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Article

Processing of a Phosphate Flotation Tails for Recovery of Rare Earths and Phosphate

1
Florida Industrial and Phosphate Research Institute, Florida Polytechnic University, Bartow, FL 33830, USA
2
Oak Ridge National Laboratory, Oak Ridge, TN 37831, USA
*
Author to whom correspondence should be addressed.
Minerals 2025, 15(9), 900; https://doi.org/10.3390/min15090900 (registering DOI)
Submission received: 26 June 2025 / Revised: 7 August 2025 / Accepted: 21 August 2025 / Published: 25 August 2025
(This article belongs to the Special Issue Circular Economy of Remining Secondary Raw Materials)

Abstract

Phosphorite, or phosphate rock, has garnered increasing attention in recent years as a promising unconventional resource for rare earth elements (REEs). This paper presents a processing scheme aimed at recovering both REEs and phosphate values from amine flotation tailings generated during phosphate beneficiation in Florida. In these tailings, REEs are primarily present as monazite and xenotime, often associated with heavy minerals. The proposed flowsheet includes gravity separation to pre-concentrate REE- and phosphate-bearing minerals, followed by flotation to further upgrade both REEs and phosphate, and finally sulfuric acid leaching to extract REEs and phosphate from the flotation concentrate. Gravity separation using a shaking table increased the total REE content from approximately 202 ppm to 657 ppm, with a concentrate yield of 12.51%, REE recovery of around 41%, and P2O5 recovery of 33%. Fatty acid flotation of the shaking table concentrate produced a final concentrate containing 1106 ppm REEs and 14.90% P2O5, with recoveries of approximately 86% for REEs and 90% for P2O5. Subsequent pyrolysis with concentrated sulfuric acid followed by water leaching achieved recoveries of about 85% for REEs and 93% for P2O5. While the process demonstrated effective concentration and leaching of REE minerals and apatite, the major challenge to further improving separation and extraction efficiency lies in the fine-grained nature of the valuable minerals and their interlocking with gangue minerals.

1. Introduction

Phosphorite, or phosphate rock, has been used as a feedstock for phosphate fertilizer production for more than 150 years. In modern manufacturing, phosphate rock is produced through beneficiation of phosphate ore, which involves removing large volumes of waste materials, including phosphatic clay and flotation tailings. The resulting phosphate concentrate then undergoes a “wet-process,” in which it is digested by mineral acids (typically sulfuric acid) to produce phosphoric acid. This process generates two major by-products: phosphogypsum and phosphoric acid sludge. Finally, phosphoric acid is converted into various phosphate fertilizers [1].
In recent years, phosphorite has attracted renewed interest for its potential as an unconventional resource for rare earth elements (REEs) [2,3,4,5,6,7,8,9,10,11,12]. Kanazawa [13] estimated that global phosphate resources contain approximately 50 million tons of REEs. Other studies [14,15] have shown that phosphate rock typically contains an average of 0.046 wt% REEs, and with around 250 million tons of phosphate rock mined globally each year, more than 110,000 tons of REEs are introduced into the phosphate industry annually—most of which are lost during processing. In Florida alone, estimated annual REE losses include approximately 5800 tons in phosphatic clay, 3400 tons in phosphogypsum, 1100 tons in phosphoric acid sludge, 780 tons in flotation tails, and 620 tons in phosphate fertilizer. If these REEs could be efficiently recovered, the total amount could meet the entire U.S. demand—and yttrium (Y) recovery alone could potentially satisfy global demand.
Central Florida hosts large sedimentary phosphate rock deposits that have been mined and processed for over a century. In current beneficiation operations, phosphate ore is first washed and classified to remove micro-fine clay (−150 mesh), producing three size fractions: pebble (+16 mesh), coarse flotation feed (−16 + 35 mesh), and fine flotation feed (−35 + 150 mesh). The coarse and fine feeds undergo the “Crago” double flotation process—direct flotation using fatty acid as a collector to concentrate phosphate minerals, followed by reverse flotation using amine to remove residual silica [16]. The resulting phosphate concentrate is then sent to the acid plant, where it is digested by sulfuric acid using the dihydrate wet-process to produce phosphoric acid. This entire process generates substantial waste streams, including phosphatic clay, fatty acid and amine flotation tails, and phosphogypsum.
Numerous studies have confirmed the presence of REEs in Florida phosphate resources [17,18,19,20]. As early as 1989, Kremer and Chokshi [21] of Mobile Research & Development Corporation conducted a comprehensive investigation, finding that phosphate ore (matrix) contained approximately 282 ppm total REEs. Their study revealed that, after beneficiation and chemical processing, REEs are distributed across various waste streams: 40% in phosphatic clay, 37.5% in phosphogypsum, 12.5% in phosphoric acid, and 10% in flotation tails. More recent work by Zhang [4] and Poul et al. [22] estimated that the total annual REE content in Florida phosphate ore exceeds 30,000 tons—significantly surpassing current U.S. demand. Notably, compared to traditional REE ores such as bastnäsite, Florida phosphate resources contain higher proportions of heavy REEs, which are both rarer and more critical for high-tech applications and clean energy technologies. Therefore, recovering REEs from Florida phosphates could play a key role in establishing a secure and resilient domestic REE supply chain.
As a founding member of the U.S. Department of Energy’s Critical Materials Innovation (CMI) Hub, the FIPR Institute has been actively conducting research on REE recovery from Florida phosphate resources. This paper presents a processing scheme aimed at recovering both REEs and phosphate values from amine flotation tailings. The proposed approach includes gravity separation to pre-concentrate REE-bearing minerals, flotation of the gravity concentrate to further upgrade REEs and phosphate, and concentrated sulfuric acid pyrolysis followed by water leaching to extract REEs and phosphate from the final concentrate.

2. Materials and Methods

2.1. Mineral Beneficiation

Investigations revealed that REEs in the flotation tailings exist primarily as monazite and xenotime, often associated with heavy minerals. Most of the phosphorus is hosted in francolite, which is typically associated with coarse gangue minerals [23,24]. Given the low concentrations of REEs and P2O5 (commonly used to represent phosphorus content in minerals) in the tailings, it is essential to employ a cost-effective and efficient beneficiation method to eliminate as much gangue material as possible.
Gravity separation using a shaking table was selected for pre-concentrating REE-bearing minerals and phosphate, based on the following considerations:
  • The specific gravities of monazite (4.6–5.7), xenotime (4.4–5.1), and phosphate rock (3.16–3.22) are significantly higher than that of quartz (~2.65), which is the predominant gangue mineral in the tailings.
  • The particle size distribution of the tailings is well-suited for gravity separation.
  • Shaking tables offer a low-cost and simple method for gravity-based separation.
Accordingly, a shaking table was used to remove a large portion of gangue minerals from the tailings. The shaking table separation was conducted at a feed rate of 50 kg/h. The resulting concentrate was then ground and further upgraded via flotation. Flotation was conducted using a Denver mechanical flotation cell with a volume of one liter and feed charge of 500 g at 1200 rpm. Slurry density in the flotation cell was 30%.

2.2. Leaching of REEs and Phosphorus

To recover REEs and phosphorus from the final concentrate, a process involving concentrated sulfuric acid pyrolysis followed by water leaching (CSAP–WL) was applied. In this method, the concentrate was first mixed with concentrated sulfuric acid, then roasted at 350 °C for 4 h. After cooling, water was added and the mixture was stirred for 2 h to leach REEs and phosphorus.
During roasting, the REE minerals and phosphate minerals (primarily fluorapatite) react with sulfuric acid according to the following reactions:
2REPO4 + 3H2SO4 → RE2(SO4)3 + 2H3PO4
Ca5(PO4)3F + 5H2SO4 → 5CaSO4 + 3H3PO4 + HF↑
where RE denotes a rare earth element.

3. Results and Discussion

3.1. Pre-Concentration with Shaking Table

A representative sample of amine flotation tails was collected from beneficiation plant at the Four Corners Mine, one of the phosphate mines in central Florida. The sample was assayed for main components and REE content, Table 1.
Quantitative mineral analysis was conducted using a Mineral Liberation Analyzer (MLA). Two rare earth minerals were identified by MLA, they are monazite and xenotime, both at very low concentrations. Major non rare earth minerals include quartz, fluorapatite, feldspar, zircon, and rutile, Table 2.
The size range of the mineral particles was also analyzed using MLA, Table 3.
A shaking table was used to concentrate the REE-bearing minerals and apatite from the amine flotation tailings. The operating conditions of the shaking table were optimized first and determined as follows:
  • Transverse angle: 5°
  • Stroke length: 7 mm
  • Stroke frequency: 310 revolutions per minute
  • Wash water: 3.5 gallons per minute
  • Tailings feed: 583 g per minute
Then, 500 kg of the sample was used in shaking table concentration testing, and the separation results are presented in Table 4.
The pre-concentration step removed up to 87.49 wt% of the solids from the feed, yielding a concentrate with significantly higher REE and P2O5 contents compared to the original tailings. However, the recoveries of REEs and P2O5 were relatively low—40.73% and 33.12%, respectively. These results suggest that a substantial portion of the valuable minerals in the amine flotation tailings exist as fine particles interlocked with gangue minerals. Without sufficient grinding to liberate these minerals, effective separation cannot be achieved.

3.2. Flotation

The shaking table concentrate was first ground in a ball mill to improve the liberation of REE-bearing minerals and apatite, followed by flotation for further beneficiation. Sodium carbonate was used to adjust the pulp pH, sodium silicate served as a gangue depressant, and sodium oleate was employed as the collector for both REE and phosphate minerals. The flotation pulp concentration was maintained at 30 wt%.
Operating conditions were optimized through a series of flotation tests, with the results presented in Figure 1, Figure 2, Figure 3 and Figure 4. Based on these tests, the optimal conditions were determined as follows: grinding fineness of 52.70 wt% passing 200 mesh (0.074 mm), pulp pH of 9.5, depressant dosage of 0.55 kg/t (based on tailings feed), and collector dosage of 1.2 kg/t (also based on tailings feed).
A rougher flotation test was then conducted under these optimized conditions, and the results are summarized in Table 5.
The results presented in Table 5 indicate that the shaking table concentrate can be further upgraded through flotation after grinding. However, the P2O5 content in the resulting flotation concentrate was only 14.90%, significantly lower than the typical 28% P2O5 found in the standard “Crago” flotation concentrate used for phosphoric acid production. Test observations revealed that the P2O5 grade of the flotation concentrate could not be substantially improved, even when grinding fineness was increased beyond 52.70 wt% passing 200 mesh (0.074 mm). This limitation is likely due to the presence of valuable minerals as micro-fine particles, tightly interlocked with gangue in the amine flotation tailings. The chemical composition and REE content of the flotation concentrate are provided in Table 6.

3.3. Leaching

The CSAP–WL leaching method described above is a widely used process for decomposing REE minerals, including monazite. To evaluate its effectiveness, the impact of sulfuric acid dosage on the leaching efficiencies of REEs and phosphorus was investigated. The results are presented in Figure 5.
As shown in Figure 5, the leaching efficiency of REEs initially increased with the rising stoichiometric ratio of H2SO4 to CaO, reaching a maximum of 85.54% at a ratio of 3.61, after which it plateaued. In contrast, the leaching efficiency of P2O5 exhibited a different trend: it remained relatively stable, ranging from 88.5% to 90.5% as the H2SO4/CaO ratio increased from 1.85 to 3.26. Beyond this point, P2O5 recovery increased steadily, reaching 95.33% at the highest tested ratio of 4.60.
The sulfuric acid consumption during leaching was significantly higher than the theoretical requirement for phosphate decomposition, primarily due to the high impurity content in the flotation concentrate.
It is important to note that the downstream processing of REEs from the leachate, including extraction and production of mixed rare earth oxides, will be detailed in separate papers to be published in the near future.

4. Conclusions

The research results presented in this paper demonstrate that REEs and phosphates in amine flotation tails can be effectively concentrated and leached using a combined separation and leaching process. Gravity separation using a shaking table increased the REE and P2O5 contents from approximately 202 ppm and 3% in the tails to 657 ppm and 8%, respectively, with a concentrate yield of 12.51%, achieving recoveries of about 41% for REEs and 33% for P2O5. After grinding, the shaking table concentrate was further upgraded by flotation, producing a concentrate containing 1151 ppm REEs and 14.90% P2O5, with recoveries of 85.91% for REEs and 89.81% for P2O5. Leaching tests showed that REEs and phosphorus could be recovered from the flotation concentrate via concentrated sulfuric acid pyrolysis followed by water leaching, resulting in approximately 85% recovery of REEs and 93% recovery of P2O5.
However, several challenges remain with the current processing approach. First, the REE minerals and apatite are finely disseminated within the amine flotation tails, with many particles interlocked with gangue minerals. This led to relatively low recoveries in the shaking table separation—40.73% for REEs and 33.15% for P2O5—and the limited upgrading of grades in the flotation concentrate. The final flotation concentrate, with 1150.82 ppm REEs and 14.90% P2O5, does not yet meet industry standards. Given that the amine flotation tails contain over 90 wt% gangue minerals (based on a P2O5 content of 3.02%), it is not economically viable to grind the entire tailings before removing the majority of gangue due to the high energy requirements.
The second issue is the high sulfuric acid consumption required in the roasting–leaching step. The acid additions tested across a stoichiometric ratio range of 1.85 to 4.60 (H2SO4 to CaO) were all substantially higher than the theoretical amount needed to decompose phosphates. The efficient leaching of REEs and P2O5 was only achieved at ratios above 3.26. This increased acid demand is primarily due to impurities such as calcite, dolomite, and Fe-bearing minerals consuming acid during roasting. Additionally, REE minerals require a high concentration of sulfuric acid for effective decomposition.

Author Contributions

Conceptualization, P.Z. and H.L.; methodology, H.L. and P.Z.; validation, H.L. and A.M.; formal analysis, A.M.; investigation, H.L. and Z.J.; resources, D.D. and P.Z.; data curation, A.M.; writing—original draft preparation, H.L. and P.Z.; writing—review and editing, P.Z.; visualization, H.L.; supervision, P.Z.; project administration, P.Z. and A.M.; funding acquisition, D.D. and P.Z. All authors have read and agreed to the published version of the manuscript.

Funding

This research was funded by the US Department of Energy through CMI Hub, grant number SC-14-392.

Acknowledgments

This research is part of a major project conducted under the Critical Materials Innovation (CMI) Hub (formerly the Critical Materials Institute), funded by the U.S. Department of Energy, Office of Energy Efficiency and Renewable Energy, Advanced Materials & Manufacturing Technologies Office. We gratefully acknowledge the guidance and leadership of Bruce Moyer, CMI focus area lead, and David DePaoli, CMI project lead. Significant matching funds were provided by the Florida Industrial and Phosphate Research Institute, Florida Polytechnic University. Special thanks are extended to The Mosaic Company for their invaluable technical support, substantial in-kind contributions, and assistance with sample collection. We sincerely appreciate the help of Mosaic employees and former employees Nicole Christiansen, Paul Kucera, Marcos Ortiz, Chris Dennis, Glen Oswald, Chaucer Hwang, Cameron Weed, and Gary Whitt.

Conflicts of Interest

The authors declare no conflict of interest.

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Figure 1. Change in floatation results with grinding fineness (pulp pH 9.3, depressor addition 0.5 kg/t tails feed, collector addition 1.2 kg/t tails feed).
Figure 1. Change in floatation results with grinding fineness (pulp pH 9.3, depressor addition 0.5 kg/t tails feed, collector addition 1.2 kg/t tails feed).
Minerals 15 00900 g001
Figure 2. Change in flotation results with pH of pulp (grinding fineness 52.70 wt% of −200 mesh particles, depressor addition 0.5 kg/t tails feed, collector addition 1.2 kg/t tails feed).
Figure 2. Change in flotation results with pH of pulp (grinding fineness 52.70 wt% of −200 mesh particles, depressor addition 0.5 kg/t tails feed, collector addition 1.2 kg/t tails feed).
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Figure 3. Change in flotation results with depressor addition (grinding fineness 52.70 wt% of −200 mesh particles, pulp pH 9.5, collector addition 1.2 kg/t tails feed).
Figure 3. Change in flotation results with depressor addition (grinding fineness 52.70 wt% of −200 mesh particles, pulp pH 9.5, collector addition 1.2 kg/t tails feed).
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Figure 4. Change in flotation results with collector addition (grinding fineness 52.70 wt% of −200 mesh particles, pulp pH 9.5, depressor addition 0.55 kg/t tails feed).
Figure 4. Change in flotation results with collector addition (grinding fineness 52.70 wt% of −200 mesh particles, pulp pH 9.5, depressor addition 0.55 kg/t tails feed).
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Figure 5. Change in leaching efficiency with sulfuric acid addition.
Figure 5. Change in leaching efficiency with sulfuric acid addition.
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Table 1. Main chemical components in amine tails sample.
Table 1. Main chemical components in amine tails sample.
ComponentTotal REEs
/ppm
P2O5
/%
CaO
/%
Fe2O3
/%
MgO
/%
Al2O3
/%
Content201.963.015.620.290.020.28
Table 2. Quantitative mineral analysis of amine tails using MLA [15].
Table 2. Quantitative mineral analysis of amine tails using MLA [15].
Mineralwt%Mineralwt%Mineralwt%
Monazite0.053Tourmaline0.353Pseudo-
rutile
1.125
Xenotime0.003Garnet0.477Rutile0.450
Zircon0.614Epidote0.119Leucoxene0.148
Apatite8.703Kyanite1.174Sphene0.337
Wavellite0.512Staurolite0.329Limonite0.016
Woodhouseite0.009Kaolin0.048Siderite0.143
Quartz80.011Pyrite0.047Zinc spinel0.003
Feldspar5.294Sphalerite0.009Others0.003
Biotite0.004Calcite0.003Total100.000
Table 3. Size distribution analysis for the major valuable minerals in amine tails [15].
Table 3. Size distribution analysis for the major valuable minerals in amine tails [15].
Size Range, mmWeight Distribution, %
ApatiteMonaziteXenotimeZirconRutilePseudo-Rutile Leucoxene
−0.32 + 0.163.32
−0.16 + 0.0844.6811.07 24.9630.1021.5624.46
−0.08 + 0.0434.6078.29 62.8363.2368.2953.88
−0.04 + 0.0212.138.5486.7110.215.959.1315.94
−0.02 + 0.013.971.985.201.470.550.783.60
−0.011.300.128.090.530.170.242.12
Total100.00100.00100.00100.00100.00100.00100.00
Table 4. Results of shaking table separation.
Table 4. Results of shaking table separation.
ProductYield/%REEs Content
/ppm
P2O5 Content/%Recovery/%
REEs P2O5
Concentrate12.51657.398.0040.7333.12
Tailings87.49136.772.3159.2766.88
Calculated feed100.00201.903.02100.00100.00
Table 5. Results of flotation.
Table 5. Results of flotation.
ProductYield/%REEs Content
/ppm
P2O5 Content/%Recovery/%
REEs P2O5
Concentrate49.231105.8214.9085.9189.81
Tailings50.77192.841.6414.0910.19
Calculated feed100.00664.468.17100.00100.00
Table 6. Main chemical components and rare earths in flotation concentrate.
Table 6. Main chemical components and rare earths in flotation concentrate.
ComponentP2O5CaOFe2O3MgOAl2O3
Content/%14.9020.262.250.160.75
ElementScYLaCePrNd
Content/ppm8.82243.06175.48336.8220.83267.05
ElementSmEuGdTbDyHo
Content/ppm0.006.4431.884.2918.164.22
ElementErTmYbLuTotal REEs
Content/ppm16.312.6412.552.291150.82
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Liang, H.; Zhang, P.; Jin, Z.; Medley, A.; DePaoli, D. Processing of a Phosphate Flotation Tails for Recovery of Rare Earths and Phosphate. Minerals 2025, 15, 900. https://doi.org/10.3390/min15090900

AMA Style

Liang H, Zhang P, Jin Z, Medley A, DePaoli D. Processing of a Phosphate Flotation Tails for Recovery of Rare Earths and Phosphate. Minerals. 2025; 15(9):900. https://doi.org/10.3390/min15090900

Chicago/Turabian Style

Liang, Haijun, Patrick Zhang, Zhen Jin, Aaron Medley, and David DePaoli. 2025. "Processing of a Phosphate Flotation Tails for Recovery of Rare Earths and Phosphate" Minerals 15, no. 9: 900. https://doi.org/10.3390/min15090900

APA Style

Liang, H., Zhang, P., Jin, Z., Medley, A., & DePaoli, D. (2025). Processing of a Phosphate Flotation Tails for Recovery of Rare Earths and Phosphate. Minerals, 15(9), 900. https://doi.org/10.3390/min15090900

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