Phosphogypsum as the Secondary Source of Rare Earth Elements
Abstract
1. Introduction
2. Industrial Production of Phosphoric Acid
3. Rare Earth Elements Usage, World Reserves, and Demand
3.1. REEs from Primary Raw Material
3.2. REEs from Secondary Raw Material
3.3. REEs and Trace Metals in Phosphate Rock
3.4. Incorporation and Occurrence of Metal Ions in PG
- ▪
- Interstitial incorporation
- ▪
- Coprecipitation
- ▪
- Isomorphous substitution
- ▪
- Diffusion from the bulk solution towards the crystal surface
- ▪
- Adsorption at the surface
- ▪
- Surface diffusion towards the steps upon the surface
- ▪
- Incorporation in the kink sites
4. Review of Existing Research on REE Recovery from Phosphogypsum
- Hydrometallurgical methods
- Solvo-metallurigical methods
- Supercritical CO2 extraction
- Hydrothermal methods
- Pyrometallurgical methods
4.1. Hydrometallurgical Methods
- ▪
- Leaching (dissolution)
- ▪
- Purification and concentration
- ▪
- Metal recovery (precipitation or electrowinning)
4.1.1. Mineral Acid Leaching
- Effect of Leaching Parameters on REE Leaching Efficiency
- Acid Type and Concentration
- Leaching Temperature
- Particle Size
- Solid-to-Liquid Ratio
- Leaching Kinetics
- Multi-Parameter Effects on REE Recovery from PG
4.1.2. Phosphogypsum Conversion
4.1.3. Bioleaching
4.1.4. Resin in Leach
4.2. Solvo-Metallurgical Leaching
4.3. Super Critical Fluidic Extraction
4.4. Hydrothermal Treatment of Phosphogypsum
4.5. Thermal Decomposition of Phosphogypsum
4.6. Comparative Overview of REE Recovery Methods
5. Summary of Current Challenges
- Origin of phosphate rock: Sedimentary versus igneous rocks have different accessory minerals and trace elements.
- Production process: Dihydrate, hemihydrate, or hemi-dihydrate processes yield PG with different crystal structures and impurity levels.
- Impurities mixed in the crystal lattice or surface: PG often holds silica, fluorides, phosphates, clays, and traces of metals (Fe, Al, Zn, Cd, Cu, etc.) and radionuclides (U, Th, Ra).
- Particle size distribution: Some PG is fine and porous, while others are coarse and crystalline, which may affect leaching and adsorption.
- Industrial bottlenecks: The very low concentration of REEs in PG (generally 0.2–1 wt% of total PG, often < 0.1% recoverable), the need to process large volumes of waste, high chemical consumption, scaling issues when moving from batch leaching to continuous systems, and the lack of standardized pretreatment methods to handle impurities and radionuclides.
- Policy support suggestions: Lack of regulatory frameworks to encourage PG valorization instead of stockpiling, international collaboration to establish safety and reuse standards for decontaminated PG residues, and government incentives (e.g., subsidies or tax credits) for pilot plants that integrate REE recovery with PG decontamination.
Future Directions
- Open-Access Global PG Characterization Database: Collaboration of international institutes and industries is necessary to create a shared database with PG characterization data (REE, impurities, particle size, etc.). A research and industry platform to standard processes and help knowledge sharing for adapting REE recovery to various PG sources.
- Pretreatment Protocol Based on PG Typology: Categorization of PG based on origin (igneous, sedimentary, weathered) and production process (dihydrate, hemihydrate, etc.), then create modular pretreatment protocols (e.g., washing, pH adjustment, thermal treatment, flotation). AI-assisted database matching PG type to develop more specific pretreatment recipes to maximize REE mobilization and impurity removal.
- Adaptive design for Resin-In-Leach Modules: Develop switchable resins/solvents (pH or ion-specific triggers) that adapt their affinity based on the PG composition, increasing selectivity across PG types.
- Co-Product Valorization for Circularity: To minimize treatment cost it will be beneficial to convert decontaminated PG residues into construction materials, agricultural additives, or fillers. Develop standards for safe reuse based on metal leachability and REE residue levels to turn waste into certified products, offsetting treatment costs.
- Utilize Machine Learning for Process Predictability: Apply machine learning models trained on datasets from various PG sources to predict best leaching, sorption, and precipitation conditions. A smart PG profiler tool that shows the ideal treatment path using real-time PG characterization data (XRF, ICP, pH, etc.).
- Modular Pilot Plant Concept for Flexible Scaling: Design a containerized pilot system with interchangeable modules for different steps (pretreatment, leaching, separation). Enables on-site testing and adaptation to local PG composition without building entirely new infrastructure.
- Techno-Economic and LCA: Techno-economic assessment (TEA) and life cycle analysis (LCA) must be included from the first design stage. A dynamic tool can be used to update environmental and cost indicators in real time as the process adopts different PG types or scales.
Author Contributions
Funding
Data Availability Statement
Conflicts of Interest
Abbreviations
| PR | Phosphate rock |
| PA | Phosphoric acid |
| PG | Phosphogypsum |
| WPA | Wet phosphoric acid |
| REE | Rare earth elements |
| S | Sedimentary |
| M | Magmatic |
| NR | Not reported |
Appendix A
| Rock Type | Sedimentary | Magmatic | Mixed | |||
|---|---|---|---|---|---|---|
| Method | HDH Hitashi | HDH Hydro | DH Jacob | DH Prayon (S) | DH Prayon (M) | HRC Kemira |
| CaO% | 30.53% | 33.65% | 33.73% | 35.20% | 35.85% | 36.42% |
| SO3% | 44.83% | 45.82% | 42.50% | 44.47% | 43.49% | NR |
| SiO2% | 4.70% | 6.82% | 4.38% | 5.13% | 2.39% | 3.25% |
| Al2O3% | 0.08% | 0.14% | 0.36% | 0.19% | 0.36% | ND |
| Fe2O3% | 0.18% | 0.06% | 0.12% | 0.18% | 0.25% | 0.03% |
| P2O5% | 1.31% | 0.28% | 0.78% | 0.52% | 0.70% | 0.38% |
| Soluble P2O5 | 0.23% | NR 1 | NR | 0.37% | 0.08% | 0.02% |
| Na2O% | 0.06% | 0.06% | 0.08% | 0.04% | 0.07% | 0.02% |
| K2O% | 0.03% | 0.02% | 0.10% | 0.01% | 0.05% | 0.02% |
| TiO2% | 0.03% | 0.02% | 0.03% | 0.03% | 0.29% | ND |
| MgO% | 0.04% | 0.07% | 0.02% | <0.01% | 0.04% | ND |
| MnO% | 0.01% | 0.01% | NR | <0.01% | 0.01% | ND |
| F% | 0.55% | 0.61% | 0.64% | 0.45% | 0.55% | 0.05% |
| Soluble F | 0.21% | 0.35% | 0.11% | 0.31% | 0.24% | 0.11% |
| CO2% | 0.21% | 0.56% | 0.39% | 0.35% | 0.28% | 0.12% |
| Organic C | 0.01% | 0.11% | 0.05% | 0.04% | 0.02% | 0.01% |
| Chloride as Cl | 1.78 mg/L | 1.86 mg/L | 0.88 mg/L | 2.4 mg/L | NR | NR |
| Nitrate | 8.70 mg/L | 6.83 mg/L | 6.86 mg/L | 10.02 mg/L | NR | NR |
| Purity | 93.27% | 94.87% | 91.83% | NR | 95.56% | NR |
| Rock Type | Sedimentary | Magmatic | Mixed | |||
|---|---|---|---|---|---|---|
| Method | HDH Hitashi | HDH Hydro | DH Jacob | DH Prayon (S) | DH Prayon (M) | HRC Kemira |
| As | NR | NR | NR | 2 ppm | NR | NR |
| Co | 1 ppm | 1 ppm | 1 ppm | <1 ppm | 5 ppm | 1 ppm |
| Zn | 49 ppm | 48 ppm | 34 ppm | 44 ppm | 14 ppm | 8 ppm |
| Pb | 6 ppm | 2 ppm | 2 ppm | 4 ppm | <1 ppm | ND |
| Cd | 1 ppm | 2 ppm | 2 ppm | 3 ppm | <1 ppm | ND |
| Ni | 3 ppm | 2 ppm | 1 ppm | 4 ppm | 1 ppm | ND |
| V | 4 ppm | 2 ppm | 4 ppm | 2 ppm | 7 ppm | 1 ppm |
| Cr | 4 ppm | 5 ppm | 7 ppm | 14 ppm | 4 ppm | 3 ppm |
| Cu | 9 ppm | 60 ppm | 1 ppm | 33 ppm | 5 ppm | 10 ppm |
| Ba | 76 ppm | 109 ppm | 150 ppm | 53 ppm | 207 ppm | 39 ppm |
| Sr | 569 ppm | 801 ppm | 402 ppm | NR | 7692 ppm | 2099 ppm |
| Th | 1 ppm | 1 ppm | 1 ppm | 3 ppm | 42 ppm | 4 ppm |
| U | 2 ppm | 5 ppm | 3 ppm | <1 ppm | 52 ppm | 8 ppm |
| Rock Type | Sedimentary | Magmatic | NR 1 | Sedimentary | Magmatic | NR |
|---|---|---|---|---|---|---|
| Method | HDH Hitashi | HDH Hydro | DH Jacob | DH Prayon | DH Prayon | HRC Kemira |
| Y | 18 ppm | 27 ppm | 72 ppm | 140 ppm | 133 ppm | 27 ppm |
| La | 9 ppm | 14 ppm | 39 ppm | 49 ppm | 1215 ppm | 165 ppm |
| Ce | 10 ppm | 15 ppm | 20 ppm | 31 ppm | 1998 ppm | 246 ppm |
| Nd | 7 ppm | 11 ppm | 25 ppm | 37 ppm | 921 ppm | NR |
| Sm | 1 ppm | 2 ppm | 6 ppm | NR | NR 2 | NR |
| Eu | <1 ppm | 1 ppm | 1 ppm | 2 ppm | NR | NR |
| Gd | 1 ppm | 2 ppm | 7 ppm | 10 ppm | NR | NR |
| Tb | <1 ppm | <1 ppm | 1 ppm | 3 ppm | NR | NR |
| Dy | 1 ppm | 2 ppm | 7 ppm | NR | NR | NR |
| Ho | <1 ppm | <1 ppm | 2 ppm | 2 ppm | NR | NR |
| Er | 1 ppm | 1 ppm | 5 ppm | 6 ppm | NR | NR |
| Tm | <1 ppm | <1 ppm | 1 ppm | NR | NR | NR |
| Yb | 1 ppm | 1 ppm | 4 ppm | 8 ppm | NR | NR |
| Lu | <1 ppm | <1 ppm | 1 ppm | 1 ppm | NR | NR |
| ∑REEs + Y | 49 ppm | 78 ppm | 180 ppm | 289 ppm | 3960 ppm | 438 ppm |
| Rock Type | Sedimentary | Magmatic | NR | Sedimentary | Magmatic | NR |
|---|---|---|---|---|---|---|
| Method | HDH Hitashi | HDH Hydro | DH Jacob | DH Prayon | DH Prayon | HRC Kemira |
| 226Ra | 211 Bq/kg | 596 Bq/kg | 414 Bq/kg | 562 Bq/kg | 267 Bq/kg | NR |
| 232Th | 5 Bq/kg | <7 Bq/kg | 5 Bq/kg | 32 Bq/kg | 185 Bq/kg | NR |
| 235U | <7 Bq/kg | <16 Bq/kg | 73 Bq/kg | <9 Bq/kg | 261 Bq/kg | NR |
| 40K | <10 Bq/kg | <26 Bq/kg | 60 Bq/kg | <12 Bq/kg | NR | NR |
| Radio activity | 0.73 | 1.99 | 1.71 | 2.03 | NR | NR |
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| Process Name | PG Type | P2O5 (%) | Eff. (%) | Key Process Steps/Outcomes |
|---|---|---|---|---|
| Dihydrate process | Dihydrate | 25–30% | 95–96% | In this process, phosphate rock is attacked with sulfuric acid and calcium sulfate as dihydrate crystals are obtained with low concentration of P2O5. |
| Di-hemihydrate | Dihydrate–Hemihydrate | 32–36% | >98% | In this process, sulfuric acid and steam are added to transform the dihydrate solids into hemihydrate. This change in structure breaks open the solid crystals, which helps to release materials that were previously trapped, such as phosphate. |
| Di attack Hemi filtration process | Dihydrate–Hemihydrate | 32–35% | 97–98% | Here, gypsum slurry recrystallizes into hemihydrate calcium sulfate in the conversion section before being filtered to produce the phosphoric acid. |
| Hemi-dihydrate Process | Hemihydrate–Dihydrate | 40–46% | >98.5% | After the first reaction in hemihydrate mode, the product acid is separated as a 46% P2O5. The remaining α-hemihydrate is further processed with sulfuric acid in conditions in which α-hemihydrate is unstable and recrystallizes as gypsum, releasing crystallized and unreacted P2O5. |
| Two stage Hemihydrate process | Hemihydrate | 39–45% | >92–95% | The reaction takes place in two stages. The first stage takes place with low sulfuric acid concentration, while the second stage works at a higher sulfuric concentration. Control of sulfate content and temperature allow the filtration of a slurry holding highly stable hemihydrate crystals that stay in slurry and do not stick easily. |
| Deposit | Mountain Pass | Byan Obo | Mount Weld | Lehat | Longnan | Xunmu | Kola Peninsula |
|---|---|---|---|---|---|---|---|
| Country | USA | China | Australia | Malaysia | China | China | Russia |
| Mineral | Bastnasite | Bastnasite | Monazite | Xenomite | High-Ylateritie | Low-Ylateritie | Loparite |
| La | 33.8 | 23.0 | 25.5 | 1.2 | 1.8 | 43.4 | 25.0 |
| Ce | 49.6 | 50.0 | 46.7 | 3.1 | 0.4 | 2.4 | 50.5 |
| Pr | 4.1 | 6.2 | 5.3 | 0.5 | 0.7 | 9.0 | 5.0 |
| Nd | 11.2 | 18.5 | 18.5 | 1.6 | 3.0 | 31.7 | 15.0 |
| Sm | 0.9 | 0.8 | 2.3 | 1.1 | 2.8 | 3.9 | 0.7 |
| Eu | 0.1 | 0.2 | 0.4 | trace | 0.1 | 0.5 | 0.1 |
| Gd | 0.2 | 0.7 | <0.1 | 3.5 | 6.9 | 3.0 | 0.6 |
| Tb | 0.0 | 0.1 | <0.1 | 0.9 | 1.3 | Trace | trace |
| Dy | 0.0 | 0.1 | 0.1 | 8.3 | 6.7 | Trace | 0.6 |
| Ho | 0.0 | trace | trace | 2.0 | 1.6 | Trace | 0.7 |
| Er | 0.0 | trace | trace | 6.4 | 4.9 | Trace | 0.8 |
| Tm | 0.0 | trace | none | 1.1 | 0.7 | Trace | 0.1 |
| Yb | 0.0 | trace | none | 6.8 | 2.5 | 0.3 | 0.2 |
| Lu | trace | trace | none | 1.0 | 0.4 | 0.1 | 0.2 |
| Y | 0.1 | trace | <0.1 | 61.0 | 65.0 | 8.0 | 1.3 |
| Waste | La | Ce | Pr | Nd | Sm | Eu | Gd | Tb | Dy | Ho | Er | Tm | Yb | Lu | Y | Sc | Ref. |
|---|---|---|---|---|---|---|---|---|---|---|---|---|---|---|---|---|---|
| Red mud | 195 | 312 | 15 | 415 | 82 | 11 | 10 | 8 | 10 | 1 | 15 | 12 | 0 | 0 | 122 | 21 | [19] |
| Fly ash | 85 | 176 | 19 | 77 | 15 | 3 | 13 | 2 | 11 | 2 | 6 | 1 | 6 | 1 | 61 | 0 | [20,21] |
| PG | 1450 | 2310 | 235 | 899 | 163 | 35 | 99 | 7 | 46 | 7 | 16 | 1 | 6 | 1 | 180 | 1 | [22,23,24] |
| E-waste 1 | 59 | 53 | 113 | 883 | 11 | 2 | 11 | 0 | 29 | 5 | 0 | 0 | 1 | 1 | 0 | 9 | [27,28] |
| PG Type | Operating Conditions | REE Content (%) | Eff. (%) | Ref. | |||
|---|---|---|---|---|---|---|---|
| Lixiviant | Temp (°C) | S/L Ratio | Time (h) | ||||
| Hemihydrate | 5–10% H2SO4 | <20 | 0.5 | 6 | NR 1 | 80 | [55] |
| Dihydrate | 10–15% H2SO4 | 40 | 0.5 | 6 | NR | 52 | |
| Dihydrate | 7% H2SO4 | NR | 0.2 | 2 | 0.37 | 59 | [56] |
| Dihydrate | 10% HCl | NR | 0.2 | 2 | NR | 49 | |
| Dihydrate | 7% H2SO4 | NR | 0.2 | 2 | NR | 62 | |
| Dihydrate | HCl | NR | 0.2 | 2 | NR | 51 | |
| Dihydrate | 7% H2SO4 | NR | 0.2 | 2 | NR | 70 | |
| Hemihydrate | 4% H2SO4 | 25 | 0.22 | NR | 0.57–0.51 | NR | [57] |
| Dihydrate | 4% H2SO4 | 25 | 0.22 | NR | 0.41–0.49 | NR | |
| Dihydrate | 96% H2SO4 + 34% H3PO4 | 72 | 0.14 | 1 | 0.0335 | 49 | [58] |
| Dihydrate | 2 M H2SO4 | 60 | 0.125 | 4 | 0.51 | 92 | [62] |
| Dihydrate | 2.5 M HCl | 45 | 0.033 | 0.3 | NR | 98.5% Y, 94.6% Nd 86.1% Dy | [63] |
| Dihydrate | 2.1 M HNO3 | 85 | 0.036 | NR | NR | 7.6% Nd, 83.5% Y, and 77.8% Dy | |
| Dihydrate | 1.3 H2SO4 | 85 | 0.033 | NR | NR | 47.7% Nd, 57.2% Y, 44.9% Dy | |
| Dihydrate | 3.3 M HNO3 | 75 | 0.111 | 0.3 | 0.036 | 87.55 | [65] |
| Dihydrate | 4% H2SO4 | NR | 0.49 | 1248 | NR | 50.5 | [67] |
| Dihydrate | 5% H2SO4 | 50 | 0.25 | 0.5 | 0.0218 | 43 | [68] |
| Dihydrate | 10–15% H2SO4 | 25 | 0.33 | 4 | 0.34–0.63 | 50.9–51.5 | [69] |
| Dihydrate | 15% H2SO4 | 100 | 0.33 | 2 | 0.041 | NR | [70] |
| Dihydrate | 1.65 M/l HNO3 | 80 | 0.1 | 2 | 0.0208 | 83 | [71] |
| Dihydrate | 3 M HNO3 | 80 | 0.1 | NR | NR | >80% Nd | [72] |
| Synthetic | 0.02 M H3PO4 | 25 | 20 | 24 | Synthetic | 85 (for Y) | [73] |
| Synthetic | 0.02 M H2SO4 | 25 | 20 | 24 | Synthetic | 80 (for Y) | [73] |
| PG Type | Operating Conditions | REE Content (%) | Leaching Eff. (%) | Ref. | |||||
|---|---|---|---|---|---|---|---|---|---|
| Lixiviant | Resin Type | Eluent | Temp (°C) | Resin/PG/Acid Ratio (v/m/v) | Time(h) | ||||
| Dihydrate | 25 g/L H2SO4 | SAC | NR | 40 | Resin/PG/Acid = 1:5:40 | 12 | 0.43–0.45 | 65% | [61] |
| Dihydrate | 10–20 g/L H2SO4 | Purolite C160 | NH4NO3 | NR | Resin/PG/Acid = 4:30:4 | 2 | NR | 15–70% | [64] |
| Dihydrate | 10 g/L H2SO4, | Purolite C150TLH, | NaCl | NR 1 | Resin/PG/Acid = 1:5:20 | 24 | 0.3387 | 15–80% | [81] |
| Dihydrate | 0.1, 1, 5 and 10 g/L H2SO4, HCl, and H3PO4 | Purolite S940, Purolite C150, Finex CS16GC | NaCl, HCl, EDTA, Na-citrate | 20 | Resin/PG/Acid = 1:5:40 | 24 | 0.17 | 45–75% from CHEL resin | [82] |
| Dihydrate | 10 g/L H2SO4 | Purolite S940, Na+ | HCl, GLDA and MGDA | 25 | Resin/PG/Acid = 1:5:40 | 20 | NR | yield 98%, purity 94% | [84] |
| Resin | Type | Key Findings | Advantages | Limitations |
|---|---|---|---|---|
| Chelating resin | Chelating (e.g., iminodiacetic acid or phosphonic acid types) | Achieved 19.2 g REE/kg resin loading and 20% purity with 1 g/L H2SO4 | High selectivity, good performance at low acid concentrations | Requires complex eluents (EDTA or conc. HCl), multi-stage processing |
| Purolite S940 | Chelating resin | Maximum loading: 0.92 eq/kg; 70% Ca-REE purity after 7 RIL cycles; REE elution with biodegradable agents | High selectivity; good for batch RIL with eco-friendly elution | Multiple cycles needed; moderate overall loading capacity |
| Macroporous sulfonated resin (DVB > 12%) | Strong acid cation (SAC) | 32.8% REE extraction; 30–40% scandium recovery | Industrial-scale tested, good for high-throughput | Low selectivity for REEs, moderate efficiency |
| Purolite C150TLH | Strong acid cation exchange | 15–80% efficiency depending on PG source; acid conc. = 10 g/L | Commercially available; effective in RIL at ambient temp | Highly variable efficiency depending on PG origin |
| AN-31 | Anion exchange | 59.7% Pr, 52.8% Sm extraction from sulfate solutions | Good for light REEs; effective at pH 2–4 | Not tested directly on solid PG; more applicable to solution-based systems |
| PG Type | Operating Conditions | REE Content (%) | Leaching Eff. (%) | Ref. | ||||
|---|---|---|---|---|---|---|---|---|
| Lixiviant | Temp (°C) | Pressure (bar) | S/L Ratio | Time (h) | ||||
| Dihydrate | (CO2) = 700–1100 g/min. | 33 | 50.6 | 0.33 | 0.16 | 0.08 | 87% (PG conversion) | [86] |
| Dihydrate | 2 M HNO3, CO2 with TBP, DEHPA | 45 | 200 | 0.7–1 | 1 | 0.42 | 93.50% | [87] |
| PG Type | Operating Conditions | REE Content (%) | Leaching Eff. (%) | Ref. | ||||
|---|---|---|---|---|---|---|---|---|
| Lixiviant | Temp (°C) | Pressure (bar) | S/L Ratio | Time (h) | ||||
| Dihydrate | 100 g/L H2SO4 | 80–200 | 0.78–14.7 | NR 1 | 6 | NR | >80 | [88] |
| Dihydrate | 0.1 mol/L HCl, 0.1 mol/L HNO3, and 0.05 mol/L H2SO4 | 100 | 9.8 | 10 | 0.08 | 0.02 | Max 98.63 (while using HCl) | [89] |
| PG Type | Operating Conditions | REE Content (%) | Leaching Eff. (%) | Ref. | ||||
|---|---|---|---|---|---|---|---|---|
| Lixiviant | Temp (°C) | Pressure (bar) | S/L Ratio | Time (h) | ||||
| Dihydrate | H2S flow 450–1350 mL/min, CO2 450–1340 mL/min | 1100 | Atm | NA | 1 | 0.19 | Enriched REE residue | [90] |
| Extraction Technique | Extraction Eff. | Environmental Impact | Technical Challenges for Scale-Up | Key Economic Cost Drivers | Profitability Outlook and Economic Feasibility |
|---|---|---|---|---|---|
| Acid Leaching | 50–85% | High (Acid waste, radionuclides) | Managing large volumes of corrosive, radioactive waste streams; low selectivity requires extensive downstream purification. | Reagent consumption (H2SO4 or HCl), neutralization costs, waste disposal. OPEX is high. | Uneconomical as a standalone process. High operational and waste management costs outweigh the value of recovered REEs unless they are integrated with PG processing for other purposes (e.g., sulfuric acid production). |
| Resin-in-Leach | 97–99% | Low-Moderate (Exhausted resin is a new waste stream) | Resin degradation/fouling over time; requires efficient solid–liquid separation; elution and regeneration cycles. | Capital cost of columns/reactors, resin cost but possible to reuse, eluent reagents. | Promising and likely feasible. High selectivity reduces downstream costs. Profitability is highly dependent on resin longevity and elution efficiency. OPEX is lower than solvent extraction for certain applications. A key candidate for commercial scale-up. |
| Carbonation | 80–86% | Moderate (Multiple waste streams) | Requires a pure, concentrated CO2 and NH3 source for in situ ammonium carbonate production, production of. Effective leaching of REEs from co-products is challenging. | Reagent cost (ammonium salts or NH3/CO2), energy cost for CO2 compression | Low feasibility. Economics are tied to the sale of main product, ammonium sulfate, and the recovery of REEs from byproducts calcium carbonate. The difficulty to penetrate in the ammonium sulphate market with existing suppliers does not offset the high process costs for REE recovery alone. |
| Bioleaching | 40–60% | Very Low (Environmentally friendly) | Extremely slow leaching kinetics (weeks/months); keeping microbial health at scale; large equipment required to meet the required residence time for scaled plants | Low reagent cost but very high operational cost due to reactor holding time (CAPEX for large tank farms), nutrient costs, aeration/agitation energy. | Uneconomical for large scale. The immense reactor volume and time required to process industrial PG volumes make the capital and operational costs prohibitive. Potential for niche, small-scale applications. |
| Solvent Extraction | 56–69.8% | High (Organic solvent loss, volatile emissions) | Requires very pure leachate (PG liquors are complex and cause crud formation); multi-stage setup is complex; solvent loss is a major operational issue. | High CAPEX for mixer-settlers, reagent cost (extractants like D2EHPA, TBP), solvent makeup cost, scrubbing/stripping reagents. | Economical only after purification. Not feasible for direct leachate. It is the necessary, established downstream step for purifying REE concentrates produced by a more selective primary extraction method (like RIL). Its economics are acceptable only after the bulk of impurities have been removed. |
| Supercritical Fluid Extraction | 93% using TBP, 88% using DHEPA | Low-Moderate (But high energy demand) | Engineering equipment for high-pressure (70–300 bar) continuous operation; material compatibility; efficient recycling of CO2 and modifiers. | Extremely high CAPEX for pressure-rated equipment, high energy cost for compression and heating, cost of CO2 and modifiers (e.g., TBP). | Uneconomical with current technology. The capital intensity and high operating energy costs are far greater than the value of the recovered REEs from PG. Significant technological breakthroughs in pressure vessel design and energy recovery are needed. |
| Hydrothermal Extraction | Almost 100% | Low-Moderate (Uses water, but high energy) | Similarly to SCFE: high-pressure/temperature reactor design; corrosion control; handling of solids in a pressurized system. | High CAPEX for autoclaves, high energy cost for maintaining T and P, reagent cost. | Uneconomical for bulk processing. The energy input required to maintain subcritical conditions for large PG tonnages is cost prohibitive. May have potential for processing specific, high-value waste streams but not for bulk PG. |
| REE Content | Plant Scale | Method | Eff. | Key Economic Metric | Ref. |
|---|---|---|---|---|---|
| REE content of 0.27–0.33 wt.% | 3.3 million tons per annum of phosphogypsum on a dry basis | Resin in leach using sulfuric acid and SAC resin Purolite C150TLH | Tested recovery from PG ranging from 15% to 80% | Economic study suggested that with mixed REE oxide product price > USD 21/kg, even low overall recovery (15%) could be economically favorable. | [81] |
| REE content > 0.5 wt.% needed for profitability | 100 tons of PG per hour | Acid leaching followed by bio-inspired adsorptive separation | 43% REE leaching efficiency | In 87% of baseline simulations and internal rate of return (IRR) > 15% was observed | [91] |
| REE content of 0.44 wt.%, focus on the recovery of Nd, Pr, Dy and Tb | 2.2 Mt/year PG, the plant is expected to produce ca. 1900 t magnet REO annually | Sulfuric acid leaching followed by continuous ion exchange | 65% recovery (magnet REE) | Operating cost USD 40.83/kg magnet REO and USD 12.91/kg total REO with post-tax IRR of 38% and a PBT of 2 years | [92] |
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Khalil, F.; Pagnanelli, F.; Moscardini, E. Phosphogypsum as the Secondary Source of Rare Earth Elements. Sustainability 2025, 17, 8828. https://doi.org/10.3390/su17198828
Khalil F, Pagnanelli F, Moscardini E. Phosphogypsum as the Secondary Source of Rare Earth Elements. Sustainability. 2025; 17(19):8828. https://doi.org/10.3390/su17198828
Chicago/Turabian StyleKhalil, Faizan, Francesca Pagnanelli, and Emanuela Moscardini. 2025. "Phosphogypsum as the Secondary Source of Rare Earth Elements" Sustainability 17, no. 19: 8828. https://doi.org/10.3390/su17198828
APA StyleKhalil, F., Pagnanelli, F., & Moscardini, E. (2025). Phosphogypsum as the Secondary Source of Rare Earth Elements. Sustainability, 17(19), 8828. https://doi.org/10.3390/su17198828

