A Targeted Approach to Critical Mineral Recovery from Low-Grade Magnesite Ore Using Magnetic and Flotation Techniques

Abstract

As a critical mineral, magnesite plays a vital role in industries such as steelmaking, construction, and advanced technologies due to its high thermal stability and chemical resistance. However, the beneficiation of low-grade magnesite ores (~38.36% MgO) remains challenging due to the presence of iron, silica, and calcium-bearing impurities. This study proposes an integrated beneficiation strategy combining medium-intensity magnetic separation and flotation techniques to upgrade a low-grade magnesite ore. After grinding to 75 µm (d₈₀), the sample was subjected to two-stage magnetic separation at 5000 Gauss to remove Fe-bearing minerals, reducing Fe₂O₃ below 0.5%. The non-magnetic fraction was then treated through a two-stage reverse flotation process using dodecylamine (350 g/t) and diesel oil (60 g/t) at pH 7 to reject silicate gangue. This was followed by a four-stage direct flotation using sodium oleate (250 g/t), sodium silicate (350 g/t), and SHMP (100 g/t) at pH 10 to selectively recover magnesite while suppressing Ca-bearing minerals. The optimized flowsheet achieved a final concentrate with MgO > 46.5%, SiO₂ ≈ 1.05%, Fe₂O₃ ≈ 0.44%, and CaO ≈ 0.73%, meeting the specifications for refractory-grade magnesite. These results highlight the effectiveness of a combined magnetic–flotation route in upgrading complex, low-grade magnesite deposits for commercial use.

1. Introduction

Magnesite (MgCO₃) is a carbonate mineral, also referred to as a salt-type mineral, with substantial commercial importance attributed to its unique refractory properties [1,2]. Recognized for its critical role in various industrial applications, magnesium, derived primarily from magnesite, is classified as a critical mineral by the U.S. Geological Survey due to concerns over supply concentration and import reliance. Globally, magnesium production is heavily concentrated in a few countries, particularly China, which dominates the market, followed by Russia, Kazakhstan, and Turkey [3]. This geographic concentration raises concerns about supply chain stability and strategic resource security. Magnesium’s applications span a wide range of industries: it plays a vital role in the production of refractory materials due to its high melting point and thermal stability, serves as a crucial component in steelmaking and construction (e.g., cement and fireproof materials), and is used in the chemical industry as a catalyst and pH regulator. In addition, magnesium is employed in the manufacturing of glass, ceramics, and heat-resistant consumer goods. These widespread and critical uses underscore the strategic importance of developing efficient beneficiation methods to upgrade low-grade magnesite ores for commercial use [4,5]. Magnesite ores typically contain key gangue minerals, including iron oxides, silicates, and carbonates, with common examples being limonite, talc, quartz, and dolomite [6]. Iron impurities can react with carbonates to form low-melting-point materials, which reduce the thermal stability and fire resistance of the final products [7].

The chemical composition of magnesite, particularly the MgO content and impurity levels (CaO, SiO₂, Fe₂O₃, Al₂O₃), critically influences its suitability for industrial applications and the quality after calcination. Commercial-grade magnesite for refractory use typically requires a MgO content between 40% and 47% (equivalent to 84.0%–99.0% MgCO₃), CaO between 0% and 3.5%, SiO₂ from 0% to 7%, and combined Fe₂O₃ and Al₂O₃ between 0% and 3% [8]. In the beneficiation process, the presence of magnesium-associated minerals with similar physical and chemical characteristics poses significant challenges [9]. Therefore, comprehensive research is essential to gain deeper insights into the physicochemical properties of the associated minerals, optimize processing conditions and technologies, and provide the necessary technical support to ensure the effective development and utilization of magnesite resources [10]. To enhance flotation efficiency and reduce costs, Zhang et al. [11] mixed collectors of sodium oleate (NaOL) and monohydric alcohols were used for magnesite and dolomite flotation. Compared to NaOL alone, mixtures with alcohols (butanol, isobutanol, octanol, isooctanol) improved mineral recovery, especially with longer carbon chains and isomeric alcohols. However, magnetic separation processes can be effectively applied in the iron removal from magnesite but processing non-magnetic minerals is difficult and in some cases is more complicated [11]. Wasmuth applied a high-intensity dry magnetic field (3.2 T) for magnesite ore (4–100 mm particle size, containing 20% SiO₂ and 4% Fe₂O₃) to achieve a substantial reduction in impurities, lowering SiO₂ to below 1.5% and Fe₂O₃ to under 0.3%. This demonstrates the effectiveness of dry magnetic separation in significantly upgrading magnesite quality for industrial applications [12]. Atasoy [13] recently investigated the application of wet high-intensity magnetic separation (WHIMS) to magnesite ore, which initially contained 77.69% MgCO₃ and 3.14% Fe₂O₃. The process increased the MgCO₃ content to 91.03% and reduced Fe₂O₃ to 0.32%, illustrating the potential of WHIMS in magnesite upgrading [13]. It is important to note that magnetization roasting is a technique commonly employed to enhance the magnetic susceptibilities of materials [13,14]. Magnetization roasting is a widely utilized technique to improve the magnetic susceptibility of materials. Magnesite middlings, initially comprising 52.70% MgO, 2.06% SiO₂, 0.61% Fe₂O₃, 1.36% CaO, 0.07% Al₂O₃, and 43.20% LOI, underwent a series of beneficiation processes including calcination, classification, and magnetic separation. These processes resulted in a weight recovery of 47%. The final caustic-calcined MgCO₃ product achieved an impressive MgO content of 89.57%, with reduced impurities: SiO₂ at 0.96%, Fe₂O₃ at 0.43%, CaO at 2.35%, Al₂O₃ at 0.05%, and LOI at 6.64% [14]. During magnetization roasting, hematite and limonite are typically reduced to magnetite, enhancing their magnetic properties. For example, limonite exhibits a significant increase in magnetism with temperature, achieving maximum saturation magnetization at 700 °C [15]. Similarly, hematite undergoes partial reduction to magnetite, further improving its magnetic susceptibility [16,17].

Flotation is considered the most cost-effective and efficient method for magnesite beneficiation, allowing for the precise separation of valuable minerals from impurities [18]. Given magnesite’s semi-soluble characteristics, the interaction between collectors and minerals during flotation is influenced by the dissolved species present in the system. Processes such as precipitation, adsorption, and hydrolysis can occur, leading to the formation of new complexes that interfere with the selective interaction between mineral surfaces and reagents [19,20]. Furthermore, the dissolution kinetics of minerals significantly influence the extent of precipitation, with faster dissolution rates promoting quicker precipitation. Bulk precipitation, in particular, can adversely affect the flotation of Mg carbonates, as the formation of bulk species reduces the availability of collectors, thereby hindering effective surface interactions [21].

Direct and reverse flotation are well-established methods for effectively separating impurities from carbonate minerals [22]. In a reverse flotation approach, separation is achieved by selectively depressing the valuable minerals. Depressants commonly utilized in this process consist of both organic and inorganic compounds, including sodium carbonate, sodium silicate (water glass), pyrophosphate, sodium hexametaphosphate (SHMP), starch, carboxymethyl cellulose (CMC), quebracho, and tannin derivatives [23]. A thorough assessment of various depressants revealed that EDTA, citric acid, and tannic acid exhibited the most significant depressant efficacy in magnesite flotation. In contrast, sodium silicate and tartaric acid showed minimal inhibitory effects, proving less effective in the flotation process of magnesite [6]. In the direct flotation of magnesite and calcium-bearing carbonates, challenges arise from the similarities in surface characteristics between these minerals. This overlap can cause depressants to target both valuable and gangue minerals, hindering effective separation. Therefore, selecting an optimal depressant is crucial for achieving efficient separation in Mg-carbonate flotation, particularly when a collector is used [24]. Effective collectors, such as oleates, hydroxamates, sulfosuccinates, and sarcosinates, are commonly employed in the flotation of calcium-bearing minerals. Additionally, various studies have shown that the combination of different collectors can exhibit a synergistic effect, enhancing the overall process efficiency [25,26]. In reverse flotation for impurity removal, cationic collectors, particularly amine-based types, are widely used to target silicon impurities like talc and quartz. The novel cationic surfactant KD-1 (Hunan Mingzhu Flotation Reagents Co.) has been shown to significantly improve reverse flotation efficiency for magnesite and quartz [19,27]. pH plays a crucial role in the beneficiation of Mg-carbonate ores. While both acidic and alkaline conditions can be employed, acidic environments tend to cause the dissolution of carbonate lattices and promote excessive collector adsorption. Due to the challenges associated with this dissolution, an alkaline pH range (7–10) is generally favored to optimize flotation performance [28].

Carbonate minerals, such as calcite and dolomite, are regarded as significant impurities in magnesite beneficiation processes. In flotation, magnesite, dolomite, and calcite exhibit similar responses to reagents, as they are classified as semi-soluble salt-type minerals containing (Ca²⁺/Mg²⁺) cations, which results in their analogous chemical compositions and behavior during the flotation process. In the absence of depressants, the collector demonstrates comparable adsorption behavior toward both magnesite and calcium carbonate gangue minerals. This leads to nearly identical recovery rates, thereby reducing the selectivity and efficiency of the flotation process [23]. Sodium hexametaphosphate is a commonly used depressant in flotation processes, particularly for suppressing calcium-containing minerals. With the chemical formula (NaPO₃)₆, SHMP is a cyclic metaphosphate featuring a P-O-P backbone structure that undergoes hydrolysis in aqueous solutions, decomposing into lower phosphate species like Na₃PO₄, Na₂HPO₄, and NaH₂PO₄ [29]. In flotation systems, the hydrolytic products, H₂PO₄⁻ and HPO₄²⁻, act as active species. These species donate electrons to metal ions such as Ca²⁺, Mg²⁺, Al³⁺, and Fe³⁺, facilitating adsorption onto mineral surfaces through electrostatic attraction or chelation. Another potential adsorption mechanism involves the formation of a strong hydrogen bonding network between the hydrolytic species of hexametaphosphate and water molecules surrounding calcium-containing mineral particles, resulting in the generation of insoluble calcium polyphosphate and calcium phosphate. The optimal pH range for maximizing the depressant function of SHMP appears to be between 2 and 12, as the active species H₂PO₄⁻ and HPO₄²⁻ dominate within this range. Consequently, flotation tests were conducted at pH 10 [30,31].

This study advances the beneficiation of low-grade magnesite ore by introducing a fully integrated and experimentally optimized processing strategy that aligns with current industrial specifications for refractory applications. Unlike prior research, which often focuses on isolated separation methods or limited reagent systems, our work proposes a multi-stage flowsheet combining medium-intensity magnetic separation, two-stage reverse flotation, and four-stage direct flotation. The proposed method achieves a MgO grade exceeding 46% while reducing SiO₂, CaO, and Fe₂O₃ levels to below 2.5%, meeting international standards for caustic-calcined magnesite. This contribution is particularly relevant for the economic exploitation of low-grade magnesite deposits, offering a scalable solution for industries facing declining availability of high-purity resources.

2. Materials and Methods

2.1. Experimental Samples

The low-grade magnesite ore was first crushed to a particle size of −2 mm using jaw and roller crushers. Subsequently, the material was ground in a ball mill to achieve the necessary degree of liberation for valuable minerals such as magnesite and gangue impurities like silica (SiO₂). This process ensured the optimization of particle size for downstream magnetic separation and flotation processes. Through comminution, the d₈₀ of the feed material was reduced to 75 μm. A representative sample was then prepared by thorough mixing and homogenization and submitted for detailed mineralogical analysis. The analysis included X-ray diffraction (Philips-Xpert Pro, Almelo, The Netherlands) and X-ray fluorescence techniques.

2.2. Experimental Procedure

2.2.1. Materials

The flotation experiments were conducted using a range of chemical reagents, each selected for its specific function within the process. Dodecylamine (Merck, Darmstadt, Germany) was employed as a cationic collector, primarily targeting silicate minerals by enhancing their hydrophobicity and promoting bubble–particle attachment. Sodium oleate (Kemcore), an anionic collector, was used to facilitate the flotation of oxide and carbonate minerals by interacting with positively charged mineral surfaces. Pine oil (Kemcore) served as a frothing agent, contributing to bubble formation and stability, which is essential for effective particle recovery in the froth phase. Gasoline (BP Australia Pty Ltd., Melbourne, Australia) was applied as a non-polar auxiliary collector or frother, enhancing the hydrophobic character of certain minerals and modifying bubble dynamics. Sodium hexametaphosphate (Merck) acted as a dispersant and depressant, preventing the aggregation of fine particles and selectively suppressing the flotation of undesired gangue minerals such as iron or calcium-bearing phases. Sodium silicate (Sigma-Aldrich, St. Louis, MO, USA) was also used as a depressant and dispersing agent, particularly effective in inhibiting the flotation of silicate gangue minerals and maintaining slurry dispersion. Tap water served as the medium for all tests conducted in the separation process. To adjust the pH during flotation, analytical-grade sulfuric acid and sodium hydroxide (Merck) were employed.

2.2.2. Methods

In this research, to eliminate iron impurities from the final concentrate and minimize disruptions in subsequent flotation stages, a medium-intensity magnetic separator with a field strength ranging from 4000 to 6000 Gauss was employed following the comminution process. Magnetic separation tests were conducted using a medium-intensity magnetic separator, model XCSQ-50*70. All flotation tests were carried out in a 3 L Denver flotation cell operating at an agitation speed of 1200 rpm, with a predetermined solid percentage and duration. The predetermined solids percentage in flotation or other mineral processing experiments is typically deduced based on a combination of the literature review, preliminary testing, and practical considerations. Sulfuric acid and NaOH were used to regulate the solution’s acidity, while pine oil was applied as a frother reagent.

The recoveries of the main components (MgO, CaO, Fe₂O₃, and SiO₂) were calculated using the following equation:

$$R(\%)={\displaystyle \frac{C\times c}{(C+T)\times f}}\times 100={\displaystyle \frac{(f-t)\times c}{(c-t)\times f}}\times 100$$

(1)

where C and T are the weights of the concentrate and tailing, respectively, while f, c, and t represent the grades of a specific element in the feed, concentrate, and tailing, respectively. The flotation process efficiency was evaluated, and optimal conditions for maximum recovery were determined. These findings were utilized to develop a beneficiation strategy for the efficient flotation of magnesite from the ore sample.

The reverse flotation process was investigated to assess the effects of pH, flotation time, dodecylamine (DDA) dosage, and gasoline dosage on the results. An L9 (3 × 4) experimental design was applied, with the parameters and levels detailed in

Table 1. Factors and levels of parameters studied in reverse flotation.

Level	Factors
	A (1)	B (2)	C (3)	D (4)
	pH	Time (minute)	Dodecylamine (DDA)	Gasoline (g/t)
			(g/t)
1	6	4	1	40
2	7	5	2	60
3	8	6	3	70

. The L9 (3 × 4) experimental design is a type of Taguchi orthogonal array used in experimental optimization, particularly when dealing with multiple factors and levels. In this study, all four operating conditions were varied simultaneously in each trial to allow for a comprehensive evaluation of their individual and interactive effects on the response. This approach, commonly used in factorial experimental designs, provides a more efficient and informative analysis compared to changing one factor at a time. By exploring the full design space, it becomes possible to identify not only the main effects of each variable but also any significant interactions between them, which are critical for process optimization and accurate modeling. Each flotation test was performed in duplicate under identical conditions, and the average values were reported. The variations between repetitions were within ±0.3% for grade and ±2% for recovery, indicating a high degree of reproducibility.

3. Results and Discussion

3.1. Characterization

The compositional analysis of the representative sample is provided in

Table 2. Chemical composition analysis of the sample (%).

Fe2O3	Al2O3	MgO	SiO2	CaO	IL *
1.05	0.26	38.36	16.53	2.44	41.15

* Ignition Loss.

. The ignition loss (IL) of 41.15% comprises volatiles such as water (H₂O), carbonates (CO₂), organic matter, sulfur, and other thermally decomposable compounds.

Table 3. Particle size distribution, chemical composition, and distribution of CaO, SiO₂, and MgO across different particle size fractions.

Size (Micron)	Wt.%	Concentration (%)			Distribution (Wt.%)
		MgO	SiO2	CaO	MgO	SiO2	CaO
+75	19.85	37.02	17.65	2.62	19.15	39.52	18.95
63	14.5	36.57	18.65	2.88	13.82	41.26	22.01
45	18.75	36.12	18.35	3.38	17.65	40.09	25.41
38	16.75	35.76	18.35	3.25	15.61	39.69	24.44
−38	30.15	42.62	8.65	1.95	33.5	22.3	6.91
	100	38.36	16.53	2.44	100	100	100

presents the distribution of the key compounds within the sample across various particle size fractions.

X-ray diffraction (XRD) analyses (Figure 1) revealed that the sample primarily consists of magnesite, which accounts for approximately 70% of its composition. Secondary minerals, including quartz (with chalcedony) and dolomite, together make up about 25%, while minor amounts of calcite (CaCO₃), lepidocrocite, and serpentine collectively constitute around 5%. In terms of physical and structural properties, the sample exhibited a predominantly white appearance, with localized areas displaying oxidation, resulting in yellow, brown, and red coloration. It featured a massive, non-crystalline structure and demonstrated a hardness of 4.1 on the Mohs scale (a medium-hard mineral).

Sample characterization also indicated that, in addition to silica, the mineral contains a relatively high concentration of the undesirable element iron (Fe), approximately 0.73%. To ensure product quality, a processing strategy is required to remove or reduce its concentration. In light of these considerations, the research incorporates fine crushing, magnetic separation, desilication through reverse flotation, and magnesite flotation as part of the overall treatment process.

3.2. Magnetic Separation

A representative sample was subjected to different medium-intensity magnetic fields. Although higher recoveries were observed at 4000 Gauss, this intensity was not selected as the optimum because it resulted in an increased level of impurities in the concentrate. As the magnetic intensity increases, more non-magnetic or weakly magnetic gangue minerals may also be captured along with the desired magnetic minerals, reducing the overall grade of the product. Therefore, a balance between recovery and purity must be considered. Lowering the intensity helps in reducing the entrainment of impurities, leading to a higher-quality concentrate, even if the total recovery is slightly reduced. This trade-off is essential in determining the most effective and economically viable operating conditions. The results indicated that a 5000 Gauss magnetic field produced a concentrate (non-magnetic product) with a Fe₂O₃ concentration below 0.8 and a 96% weight recovery of non-magnetic material (

Table 4. Test conditions and results obtained (Fe₂O₃ (%)) from magnetic separators (different fields).

Product	Weight Percentage (%)	Fe2O3 (%)	Magnetic Field (Gauss)	Recovery (%)
Magnetic	2.44	5.17	4000	11.99
Non-Magnetic	97.56	0.93		88.01
Magnetic	3.89	6.86	5000	25.37
Non-Magnetic	96.11	0.8		74.63
Magnetic	6.01	6.88	6000	28.16
Non-Magnetic	93.99	0.8		71.84

). Consequently, this intensity was chosen as the optimal magnetic field intensity.

Table 5. Results from two stages of magnetic separation (field intensity = 5000 Gauss).

Stages	Products	Weight (%)	Fe2O3 (%)	Recovery (%)
Stage 1	Conc.	3.89	6.86	25.37
	Tailing	96.11	0.8	74.63
Stage 2	Conc.	2.53	11.56	27.80
	Tailing	94.58	0.5	72.20

presents the results of two stages of magnetic separation using the optimal field intensity of 5000 Gauss. The results demonstrated that both stages are essential to reduce the iron content to below 0.5%. The magnetic concentrate contains 34% MgO (recovery = 5.72%) and 16.11% SiO₂ (recovery = 6.53%). The magnetic separation stage effectively reduced Fe₂O₃ to below 0.5%, providing a cleaner feed for flotation and increasing downstream flotation selectivity.

3.3. Flotation Tests

The non-magnetic products were subjected to sequential reverse and direct flotation trials to separately remove impurities, such as silicates and CaO, from magnesium minerals.

3.3.1. Reverse Flotation

Following the removal of iron-bearing impurities, the non-magnetic fraction was subjected to reverse flotation to target the separation of silicate gangue minerals. The results and optimal levels are presented in

Table 6. Experimental conditions, factor levels, and results.

Tests No.	Factors				Results
	A	B	C	D	Yield (Wt.%)	Grade (%) (MgO)	Recovery (%) (MgO)
1	1	1	1	1	68.59	37.84	68.3
2	1	2	2	2	69.00	37.72	68.49
3	1	3	3	3	69.92	38.14	70.10
4	2	1	2	3	71.45	38.56	72.50
5	2	2	3	1	72.60	39.00	74.51
6	2	3	1	2	74.64	39.58	77.70
7	3	1	3	2	71.09	38.93	72.83
8	3	2	1	3	70.98	38.05	71.07
9	3	3	2	1	72.00	38.99	73.88

and

Table 7. Selection of optimal factors and levels.

Factor		Optimum Level	Optimum Amount
A	pH	2	7
B	Time (minute)	3	6
C	Dodecylamine (DDA) (g/t)	2	350
D	Gasoline (g/t)	2	60

. The test results with optimal values and levels are given in Figure 2.

In the flotation of silicate minerals, particularly quartz, amine collectors like dodecylamine (DDA) play a crucial role in enhancing hydrophobicity, thereby facilitating selective separation from gangue minerals. At lower pH values, amines, including DDA, are protonated to form cationic species (RNH₃⁺), which are highly reactive and readily interact with the negatively charged quartz surface. This interaction occurs primarily through electrostatic attraction between the positively charged amine molecules and the negatively charged silanol groups (Si-OH) on the quartz surface [32,33]. Furthermore, the adsorption of amine collectors on quartz is facilitated by hydrogen bonding, particularly between the amino group (-NH₂) of the amine and the hydroxyl groups (-OH) on the mineral surface. This hydrogen bonding further strengthens the adsorption of the collector, enhancing the hydrophobicity of the quartz particles. The combination of DDA with surfactants like gasoline, on the other hand, generates synergistic effects, leading to co-adsorption on the mineral surface. This interaction reduces surface tension, promoting better bubble formation and enhancing particle–bubble attachment, ultimately improving flotation efficiency [32].

Figure 2 illustrates the results of a separation process with optimized factor levels designed to recover MgO and SiO₂ from raw material. The bar chart displays the results in three categories: weight percentage, grade, and recovery of MgO and SiO₂. The concentrate (tailings under reverse conditions) demonstrates a considerably higher weight percentage (exceeding 80%) compared to both the tailings and magnetic concentrate, indicating effective separation of the desired components. The MgO grade and recovery in the concentrate are 39.37% and 82.67%, respectively, confirming its enrichment, while the tailings exhibit lower values, indicating the presence of impurities or less-targeted material. The SiO₂ concentration (13.92%) in the final concentrate is lower than that of MgO (39.37%), suggesting partial rejection or selective separation of SiO₂ from the desired components.

To optimize the number of reverse flotation stages and ensure the quality of the final concentrate, half of the reagents used in the initial stage—DDA (175 g/t) and gasoline (30 g/t)—were also applied in the second stage. The results of both flotation stages are presented in Figure 3.

The flotation concentrate (from two stages) exhibits a high weight percentage (exceeding 73%), indicating effective separation. The MgO grade in the flotation concentrate is significantly enriched, reaching approximately 40%, which highlights the success of reverse flotation in selectively isolating MgO. The SiO₂ grade in the final concentrate is reduced by more than 1% compared to the first concentrate, with its presence more prominent in the tails, suggesting effective rejection of SiO₂ during the process. MgO recovery is highest in the flotation concentrate, reflecting the success of the two-stage reverse flotation process. In contrast, SiO₂ recovery in the concentrate is lower, at less than 60%, which is a 10% improvement over stage one, consistent with its rejection of around 12% in the final tailings.

3.3.2. Direct Flotation

To enhance the MgO grade and minimize residual impurities like other Ca-bearing minerals and silicates, the concentrate resulting from a two-stage reverse flotation experiment was further treated through the direct flotation method. In the direct flotation experiments, sodium oleate was used as a collector, and pine oil was employed as a frother. Sodium silicate and SHMP were utilized as dispersants and depressants. Sodium silicate functioned as a dispersant by preventing the re-agglomeration of fine particles and reducing the tendency of silicates to form clusters. This helped ensure that unwanted silicate minerals remained well-dispersed in the slurry, preventing their flotation and contributing to a higher MgO grade in the concentrate. Additionally, sodium silicate increased the hydrophilicity of silicate minerals, making them more resistant to attachment to air bubbles, thus reducing their flotation recovery. On the other hand, SHMP acted as a selective depressant, specifically targeting Ca-bearing minerals and other silicates. SHMP adsorbed onto the surface of these minerals through mainly electrostatic or chelation interactions, decreasing their hydrophobicity and making it more difficult for them to adhere to the bubbles. The selective depression effect of SHMP on Ca-bearing minerals and silicates was determined through a series of flotation experiments using varying SHMP dosages (100, 150, and 200 g/t). The results showed that increasing the SHMP dosage led to a significant reduction in CaO content in the concentrate, confirming its selective action. These findings, in combination with well-established mechanisms in the literature—such as SHMP’s ability to form complexes with calcium ions through electrostatic and chelation interactions—support the conclusion that SHMP selectively adsorbs onto Ca-bearing minerals, reducing their hydrophobicity and flotation tendency. This selective depression of impurities ensured that the final concentrate contained fewer residual Ca-bearing minerals and silicates, thereby improving the overall quality and purity of the MgO concentrate.

-

Sodium silicate tests

Sodium silicate is a widely used modifier reagent in flotation, primarily acting as a depressant for silicate gangue minerals such as quartz, as well as carbonate minerals like dolomite and calcite, especially in systems with carboxylic collectors. Upon dissolution in water, sodium silicate dissociates to form a range of species, with their distribution being strongly influenced by the pH of the solution. At pH values below 9.5, tetrahydroxysilane (Si(OH)₄) predominates, while within the pH range of 9.5 to 12.4, the monosilicic acid anion (SiO(OH)₃⁻) becomes the dominant species. At pH levels exceeding 12.5, the polysilicic acid anion (SiO₂(OH)₂²⁻) prevails [34,35]. In the flotation tests at pH 10, SiO(OH)₃⁻ is the most active species, playing a crucial role in depressing silicate minerals by adsorbing onto their surfaces and forming a hydrophilic layer that hinders collector attachment. The depression mechanism involves both physical and chemical adsorption, where the silicate species act as a Lewis base, sharing electron pairs with the Si-O groups on quartz and Ca-O groups on calcite and dolomite, disrupting the minerals’ hydrophobicity and making them less likely to float; primarily SiO(OH)₃⁻ and calcium ions (Ca²⁺) are on the mineral surfaces, resulting in the formation of a stable complex, such as Ca²⁺-OSi(OH)₃<surf.>, which further enhances the depression of these minerals and ensures selective flotation [35,36].

To further assess the depression effectiveness of sodium silicate, experiments were conducted using three different doses, 300, 350, and 400 g/t, to determine the optimal level for suppressing silicates and other gangue minerals carried over during the reverse flotation process. Based on the test results presented in Figure 4, a dosage of 350 g/ton was found to be optimal. The data show that increasing the sodium silicate dosage led to a decrease in silica content in the concentrate, while the MgO grade slightly increased, reaching approximately 44%. The effect on concentrate yield was also significant, with the highest yield of around 50% achieved at 350 g/t. These quantitative outcomes substantiate the selection of 350 g/ton as the most favorable dosage under the tested conditions.

A comparison of the results from the reverse and direct flotation tests reveals a significant improvement in the MgO grade of the final concentrate, exceeding 4%, with a recovery rate above 55%. Additionally, the SiO₂ content showed a reduction, achieving an approximate 8% improvement.

-

Sodium hexametaphosphate (SHMP, (NaPO₃)₆) tests

The results from three comparable tests using different SHMP dosages (100, 150, and 200 g/t) revealed that the most favorable outcome was achieved with a reagent consumption of 100 g/t (Figure 5). The intermediate product, which primarily contains depressed calcium-bearing gangues, showed a recovery of over 51% CaO with a grade of 7.6%. This led to the removal of approximately 45% of CaO impurities compared to the fresh feed material, which had a CaO content of 2.44%, resulting in a final concentrate with an acceptable CaO grade of 1.1%.

-

Sodium oleate (NaOL) tests

Sodium oleate (NaOL) is a widely utilized anionic collector in flotation processes, particularly for magnesite. In aqueous environments, the hydrolysis of sodium oleate involves a complex equilibrium, leading to the formation of oleic acid molecules, oleate ions, and oleate–molecule complexes, with the distribution of these species being pH-dependent. The adsorption mechanism is driven by the interaction of the mainly carboxylate groups of oleate with magnesium ions (Mg²⁺) on the magnesite surface, facilitating the formation of a hydrophobic metal–carboxylate complex. This hydrophobic layer significantly improves the flotation of magnesite by enhancing its water-repellent properties [37,38].

The influence of sodium oleate dosages on flotation performance was evaluated through three experimental sets, utilizing optimized concentrations of sodium silicate and SHMP, with varying levels of oleic acid, as shown in

Table 8. Influence of sodium oleate dosage on magnesite flotation under optimal conditions: pH 10, 100 g/t pine oil, 100 g/t SHMP, and 350 g/t SS.

Sodium Oleate (g/t)	Product	Weight Percentage (%)	Grade (%)		Recovery (%)
			MgO	SiO2	MgO	SiO2
200	Concentrate	46.45	43.97	5	52.52	13.78
	Middle conc.	20.83	38	29.47	20.35	36.41
	Tails	26.26	32.03	27.95	21.63	43.53
	Concentrate (magnetic)	6.16	33.11	16.39	5.5	6.28
	Feed	100	38.79	16.01	100	100
250	Concentrate	51.3	43.76	5.04	57.79	16.14
	Middle conc.	16.6	37.06	31.99	15.83	32.79
	Tails	25.79	32	27.98	21.25	44.54
	Concentrate (magnetic)	6.31	32.04	16.96	5.2	6.61
	Feed	100	38.84	16.02	100	100
300	Concentrate	52.86	42.1	5.1	57.55	16.74
	Middle conc.	15.4	40.4	34.66	16.09	33.15
	Tails	25.55	32.1	27.67	21.21	43.91
	Concentrate (magnetic)	6.19	32.16	16.12	5.15	6.2
	Feed	100	38.67	16.01	100	100

. According to the species distribution diagrams of oleate collectors, the highest activity of ionic oleate (R-COO⁻) and its dimer ((R-COO)₂²⁻) occurs at pH values above 8, so to leverage this optimal activity, the experiments were conducted at a pH of 10. The results indicated that increasing the oleic acid dosage improved recovery, though it may adversely affect the concentrate quality. As a result, an optimal oleic acid dosage of approximately 250 g/t was determined.

3.3.3. Direct Flotation Optimal Conditions

Based on preliminary experiments, the optimal conditions for direct flotation were identified as 100 g/t pine oil, 100 g/t SHMP, 350 g/t sodium silicate, 250 g/t sodium oleate, a solids concentration of 25%, and a pH of 10. Under these conditions, a product with a MgO grade of approximately 43.76%, silica content of about 5.04%, and a MgO recovery exceeding 57.79% was achieved. A comprehensive series of flotation experiments was conducted to assess the quantity and quality of the final concentrate, as well as to outline the key stages of the direct flotation process. The tests included a four-stage flotation process consisting of rougher flotation followed by three stages of recleaning, all optimized for maximum performance.

Table 9. Chemical composition analysis of samples from critical points in beneficiation circuit of magnesite processing.

Product	Yield (Wt%)	Grade (%)				Recovery %
		MgO	SiO2	CaO	Fe2O3	MgO	SiO2	CaO	Fe2O3
Final concentrate (direct flotation—cleaner 3)	15.83	46.52	1.05	0.73	0.44	19.98	0.99	0.91	6.21
Intermediate 3 (direct flotation tailings—cleaner 3)	9.85	45.5	1.41	1.32	0.44	10.12	0.75	2.78	3.71
Intermediate 2 (direct flotation tailings—cleaner 2)	12.99	43.87	3.1	2.6	0.47	14.95	2.44	9.11	5.81
Intermediate 1 (direct flotation tailings—cleaner 1)	15.13	42.17	4.03	2.02	0.44	15.23	8.47	18.73	0.34
Tailings (direct flotation—rougher)	17.34	32.05	31.21	4.7	0.53	14.47	31.94	43.36	8.74
Tailings (reverse flotation—second stage)	8.19	34.08	28.66	1.99	0.63	6.09	12.98	10.76	7.91
Tailings (reverse flotation—first stage)	14.23	33.51	27.58	1.76	0.65	11.72	33.55	8.24	6.8
Magnetic conc. (second attempt)	3.95	31.18	16.9	2.43	11.56	2.03	2.55	2.48	31.54
Magnetic conc. (first attempt)	2.49	33.57	17.08	2.1	6.86	5.41	6.33	3.63	28.94
Feed	100	38.36	16.53	2.44	1.05	100	100	100	100

presents the results, detailing the grade and recovery percentages for MgO, SiO₂, Fe₂O₃, and CaO, as well as their respective weight percentages across different product streams.

The final concentrate (direct flotation—cleaner 3), which accounts for 15.83% of the total feed mass, demonstrates a high MgO grade of 46.52%. This indicates the successful enrichment of magnesium oxide through the flotation process. The intermediate products, with a grade above 42%, contributed about 40% of the total feed mass. The low SiO₂ content in the final concentrate (1.05%) corresponds to a 94% removal of silica from the feed, highlighting the process’s efficiency in eliminating silica. Similarly, CaO is effectively reduced in the final concentrate, with a grade of 0.73% and recovery of just 0.91%, underscoring the selective removal of calcium oxide.

The four-stage direct flotation process demonstrates a progressive increase in the MgO grade, emphasizing the effectiveness of the applied strategy. Starting with intermediate 1, which has a MgO grade of 42.17%, the grade steadily improves through intermediate 2 and intermediate 3, ultimately reaching 46.52% in the final concentrate. This upward trend illustrates the efficient selective concentration of MgO at each flotation stage, ensuring the production of a high-quality concentrate. The sequential separation of MgO across the flotation stages supports the optimization of flotation parameters, reinforcing the strategy’s effectiveness. In contrast, the preliminary stages show a notable increase in impurity grades, such as SiO₂ and CaO, from the magnetic concentrate to the tailings. For instance, the SiO₂ content increases from 16.9% in the magnetic concentrate (second-stage feed) to 31.21% in the rougher flotation tailings. This trend highlights the successful isolation of impurities into the tailings while retaining MgO in the concentrate. Additionally, the use of magnetic separation as a pre-concentration step further enhances flotation performance by improving feed quality, contributing to the overall efficiency of the process. To clarify the global outcome of the process, the final concentrate represents 15.83% of the total mass, containing 46.52% MgO, which corresponds to an overall MgO recovery of 19.98%. The rest of the MgO is distributed across intermediate streams (approx. 40% recovery) and flotation/magnetic tailings. This demonstrates the selectivity of the flotation process and the potential for further optimization to improve recovery without compromising product quality.

To draw more conclusive comparisons, a flotation test should have been conducted under optimum conditions using the feed material as is, without any pre-concentration. This would provide a direct benchmark to evaluate the actual benefits of the pre-concentration stage. Without such a test, it is difficult to definitively attribute improvements in recovery or the grade solely to the pre-concentration process, as opposed to natural variability in the feed or flotation behavior. Including this control would strengthen the validity of the deductions made regarding the effectiveness of the pre-concentration step.

Figure 6 illustrates the beneficiation process following the completion of all relevant tests under optimal conditions, showcasing the effectiveness of each stage in achieving the desired product quality. Although a detailed cost–benefit analysis was beyond the scope of this study, the economic viability of the proposed beneficiation flowsheet (Figure 6) can be qualitatively supported by several key observations. The process effectively upgrades low-grade magnesite ore (~38.36% MgO) to a high-grade concentrate (46.52% MgO, with SiO₂ < 1.1%, Fe₂O₃ < 0.5%, and CaO < 0.75%), meeting the specifications required for refractory and metallurgical applications, thereby increasing its market value. The reagents used in the process, sodium silicate, sodium hexametaphosphate (SHMP), and sodium oleate, are not only cost-effective and readily available but also applied at optimized dosages, helping reduce chemical consumption and operational expenses. Additionally, the design of the flowsheet emphasizes efficient impurity removal and high product recovery, which reduces tailings disposal and maximizes resource utilization. These factors, taken together, suggest that the proposed process is both economically favorable and resource-efficient. A comprehensive techno-economic assessment will be undertaken in future work to provide quantitative validation of its commercial potential. The process begins with crushing, which reduces the ore size and facilitates the liberation of valuable minerals for subsequent processing. This is followed by grinding, which reduces the particle size to d₈₀ < 75 µm, enhancing surface contact for more efficient flotation. Magnetic separation (at 5000 Gauss) is then employed to remove magnetic impurities, improving the overall flotation efficiency. The process continues with a two-stage reverse flotation to remove silicates, followed by a flotation sequence consisting of rougher flotation and three cleaning stages, which selectively concentrate magnesite by separating it from gangue materials. The final product, with MgO: 46.52%, SiO₂: 1.05%, Fe₂O₃: 0.44%, and CaO: 0.73%, meets internationally recognized commercial specifications for caustic calcined magnesia, which serves as the primary feedstock for refractory applications. Each stage of the process plays a crucial role in enhancing the magnesite content and eliminating undesirable impurities. The proposed beneficiation process is readily scalable, with all unit operations and process conditions deliberately chosen to match conventional industrial practices. It employs standard equipment, including medium-intensity magnetic separators and Denver-type flotation cells, that are commonly used in commercial mineral processing plants. Moreover, the operating parameters, such as pH range, reagent type, and flotation residence time, reflect well-established industry norms, which support seamless scale-up from laboratory to pilot and full-scale operations. Because the flowsheet avoids complex or specialized techniques (e.g., roasting or acid leaching), it is operationally straightforward and compatible with existing plant infrastructure. This simplicity reduces the need for major capital investment or equipment retrofitting, further reinforcing its industrial applicability.

3.4. Comparative Evaluation with Previous Studies

To assess the effectiveness of the proposed beneficiation approach, a comparison was conducted with selected studies reported in the literature (

Table 10. Comparative summary of magnesite flotation performance under various feed grades and conditions.

Sample Type	Feed Composition (MgO/SiO2/Fe2O3) (%)	Final Concentrate Composition (MgO/SiO2/Fe2O3) (%)	MgO Recovery (%)	pH	Collectors	Depressants/Notes	References
Ore	38.36/16.53/1.05	46.52/1.05/0.44	~60	7 (rev.), 10 (dir.)	DDA, NaOL	SHMP, SS	This study
Ore	46.34/0.76/0.40	47.15/0.16/N.A.	73.28	5.5	Mixed amines	SHMP, SS	[1]
Ore	<46/~3.74/N.A.	>47.42/<0.18/N.A.	>78.36	5.5–6	DDA, ODA, NaOL, Mmm-10	SHMP, SS	[2]
Pure + ore	46.01/2.38/0.23	>47/N.A./N.A.	~70	5	EAH	SH, SS, TP, TN, OA	[6]
Pure	47.20/0.19/0.32	N.A.	74.35	11.6	KD	SH, SS, CMC	[7]
Ore	N.A.	46.85/1.21/N.A.	71.64	N.A.	NaOL, Lauric amine	SS	[39]
Ore	95.53/0.85/0.82	97.31/0.17/0.69	78.86	Neutral	LDK	SHMP, SS	[40]
Ore	43.52/2.4/N.A.	47.02/0.29/N.A.	71.64	2-stage	BK428, BK420	Sequential flotation	[41]
Ore	45.85/3.74/N.A.	46.81/0.54/N.A.	80.78	7	BK428, BK419B	SS, SHMP	[42]
Ore	41.78/4.21/N.A.	46.87/0.41/N.A.	68.21	Basic	NaOL, amine mix	SS + SHMP	[43]
Ore	32.36/17.7/N.A.	42.25/6.73/N.A.	Low	N.A.	NaOL, DDA	LM1/LM2, SHMP	[44]
Pure	48.35/0.15/0.12	N.A.	>98.98	10–11	NaOL + alcohols	Alcohol chain effect	[11]
Pure + mix	47.72/0.42/N.A.	N.A.	>80	~9.5	FA-1	SHMP, TSPP, CMC	[21]
Tailings	Mag: 52.3–60/Dol: 33–37	77.5/N.A./N.A.	45.5–53.2	10	Resanol A100	SHMP, HEDP	[23]
Pure	44.88/0.87/N.A.	45.73/N.A./N.A.	~85	10.5	NaOL + Tween 80	SHMP	[27]
Pure	47.27/0.21/N.A.	N.A.	96.78	9.8	NaOL:BS-12 (3:2)	SHMP	[38]

NaOL: sodium oleate; DDA: dodecylamine; SHMP: sodium hexametaphosphate; CMC: carboxymethyl cellulose; EAH: ether amine hydrochloride; KD: a commercial cationic amine collector; SH: sodium hexametaphosphate; SS: sodium silicate (water glass); TP: tetrasodium pyrophosphate; TN: sodium tripolyphosphate; OA: oxalic acid; BS-12: dodecyl dimethyl betaine (amphoteric surfactant); FA-1: tall oil fatty acid (mainly oleic and linoleic acids); TSPP: tetrasodium pyrophosphate; HEDP: 1-Hydroxyethylidene-1,1-diphosphonic acid; Resanol A100: commercial anionic collector for magnesite; LM1/LM2: proprietary depressants (composition undisclosed); BK428, BK420, and BK419B: proprietary flotation collectors; LDK: mixed amine collector (proprietary formula).

). The results demonstrate that our method delivers competitive performance, even when starting from significantly lower-grade feed material. Whereas many previous studies processed ores with MgO contents exceeding 45%, our study focused on a low-grade magnesite ore containing 38.36% MgO, 16.53% SiO₂, and 1.05% Fe₂O₃. Despite the lower feed grade, the integrated magnetic separation and multi-stage flotation flowsheet produced a final concentrate with 46.52% MgO, 1.05% SiO₂, and 0.44% Fe₂O₃, comparable to or better than results reported in the literature using higher-grade inputs. For instance, Tan et al. obtained 46.87% MgO and 0.41% SiO₂ from a feed with 41.78% MgO, while Wei et al. reported a concentrate containing 46.85% MgO and 1.21% SiO₂, although their feed grade was not specified.

4. Conclusions

Magnesium is increasingly recognized as a critical metal for high-tech industries due to its unique properties and broad range of advanced applications. However, the depletion of high-grade, high-purity magnesite deposits, combined with rising demand, highlights the urgent need to process lower-grade magnesite resources.

This study focused on the beneficiation of low-grade magnesite ore (~70% MgCO₃), which contains high levels of impurities (Fe₂O₃ + Al₂O₃ + SiO₂ + CaO > 20%). The major challenge was the separation of magnesite from gangue minerals with similar physico-chemical characteristics.

To address this, magnetic separation and flotation methods were systematically evaluated. Magnetic separation at 5000 Gauss was found to be optimal, reducing Fe₂O₃ content in the non-magnetic product to 0.8% while achieving a high MgO recovery rate of over 94%. These results were consistent across repeated trials, confirming process reliability.

In parallel, flotation experiments using a combined approach—two-stage reverse flotation with dodecylamine (DDA) for silicates, followed by four-stage direct flotation for magnesite—achieved effective removal of over 94% of silica and other impurities. This process yielded a magnesite concentrate with MgO content exceeding 46%, meeting commercial standards for refractory applications.

Overall, this study demonstrates that a combination of optimized magnetic separation and flotation techniques can successfully upgrade low-grade magnesite to a high-purity concentrate, supporting more sustainable utilization of magnesite resources.