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Recovery of an yttrium europium oxide phosphor from waste fluorescent tubes using a
Brønsted acidic ionic liquid, 1-methylimidazolium hydrogen sulfate
Nicolas Schaeffer, Xiaofan Feng, Sue Grimes*, Christopher Cheeseman
Department of Civil and Environmental Engineering, Imperial College, London SW7 2AZ
United Kingdom,
*Corresponding Author: [email protected]
Telephone: +44-0207-594-5966
Abstract
Background: Spent fluorescent lamps, classified as hazardous waste in the EU, are segregated
at source. Processes for the recovery of critical rare-earth (RE) elements from the phosphor
powder waste, however, often involve use of aggressive acid or alkali digestion, multi-stage
separation procedures, and production of large aqueous waste streams which require further
treatment.
Results: To overcome these difficulties phosphor powder pre-treated with dilute HCl was
leached with a 1:1 wt. [Hmim][HSO4]:H2O solution at a solid:liquid ratio of 5 %, at 80 oC for
4 h with stirring at 300rpm to recover 91.6 wt.% of the Y and 97.7 wt.% of the Eu present.
The yttrium-europium oxide (YOX), (Y0.95Eu0.05)2O3, recovered by precipitating the dissolved
RE elements from the leach solution with oxalic acid and converting the oxalate to an oxide
phase by heating, was characterised by FTIR, XRD and luminescence analysis. The analyses
suggest the recovered oxide has the potential to be directly reused as YOX phosphor.
Regeneration and reuse of the ionic liquid is achieved with only minor leaching efficiency
losses found over four leaching/recovery cycles.
Conclusion: The recovery of yttrium europium oxide from waste fluorescent tube phosphor
by a simple efficient low cost ionic liquid process has been developed.
Keywords: rare earth elements, strategic material recovery, hydrometallurgy, fluorescent
lighting phosphors, spent fluorescent lamps.
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1. Introduction
The rare earth (RE) elements (lanthanides, Y and Sc) have distinct metallurgical,
chemical, catalytic, electrical, magnetic and optical properties that are exploited in electrical
and electronic applications such as permanent magnets, rechargeable NiMH batteries, fibre
optics, light-emitting diodes (LEDs), catalysts and phosphors. Alternatives to rare earth
elements are limited and demand for these metals has increased dramatically in recent years1
and is projected to continue increasing at a rate of 3.7-8.6% per annum.2,3 The European Union
classify RE elements as critical materials for the future development of key technologies that
are subject to a high supply risk.3
Waste Electrical and Electronic Equipment (WEEE) is the fastest growing global waste
stream. It contains a variety of RE elements of which only about 1% are currently recovered
for re-use.4 Phosphor powders used in fluorescent lamps are chemically complex with some
compositional variation between manufacturers, but typically consist of a calcium
halophosphate matrix doped with RE elements (La, Ce, Eu, Gd, Tb and Y). They also contain
Al, Si, P, Ca and Ba at relatively high concentrations with Sr, Mg, Mn, Sb, Cl, F, Hg, Pb and
Cd present at trace levels.5 It is estimated that waste fluorescent lamps will contain up to 25,000
tonnes of RE elements by 2020 making waste phosphor a potential secondary source of these
critical metals.1
Spent fluorescent lamps, classified as hazardous waste in the EU, are segregated at
source through a well defined collection and treatment chain. Processing the waste typically
involves separating the Al end caps from the phosphor-coated glass, with the Hg present being
recovered by thermal desorption and distillation.6 The surface layer of the fluorescent tubes
containing the RE elements is separated from the bulk glass substrate, powdered and seived to
< 50µ to concentrate the RE elements in this size fraction. Conventional processes to solubilise
the RE elements in the phosphor powder and to recover them from solution generally involve
the use of aggressive acid or alkali digestion, multi-stage separation procedures, and the
production of large aqueous waste streams which require further treatment.7-11 Selective
recovery of RE elements from the acid or alkaline digestates can be achieved by a number of
processes12 including: precipitation, solvent extraction, ion exchange, electro-winning, and
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chromatography13 with solvent extraction and oxalate precipitation being the two most
commonly reported methods. An acid digestion process has been used, for example, to recover
yttrium oxide from a waste phosphor powder in 20% H2SO4 at 90oC using a solid:acid w/v
ratio of 1:5 followed by precipitation of the yttrium as yttrium oxalate.
Both room temperature ionic liquids (ILs), that usually consist of organic cations and
inorganic anions with melting points below 100 oC, and deep eutectic liquids can, if properly
selected14-17, offer environmentally beneficial alternatives to both organic and inorganic
solvents in selective leaching and recovery processes. The interactions between ionic liquids
and lanthanide and actinide elements have, for example, been reviewed by Binnemans.18
Exploitation of these interactions to RE element separation and recovery processes have been
described in a number of studies including: the homogeneous liquid–liquid extraction of
Nd(III) by choline hexafluoroacetylacetonate in the ionic liquid choline
bis(trifluoromethylsulfonyl)imide19; the use of functionalized ionic liquids in the solvent
extraction of RE elements20; the selective extraction and recovery of rare earth metals from
phosphor powders from waste fluorescent lamps using an ionic liquid system21; the solid–liquid
extraction of yttrium22 and of RE elements from waste fluorescent tube phosphors on ionic
liquid impregnated resins13; use of a functionalized ionic liquid in the solvent extraction of
trivalent RE metal ions23; the use of [Hbet][Tf2N]:H2O systems in the development of a one-
step process for the separation of light from heavy rare earth elements24; the recovery of rare
earth elements from simulated fluorescent powder using bifunctional ionic liquid extractants25
and the use of the mixed IL system ([DMAH][NTf2]):[Bmim][Tf2N] to dissolve the RE-
containing mineral, bastnaesite.26
The recovery efficiency of phosphors from their thin film oxide layers present in waste
fluorescent tubes is limited7 by a number of factors including, the grade of the rare earth oxide
product, reagent costs, energy consumption, and the environmental impacts of the process used.
Conventional acid- or alkali-based methods such as that used in the sulfuric acid leaching
processes (Figure 1) are limited by their use of aggressive leaching agents, the need to use
higher temperatures and the large volumes of aqueous waste streams produced. The use of
ionic liquids in recovering rare earth oxides from phosphors has been limited by factors such
as low metal loading capacity, long leaching times and high reagent costs. The fluorinated ionic
liquids such as [Hbet][Tf2N] can, for example, be 5-20 more expensive than alternative leach
media27. We now report on the use of a low-temperature recyclable low-cost Brønsted acidic
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ionic liquid leach step to replace sulfuric acid in a typical process (Figure 1) for the recovery
of RE element phosphors from waste fluorescent tubes.
Figure 1. Typical process for recovery of Y and Eu oxides.
2. Materials and methods
2.1 Materials
Waste phosphor was obtained from a leading UK fluorescent tube recycler (Balcan
Engineering Limited, UK). The X-ray data for the as-received waste phosphor are shown in
Figure 2. The Brønsted acid ionic liquid [Hmim][HSO4] (Figure 3) was synthesised following
the method described by Chen et al.27 All chemicals used were of reagent grade purchased
from either Sigma-Aldrich or Fisher Scientific and used without further purification.
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Figure 2. XRD analysis of the as-received waste phosphor with the main crystalline phases
identified
Figure 3. Schematic of 1-methyl imidazolium hydrogen sulfate IL - [Hmim][HSO4]
2.2 Experimental Techniques
The as-received phosphor fractions were characterised by particle size analysis using a
Beckman Coulter LS-100 Series with a particle size detection range between 0.4 and 900μm,
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by X-ray fluorescence (XRF) analysis using a Bruker S4 Explorer spectrometer and by high
resolution scanning electron microscopy using a field emission gun scanning electron
microscope (FEGSEM) fitted with an Oxford Instruments INCA energy dispersive (EDS) X-
ray spectrometer (Gemini 1525, USA). The main crystalline phases in both the as-received
phosphor and in the final product of the recovery process were identified by X-ray diffraction
(XRD) using a Philips X’pert PRO PANalytical with an automatic divergence slit, a graphite
monochromator and Cu-Kα radiation and PANalytical X'Pert HighScore Plus software. Metal
analyses on phosphor leach solutions and the final recovered product were obtained, in
triplicate, using inductively coupled plasma-optical emission spectroscopy (ICP-OES, Perkin
Elmer Optima 7000 DV). All sample solutions were diluted to avoid potential interference in
the analyses by the ionic liquid. The recycled ionic liquid and the recovered oxide were
characterised, by comparison with pure samples, using Fourier transform infrared (FTIR)
spectra on a Thermo Scientific Nicolet 6700 spectrometer with samples examined directly
using a Quest single reflection diamond attenuated total reflection (ATR) accessory.
Luminescence measurements on the recovered RE oxide phase in the range 350-750 nm were
performed on a Jasco FP-6000 Series spectrofluorometer equipped with a 150 W xenon lamp
with an excitation of 254 nm.
The ionic liquid-based methodology developed for the recovery of a YOX phosphor
from waste fluorescent tubes in this work involved five stages – sample characterisation and
pre-treatment; optimisation of leaching with an ionic liquid; rare-earth element concentration
as oxalates; YOX product characterisation; and ionic liquid recovery for recycle, with the
characterisation techniques used in each stage being shown in the following flow diagram:
2.3 Phosphor Pre-Treatment
XRF analysis of the as-received phosphor from waste fluorescent tubes (Table 1) has
the particle size distribution that shows that RE elements account for about 18.3 wt.% of the
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as-received waste phosphor powder with Ca, P, Si, Al and Ba arising mainly from the glass
matrix elements being the other major elements present. SEM micrographs of the as-received
phosphors indicate that the larger (≥25 µm) are composed mainly of the glass matrix (identified
by the presence of strong Si peaks in EDS analysis) whilst the smaller diameter particles
contain most of the phosphor powder (identified by the presence of strong yttrium EDS peaks)
(Figure 4). The RE elements are concentrated in the waste phosphor powder by two pre-
treatment stages: (i) removal of the > 25 µm particles from the as-received material by sieving
to leave the ≤ 25 µm fraction in which the phosphor is concentrated. The ≤ 25 µm fraction
represented only 10% of the total waste sample but contained 63% of the rare earth elements
present and for this reason the ≤ 25 µm fraction was used to demonstrate the methodology
developed in this work. The largest size fractions (>50 µm) contained only 15% of the rare
earth elements but could be further crushed to increase the mass of the ≤25 µm fraction; and
(ii) leaching the sieved powder (≤ 25 µm fraction) with 0.5M HCl. The data in Table 1 show
that (a) the ≤ 25 µm fraction of the waste phosphor has a greater Y content (17.6 %), and an
increased total RE content (26.5%) than the as-received powder, and (b) contacting the ≤ 25
µm fraction of the waste phosphor for 1 hour at room temperature with 0.5 M HCl at a
solid:liquid ratio of 1:10 with mixing at 500 rpm. The HCl treatment results in the removal of
relatively soluble glass matrix elements to give a final RE content in the treated waste phosphor
of >22%.
Table 1. Chemical analysis of the phosphor powder and percentage leached during the HCl
pre-treatment step
Concentration (wt.%)
As-received†
≤ 25 µm fraction†
≤ 25 µm fraction after
HCl pre-treatment‡
Y 11.7 17.6 26.5
Eu 1.1 1.3 1.9
La 2.6 3.5 5.4
Ce 1.2 2.4 3.7
Tb 1.1 1.2 1.9
Gd 0.6 0.5 0.8
Ca 21.2 23.8 20.9
Ba 4.5 4.4 2.5
Al 4.5 7.4 11.1
Na 1.0 1.9 2.4
Fe 1.0 0.9 0.7
Others 49.5 35.1 22.2
† XRF analysis, ‡ ICP-OES analysis
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Figure 4. SEM images of the as-received phosphor powder at a) 500× magnification and b)
2500× magnification. c) EDS spectrum of three regions in Figure 4b highlighted in white
2.4 Optimisation of waste phosphor leaching using [Hmim][HSO4]
All leaching tests were performed on the HCl pre-treated ≤ 25 µm fraction of the waste
phosphor to limit the amount of residual glass particles present. Initial leaching conditions in
50 wt.% aqueous [Hmim][HSO4] with stirring at a mixing velocity of 300 rpm were: leaching
time, 2 h, leaching temperature, 80 oC, and a solid:liquid ratio, 10%. Leaching parameters were
varied individually whilst maintaining all others at the initial test conditions to obtain the
optimum leaching conditions with respect to [Hmim][HSO4] concentration (0-100 wt.%),
temperature (20-100 oC), leach time (0.5-24 hr), and solid:liquid ratio (1:20-1:5). The leach
solutions obtained under different conditions were filtered through a 0.22 µm cellulose nitrate
filter prior to analysis of RE element content by ICP-OES. The variations in leaching
efficiencies (wt %) of Y, Eu, Tb, Ca and Al with, [Hmim][HSO4] concentration, temperature,
contact time and solid:liquid leach ratio, are in Figures 5a-d respectively.
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Figure 5. Leaching efficiency of Y, Eu, Tb, Ca and Al as a function of (a) [Hmim][HSO4]
concentration, (b) temperature, (c) contact time and (d) solid:lquid ratio.
2.5 RE oxide recovery from aqueous [Hmim][HSO4] leachate by oxalic acid precipitation
followed by conversion to oxide
The recovery of REEs from the optimised aqueous [Hmim][HSO4] leach solution by
oxalic acid precipitation was optimised (Figures 6a-d) with respect to temperature (20-60 oC),
contact time (15-45 min), pH (0.9-1.4) and oxalic acid:REE molar ratio (0.8-1.75). The leachate
pH was adjusted using a 2 M NH3 solution. The final product from the oxalate precipitation
was filtered off, washed with deionised water and dried at 105 oC before conversion to a RE
oxide phase by raising the temperature of the solid to 650 oC for 1 hour. The final oxide product
was washed with deionized water and dried at 105 oC, and characterised using ICP-OES, FTIR,
XRD and luminescence measurements.
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Figure 6. The recovery efficiency of the main leached element as a function of (a) temperature,
(b) contact time, (c) solution pH, and d) OA:REE molar ratio.
2.6 [Hmim][HSO4] recycling and re-use over four leaching and recovery cycles
Following the optimised leaching and recovery process, the aqueous IL solution was
diluted with 5 times the quantity of distilled water and shaken with an excess of XAD-DEHPA
solvent impregnated resin for one hour to extract the non-precipitated metal components onto
the resin28. After filtering off the solvent impregnated resin, [Hmim][HSO4] was recovered by
evaporating the water present under vacuum at 60 oC for 4 hr. The recovered [Hmim][HSO4]
was then re-used in a further three cycles of recovering RE oxides from the phosphor in the
optimised leaching process.
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3. Results and Discussion
3.1. Pre-treatment
The main crystalline components identified by XRD (Figure 2) in the as-received
phosphors in addition to the yttrium europium oxide ((Y0.95Eu0.05)2O3) known to be present are
fluorapatite (Ca5(PO4)3F), cerium phosphate (CePO4) and barium silicate (BaSi2O5), and
consistent 8-9,21 with Y and Eu being present in the phosphors as oxide phases and Ce, Gd and
Tb being present as phosphates.
Removal of the large particles (>25µm) from the as-received phosphor followed by
leaching with dilute hydrochloric acid results in the concentration of the Y present in the waste
phosphor from 11.7 to 17.6 and finally to 26.5%,. and of the other REEs from 6.6 to 8.9 and
finally to 13.7% (Table 1). Leaching of the ≤ 25µm fraction with 0.5M HCl further removes
35.5 wt.% of the non-RE element content including over 40% of the Ca (the main component
of the <25µm fraction) present while leaching out only 2.8 wt.% of the RE elements (Table 1).
3.2. Optimisation of the [Hmim][HSO4]:H2O phosphor leaching process
The data in Figures 5 (a-d) showing the variations in the solubility of Y, Eu, Tb, Ca,
and Al in the pre-treated phosphor waste with [Hmim][HSO4] concentration, temperature,
phosphor:leach solution ratio, and contact time, have been used to determine the optimum leach
conditions for the recovery of RE elements from the phosphors.
The solubilities of both of Y and Eu increase with increasing [Hmim][HSO4]
concentration in the aqueous leach solution to reach a maximum at 1:1 [Hmim][HSO4]:H2O
wt% (Figure 5a). In addition to the lower solubilities of the target elements Y and Eu in
solutions containing higher percentages of [Hmim][HSO4], the viscosities of the solutions
increase making them less suitable leach media - the viscosity of the 1:1 [Hmim][HSO4]:H2O
mixture at 20 oC (4.2 cP), for example, increases to 6.9 cP for a 98 wt.% [Hmim][HSO4]
solution. The efficiency of the leaching of Y and Eu increases by 50.6 wt.% and 73.7 wt.%
respectively as the temperature is increased from 22 oC to a maximum at about 80 oC (Figure
5b); and the leaching of both Y and Eu from phosphor wastes increases with time up to about
4 hours (Figure 5c). Higher temperatures are required to overcome the low solubility of RE
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phosphates present in the phosphor and at 100 oC, 36.0 wt% of Gd, 3.7 wt.% of Tb, 2.1 wt.%
of Ce and 0.14 wt.% of La are leached into the aqueous [Hmim][HSO4] solution. The recovery
of Eu (92.4%) and Y (81.4%) is high when phosphor:liquid ratios of 5% are used but decreases
with increasing solid: liquid ratio (Figure 5d) to 46.5 and 32% respectively with a solid:liquid
ratio of 1:5.
The results show that the optimum conditions for the use of [Hmim][HSO4]:H2O mixtures to
leach the rare earth elements Y and Eu from treated phosphor samples are: use of a 1:1 wt%.
[Hmim][HSO4]:H2O solution at 80 oC for 4 hours with a solid: liquid ratio of 5 %, and a mixing
speed of 300 rpm. Under the optimised leaching conditions, high recoveries of Y of 92% and
of Eu of 98% are achieved (Table 2).
Table 2. Percentage leached of elements present in the phosphor powder under the optimised
[Hmim][HSO4] leach conditions
Element wt.% leached
Y 91.6 ± 2.06
Eu 97.7 ± 2.23
Gd 39.5 ± 0.81
Tb 4.05 ± 0.13
Ce 2.40 ± 0.07
La 0.20 ± 0.01
3.3. Optimisation of Y and Eu recovery by oxalic acid precipitation followed by conversion
to oxide.
RE elements can be selectively recovered from acid or ionic liquid solutions by
precipitation with oxalic acid. Use of solid oxalic acid allows for (i) an efficient one-step in-
situ recovery process, (ii) recovery of [Hmim][HSO4] for re-use, and, (iii) easy conversion of
the RE oxalate product to an oxide phase by calcination. The data in Figure 6 show that the
optimum conditions for the precipitation of rare earth elements from the leach solutions with
oxalic acid are: use of a slightly greater than stoichiometric oxalic acid: RE ratio of 1.6:1; at
40oC and pH 0.9 with a contact time of 0.5h. The product from the oxalate precipitation was
filtered off, washed with deionised water and dried at 105 oC before conversion to an RE oxide
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phase by calcination at 650 oC to give an oxide product with an average particle diameter of
1.4 ± 0.14 µm that was characterised by, XRD and photoluminescence analysis.
3.4. Characterisation of the recovered oxide product
The recovered oxide product has been characterised using ICP-OES, FTIR, XRD and
luminescence measurements.
The XRD of the recovered product (Figure 7) is consistent with (Y0.95Eu0.05)2O3. This
is chemically similar to the most widely used YOX phosphor for the emission of red light
(Y2O3:Eu3+). The FTIR spectrum shows the recovered oxide product to be identical to a
commercially YOX phosphor, with the characteristic peaks at 555, 461 and 415 cm-1.
Figure 7. XRD analysis of the recovered oxide product after calcination
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The luminescence spectrum of the recovered powder under 254 nm wavelength
excitation (Figure 8) exhibits the optical properties of Eu3+ ions in a cubic Y2O3 host structure29-
31. The emission spectrum displays the characteristic 5D0 7Fj (j=0, 1, 2, 3) and 5D1 7F0 of
Eu3+ transitions. The high intensity red hypersensitive 5D0 7F2 transition at 612 nm is the
dominant emission with the 5D0 7F3, 5D0 7F1,
5D0 7F0, and 5D1 7F0 emission peaks
located at 631 nm, 587 – 600 nm, 582 nm and 533-539nm respectively The SEM image in
Figure 8 shows the final oxide product of the process under excitation at 254 nm. The
luminescent life-time of the sample was measured by following the decay of the 612 nm
emission for 25 ms after excitation at 254 nm. The resulting decay curves were fitted with a
monoexponential decay equation (R2=0.99)31 to give: It = 8.97et/0.72, where It is the intensity at
time t (ms), 8.97 the intensity at t = 0 (I0) and 0.72 ms is the luminescence lifetime (τ). The
quality of the luminescence spectrum and luminescence lifetime (0.72 ms) of the recovered
oxide product confirms that it has potential for direct re-use as YOX phosphor.
Figure 8. Luminescence spectra of the recovered powder under 254 nm wavelength excitation.
Inset photograph: Recovered oxide product under 254 nm light.
3.5. [Hmim][HSO4] recycling and reuse and a process flow
To be considered as a viable economic and environmental alternative to inorganic acids
such as H2SO4, [Hmim][HSO4] must possess a high selectivity and extraction capacity for Y
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and Eu, incur minimal losses and possess excellent chemical stability. To assess the stability
of [Hmim][HSO4], the IL was recycled and re-used over four leaching and recovery cycles
under optimised conditions and the results are presented in Figure 9.
Figure 9. FTIR spectra of virgin and four time recycled [Hmim][HSO4]
Leaching of the REEs decreased slightly over the four leaching/stripping cycles with
decreases of 1.3 wt.% for Y, 5.6 wt.% for Eu and 11.3 wt.% for Gd. Complete removal of all
added oxalic acid is ,however, essential prior to re-use as it has a detrimental effect on the
leaching efficiency of RE oxides. The small decrease in Eu and Gd leaching efficiency is most
likely due to (1) a lower proton content in [Hmim][HSO4] and/or (2) partial saturation of the
IL. This is easily remediated by addition of H2SO4 in [Hmim][HSO4] to compensate for the
loss of H+ ions.
The recovery of yttrium europium oxide (Y0.95Eu0.05)2O3 from waste fluorescent tube
phosphor by a simple 4-step efficient low cost ionic liquid process, that includes recycle and
reuse of the ionic liquid [Hmim][HSO4] is presented in Figure 10.
Recently, Chen and co-workers27 demonstrated that the Brønsted acidic IL, 1-
methylimidazolium hydrogen sulfate ([Hmim][HSO4]) can be produced on an industrial scale
at prices between $2.96 and $5.88 kg−1 and it therefore provides a less expensive material for
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use as a solvent particularly if it can be used in closed cycle applications like the recovery of
the YOX phosphor.
Figure 10. Schematic diagram of a process for the recovery of (Y0.95Eu0.05)2O3 from waste
fluorescent tubes using [Hmim][HSO4]
4. Conclusions
The recovery of yttrium europium oxide [(Y0.95Eu0.05)2O3] from waste fluorescent tube
phosphor by a simple efficient low cost IL process is described; a process which replaces a
sulfuric acid leach in the standard process. The optimum conditions for solubilising RE oxides
from acid pre-treated phosphor powder involves leaching with a 1: 1 wt. [Hmim][HSO4]:H2O
solution at a solid: liquid ratio of 5 %, 300 rpm, at 80 oC for 4 h. The percentage Y, Eu and Ca
leached under these conditions is 91.6 wt. %, 97.7 wt.% and 24.9 wt.% respectively. The REEs
in the leachate were precipitated by addition of oxalic acid at a molar ratio (OA: REE) of 1.5,
a stripping temperature of 60 oC, mixing time of 15 minutes and a solution pH of 0.9. Calcining
the oxalates at 650 oC for 1 h gives an 86.0 wt.% mixed rare earth oxide (Y0.95Eu0.05)2O3
recovered product with Ca being the major residual impurity. Luminescence analysis indicates
that the recovered yttrium europium oxide has the potential to be directly reused as YOX
phosphor. The replacement of the sulfuric acid leach stage with the ionic liquid [Hmim][HSO4]
leach has the following process advantages; it avoids both the use of aggressive acid leaching
that may also involve high temperatures and the production of large volumes of aqueous waste
water. The economics of the ionic liquid process are also favourable because [Hmim][HSO4]
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can now be synthesised in bulk at less than $6/kg27 and can be recovered for reuse in the process
over at least four recycle steps.
Acknowledgement
We wish to acknowledge an EPSRC DTA Scholarship for N.S. and the EPSRC UK National
Mass Spectrometry Facility at Swansea University, United Kingdom, for mass spectrometry
analysis. The authors would like to thank the Chemistry Department at Warwick University for
performing the luminescence analysis.
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