Article Optimization of Acid Leaching of Rare-Earth Elements from Mongolian -Based

Rina Kim 1, Heechan Cho 1,*, Kenneth N. Han 2, Kihong Kim 1 and Myoungwook Mun 1

1 Department of Energy Resources Engineering, Seoul National University, Seoul 08826, Korea; [email protected] (R.K.); [email protected] (K.K.); [email protected] (M.M.) 2 Department of Materials and Metallurgical Engineering, South Dakota School of Mines and Technology, Rapid City, SD 57701-3995, USA; [email protected] * Correspondence: [email protected]; Tel.: +82-2-880-8271

Academic Editor: William Skinner Received: 31 March 2016; Accepted: 24 June 2016; Published: 30 June 2016

Abstract: Optimization of the acid leaching process for Mongolian apatite-based ore containing rare-earth elements (REEs) was studied. The ore contained approximately 10% of REEs as total rare earth oxides, and the major impurities were Ca (33% as CaO) and Fe (23% as Fe2O3). Fe bearing minerals could be removed by passing the sample through a wet high-intensity magnetic separator before leaching. After magnetic separation, basic leaching tests were conducted to investigate the influence of the acid type and concentration, temperature, and the pH on the REE leaching level and kinetics. was found to be the most effective leaching agent, leaching more than 90% of REEs in an hour. However, the concentrations of Ca ions in the leachate were also high, which would complicate recovery of the REEs. Therefore, to reduce the amount of Ca ions in the leachate, a two-stage leaching procedure was attempted. In stage 1, the sample was preleached using 1.0 M hydrochloric acid to dissolve Ca. In stage 2, the solid residue of stage 1 was leached using 2.0 M hydrochloric acid to dissolve REEs. Consequently, this two-stage leaching significantly reduced the Ca concentration in the final leachate without affecting the leaching levels of REEs.

Keywords: rare-earth element; acid leaching; apatite; magnetic separation; two-stage leaching; Ca removal

1. Introduction With increasing demand for rare-earth elements (REEs) and the corresponding recent surge in their prices, many countries have begun to reevaluate old rare-earth mining prospects and explore new ones. According to a 2009 estimate by the U.S. Geological Survey, Mongolia has 31 million tonnes of rare-earth reserves (approximately 17% of the world’s total), exceeded only by China [1]. In light of Chinese restrictions on REE exportation, Mongolia has become a leading contender in the current rush for REE resources [1]. At present, Mongolia plans to increase its transportation infrastructure to ship resources to Russia and the Far East [1]. Therefore, Mongolian resources potentially represent a major supply of REEs for Korea. The REE ore deposit currently under investigation is located in southern Mongolia and contains over 300,000 t of rare-earth oxides (REOs) as combined REOs according to a recent exploration campaign [2]. The average REO grade in this area is 1.36% but reaches 6.15% in a newly discovered high-grade zone. This deposit consists of sheet-like magnetite–apatite ore bodies [3]. REE are usually processed in two steps: physical concentration and leaching. The first step typically includes crushing the ore and physically separating the REOs from the ore. This separation process dramatically increases the percentage of REOs in the working material but is successful only if the REEs are largely concentrated in a single mineral phase such as a form of bastnäsite (REECO3F)

Minerals 2016, 6, 63; doi:10.3390/min6030063 www.mdpi.com/journal/minerals Minerals 2016, 6, 63 2 of 15

or monazite (REEPO4)[4]. Unfortunately, this is not the case for the REE ore under investigation, which contains primarily apatite [Ca10(PO4)6(OH,F,Cl)2] as the major source mineral. REEs in apatite usually occur as ionic substitutions for Ca within the lattice rather than in discrete mineral form [5–7]. Intuitively, it is possible to obtain REE concentrates by selective separation of REE-bearing apatite from the ore. Froth flotation is a common process applied to ores. Consequently, various flotation tests were conducted in an attempt to obtain REE concentrates. However, the results were not satisfactory, yielding either a low recovery and/or little REE enrichment in the concentrates. Other physical concentration methods such as gravity separation and magnetic separation were also found to be ineffective in concentrating REEs, but magnetic separation showed positive effects on impurity removal. The extraction of REEs from ore usually involves decomposition with acidic or alkaline reagents. In the acid route, acid baking is commonly used for the major REE minerals including monazite, bastnäsite, and . The treatment involves mixing REE concentrates with followed by baking at around 200 ˝C. The resulting cake is leached with water to dissolve REEs. Alkaline digestion is typically used for monazite. The ore is decomposed with NaOH to produce rare earth hydroxides. The resulting products are then leached using hydrochloric acid to dissolve REEs. In addition, direct decomposition by HCl is another common method used for carbonate minerals such as bastnäsite. Several studies have examined the acid leaching of REE apatite [8–12]. It was reported that apatite leached well in hydrochloric or nitric acid [8,10,11], but it was very difficult to attack with sulfuric acid unless the concentration and/or temperature were very high [8,10,12]. However, owing to the complex metallurgy of REEs, there is no standard process for leaching REE-bearing minerals. The choice of decomposition method depends on the mineralogy of the REE-containing phases as well as other non-REE minerals present in the ore that often affect the effectiveness of the leaching process. This study investigates the leaching behavior of a Mongolian REE ore in response to various acids. To reduce impurities in the leachate, pre-removal of Fe by magnetic separation and a two-stage leaching process for selective removal of Ca were studied. The specific objectives were to examine the mineralogical characteristics of the sample, determine the effects of the acid type and concentration, temperature, and the pH on the leaching efficiency, and assess the optimum conditions of the leaching process for this apatite-based ore.

2. Materials and Methods

2.1. Mineralogical Characteristics of the Sample The sample was collected from the southern region of Mongolia. Previous studies reported that the REE deposit consists of apatite ore bodies including apatite, barian celestite, magnetite, hematite, goethite, fluorite, gypsum, phlogopite, carbonate (mostly calcite), pyrite, and monazite-(Ce) [3,13]. The chemical composition of the sample was analyzed by X-ray fluorescence (XRF) and summarized in Table1. The total level of REOs is approximately 10%. The most abundant REE is Ce, followed by La, Nd, Pr, and Y. X-ray diffraction (XRD) (D/MAX-2500V/PC, Rigaku, Tokyo, Japan) was used to characterize the mineral phases in the sample. As shown in Figure1, the main minerals were identified as apatite and silicate minerals. Minor minerals were goethite, hematite, maghemite, quartz, monazite, and so on. Detailed mineralogical characterization of the ore was conducted by Enkhbayar et al. [14], who found that REEs were mostly associated with fluorapatite and . As shown in Figure2, the apatite grains have hexagonal structures with a layered arrangement of outer grown upon an inner layer of crystals. Electron probe microanalysis revealed that the REE (Ce2O3 + La2O3 + Nd2O3) content increases from the center to the edge of the specimen (9.45 wt % in mh21, 12.51 wt % in mh23, 45.57 wt % in mh24). The Si and Na contents also increased from the center to the edge of the structure. Conversely, the amounts of P and Ca decreased from the center to the edge of the sample. These trends indicate that more REE was substituted for Ca in apatite as the crystals Minerals 2016, 6, 63 3 of 15 Minerals 2016, 6, 63 3 of 16 grew.grew. Monazite Monazite crystals crystals were were alsoalso foundfound betweenbetween th thee two two outer outer layers. layers. Therefore, Therefore, it itcan can be beexpected expected Minerals 2016, 6, 63 3 of 16 thatthat the the leaching leaching rate rate would would initially initially bebe highhigh owin owingg to to the the presence presence of of a ahigh-grade high-grade REE REE layer layer in inthe the outer rim of the apatite particles and then decrease because monazite is less reactive than apatite [15]. outergrew. rim of Monazite the apatite crystals particles were also and found then decreasebetween th becausee two outer monazite layers. isTherefore, less reactive it can than be expected apatite [15]. that the leaching rate would initially be high owing to the presence of a high-grade REE layer in the Table 1. X-ray fluorescence (XRF) analysis of the sample. outer rim of the apatiteTable particles 1. X-ray and then fluorescence decrease (XRF) because analysis monazite of the is less sample. reactive than apatite [15]. Formula Concentration (%) Formula Concentration (%) FormulaCaO Table Concentration 1. X-ray32.75 fluorescence (%) (XRF) Formula MgOanalysis of the Concentrationsample.0.53 (%) FeFormulaCaO2O3 Concentration22.91 32.75 (%) Formula MgOSrO Concentration 0.47 0.53 (%) FeP22CaOOO53 18.3022.9132.75 NaMgO SrO2O 0.530.45 0.47 P O 18.30 Na O 0.45 SiO2Fe25 O3 9.6522.91 Pr SrO6O2 11 0.470.42 SiO2 9.65 Pr6O11 0.42 CeOP2O2 5 5.8018.30 MnONa2O 0.450.29 CeO2 5.80 MnO 0.29 La2SiOO3 2 3.019.65 Pr Y26O113 0.420.28 La2O3 3.01 Y2O3 0.28 CeO2 5.80 MnO 0.29 SO3 1.801.80 K K22OO 0.25 0.25 La2O3 3.01 Y2O3 0.28 Nd22O3 1.481.48 ThO ThO2 0.060.06 SO3 1.80 K2O 0.25 Al22O3 0.870.87 ZnO ZnO 0.05 0.05 TiONd2O3 0.621.48 ThO CuO2 0.06 0.04 TiO22 0.62 CuO 0.04 Al2O3 0.87 ZnO 0.05 TiO2 0.62 CuO 0.04

FigureFigure 1. 1.X-ray X-ray diffraction diffraction (XRD)(XRD) analysis of the the sample. sample. (a ()a )(Ce, (Ce, La, La, Nd) Nd) hydroxyapatite, hydroxyapatite, , fluorapatite, chlorapatite;Figure 1. X-ray (b) Nd, diffraction Mn silicates. (XRD) analysis of the sample. (a) (Ce, La, Nd) hydroxyapatite, fluorapatite, chlorapatite; (b) Nd, Mn silicates. chlorapatite; (b) Nd, Mn silicates.

FigureFigureFigure 2. 2. The2. The The layeredlayered layered structure structure of of apatiteapatite inin in thethe the samplesample sample [14]. [14]. [14 ].

2.2.2.2. Experimental2.2. Experimental Experimental Procedure Procedure Procedure Lumps of the ore sample were crushed into smaller pieces with a hammer and were crushed LumpsLumps of of the the ore ore sample sample werewere crushedcrushed into smaller pieces pieces with with a ahammer hammer and and were were crushed crushed again to −1 mm in stages by a jaw crusher and disk mill. They were then ground to 90% of the passing againagain to to´ 1−1 mm mm in in stages stages byby aa jawjaw crusher and disk disk mill. mill. They They were were then then ground ground to to90% 90% of ofthe the passing passing size of 0.3 mm using a ball mill. Figure3 shows the cumulative size distribution of the sample.

Minerals 2016, 6, 63 4 of 16

size of 0.3 mm using a ball mill. Figure 3 shows the cumulative size distribution of the sample. Because the chemical analysis indicated that the sample contained a large amount of Fe (23% Fe2O3), Fe-bearing minerals were removed from the sample before leaching. A mixture of water and the sample, bearing 10 wt % solid materials, was prepared and fed through a laboratory wet high-intensity magnetic separator (Eriez Series L-4 Model, Eriez, Erie, PA, USA) operating at 0.2 T. The chemical compositions of the resulting magnetic and nonmagnetic streams were analyzed using XRF. Basic leaching tests were conducted using sulfuric, hydrochloric, and nitric acids at various concentrations in the ranges of 0.1–13.0 M for sulfuric acid and 0.5–2.0 M for hydrochloric and nitric acid. All of the leaching tests were replicated 4–5 times, and average values obtained after removing significant outliers were used for analysis. Ten grams of the sample were added to a beaker containing 90 mL of the acid solution. The slurry was then agitated with an impeller-type stirrer (Daehan Scientific, Wonju, Korea). The slurry samples were collected at regular intervals and the leachate was separated from the solid particles using a filter for lower acid conditions (0.1–2.0 M acids) or a centrifuge for higher acid conditions (5.0–13.0 M acids). The separated leachate was diluted with 1.0 M hydrochloric acid to prevent precipitation of REEs and analyzed using an inductively coupled plasma optical emission spectrometer. As the leachate contained a considerable amount of Ca, a two-stage leaching method was employed to reduce the amount of Ca in the final leachate. In stage 1, the sample was preleached using 1.0 M hydrochloric acid to dissolve Ca. In stage 2, the leachate of stage 1 was separated from the residue, Minerals 2016, 6, 63 4 of 15 and only the solid residue was leached using 2.0 M hydrochloric acid to dissolve the REEs.

FigureFigure 3. 3.Cumulative Cumulative size size distribution distribution of of the the sample sample after after crushing crushing and and milling. milling.

3. Results and Discussion Because the chemical analysis indicated that the sample contained a large amount of Fe (23% Fe3.1.2O Magnetic3), Fe-bearing Separation minerals were removed from the sample before leaching. A mixture of water and the sample, bearing 10 wt % solid materials, was prepared and fed through a laboratory wet high-intensity magneticTable separator 2 presents (Eriez a chemical Series L-4analysis Model, of both Eriez, the Erie, feed PA, sample USA) and operating the magnetic at 0.2 and T. The nonmagnetic chemical compositionsstreams collected of the resultingfrom the magneticmagnetic and separator. nonmagnetic The streamsFe2O3 content were analyzed in the usingnonmagnetic XRF. portion decreasedBasic leachingsignificantly tests from were the conducted feed value using of 22.9% sulfuric, to 10.6%. hydrochloric, In addition, and the nitricREE contents acids at increased various concentrationscorrespondingly, in the as rangesREE-containing of 0.1–13.0 apatite M for is sulfuric nonmagnetic. acid and Some 0.5–2.0 REEs M for were hydrochloric lost to the and magnetic nitric acid.stream, All but of the the leaching recovery tests of the were REEs replicated in the nonm 4–5 times,agnetic and stream average was valuesstill greater obtained than after90%. removing The Fe2O3 significantcontent of outliersthe magnetic were used stream for was analysis. very high, Ten grams and the of theFe grade sample was were compatible added to with a beaker typical containing iron-ore 90concentrates. mL of the acid Therefore, solution. magnetic The slurry separation was then agitatedbefore leaching with an impeller-typeoffers a potential stirrer opportunity (Daehan Scientific, not only Wonju,to reduce Korea). Fe impurities The slurry in samples the leachate were but collected also to at generate regular intervalsrevenue by and selling the leachate the Fe-rich was separatedbyproduct fromto smelters. the solid particles using a filter for lower acid conditions (0.1–2.0 M acids) or a centrifuge for higher acid conditions (5.0–13.0 M acids). The separated leachate was diluted with 1.0 M hydrochloric acid to prevent precipitation of REEs and analyzed using an inductively coupled plasma optical emission spectrometer. As the leachate contained a considerable amount of Ca, a two-stage leaching method was employed to reduce the amount of Ca in the final leachate. In stage 1, the sample was preleached using 1.0 M hydrochloric acid to dissolve Ca. In stage 2, the leachate of stage 1 was separated from the residue, and only the solid residue was leached using 2.0 M hydrochloric acid to dissolve the REEs.

3. Results and Discussion

3.1. Magnetic Separation Table2 presents a chemical analysis of both the feed sample and the magnetic and nonmagnetic streams collected from the magnetic separator. The Fe2O3 content in the nonmagnetic portion decreased significantly from the feed value of 22.9% to 10.6%. In addition, the REE contents increased correspondingly, as REE-containing apatite is nonmagnetic. Some REEs were lost to the magnetic stream, but the recovery of the REEs in the nonmagnetic stream was still greater than 90%. The Fe2O3 content of the magnetic stream was very high, and the Fe grade was compatible with typical iron-ore concentrates. Therefore, magnetic separation before leaching offers a potential opportunity not only to reduce Fe impurities in the leachate but also to generate revenue by selling the Fe-rich byproduct to smelters. Minerals 2016, 6, 63 5 of 15

Minerals 2016, 6, 63 5 of 16 Table 2. Result of XRF analysis of raw sample and magnetically separated samples. Table 2. Result of XRF analysis of raw sample and magnetically separated samples Content (wt %) Magnetic Intensity Sample Yield (%) Magnetic Content (wt %) Sample Yield (%) Fe2O3 CeO2 La2O3 Nd2O3 Pr6O11 Y2O3 Intensity Fe2O3 CeO2 La2O3 Nd2O3 Pr6O11 Y2O3 - Raw 100 22.9 5.8 3.0 1.5 0.4 0.3 - Raw 100 22.9 5.8 3.0 1.5 0.4 0.3 Magnetic 19.5 72.5 1.5 0.9 0.4 0.1 0.1 0.2 T Magnetic 19.5 72.5 1.5 0.9 0.4 0.1 0.1 0.2 T Nonmagnetic 80.5 10.6 6.8 3.6 1.6 0.5 0.3 Nonmagnetic 80.5 10.6 6.8 3.6 1.6 0.5 0.3 RecoveryRecovery to Nonmagnetic (%)(%) 37.637.6 95.095.0 94.594.5 94.994.9 96.796.7 96.196.1

3.2.3.2. Effect Effect of of Acid Acid Concentration Concentration and andType Type FigureFigure4 shows4 shows the the changechange in the the leaching leaching effici efficiencyency with with the thesulfuric sulfuric acid acidconcentration. concentration. The Theexperiments experiments were were run run for for5 h 5at h20 at °C. 20 At˝C. low At acid low acidconcentrations, concentrations, the leaching the leaching level increased level increased with withincreasing increasing acid acid concentration. concentration. However, However, the leaching the leaching level decreased level decreased at 5.0 M. at This 5.0 M. may This result may from result fromreprecipitation reprecipitation of REEs of REEs via viacalcium calcium sulfate sulfate formation. formation. It is It known is known that that calcium calcium sulfate sulfate hydrates hydrates (gypsum,(gypsum, hemihydrate, hemihydrate, and and anhydrite) anhydrite) areare readilyreadily formed wher whereverever calcium calcium and and sulfate sulfate are are present present together in aqueous solutions [16]. together in aqueous solutions [16].

100

80

60

40 Percent leaching Percent

20

0 02468101214 Sulfuric acid concentration (M) Ca Ce La Nd Pr

˝ FigureFigure 4. 4.Leaching Leaching levels levelsat at variousvarious initial sulfuric acid acid concentrations concentrations at at 20 20 °C.C.

Indeed,Indeed, the the XRD XRD analysisanalysis ofof thethe residueresidue af afterter leaching leaching shows shows the the presence presence of ofgypsum gypsum 4 2 4 (CaSO(CaSO4¨2H∙2H2O)O) and and anhydrite anhydrite (CaSO(CaSO4) (Figure 55).). This residue may may contain contain REEs REEs by by isomorphous isomorphous substitutionssubstitutions for for Ca Ca2+2+[ 8[8,12,17,18],,12,17,18], becausebecause Ca and REEs REEs have have similar similar ionic ionic radii. radii. To To confirm confirm the the existence of REEs in the precipitated gypsum and CaSO4, the leaching residue was dried in existence of REEs in the precipitated gypsum and CaSO4, the leaching residue was dried in atmosphere andatmosphere was washed and with was water.washed The with solution water. The phase soluti fromon thisphase washing from this was washing found was to contain found ato significant contain a significant amount of Ca and REEs (1.03 g/L Ca, 0.08 g/L Ce, 0.03 g/L La, 0.03 g/L Nd, amount of Ca and REEs (1.03 g/L Ca, 0.08 g/L Ce, 0.03 g/L La, 0.03 g/L Nd, 0.02 g/L Pr, and 0.02 g/L Pr, and 0.009 g/L Y), which indicates that the precipitated calcium sulfate hydrates contain 0.009 g/L Y), which indicates that the precipitated calcium sulfate hydrates contain REEs by ion REEs by ion substitution. The Ca concentration seems to be higher than the solubility of gypsum (2.4 substitution. The Ca concentration seems to be higher than the solubility of gypsum (2.4 g/L), but g/L), but the solution may contain free calcium ions and associated calcium sulfate neutral species the solution may contain free calcium ions and associated calcium sulfate neutral species [CaSO4(aq)] [CaSO4(aq)] dissolved from anhydrite [16,19]. dissolved from anhydrite [16,19]. On the other hand, when the concentration of sulfuric acid further increased above 5.0 M, the leachingOn the level other increased hand, when again. the It can concentration be postulated of sulfuricthat the acidincrease further in the increased acidity may above change 5.0 M, the the leachingsolution level chemistry increased to limit again. the precipitation It can be postulated of calcium that sulfate the increasehydrates. in To the examine acidity more may closely change the the solutionrelationship chemistry between to limit the leaching the precipitation patterns and of calciumH2SO4 concentration, sulfate hydrates. the Ca To speciation examine was more calculated closely the relationshipfor the Ca-SO between4-H2O thesystem leaching using patterns thermodynamic and H2SO data4 concentration, available in the the literature Ca speciation [20–22]: was calculated for the Ca-SO4-H2O system using thermodynamic data available in the literature [20–22]:

Minerals 2016, 6, 63 6 of 15

´ ´ ´ rH2SO4s OH ´26 HSO4 ` H2O “ H2SO4 ` OH K1 “ ´ “ 3.002 ˆ 10 , (1) HSO“4 ‰ “ 2´ ‰ SO4 HSO´ ` OH´ “ SO2´ ` H O K “ “ 1.034 ˆ 1012, (2) 4 4 2 2 ” ´ ı ´ HSO4 OH ` Minerals 2016, 6, 63 2` ´ ` “ rCa ‰“OHq ‰ 6 of 16 Ca ` OH “ CapOHq K3 “ “ 25.119, (3) Ca2`` OH´‰ HSO + HO= HSO +OH ”= ı = 3.002 10 , (1) “ ‰ CaSO4 paqq 2` 2´ 2 Ca ` SO “ CaSO K4 “ “ 1.157 ˆ 10, (4) HSO +OH4 = SO + H4 paqO q = = 1.034 10 , (2) ” 2` 2´ı Ca SO4 () Ca + OH =Ca(OH) ” ı= ” ı = 25.119, (3) 2` ´ 2` ´2 ´6 Ca ` 2OH “ CapOHq2 Ks1 “ Ca rOH s “ 3.890 ˆ 10 , (5) () Ca + SO =CaSO () ” =ı = 1.157 10 , (4) 2` 2´ 2` 2´ ´5 Ca ` SO4 “ CaSO4 psq Ks2 “ Ca SO4 “ 4.930 ˆ 10 , (6) Ca +2OH =Ca(OH) = CaOH = 3.890 10, (5) ” ı ” ı 2` 2´ 2` 2´ ´5 Ca Ca` SO+ SO4 ` 2H=CaSO2O “ ()CaSO 4 ¨ 2H2O Ks3“= CaCa SOSO 4= 4.930“ 3.140 10ˆ 10, .(6) (7) ” ı ” ı Ca + SO +2HO=CaSO ∙2HO ´ = Ca ´SO 2´= 3.1402+ 10 . + (7) There are seven species in the solution: OH ,H2SO4, HSO4 , SO4 , Ca , Ca(OH) , CaSO4(aq). Therefore,There three are more seven equationsspecies in the in additionsolution: OH to− Equations, H2SO4, HSO (1)–(4)4−, SO4 are2−, Ca required2+, Ca(OH) to+, CaSO calculate4(aq). the equilibriumTherefore, composition. three more Twoequations additional in addition equations to Equations used were (1)–(4) mass are balance required equations to calculate for Ca the species and sulfateequilibrium species. composition. One additional Two additional equation equations could used be were the mass charge balance balance equations equation, for Ca species but in our calculations,and sulfate the species. OH´ concentrationsOne additional equation were assumed could be to the be charge known balance and were equation, calculated but in fromour the calculations, the OH− concentrations were assumed to be known and were calculated from the final final pH of the solutions in the actual experiment. The total Ca concentration was assumed to be pH of the solutions in the actual experiment. The total Ca concentration was assumed to be 0.5 M considering the number of moles of Ca in the ore. The equilibrium composition was then 0.5 M considering the number of moles of Ca in the ore. The equilibrium composition was then calculatedcalculated under under different different levels levels of of total total sulfate sulfate concentrations (1.0, (1.0, 2.0, 2.0, 5.0, 5.0, 9.8, 9.8, and and 13.0 13.0M). If M). the If the 2+ calculatedcalculated values values were were greater greater than than any any of of the theK Kspsp valuesvalues [Equations [Equations (5)–(7)], (5)–(7)], the the Ca2+ Ca concentrationconcentration was recalculatedwas recalculated using using the the Equations Equations (5)–(7) (5)–(7) andand the calculations we werere repeated repeated until until the calculated the calculated valuesvalues satisfied satisfied all conditions.all conditions. The The exact exact solution solution was obtained obtained by by an an iterative iterative method method using using “Solver” “Solver” functionfunction in MS in ExcelMS Excel (Microsoft, (Microsoft, Redmond, Redmond, WA, WA, USA).USA).

FigureFigure 5. XRD 5. XRD analysis analysis of of residue residue fromfrom le leachingaching with with 13.0 13.0 M Msulfuric sulfuric acid. acid.

Figure 6 shows the calculation results, showing the distribution of Ca species as a function of the Figureinitial sulfuric6 shows acid the concentration. calculation results,In this highly showing acidic the condition, distribution Ca was of present Ca species mainly as in a the function form of the initialof Ca sulfuric2+ and CaSO acid4(aq) concentration.. The concentration In this of highly CaSO acidic4(aq) does condition, not change Ca waswith presentthe sulfuric mainly acid in the 2+ formconcentration, of Ca and CaSOas the 4(aq)solution. The become concentration saturated ofowing CaSO to4(aq) the highdoes concentrations not change with of Ca the2+ and sulfuric SO42−. acid 2+ 2+ 2´ concentration,On the other as hand, the solution the concentration become saturatedof Ca changes owing with to the sulfuric high concentrations acid concentration, of Ca showingand SOa 4 . On theminimum other hand, at 5.0 theM. The concentration final pH of the of Casolution2+ changes decreased with initially the sulfuric with increasing acid concentration, the sulfuric showingacid a minimumconcentration, at 5.0 M.but Theremained final pHthe same of the from solution 2.0 to 5.0 decreased M of sulfuric initially acid. withHowever, increasing with more the addition sulfuric acid concentration,of sulfuric butacid, remained the final pH the decreased same from again, 2.0 towhich 5.0 Maccompanies of sulfuric the acid. increases However, in the with concentration more addition of Ca2+. It indicates that the concentration of Ca2+ species changed in a complicated manner due to the of sulfuric acid, the final pH decreased again, which accompanies the increases in the concentration competing balance between the precipitation and dissolution of Ca, which is affected by H+, SO42− and Ca2+ in the solution. Accordingly, the total soluble Ca species varies with the sulfuric acid concentration in a pattern very similar to the leaching pattern. This result suggests that sulfuric acid may not be a good leaching agent for this ore, as the side reaction, Ca(REE) sulfate precipitation, limits the REE leaching level, which did not exceed 80%, even when a large excess of acid (13.0 M) was used.

Minerals 2016, 6, 63 7 of 15 of Ca2+. It indicates that the concentration of Ca2+ species changed in a complicated manner due to the competing balance between the precipitation and dissolution of Ca, which is affected by H+, 2´ 2+ SO4 and Ca in the solution. Accordingly, the total soluble Ca species varies with the sulfuric acid concentration in a pattern very similar to the leaching pattern. This result suggests that sulfuric acid may not be a good leaching agent for this ore, as the side reaction, Ca(REE) sulfate precipitation, limits the REEMinerals leaching 2016, 6 level,, 63 which did not exceed 80%, even when a large excess of acid (13.0 M)7 of was16 used.

Minerals 2016, 6, 63 7 of 16 0.007 1.0 2+ + Ca CaOH CaSO4 (aq) Ca Tot Final pH 0.0060.007 1.0 2+ + 0.5 Ca CaOH CaSO4 (aq) Ca Tot Final pH 0.006 0.005 0.5 0.0

0.0040.005 0.0 -0.5

0.0030.004 Final pH -0.5 -1.0

0.0020.003 Final pH

Concentration of Ca species (M)Concentration -1.0 -1.5 0.0010.002

Concentration of Ca species (M)Concentration -1.5 0.0000.001 -2.0 02468101214 0.000 Initial sulfuric acid concentration (M) -2.0 02468101214 ˝ FigureFigure 6. Ca 6. Ca speciation speciation under under different differentInitial initial sulfuricinitial sulfuric acid concentration sulfuricacid concentrations acid (M) concentrations at 25 °C, 1 atm. Equilibrium at 25 C, 1 atm. Equilibriumcomposition composition calculated calculated by solving byEquations solving (1)–(7). Equations (1)–(7). Figure 6. Ca speciation under different initial sulfuric acid concentrations at 25 °C, 1 atm. Equilibrium Figurecomposition 7 shows calculated the variation by solving in theEquations leaching (1)–(7). levels with the initial concentration of hydrochloric Figureacid. 7Almost shows no the REEs variation were inleached the leaching when the levels concentration with the was initial less concentration than 1.0 M. At of hydrochlorichigher acid. AlmostconcentrationsFigure no 7 REEsshows of hydrochloric the were variation leached acid, in the when leaching the levels concentration withincreased the initial sharply. was concentration At less an thanacid of concentration hydrochloric 1.0 M. At higher concentrationsofacid. 2.0 M,Almost ofnearly hydrochloric no 100% REEs of thewere acid,REEs leached thewere leaching leached.when the Ho levels wever,concentration increased the amount was sharply. ofless Ca inthan At the an 1.0leachate acid M. concentrationAt was higher also of concentrations of hydrochloric acid, the leaching levels increased sharply. At an acid concentration 2.0 M, nearlyvery high 100% (27,480 of the mg/L), REEs which were would leached. undoubtedl However,y cause the amountcomplications of Ca during in the REE leachate recovery was by also very eitherof 2.0 precipitationM, nearly 100% or solventof the REEs extraction. were leached. However, the amount of Ca in the leachate was also high (27,480very high mg/L), (27,480 which mg/L), would which undoubtedly would undoubtedl causey cause complications complications during during REE REE recoveryrecovery by by either precipitationeither orprecipitation solvent extraction. or100 solvent extraction.

100 80 Fe P 80 CaFe 60 CeP LaCa 60 NdCe 40 La

Percent leaching Pr Nd 40 Y

Percent leaching 20 Pr Y 20 0 0123 0 Hydrochloric acid concentration (M) 0123 Figure 7. Leaching levels at various initial hydrochloric acid concentrations at 20 °C. Leaching time: Hydrochloric acid concentration (M) 2 h; solid/liquid (S/L) ratio (w/v): 1:9. Figure 7. Leaching levels at various initial hydrochloric acid concentrations at 20 °C. Leaching time: Figure 7. Leaching levels at various initial hydrochloric acid concentrations at 20 ˝C. Leaching time: Figure2 h; solid/liquid 8 shows (S/L)the changeratio (w /vin): the1:9. leaching levels with time when 2.0 M hydrochloric acid was w v 2 h;used solid/liquid as the leaching (S/L) agent. ratio (More/ ): than 1:9. 90% of the REEs were leached in 10 min. At 1 h, almost 100% of theFigure REEs were 8 shows leached, the change and the in reaction the leaching reached levels equilibrium with time in when 2 h. It 2.0 could M hydrochloricbe noted that acid the ratewas Figureofused the8 as leachingshows the leaching thereaction change agent. was More inquite the thanfast. leaching 90%Thus, of it the levelswas REEs sufficient with were timeleached to leach when in the 10 2.0 oremin. Mat At room hydrochloric 1 h, temperaturealmost 100% acid was of the REEs were leached, and the reaction reached equilibrium in 2 h. It could be noted that the rate used asfor the a leachingrelatively short agent. time. More than 90% of the REEs were leached in 10 min. At 1 h, almost 100% of of the leaching reaction was quite fast. Thus, it was sufficient to leach the ore at room temperature for a relatively short time.

Minerals 2016, 6, 63 8 of 15 the REEs were leached, and the reaction reached equilibrium in 2 h. It could be noted that the rate of the leaching reaction was quite fast. Thus, it was sufficient to leach the ore at room temperature for a relativelyMinerals short 2016 time., 6, 63 8 of 16

100

Minerals 2016, 6, 63 80 8 of 16

100 60

80 40 Percent leaching Percent

60 20

40 0 Percent leaching Percent 0 20406080100120

20 Time (min) Fe P Ca U Th Ce La Nd Pr Y 0 Figure 8. Change in leaching0 levels 20406080100120 with time using hydrochloric acid at 20 °C. Initial acid Figure 8. Change in leaching levels with time using hydrochloric acid at 20 ˝C. Initial acid concentration: concentration: 2.0 M; leaching time: 2 h; S/L ratio:Time 1:9. (min) 2.0 M; leaching time: 2 h; S/L ratio: 1:9. Fe P Ca U Th Unfortunately, the radioactiveCe elementsLa U and ThNd were alsoPr leached withY the REEs, although the level of radioactivity was statistically insignificant for the ore treated in this study (U: 45 mg/L, Figure 8. Change in leaching levels with time using hydrochloric acid at 20 °C. Initial acid Unfortunately,Th: 24 mg/L). the Additionally, radioactive the elementsTh leaching Ulevel and decreased Th were with also time leached because with Th precipitated the REEs, as although concentration: 2.0 M; leaching time: 2 h; S/L ratio: 1:9. the level ofphosphate radioactivity in the relatively was statistically low pH region. insignificant Thus, the radioactivity for the ore of treatedthe leaching in thissolution study could (U: be 45 mg/L, further decreased. Note, however, that if the radioactivity of the ore was significant, the removal of Th: 24 mg/L).Unfortunately, Additionally, the theradioactive Th leaching elements level U and decreased Th were also with leached time with because the REEs, Th although precipitated as radioactive elements from the leachate would require the use of separation techniques such as phosphatethe in level the of relatively radioactivity low was pH statistically region. Thus,insignificant the radioactivity for the ore treated of thein this leaching study (U: solution 45 mg/L,could be selective precipitation or the solvent extraction method. further decreased.Th: 24 mg/L). Note, Additionally, however, the that Th leaching if the radioactivity level decreased of with the time ore wasbecause significant, Th precipitated the removalas of Leaching tests were conducted at elevated temperatures to determine whether the leaching phosphate in the relatively low pH region. Thus, the radioactivity of the leaching solution could be radioactivelevels elements and reaction from rate the could leachate be improved. would requireBecause almost the use 100% ofseparation of the REEs were techniques leached using such 2.0 as selective further decreased. Note, however, that if the radioactivity of the ore was significant, the removal of precipitationM hydrochloric or the solvent acid extractionin an hour, method.tests of the temperature dependence were conducted using radioactive elements from the leachate would require the use of separation techniques such as 1.0 M hydrochloric acid at 20, 50, and 80 °C. As shown in Figure 9, the overall leaching levels for all Leachingselective tests precipitation were conducted or the solvent at extraction elevated method. temperatures to determine whether the leaching REEs did not increase significantly; however, other impurity ions (Fe, P, Ca, and U) were leached more levels and reactionLeaching rate tests could were beconducted improved. at elevated Because temperatures almost 100%to determine of the whether REEs were the leaching leached using as the temperature increased. Therefore, it is preferable to operate with a higher concentration of levels and reaction rate could be improved. Because almost 100% of the REEs were leached using 2.0 2.0 M hydrochlorichydrochloric acidacid at in room an hour, temperature tests ofthan the to temperatureoperate at elevated dependence temperatures. were conducted using 1.0 M hydrochloricM hydrochloric acid at 20, acid 50, andin an 80 hour,˝C. Astests shown of the intemperature Figure9, thedependence overall leachingwere conducted levels using for all REEs 1.0 M hydrochloric acid at 20, 50, and 80 °C. As shown in Figure 9, the overall leaching levels for all did not increaseREEs did significantly;not increase significantly;100 however, however, other other impurity impurity ions ions (Fe, P, P, Ca, Ca, and and U) were U) wereleached leached more more 20 ℃ as the temperatureas the temperature increased. increased. Therefore, Therefore, it it is is preferable preferable to to operate operate with50 with ℃a higher a higher concentration concentration of of hydrochlorichydrochloric acid at acid room at room temperature90 temperature than than to to operateoperate at at elevated elevated temperatures. temperatures.80 ℃

10080 20 ℃ 50 ℃ 80 ℃ 7090 Percent leaching

6080 20

70 0 Percent leaching Fe P Ca Ce La Nd Pr Y U Th 60 Figure 9. Comparison20 of final leaching levels using hydrochloric acid at 20, 50 and 80 °C. Initial acid concentration: 1.0 M; leaching time: 3 h; S/L ratio: 1:9. 0 Fe P Ca Ce La Nd Pr Y U Th

Figure 9. ComparisonFigure 9. Comparison of final of leachingfinal leaching levels levels using using hydrochlorichydrochloric acid acid at 20, at 50 20, and 50 80 and °C. 80Initial˝C. acid Initial acid concentration:concentration: 1.0 M; leaching1.0 M; leaching time: time: 3 h; 3 S/Lh; S/L ratio: ratio: 1:9.

Minerals 2016, 6, 63 9 of 15

Minerals 2016, 6, 63 9 of 16 Figure 10 shows the leaching levels at varying concentrations of nitric acid. No REEs, aside from Y, Figure 10 shows the leaching levels at varying concentrations of nitric acid. No REEs, aside from were leached at 1.0 M, whereas impurities such as P and Ca were leached quite well. However, as the Y, were leached at 1.0 M, whereas impurities such as P and Ca were leached quite well. However, as acid concentrationthe acid concentration was increased was increased to 2.0 toM, 2.0 almostM, almost 100% 100% of of the the REEsREEs were leached. leached. These These results results were similarwere to thesimilar hydrochloric to the hydrochloric acid case. acid Because case. Becaus nitrice acid nitric is generallyacid is generally more expensivemore expensive than than hydrochloric acid, ithydrochloric was not consideredacid, it was not in laterconsidered sections. in later Consequently, sections. Consequently, hydrochloric hydrochloric acid was acid chosen was as the optimumchosen leaching as the optimum agent forleaching this apatiteagent for REE this apat ore,ite and REE the ore, effects and the of effects other ofvariables other variables on the on leaching efficiencythe leaching were investigated efficiency were using investigated hydrochloric using hydrochloric acid in later acid experiments. in later experiments.

100

80 Fe P Ca 60 Ce La Nd 40

Percent leaching Percent Pr Y 20

0 0123 Nitric acid concentration (M)

˝ FigureFigure 10. LeachingLeaching levels levels at various at various initial nitric initial acid nitric concentrations acid concentrations at 20 °C. Leaching at 20 time:C. 2 Leaching h; S/L ratio: time: 1:9. 2 h; S/L ratio: 1:9. 3.3. Two-Stage Leaching 3.3. Two-StageAs mentioned Leaching previously, the leachate obtained using 2.0 M hydrochloric acid contained a significant amount of Ca, as the major mineral phase of the sample is apatite. Therefore, a two-stage Asleaching mentioned process was previously, attempted theto reduce leachate the amount obtained of Ca usingin the leachate. 2.0 M hydrochloric acid contained a significant amount of Ca, as the major mineral phase of the sample is apatite. Therefore, a two-stage leaching3.3.1. process Stage 1 wasLeaching attempted to reduce the amount of Ca in the leachate. In stage 1, 1.0 M hydrochloric acid was used to dissolve the apatite. Figure 11 shows the change 3.3.1.in Stage the leaching 1 Leaching levels of P and Ca ions together with the REEs over time. Within the first 10 min of leaching, the REE leaching levels were higher than those of P and Ca. This may result from the In stage 1, 1.0 M hydrochloric acid was used to dissolve the apatite. Figure 11 shows the change mineralogical structure of the ore. As mentioned earlier, the REE concentration was higher at the in theouter leaching edge of levels the apatite of P andgrains Ca than ions at the together center. The with monazite the REEs grains, over which time. bore Within a much the higher first 10 min of leaching,percentage the of REE REEs, leaching were also levels present were at the higher surface thanof the thoseapatite. of Conversely, P and Ca. the This amounts may of result P and from the mineralogicalCa increase structure gradually of from the ore.the edge As mentionedto the center earlier,of the apatite the REE grains. concentration Thus, a high wasinitial higher leaching at the outer edge oflevel the of apatite REEs is grainsexpected, than as the at theleaching center. solution The monazitereacts first with grains, the whichouter layer bore of athe much solid higherparticles. percentage of REEs,After were 10 min, also the present REE levels at the in surface the leachate of the decreased. apatite. It Conversely, seemed that thethis amounts could be attributed of P and Cato increase precipitation of REEs as . gradually from the edge to the center of the apatite grains. Thus, a high initial leaching level of REEs is To analyze the leaching behavior more specifically during stage 1, a thermodynamic analysis expected,was performed as the leaching to calculate solution the leaching reacts firstpH values with for the apatite, outer layerREE phosphates, of the solid and particles. Ca phosphate After 10 min, the REEusing levels thermodynamic in the leachate data available decreased. in the Itliteratur seemede [22–27]. that this It was could assumed be attributed that the concentration to precipitation of REEsof as the phosphates. substance reaches 1.0 mol/L for apatite and Ca phosphate, and 0.1 mol/L for REE phosphates. ToThe analyze leaching the pH leaching values behaviorwere calculated more specificallyon the basis duringof the stagechemical 1, areactions thermodynamic given in analysis Equations (8)–(10). The results are summarized in Tables 3 and 4. Fluorapatite, hydroxyapatite, and was performed to calculate the leaching pH values for apatite, REE phosphates, and Ca phosphate chlorapatite leach at pH values of 1.0, 3.2, and 2.3, respectively. Therefore, in the early stages of usingleaching thermodynamic with 1.0 M hydrochloric data available acid, in all the three literature forms of apatite [22–27 could]. It wasbe leached assumed from the that ore the because concentration of thethe substance pH of the acid reaches solution 1.0 was mol/L 0: for apatite and Ca phosphate, and 0.1 mol/L for REE phosphates. The leaching pH values were calculated on the basis of the chemical reactions given in Equations (8)–(10). The results are summarized in Tables3 and4. Fluorapatite, hydroxyapatite, and chlorapatite leach at pH values of 1.0, 3.2, and 2.3, respectively. Therefore, in the early stages of leaching with 1.0 M hydrochloric acid, all three forms of apatite could be leached from the ore because the pH of the acid solution was 0: Minerals 2016, 6, 63 10 of 15

Minerals 2016, 6, 63 ` 2` 10 of 16 Ca10pPO4q6X2 ` 20H “ 10Ca ` 6H3PO4 ` 2HX pX “ F, OH, Clq , (8)

` 2` ( ) Ca(POCa)3XpPO+ 20H4q2 `=6H 10Ca “+6H3CaPO`+2HX2H3POX=F,OH,Cl4,, (8) (9) Ca(PO) +6H` =3Ca3`+2HPO, (9) REEPO4 ` 3H “ REE ` H3PO4. (10) REEPO +3H =REE +HPO. (10)

100

80

60

40 Percent leaching

20

0 0 50 100 150 200 250 Time (min) Fe P Ca Ce La Nd Pr Y

Figure 11.FigureChange 11. Change in leaching in leaching levels levels over the the course course of stage of 1 stage using hydrochloric 1 using hydrochloric acid at 20 °C. Initial acid at 20 ˝C. acid concentration: 1.0 M; leaching time: 4 h; S/L ratio: 1:9. Initial acid concentration: 1.0 M; leaching time: 4 h; S/L ratio: 1:9. During leaching, it was found that the pH increased from 0 to approximately 1.5. In this range, Duringthe leaching,dissolved REEs it was and foundphosphate that ions the can pH react increased to form REE from phosphates, 0 to approximately as the leaching pH 1.5. for REE In this range, phosphates is less than or around 0 (Table 4). In contrast, Ca does not precipitate because both apatite the dissolvedand Ca REEs phosphate and phosphate are still leachable ions in can this react pH regi toon, form as shown REE phosphates,in Table 3. Therefore, as the Ca leaching dissolution pH for REE phosphatescontinues is less thanduring or this around stage, whereas 0 (Table the4). amount In contrast, of REEs Ca in doesthe liquid not phase precipitate decreases because as the REE both apatite and Ca phosphatephosphates are precipitate. still leachable in this pH region, as shown in Table3. Therefore, Ca dissolution continues during this stage, whereas the amount of REEs in the liquid phase decreases as the REE Table 3. Thermodynamic data and leaching pH of apatite and calcium phosphate according to phosphates precipitate.Equations (8) and (9), respectively (ΔG°f of H+: 0 kJ/mol; Ca2+: −553.58 kJ/mol; H3PO4: −1142.54 kJ/mol).

ΔG°f(kJ/mol) Table 3. ThermodynamicApatite data and leaching pH of apatite and calcium phosphateLeaching pH according to Apatite HX1 ΔG°rxn ˝ + 2+ Equations (8) and (9),Ca10 respectively(PO4)6F2 Fluorapatite (∆G f of H : 0– kJ/mol;12,885.9 Ca –296.82: ´553.58 kJ/mol;–98.76 H3PO41.00: ´ 1142.54 kJ/mol). Ca10(PO4)6(OH)2 Hydroxyapatite –12,503.5 –237.18 –361.93 3.24 Ca10(PO4)6Cl2 Chlorapatite –12,403.0 –˝131.23 –250.50 2.33 ∆G f(kJ/mol) ApatiteCa3(PO4)2 –3884.82 - –61.00 1.84Leaching pH 1 ˝ 1 X = F, OH, Cl for fluor, Apatitehydroxy and chlorapatite,HX respectively.∆G rxn

Ca10(PO4)6F2 Fluorapatite –12,885.9 –296.82 –98.76 1.00 Table 4. Thermodynamic data and leaching pH of REE phosphates according to Equation (10) (ΔG°f Ca10(PO4)6(OH)2 Hydroxyapatite –12,503.5 –237.18 –361.93 3.24 of H+: 0 kJ/mol; H3PO4: −1142.54 kJ/mol). Ca10(PO4)6Cl2 Chlorapatite –12,403.0 –131.23 –250.50 2.33 Ca3(PO4)2 Δ–3884.82Gf° (kJ/mol) - –61.00 1.84 REE Phosphate Leaching pH 3+ 1 X = F, OH, Cl for fluor,REEPO hydroxy4 REE andΔ chlorapatite,G°rxn respectively. YPO4 –1846.96 −686.92 17.50 –0.36 LaPO4 –1898.43 –730.46 25.43 –0.82 ˝ Table 4. Thermodynamic dataCePO and4 leaching–1827.32 pH of–677.33 REE phosphates7.45 according 0.23 to Equation (10) (∆G f of + PrPO4 –1880.33 –712.88 24.92 –0.79 H : 0 kJ/mol; H3PO4: ´1142.54 kJ/mol). NdPO4 –1870.92 –704.09 24.29 –0.75

∆G ˝ (kJ/mol) REE Phosphate f Leaching pH 3+ ˝ REEPO4 REE ∆G rxn

YPO4 –1846.96 ´686.92 17.50 –0.36 LaPO4 –1898.43 –730.46 25.43 –0.82 CePO4 –1827.32 –677.33 7.45 0.23 PrPO4 –1880.33 –712.88 24.92 –0.79 NdPO4 –1870.92 –704.09 24.29 –0.75 Minerals 2016,, 6,, 6363 11 of 1516

In Figures 7 and 11, it can be noted that the percentage of dissolved P is always less than that of In Figures7 and 11, it can be noted that the percentage of dissolved P is always less than that of Ca Ca for 1.0 M hydrochloric acid. This also implies that the dissolved P is consumed as the REE for 1.0 M hydrochloric acid. This also implies that the dissolved P is consumed as the REE phosphates phosphates form. Otherwise, the leaching levels of P, REEs, and Ca would be similar over time. form. Otherwise, the leaching levels of P, REEs, and Ca would be similar over time. Table5 presents Table 5 presents the hypothetical P leaching levels if no precipitation occurred. In these calculations, the hypothetical P leaching levels if no precipitation occurred. In these calculations, it was assumedREE + it was assumed that the REEs disappeared from the solution according3 `to the 3reaction´ that the REEs disappeared from the solution according to the reaction REE ` PO “ REEPO4; i.e., PO = REEPO; i.e., moles of phosphate ions and REEs were removed from the solution4 in a 1:1 ratio moles of phosphate ions and REEs were removed from the solution in a 1:1 ratio as precipitation as precipitation progressed. When the number of moles of phosphate ions consumed by the progressed. When the number of moles of phosphate ions consumed by the precipitation reaction is precipitation reaction is calculated and added to the experimentally measured phosphate level, the calculated and added to the experimentally measured phosphate level, the calculated leaching levels calculated leaching levels of P are very close to those of Ca throughout the process. This confirms that of P are very close to those of Ca throughout the process. This confirms that the REE phosphates the REE phosphates precipitate out of the solution during this stage. precipitate out of the solution during this stage. Table 5. Comparison of P and Ca leaching level during stage 1 leaching. Table 5. Comparison of P and Ca leaching level during stage 1 leaching. P Leaching Level Leach Time (min) Ca Leaching Level ActualP LeachingRe-Calculated Level Leach Time (min) Ca Leaching Level 1 Actual52.25 Re-Calculated52.26 52.93 3 60.26 60.28 60.86 1 52.25 52.26 52.93 37 60.2660.29 60.31 60.28 61.01 60.86 710 60.2959.29 60.23 60.31 62.07 61.01 1020 59.2959.83 74.53 60.23 76.31 62.07 2030 59.8354.33 69.30 74.53 72.41 76.31 30 54.33 69.30 72.41 4040 59.83 76.87 76.87 80.74 80.74 5050 58.44 75.20 75.20 79.01 79.01 6060 63.50 81.83 81.83 86.16 86.16 120120 60.24 77.70 77.70 81.80 81.80 180 63.42 81.83 86.16 180 63.42 81.83 86.16 240 62.21 80.25 84.37 240 62.21 80.25 84.37

Indeed, some REE phosphates were identified identified by XR XRDD analysis of the residue from stage 1, as shown in Figure Figure 12.12. This This indicated indicated that that it itwould would be be possible possible to todissolve dissolve Ca Caand and P selectively P selectively from from the theore orewhile while keeping keeping the theREEs REEs in inthe the solid solid residue residue by by using using 1.0 1.0 M M hydrochloric hydrochloric acid acid for for a reasonable leaching time, approximately 3–43–4 h.h.

Figure 12. XRD analysis of the residu residuee from stage 1, showing the presence of REE phosphates.

3.3.2. Stage 2 Leaching The solid residueresidue fromfrom stagestage 1 1 was was subjected subjected to to stage stage 2 2 leaching leaching using using 2.0 2.0 M M hydrochloric hydrochloric acid acid as aas leaching a leaching agent. agent. Figure Figure 13 shows 13 shows that thethat leaching the leaching of REEs of REEs reaches reaches nearly nearly 100% in100% a short in a time. short Table time.6 comparesTable 6 compares the leaching the leaching levels of levels REEs, of calculated REEs, calc asulated the ratio as the of theratio amount of the ofamount REEs inof theREEs leachate in the leachate after 120 min of leaching to that in the raw sample, with those obtained by one-stage leaching

Minerals 2016, 6, 63 12 of 16 under the same conditions. The amounts of the major impurities (Ca, P, and Fe) in the final leachate are also shown. The leaching levels of the REEs were not affected by the use of two-stage leaching, indicating that only the impurities were selectively leached during the first stage of leaching. As a consequence, the amounts of Ca and P in the final leachate were considerably reduced from 27,480 Mineralsand 11,2602016, mg/L6, 63 to 5030 and 4400 mg/L, respectively. 12 of 15

Table 6. Comparison of REE leaching levels in the leachate after 120 min and the amount of some afterimpurities 120 min of in leaching the final leachate to thatin be thetween raw one sample, and two-stage with those processes. obtained by one-stage leaching under the same conditions. The amounts of the major impurities (Ca, P, and Fe) in the final leachate are also REE Leach Level (%) Amount of Impurity (mg/L) shown. TheType leaching of Leaching levels of the REEs were not affected by the use of two-stage leaching, indicating that only the impurities were selectivelyCe La leachedNd duringPr theY first stageCa of leaching.P As aFe consequence , the amounts ofOne-stage Ca and P in the98.1 final leachate97.7 98.0 were 98.0 considerably 98.7 reduced27,480 from11,260 27,480 and190 11,260 mg/L to 5030 and 4400Two-stage mg/L, respectively.98.4 97.9 98.8 90.7 90.0 5030 4400 147

100

80

60

40 Percent leaching Percent

20

0 0 20 40 60 80 100 120 140 160 180 Time (min) Fe P Ca Ce La Nd Pr Y

FigureFigure 13.13. Leaching levels for various elements over th thee course of stage 2 usingusing hydrochloric acid. InitialInitial acidacid concentration:concentration: 2.02.0 M;M; leachingleaching time:time: 3 h; S/L ratio: ratio: 1:9. 1:9.

3.4. OptimalTable 6. ProcessingComparison Route of REE leaching levels in the leachate after 120 min and the amount of some impurities in the final leachate between one and two-stage processes. On the basis of the experimental results of this study, we propose a route for optimal processing and leaching of REEs from this ore, as shown in Figure 14. In this process, the ore is crushed and REE Leach Level (%) Amount of Impurity (mg/L) ground Typeto 90% of passing Leaching 0.3 mm. Then, magnetic separation is applied to separate Fe bearing minerals, which results in 60% removalCe of Fe with La a 5%–6% Nd loss Pr of REEs. Y The Fe Ca grade of the P magnetic Fe product is very highOne-stage (over 70%), so it 98.1 can be 97.7 sold to98.0 the steel-making 98.0 98.7 industry. 27,480 The 11,260nonmagnetic 190 stream is then subjectTwo-stage to a two-stage leaching 98.4 97.9 process. 98.8 In stage 90.7 1, the 90.0 ore is leached 5030 with 4400 1.0 M 147HCl, where approximately 90% of the Ca is removed with no loss of REEs. The residue from stage 1 leaching is 3.4.then Optimal subject Processingto stage 2 Routeleaching with 2.0 M HCl, where most of the REEs are leached out in an hour. The finalOn the recovery basis of of the REEs experimental after leaching results is ofaround this study, 92%–94% we propose for each a REE—a route for considerably optimal processing higher andamount leaching than that of REEs yield from by other this ore,current as shown processing in Figure practices. 14. In Table this process,7 summarizes the ore the is metallurgical crushed and groundbalance toof 90%the proposed passing 0.3 processing mm. Then, flow. magnetic separation is applied to separate Fe bearing minerals, which results in 60% removal of Fe with a 5%–6% loss of REEs. The Fe grade of the magnetic product is very high (over 70%), so it can be sold to the steel-making industry. The nonmagnetic stream is then subject to a two-stage leaching process. In stage 1, the ore is leached with 1.0 M HCl, where approximately 90% of the Ca is removed with no loss of REEs. The residue from stage 1 leaching is then subject to stage 2 leaching with 2.0 M HCl, where most of the REEs are leached out in an hour. The final recovery of REEs after leaching is around 92%–94% for each REE—a considerably higher amount than that yield by other current processing practices. Table7 summarizes the metallurgical balance of the proposed processing flow. Minerals 2016, 6, 63 13 of 16

Table 7. Metallurgical balance of the optimal process.

Elements (%) Distribution (%) Process Sample Yield (Wt %) Ca Fe Ce La Nd Pr Y Ca Fe Ce La Nd Pr Y Magnetic Magnetic 19.51 8.08 72.52 1.48 0.87 0.36 0.07 0.05 5.04 62.37 5.04 5.49 5.09 3.30 3.89 separation Nonmagnetic 80.49 36.93 10.61 6.76 3.64 1.65 0.53 0.32 94.96 37.63 94.96 94.51 94.91 96.70 96.11 Leachate 46.06 54.44 0.02 0.01 0.01 0.00 0.01 0.01 80.12 0.05 0.06 0.16 0.06 1.00 2.11 1st stage leaching Residue 34.42 13.50 24.78 15.80 8.51 3.85 1.22 0.73 14.84 37.59 94.90 94.35 94.85 95.70 94.00 2nd stage Leachate 17.14 27.01 2.20 31.11 16.70 7.64 2.40 1.45 14.79 1.66 93.05 92.24 93.70 93.71 92.76 leaching Residue 17.28 0.09 47.18 0.62 0.38 0.09 0.05 0.02 0.05 35.92 1.85 2.10 1.15 1.99 1.24 Calculated feed 100.0 31.30 22.69 5.73 3.10 1.40 0.44 0.27 100.00 100.00 100.00 100.00 100.00 100.00 100.00 Minerals 2016, 6, 63 13 of 15 Measured feed 100.0 32.75 22.91 5.80 3.01 1.48 0.42 0.28 ------

Figure 14. Optimal processing of the ore.

Table 7. Metallurgical balance of the optimal process.

Elements (%) Distribution (%) Process Sample Yield (Wt %) Minerals 2016, 6, 63; doi:10.3390/min6030063 Ca Fe Ce www.mdpi.com/journal/minerals La Nd Pr Y Ca Fe Ce La Nd Pr Y Magnetic 19.51 8.08 72.52 1.48 0.87 0.36 0.07 0.05 5.04 62.37 5.04 5.49 5.09 3.30 3.89 Magnetic separation Nonmagnetic 80.49 36.93 10.61 6.76 3.64 1.65 0.53 0.32 94.96 37.63 94.96 94.51 94.91 96.70 96.11 Leachate 46.06 54.44 0.02 0.01 0.01 0.00 0.01 0.01 80.12 0.05 0.06 0.16 0.06 1.00 2.11 1st stage leaching Residue 34.42 13.50 24.78 15.80 8.51 3.85 1.22 0.73 14.84 37.59 94.90 94.35 94.85 95.70 94.00 Leachate 17.14 27.01 2.20 31.11 16.70 7.64 2.40 1.45 14.79 1.66 93.05 92.24 93.70 93.71 92.76 2nd stage leaching Residue 17.28 0.09 47.18 0.62 0.38 0.09 0.05 0.02 0.05 35.92 1.85 2.10 1.15 1.99 1.24 Calculated feed 100.0 31.30 22.69 5.73 3.10 1.40 0.44 0.27 100.00 100.00 100.00 100.00 100.00 100.00 100.00 Measured feed 100.0 32.75 22.91 5.80 3.01 1.48 0.42 0.28 ------Minerals 2016, 6, 63 14 of 15

4. Conclusions The REE ore deposit currently under investigation can become a new major source of REEs judging from its magnitude and grade. Furthermore, test results show that this ore can be processed by a rather simple route with a recovery of all REEs of over 90% REEs. It seems that physical separation of REE-containing minerals may not be easy. However, removal of Fe-bearing minerals before leaching is possible using magnetic separation with minimal loss of REEs. This would reduce the operation cost of the leaching process because much of the acid could be consumed by Fe minerals in the ore. At the same time, additional profit can be realized by the sale of Fe concentrate as the magnetic product has a smelter grade. Leaching tests reveal that REEs can be effectively extracted using HCl at ambient temperature and pressure. The U and Th content of the leachate was not at a significant level that demands additional treatment. However, the Ca level was very high when direct leaching with HCl was performed. Consequently, a two-stage leaching process was developed. In stage 1, leaching is conducted under a mild condition (1.0 M HCl) to elute Ca from the ore. In stage 2 leaching, 90%–100% of the REEs were recovered using 2.0 M HCl. The final recovery of REEs after leaching is very high around 92%–94% for each REE—a considerably higher amount than that yield by other current processing practices.

Acknowledgments: This work was supported by a grant from the Korea Institute of Energy Technology Evaluation and Planning (KETEP) funded by the Ministry of Trade, Industry, and Energy (MOTIE) of the Korean government (No. 20122010300041). Author Contributions: Rina Kim performed the experiments, and analyzed and interpreted the data. Heechan Cho was the Principal Investigator of this work and edited the paper. Kenneth N. Han contributed to interpretation of the data. Kihong Kim and Myoungwook Mun performed magnetic separation tests and helped to analyze the data. Conflicts of Interest: The authors declare no conflict of interest.

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