller.

For the purposes of the present invention it will be understood that the term "ore" includes "ore concentrates" produced from ore that has been treated to increase the concentration of the rare earth element. The ore concentrates may have a minimum grade of 30wt% rare earth element, preferably 40wt% rare earth element, more preferably greater than 50wt% rare earth element. Ore concentrate may be formed by subjecting ore, such as comminuted ore, to magnetic separation, sorting, classification, gravity and/or flotation.

Preferably, the rare earth element is a lanthanide.

Preferably, the rare earth element is zirconium, hafnium, niobium, tantalum, yttrium or scandium.

The flux may comprise one or more of an alkali metal oxide, an alkali metal hydroxide, or an alkali metal borate or perborate. Examples of fluxes include sodium peroxide, sodium hydroxide, lithium tetraborate, sodium perborate or lithium metaborate or mixtures thereof. Preferably, the flux comprises sodium peroxide, sodium hydroxide or a mixture thereof.

In one form, the ratio of flux to ore or ore concentrate is a minimum of 1 :1 , 2:1 or 3:1 , preferably at least 4:1 , 5:1 , 6:1 , 7:1 or 8:1. Without wishing to be bound by theory, it is understood that, generally, the greater the relative amount of flux that is mixed with the ore in step (i), the less energy is required to fully fuse the mixture in step (ii) because a lower heating temperature is required.

In one form, the heating step (ii) is conducted at a temperature of at least 300°C, preferably no more than 600 or 65CTC, more preferably 300°C - 650°C, 350°C - 600°C, 400°C - 600°C, 450°C - 550°C, 300°C - 450°C, 300°C - 400X, 350°C - 450°C, 350°C - 400°C, 450°C - 500°C or 450°C - 650°C. The heating step (ii) may be conducted in a nickel, graphite or platinum receptacle, such as crucible. Alternatively the heating step may be conducted in a rotary kiln or a fluidized bed reactor. The heating step may also comprise the use of an intermediate heat transfer unit or other means to recycle heat from the waste streams of the process. The heating step may also further comprise a pre-heating step such that the heating is performed in stages.

The heating step (ii) may also be conducted in a material that is resistant to high temperatures, up to 650°C. Such materials include graphite, alumina and platinum.

The heating step (ii) may be conducted in a reducing or inert atmosphere, for example in an atmosphere of carbon dioxide or an inert gas. The heating step (ii) may be conducted for a range of times depending, at least partially on the heating temperature and whether the intention is to produce a partially or fully fused mixture. Preferably, the heating step (ii) is conducted for 10 - 120 or 180 minutes or for at least 10, 5, 30, 45, 60, 90, 120 or 180 minutes. The fused mixture created in the heating step may be fully or partially fused. In this regard, the purpose of the heating step is to liberate or solubilise the metal values from the ore or ore concentrate. With this in mind, heating for longer and/or at higher temperatures will tend to result in a fuller or more complete fusion of the mixture. However, there are also advantages in heating steps that create partially fused mixtures, provided they do not unduly compromise the efficiency or effectiveness of the process, because these can be carried out at lower temperatures and for less time than heating steps that create fully fused mixtures. Heating steps carried out at lower temperatures and/or for less time require less energy. Partially fused mixtures are formed when there has been sufficient contact between the ore or ore concentrate and the molten flux to solubilise or liberate the rare earth element but insufficient contact to result in the formation of a fully fused mixture consisting essentially of a molten glass phase. In any event, it will be appreciated that, within the ranges of operating parameters recited herein, a skilled person can adapt and manipulate parameters and reaction variables to achieve a mixture with the desired level of fusion.

The fused solid may be fully or partially fused and it will be appreciated that whether the fused solid is fully or partially fused is dependent, at least partially, on whether the fused mixture created in the heating step (ii) is fully or partially fused.

Applicant has surprisingly discovered that partially fused products can be used to efficiently recover the rare earth element as well as fully fused solids. Applicant has also found that partially fused solids are reactive and can confer other advantages on the process such as easier material handling relative to fully fused solids. These material handling advantages include increased service life of plant and machinery and the potential for the partially fused solids to require much simpler downstream processing to recover the rare earth elements. For the purposes of the present invention the term "partially fused solid" means a solid that includes solubilised rare earth element but has not become a fully fused solid. Preferably, a "partially fused solid" does not include a glass phase or only includes trace amounts of a glass phase. Preferably, the partially fused solid comprises a frit, such as a completely fused frit.

For the purposes of the present invention the term "fully fused solid" means a solid that includes solubilised rare earth element and a glass phase. Preferably the glass phase is one or more of homogenous, uniform and/or free of discrete particles. In another form the glass phase is amorphous and/or hard. The cooling step (iii) may comprise exposing the fused mixture to ambient temperature.

The cooling step (iii) may be conducted in a dry nitrogen (N ₂ ) atmosphere. This will enhance the speed with which the fused mixture is cooled and ease the course of further processing. The cooling step (iii) may comprise cooling the fused mixture to a temperature that will not cause boiling of the acidic solution in step (iv) upon contact with the fused solid.

The acidic solution used in the treating step (iv) may comprise a weak hydrochloric acid solution such as 10% HCI or a mixture of hydrochloric and nitric acid (aqua regia). The acidic solution used in the treating step (iv) may also comprise a stronger acid solution of at least 20%-35% such as hydrochloric acid. In this regard, when the heating step (ii) results in a partial fusion, a stronger acidic solution is likely to be required to dissolve the rare earth elements during step (iv).

The treating step (iv) may also comprise dissolving the fused solid in water to form an aqueous solution and acidifying the aqueous solution. In another arrangement, however, the fused solid may be solubilised by an acid such as a weak or strong acid solution. The treating step (iv) may be conducted in a vessel that is resistant to acid attack, such as a suitable plastic vessel for example. In one embodiment, the vessel may be formed from Teflon®.

The rare earth element may be recovered from the ore treated according to the first aspect of the present invention. Thus, according to a second aspect, the present invention provides a process for recovering a rare earth element from an ore containing at least one refractory mineral comprising the rare earth element, the process comprising the steps:

(i) combining the ore with a solid flux to form a mixture;

(ii) heating the mixture formed in step (i) to form a fused mixture;

(iii) cooling the fused mixture to form a fused solid;

(iv) treating the fused solid with an acidic solution to dissolve the rare earth element; and

(v) recovering the rare earth element. Advantageously, the process of the present disclosure results in substantially more efficient rare earth element recovery from the refractory minerals and furthermore requires the use of significantly smaller volumes of acids and/or less energy. In at least some embodiments, the process may achieve up to 80% recovery of the rare earth element from the ore or ore concentrate. Preferably, the process achieves at least 81 %-95% recovery of the rare earth element.

The recovering step (v) may comprise subjecting the dissolved the rate earth element to chromatographic exchange column techniques to recover the rare earth element as chlorides.

The recovering step (v) may also comprise converting the chlorides to oxides. Subsequently, the oxides may be converted to metals or alloys by electrolysis or metallothermic reduction. The metals/alloys may be high purity elemental metals (for example neodymium metal) or may be alloys such as mischmetal or ferro-alloys.

The processes according to the first and second aspect of the present invention may also involve the treatment of ore prior to its combination with the solid flux. Thus, the present invention provides a process for treating an ore containing at least one refractory mineral comprising a rare earth element, the process comprising the steps:

(i) comminuting the ore;

(ii) combining the ore from step (i) with a solid flux to form a mixture; (iii) heating the mixture formed in step (i) to form a fused mixture;

(iv) cooling the fused mixture to form a fused solid; and

(v) treating the fused solid with an acidic solution to dissolve the rare earth element.

The at least one refractory mineral may be fine grained. The present invention also provides a process for recovering a rare earth element from an ore containing at least one refractory mineral comprising the rare earth element, the process comprising the steps:

(i) comminuting the ore;

(ii) combining the ore with a solid flux to form a mixture; (iii) heating the mixture formed in step (i) to form a fused mixture;

(iv) cooling the fused mixture to form a fused solid;

(v) treating the fused solid with an acidic solution to dissolve the rare earth element; and

(vi) recovering the rare earth element. Without wishing to be bound by theory, the grain size of many minerals containing rare earth elements (including pyrochlore, an ore of niobium) in laterite ores are approximately 5-1 Opm in size. Accordingly, those minerals may be liberated by comminuting the ore to approximately the grain size of the minerals, which can substantially enhance recovery of the rare earth element. Preferably, the step of comminuting the ore or ore concentrate results in an ore or ore concentrate with a particle size that equates to the grain size of the mineral containing the rare earth element. Even more preferably, the step of comminuting the ore or ore concentrate results in an ore or ore concentrate with a particle size less than 150μηη, 125μηι, 100μηι, 75μιη, 50μιη, 30μιη θΓ 20μιη, more preferably less than 10pm, even more preferably between 5 m and 10μηι.

The comminution step may be carried out using, for example, the Isamill® process.

The processes described herein may be applied to ore with relatively high levels of unwanted metals such as iron, manganese, uranium, thorium or molybdenum. Thus, the processes may further comprise the step of removing one or more unwanted metals.

The unwanted metals may be removed from the ore prior to the step of combining the ore or ore concentrate with a solid flux. Examples of removal steps suitable at this stage of the method are the use of magnetic separation, flotation and/or gravity separation such as ultra fine gravity separation.

The unwanted metals may also be removed later in the process such as after the step of treating with an acidic solution. An example of a removal step suitable at this stage of the method is the precipitation of the unwanted metals. With particular reference to the removal of iron, a precipitation step involving the adjustment of pH to precipitate the iron could be used.

Other processes for removing the unwanted metals include ion exchange and solvent extraction. Suitable extraction may involve the use of organic solvents such as tributyl phosphate and kerosene or combinations thereof.

In another embodiment of the invention, unwanted metals can be removed during the recovering step. For example, chromatographic exchange column techniques, such as anion and cation exchange columns, can be used to select for the target metal values in preference to the iron or other unwanted metals. Brief Description of the Figures

Embodiments of the present disclosure will now be described, by way of example only, with reference to the accompanying Figures, in which:

Figure 1 A is a flow chart of a process for concentrating rare earth elements from an ore;

Figure 1 B is a flow chart of another process for concentrating rare earth elements from an ore;

Figure 2 is a chromatogram demonstrating how rare earth elements as rare earth element chlorides can be efficiently separated using chromatography; Figure 3 is a graph showing the effect of flux composition on recovery of total REE+Y. Key: diamond = 500°C; square = 600°C; and circle = 650°C;

Figure 4 is a graph showing the effect of flux composition on recovery of Nb. Key: diamond = 500°C; square = 600°C; and circle = 650°C;

Figure 5 is a graph showing the effect of temperature and flux on recovery of total REE+Y. Key: diamond = 100% NaOH; square = 50:50 mixture of NaOH and Na ₂ 0 ₂ ; and circle = 100%Na ₂ O ₂ ;

Figure 6 is a graph showing the effect of temperature and flux composition on recovery of Nb. Key: diamond = 100% NaOH; square = 50:50 mixture of NaOH and Na ₂ 0 ₂ ; and circle = 100%Na ₂ O ₂ ; Figure 7 is a graph showing the effect of flux composition on recovery of total REE+Y with a heating step carried out at 650°C;

Figure 8 is a graph showing the effect of flux composition on recovery of Nb with a heating step carried out at 650°C;

Figure 9 is a graph showing the recovery of total REE+Y using a heating step of 650°C and a 10% HCI lixiviant. Key: square = 100% NaOH; diamond = 50:50 mixture of NaOH and Na ₂ 0 ₂ ; and triangle = 100%Na ₂ O ₂ ; Figure 10 is a graph showing the recovery of Nb using a heating step of 650°C and a 10% HCI lixiviant. Key: square = 100% NaOH; diamond = 50:50 mixture of NaOH and Na ₂ (-½; and triangle = 100%Na2O ₂ ;

Figure 11 is a graph showing the recovery of total REE+Y and Nb using a sample:flux ratio of 1 :6 and varying volumes of 32% HCI lixiviant. Key: square = Nb; diamond = total REE + Y;

Figure 12 is a graph showing the recovery of total REE + Y and Nb using a sample flux ratio of 1 :30 and varying volumes of 32% HCI lixiviant. Key: square = Nb; diamond = total REE + Y; Figure 13 is a graph showing the recovery of total REE+Y, Nb and Fe across a range of sample:flux ratios (1 :0 - 1 :12). Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 14 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of flux compositions with a sample to flux ratio of 1 :4 and heating for 1 hour at 650°C. Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 15 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of heating times with a sample to flux ratio of 1 :6 and heating at 650°C. Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 16 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of heating temperatures with a sample to flux ratio of 1 :6 and heating time of 1 hour. Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 17 is a graph showing the recovery of Fe, total REE+Y and Nb using two different crucible materials and with heating at 650°C for 60 minutes across a range flux:sample ratios. Key: square/dotted line (lowermost) = Nb/graphite; square/solid line = Nb/ceramic; triangle/solid line = total REE+Y/ceramic; circle/dotted line = total REE+Y/graphite; square/dotted line (uppermost) = Fe/graphite; diamond/solid line = Fe/ceramic;

Figures 18A-18D are photographs of the fused solid from the ceramic crucibles for sample:flux ratios of 1 :1 , 1 :6, 1 :8 and 1 :12, respectively. Figure 19 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of heating temperatures with a sample to flux ratio of 1 :6 and heating time of 15 minutes. Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 20 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of sample:flux ratios with a heating temperature of 400°C and heating time of 15 minutes. Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 21 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of heating times with a heating temperature of 400°C and 1 :1 sample:flux ratio. Key: square = Nb; triangle = total REE+Y and diamond = Fe; Figure 22 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of heating times with a heating temperature of 400°C and 1 :6 sample:flux ratio. Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 23 is a graph showing the recovery of Fe, total REE+Y and Nb across a range of heating times with a heating temperature of 350°C and 1 :6 sample:flux ratio. Key: square = Nb; triangle = total REE+Y and diamond = Fe;

Figure 24 is a graph showing the recovery of Fe, total REE+Y and Nb across a range sample:flux ratios (using a 50:50 mixture of sodium hydroxide:sodium perborate as the flux) with a heating time of 15 minutes and a heating temperature of 450°C. Key: square = Nb; triangle = total REE+Y and diamond = Fe; Figure 25 is a graph comparing the recovery of Fe, total REE+Y and Nb using a 50:50 mixture of sodium hydroxide:sodium peroxide (dotted lines) or a 50:50 mixture of sodium hydroxide:sodium perborate (solid lines) flux at a sample:flux ratio of 1 :6 and across a range of heating temperatures for a heating time of 15 minutes. Key: square = Nb; triangle = total REE+Y and diamond = Fe; Figure 26 is a graph demonstrating the selective extraction of Fe from liquor containing total REE+Y, Nb and Fe using 100% TBP across a range of organic/aqueous ratios. Key: square = Fe; cross (lowermost) = total REE+Y; and cross (uppermost) = Nb; and Figure 27 is a graph demonstrating the selective extraction of Fe from liquor containing total REE+Y, Nb and Fe using 50% TBP in kerosene across a range of organic/aqueous ratios. Key: square = Fe; triangle = total REE+Y; and diamond = Nb. Detailed Description of Embodiments

Referring to the Figures, embodiments of the present disclosure provide processes for the recovery of one or more metal values from one or more refractory minerals in ore. The processes are especially advantageous in recovering rare earth elements that have previously proven difficult to recover from refractory minerals, such as those minerals found in laterite ores

The processes may be applied to any rare earth element containing ores including ores having any one or more of the minerals listed in Table 1 below.

Minerals that contain rare earth elements and occur in economic or potentially economic deposits (Ln = lanthanides)

In general, the process according to embodiments of the present disclosure involves the following steps: (1 ) Comminuting the ore to liberate metal value containing minerals, in particular minerals containing specialty metal values, from composite grains.

(2) Separating the target metal value containing minerals from the bulk to form an ore concentrate by techniques such as magnetic separation and/or flotation.

(3) Combining the ore concentrate with a solid flux and fusing this mixture at elevated temperatures. The fused mixture is cooled to produce a fused solid or glass that incorporates all of the mineral phases in the ore concentrate, even the refractory minerals.

(4) The fused solid or glass is subsequently dissolved in a weak acid.

(5) Once into solution, chromatographic exchange techniques are used to separate fractions of the rare earth elements as chlorides.

(6) The chlorides can be sold or alternatively can be refined further and sold as oxides or as metals or alloys.

It is to be appreciated, however, that in some embodiments, the process may begin with an ore or ore concentrate that is already sufficiently comminuted and/or concentrated to enable steps (3)-(6) to be conducted thereon. In such embodiments, one or both of the steps of comminution (1 ) and separation (2) need not be performed.

In some embodiments, the ore or ore concentrate from which the rare earth elements are recovered is the tailings from another metallurgical process. For example, the ore or ore concentrate may be tailings from nickel recovery process from a laterite ore developed over ultramafic rocks which often contains significant values of scandium for example. Depending on the condition of the tailings, the tailings may be subjected to the comminution step (1 ) and/or the separation step (2) above or neither.

Example 1 A - Concentration of Laterite Ore

Figure 1A provides a schematic of a process, generally indicated by the numeral 10, for concentrating the rare earth elements, in a laterite ore. Crushing

Four tonnes of ore is trucked 12 to a hopper 14 and then crushed and screened 16 to produce a particle size distribution of 100% passing 20 mm. The sub 20 mm material is fed into the grinding circuit 18. Grinding

The grinding circuit 18 consists of a single ball mill in closed circuit to produce a target grind size of at least 38 pm. The fine grind is required to readily enable separation of the valuable rare earth elements from the other minerals in the ore prior to flotation. In one example, an Isamill® grinding process is used to achieve a target grind size of 5-10pm. This provides even better separation of the rare earth elements because the grain size of many rare earth elements (such as niobium) crystals in laterite ores are approximately 5-10pm in size. Furthermore, this will substantially reduce the amount of flux required for the fusion process described below. As a result, smaller volumes of acid will be required to dissolve all of the specialty metal bearing minerals and produce a solution of appropriate normality for chromatographic separation of the rare earth elements described below.

Low intensity magnetic separation

To reduce the volume of material entering the flotation circuit 20 (described below) the bulk material from the grinding circuit 18 is first subjected to magnetic separation to separate out iron (Fe) and manganese (Mn) oxide laterite phases. It is noted that the particle size exiting the grinding circuit 18 needs to be sufficiently small for the magnetic separation to be successful. If the particles entering the magnetic separation unit are too big then rare earth minerals may be pulled out with the Fe and Mn oxide phases.

In one form, the magnetic separation involves flowing the material under gravity through a vertical column. Magnetic fields are generated at the sides of the column, pulling the Fe and Mn oxide particles to the side whilst the rare earth elements are collected at the bottom of the column. The column is preferably filled with a liquid such as water to slow the flow of minerals through the column and enable the magnetic field to act on the particles. As an alternative or in addition to the use of magnetic separation, unwanted metals such as Fe and Mn could be removed at later stages of the process by (pH) precipitation and/or chromatography.

Flotation

The ore from the grinding circuit 18 is then added to water to form a pulp and subjected to froth flotation 20 to further separate the rare earth element -bearing minerals from the bulk.

Suitable chemical additives are added to the pulp to make the rare earth elements hydrophobic. Once hydrophobic, the rare earth element minerals will attach themselves to air bubbles that are bubbled through the pulp and are lifted from the pulp. The mineralised froth that overflows from the top of the flotation cell will then be collected. Suitable chemicals can also be added to the pulp to target the non-rare earth element minerals to make them hydrophilic and hence be depressed, sinking to the bottom and thus allowing them to be extracted to the tailings dam. It is expected that approximately 70% of the rare earth element bearing minerals are recovered in the flotation circuit to produce a flotation concentrate of at least 40wt% rare earth oxides, niobium oxide and/or other specialty metal oxides.

In some embodiments, the flotation circuit is a two-circuit process in which the first circuit of flotation termed roughing is followed by several stages of scavenging. The objective of the rougher/scavenger circuit is to recover as much of the rare earth elements as possible.

The second circuit of the flotation process termed cleaning involves adding depressants to the concentrate recovered in the rougher/scavenger circuit to remove tailings entrained in the froth. It is expected that this will achieve a target concentrate grade of greater than 40wt% rare earth oxides, niobium oxide and/or other specialty metal oxides.

Dewate ing prior to fusion

The final flotation concentrate is pumped into a tank of suitable volume where as a result of the addition of hydrated lime and flocculent, the solids will settle. This will cause the solid particles to form floccs that are heavy and thus settle to the bottom of the thickener cone. The thickened pulp will then be collected using pressure filtration 22. This yields a cake 24 that following desiccation to remove moisture is then subjected to the fusion process described below. The starting four tonnes of ore yields one tonne of concentrate for further treatment.

Tailings Storage

Following common practice the flotation tailings are stored in a tailings dam where the supernatant liquid will be collected and pumped back to the plant for recycling.

Flux addition and Fusion

The concentrated ore 24 produced in the above grinding and flotation processes is then mixed with a solid flux of sodium peroxide and/or sodium hydroxide. Other suitable alkali oxides and/or alkali hydroxides or lithium tetraborate or lithium metaborate may be used as the solid flux. However, sodium peroxide and sodium hydroxide are preferred because of their low fusing temperatures. The amount of solid flux added is in a flux to concentrate weight to weight ratio of at least 3:1 , but preferably at least 4:1 or 6:1. Generally, the more flux that is added the lower the fusion temperature of the mixture and hence the less energy required in the fusion step. Accordingly, the amount of flux added is generally a balance between the cost of the flux and the cost of energy. Using sodium peroxide and/or sodium hydroxide as the flux, the mixture is heated to between 450°C and 500°C in a graphite, nickel or platinum crucible. For cost effectiveness, a graphite crucible is preferred. This heating fuses the mixture to produce a fused solid in the form of a glass that incorporates all of the mineral phases in the concentrate, even the refractory specialty metal minerals. In some examples, the fusing step may be carried out in a reducing (carbon dioxide) or inert (inert gas such as nitrogen) atmosphere. Where heating up to at least 650°C is required, crucibles of highly temperature resistant material may be utilised.

The molten glass is then quenched by exposure to ambient temperature to solidify the molten glass. This may involve simply exposing the molten glass to ambient atmospheric air or in some embodiments may involve applying a vapour layer of dry nitrogen (N ₂ ) gas to the molten glass for more rapid cooling. The glass is cooled to a sufficient temperature that will not result in boiling of the acid in the subsequent process step described below.

Dissolving Glass

The glass is harvested and transferred to an acid resistant dissolution vessel (formed of a suitable plastic material such as Teflon®) where it will be dissolved in a weak acid.

The acid used is preferably a weak hydrochloric acid, but may be a mixture of hydrochloric acid and aqua regia (nitric acid). Notably, the hydrochloric acid converts the rare earth elements into rare earth chlorides, which will suitably enable separation of the individual rare earth elements by chromatographic separation.

Chromatographic separation of rare earth elements

Chromatographic separation is then applied to aqueous acid solution of the specialty metals to separate the entire spectrum of rare earth elements. Chromatographic separation is particularly advantageous because it enables separation of the light rare earth elements (La, Ce, Pr, Nd and Sm) as well as the high value heavy rare earth elements (Eu, Gd, Tb, Dy, Ho, Er, Tm, Yb, and Lu) and yttrium (Y) to purities of >99.99wt%.

The percentage extraction of rare earths decreases in the order: Lu3+>Yb3+>Tm 3+> Er3+>Y 3+>Ho 3+ >Dy 3+ and so on down to La. The separation of these elements generally increases with atomic number. This is likely the result of the increase in strength of the electrostatic interaction between the eluant anion and the rare earth element cation as the size of the latter progressively decreases.

Importantly, the Y3+ elution peak lies between that of Ho3+ and Er 3+ as would be expected on the basis of its ionic radius. Figure 2 is a chromatogram of the rare earth elements (as chlorides) and demonstrates how the rare earth elements can be readily separated using this technique. It is noted that Figure 2 shows the time taken to elute the various rare earth elements under gravity feed. The elution speed could be significantly increased by conducting the chromatographic separation under elevated pressure, such as by the application of nitrogen gas to the columns. Alternatively, simply increasing the molarity of the HCI will also increase the speed of recovery of the different heavy rare earth elements.

Laboratory scale separation

The following is a description of laboratory scale chromatographic separation of rare earth elements.

The solution containing the rare earth elements is loaded onto a First Column being a cation-exchange column of 3mL AG 50WX8 resin (200-400 mesh) packed in a polyethylene tube (inner diameter 7 mm; height 80 mm).

Prior to loading the column was back-washed with 60 ml 6M HCI and preconditioned with 4M HCI to remove any memory effects. The major elements are then eluted with 4M HCI in the first fraction of 10mL This fraction also contains some heavy rare earth elements. With careful column calibration and adjusting the normality of the eiuant as required, the heavy rare earth elements can also be recovered for further processing. The next 20mL fraction contains the light rare earth elements and barium (Ba).

If the ore concentrate contains nickel and scandium these elements can be separated as follows. Nickel is separated first by eluting with 1 M NaCI. However, scandium is tightly bound to the resin and remains in the column. Scandium is finally eluted using 1.5 M H ₂ S0 ₄ (sulphuric acid) where it is dried and collected as a sulphate.

The light and heavy rare earth element fractions obtained are combined and evaporated to dryness then, re-dissoived in a small volume of very weak 0.3M HCI.

This redissolved solution is then loaded onto a Second Column for separation of the individual rare earth elements. The preparation of the Second Column involves the following process: 100 mg HDEHP [di(2-ethylhexyl) orthophosphoric acid] is mixed with 1 g teflon powder (Voltalef 300LD PL micro). An acetone-water mixture is then added to keep the teflon/HDEHP mixture moist. When the acetone evaporates the mixture is then transferred to a teflon tube fritted with a teflon wool (TFE shrinkable teflon tube; inner diameter after shrinkage is 5mm). In a 24cm high column the flow rate is 3 drops/minute. The column is cleaned by back-washing with 20mL 6M HCI and then is preconditioned with 0.3 HCI. The rare earth elements are then eluted with 0.3M HCI at atmospheric pressure. This process can be sped up using a peristaltic pump.

Typical rare earth element separation with the ion-exchange procedure is shown in Figure 2. Free column volume is generally indicated by the numeral 11 and Ca and Ba elute first (12) followed by La (14), Ce (16), Pr (18), Nd (20) and then progressive separation of Eu, Gd, Tb, Dy, Ho, Er, Tm, Yb and Lu (22) and then Sm (24). The total blank (contamination introduced during processing from the acids and the environment) of the chemical procedure is negligible.

The second column is easily calibrated using an inductively coupled plasma mass spectrometer (ICPMS). This allows collection of individual fractions of all of the rare earth elements in chloride form.

An alternative resin that could be used is termed RE-Spec resin. This rare earth element selective resin is produced by EIChroM Industries for extraction of rare earth elements using chromatographic techniques. The acid used for this elution procedure is HN0 ₃ . Rare Earth Element Metals

The individual rare earth element chlorides produced by the chromatographic separation can subsequently be converted to rare earth oxides. The rare earth oxides in turn may be converted into high purity metals (e.g. Nd metal) or alloys of rare earths (e.g. mischmetal or ferro-alloys) by any suitable methods such as electrolysis or metallothermic reduction.

Specialty Metal Recovery

The chromatographic separation also outputs a stream which is substantially free of rare earth elements and is substantially concentrated in specialty metals (as chlorides). This specialty metal concentrate is then subjected to any known separation technique to separate the individual specialty metals from each other (i.e. niobium, tantalum, zirconium, hafnium). Example 1 B - Recovery of rare earth elements from ore

Figure 1 B provides a schematic of another process, generally indicated by the numeral 100, for concentrating the rare earth elements, in an ore such as a laterite ore 112. Typical feed rates for the plant would be 3-30 tonnes ore/hr. Comminution

At 114, the ore 112 is comminuted to a particle size of about 6mm and then fed to a grind circuit 1 16 to reduce the particle size of the ore further to about 106pm and render it suitable for combination with flux and further treatment (see below).

Optionally, and prior to entering the grind circuit 116, the comminuted ore from step 114 can be subjected to beneficiation step 114A to further concentrate the ore containing mineral from the gangue and the gangue can be discarded to tailings 118. Similarly and optionally, prior to being subjected to further treatment, the comminuted ore from step 116 can be subjected to beneficiation step 116A to further concentrate the ore containing mineral from the gangue and the gangue can be discarded to tailings 118.

Furnace Feed Preparation

The comminuted ore 120 resulting from the above treatments is then combined with a solid flux 122 in a suitable mixing device such as a screw feed mixer 124 in a 6:1 ratio of flux to ore. The solid flux 122 comprises a 50:50 mixture of sodium hydroxide and sodium perborate 124 that is combined from stock supplies of each reactant 126 in a suitable mixing device such as an augur including a screw feed mixer 128.

Heating

The 6:1 flux to ore mixture passes, on a continuous feed basis, to rotary kiln 130 where it is heated at between 350-400°C for 15 to 30 minutes. The heating is carried out to fuse the mixture and produce a partially fused solid that incorporates the rare earth element mineral. Following sufficient heating, the partially fused solid is removed from the rotary kiln 130 and grate cooled at ambient temperature 132 before being passed to mill/auger 134 for de-conglomeration.

The partially fused solid 136 is then treated with 32% hydrochloric acid 138 in an agitated leach tank 140 for a predetermined time before being filtered 142 to produce leachate 144 containing relatively high concentrations of rare earth elements. To the extent that the residue from the filtration process 142 contains sufficiently high concentrations of rare earth elements, it can be subjected to further washing 146 with acid to produce washate 148 and a solid fraction 150 to be passed to tailings 118. Washate 148 can be combined with the leachate 144.

Recovery

Reactants can be recovered from the washate 148 and leachate 144 and recycled for re-use and the concentrated rare earth elements are recovered via chromatographic separation techniques 152 as described elsewhere herein, such as Example 1A.

Examples 2 - 5

(a) General Methods

(i) Mixing Method

All samples were prepared as a fine powder, ground to a minimum of 80% passing 106μιη. Unless otherwise specified, the sample used was ground to P80 10pm. The NaOH reagent was used as provided, in pellet form. Similarly, the Na20 ₂ was provided as 1-2mm round beads. The mixing method involved combining the ground sample and each flux at the specified ratios.

(ii) Fusion Method

The furnace used for fusion of the sample mixture was preheated to the desired temperature. Once stable at the temperature the crucible was placed in to the furnace and the temperature again stabilised. Once stable, the sample was left to fuse for the specified amount of time. The sample was then removed and allowed to cool. (iii) Extraction of fused sample and lixiviant addition

To the extent that the fused material is too viscous to cast, the fused material was physically liberated from the crucible and, where required, this was assisted with the addition of the lixiviant. The lixiviant was added in stages and used conservatively as to achieve a more concentrated solution to pass to the subsequent stages of separation.

(iv) Analysis of Rare Earth Elements

The key analysis suite (XRF/ICP-MS) was conducted for Si, Fe, Mn, Mg, Ta, Nb, Sn, P, Ca , K, S, Na, Ba, Ti, Al, Pb, Sr, Ce, La, Nd, Th, U, Y. Total REE+Y was calculated based on the sum of Ce, La, Nd and Y.

(b) Materials

(i) Sample

The head analysis of the sample used in this example is presented in Table 2.

Table 2

Example 2 - Fusion Temperature and Flux Composition

Materials/Methods

Tests were conducted to recover Total REE+Y and Nb from 1 g of sample using varying flux compositions (Sample:Flux Ratio of 1 :6) and at temperatures of 500°C, 600°C and 650°C.

Washing of the fused sample to recover the targets was conducted using 10% HCI. Results

The results covering the effect of flux composition at various temperatures are presented graphically in Error! Reference source not found.3 and 4.

The results covering the effect of temperature at various flux compositions are presented graphically in Error! Reference source not found.5 and 6. Example 3a -Sample:Flux Ratio

Materials/Methods

Tests to recover Total REE+Y and Nb were conducted using the same flux compositions in Example 2 and at a temperature of 650°C but with 1g of sample and an increased Sample:Flux Ratio of 1 :30.

As in Example 2 washing of the fused sample to recover the targets was conducted using 10% HCI.

Results

The results are presented graphically in Error! Reference source not found.7 and 8. Example 3b- Lixiviant

Materials/Methods

To assess the effect of the volume of lixiviant on the recovery, three stages of lixiviant (10% HCI) addition were also assessed during the tests carried out in Example 3a. Results

The results are presented graphically in Error! Reference source not found.9 and 10.

Example 4 -Sample:Flux Ratio and Lixiviant Volume

Materials/Methods

Tests were conducted to further investigate the impact of a 1 :6 versus a 1 :30 sample:flux ratio on recovery using 10g and 2g of sample, respectively, in graphite crucibles. The fusion was carried out at 650°C.

To also assess the impact of lixiviant volume, washing of the fused sample to recover the targets was conducted using different volumes of 32% HCI.

Results

The results with respect to the impact of lixiviant volume on recovery of total REE+Y and Nb are presented graphically in Error! Reference source not found.11 and 12.

Example 5 -Sample:Flux Ratio

Materials/Methods Tests were conducted to further investigate the impact of sample:Flux ratio using varying amounts of sample to a total flux/sample mixture of 60g and a fusion temperature of 650°C across a range of sample:Flux ratios.

A baseline test was conducted at a sample:Flux ratio of 1 :0 to examine the recovery in the absence of flux. This baseline tests is treated as a comparison between a high temperature process combined with an acid leach and a fusion at high temperature followed by acid dissolution.

Results

The results for these tests are presented in Error! Reference source not found.13 with respect to Fe, Nb and Total REE+Y.

Example 6 - Flux Composition

Materials/Methods

Five (5) tests were conducted to confirm the effect of Flux Composition between a binary mixture of NaOH and Na2C½. Tests were conducted at a Sample:Flux ratio of 1 :4 to observe any increase in recovery. Tests were conducted at 650°C for 1 hour.

Results

The results for these tests are presented in Figure 14 with respect to Fe, Nb and Total REE+Y. More than 95% of the Fe and Total REE+Y was recovered at a flux composition with <50% NaOH >50% Na2C>2. The Nb recovery was reported at -85% at 100% Na ₂ 0 ₂ and drops as the NaOH % increases.

Example 7 - Heating Time

Materials/Methods

Five (5) tests were conducted to examine the effect of Fusion Time on the recovery of the elements of interest. The tests were conducted at a Sample:Flux Ratio of 1 :6 at 650°C for varied bake times.

Results

The results for these tests are presented in Figure 15 with respect to Fe, Nb and Total REE+Y. The Fe recovery was reported at -100% across all tested heating times, while the Total REE+Y and Nb had a varied response. The maximum Nb recovery was reported at 1 hour Fusion Time (94%) while the maximum Total REE+Y recovery was reported at 0.25 hour (98%).

Example 8 - Heating Temperature

Materials/Methods

Five (5) tests were conducted to confirm the effect of temperature on recovery of Fe, Nb and total REE+Y. Temperatures tested from 450°C to 650°C were examined at Fusion Times of 1 hour at 1 :6 sample:flux ratio.

Results

The results are presented in Figure 16. The best results were achieved at 450°C with >95% recovery of the Total REE+Y, Nb and Fe.

Example 9 - Fused Solid

Materials/Methods

To examine the effect of the crucible material on recovery and the properties of the fused solid a number of tests were carried out using different crucibles.

Tests were conducted with heating at 650°C for 60 minutes using a range flux:sample ratios and graphite or ceramic crucibles. The flux was 50:50 NaOH to a2C>2 at ratios of 1 :1 through to 1 :12.

Each of the tests conducted in ceramic crucibles were cracked prior to lixiviant treatment to examine the cross section of the fused solid.

Results

The effect of the crucibles on recovery is depicted in Figure 17.

The images in Figures 18A-18D show the cross section of the fused solid from the ceramic crucibles for sample:flux ratios of 1 :1 , 1 :6, 1 :8 and 1 :12, respectively. Figure 18A shows that limited fusion was observed at the 1 :1 Sample:Flux ratio. This level of fusion is unlikely to be reproducible and is otherwise unsuitable for further processing.

For the Sample:Flux ratio of 1 :6 (Figure 18B) and higher (Figure 18C and 18D), the fused solid showed 'pits' on the surface and throughout. Whilst not wishing to be bound by this Applicant believes that a fused solid (frit) that is beyond that shown in Figure 18A but has not progressed to the point depicted in Figure 18B are the most desirable because they are achieved prior to the melt stage and, provided recovery is not overly compromised, would be more cost effective and less intensive for safer material handling. This was investigated further in the example described hereafter.

Example 10 - Recovery from Fused Solid (Frit)

Materials/Methods

Tests were conducted using ceramic crucibles at temperatures between 350°C and 450°C, for times ranging from 10 to 60 minutes. One test was selected and run at various sample:flux ratios to again examine the effect on the fused solid formation. The flux was 50:50 IMaOH to Na ₂ 0 ₂ .

Results

The results are summarised in Figures 19-23 and the best results obtained are set out in Table 3.

Table 3

Example 11 - Use of Sodium Perborate Flux

Materials/Methods

Sodium perborate was assessed as an alternative flux material. Seven (7) tests were conducted in order to compare against the sodium peroxide flux. These were conducted with heating at various temperatures for 15 minutes and across a range of Sample:Flux Ratios.

Results

The results obtained using sodium perborate as flux are summarised in Figure 24. Fe, Nb and Total REE+Y recovery increased with increasing Sample:Flux ratio and maximum Recovery >90% was achieved at a Sample: Flux Ratio of 1 :5.

A comparison between the results obtained using sodium perborate or sodium hydroxide flux showed that the perborate delivered slightly higher recoveries at lower temperatures (Figure 25). Example 12 - Removal of Fe

In the examples above, Fe is a co-recovered with total REE+Y and Nb. Work was conducted to assess the use of tributyl phosphate (TBP) for extracting Fe to further concentrate the total REE+Y and Nb. Materials/Methods

A series of solvent extraction tests were conducted on a combined liquor generated from the previous examples. The testwork conducted examined the Organic/Aqueous Ratio for Tributyl Phosphate (100%) and Tributyl Phosphate (50%) in Kerosene.

Results

The results (Figures 26 and 27) show that the solvent extraction is selective for the Fe.