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Article

Recovery of High-Alkali-Grade Feldspar Substitute from Phonolite Tailings

by
Savas Ozun
1,*,
Semsettin Ulutas
2 and
Sema Yurdakul
3
1
Department of Mining Engineering, Süleyman Demirel University, Isparta 32260, Türkiye
2
General Directorate of Mining and Oil Affairs, Ministry of Energy and Natural Resources, Ankara 06560, Türkiye
3
Department of Environmental Engineering, Süleyman Demirel University, Isparta 32260, Türkiye
*
Author to whom correspondence should be addressed.
Processes 2025, 13(8), 2334; https://doi.org/10.3390/pr13082334
Submission received: 11 May 2025 / Revised: 18 June 2025 / Accepted: 23 June 2025 / Published: 23 July 2025
(This article belongs to the Section Separation Processes)
Browse Figures
Versions Notes

Abstract

Phonolite is a fine-grained, shallow extrusive rock rich in alkali minerals and containing iron/titanium-bearing minerals. This rock is widely used as a construction material for building exteriors due to its excellent abrasion resistance and insulation properties. However, during the cutting process, approximately 70% of the rock is discarded as tailing. So, this study aims to repurpose tailings from a phonolite cutting and sizing plant into a high-alkali ceramic raw mineral concentrate. To enable the use of phonolite tailings in ceramic manufacturing, it is necessary to remove coloring iron/titanium-bearing minerals, which negatively affect the final product. To achieve this removal, dry/wet magnetic separation processes, along with flotation, were employed both individually and in combination. The results demonstrated that using dry high-intensity magnetic separation (DHIMS) resulted in a concentrate with an Fe2O3 + TiO2 grade of 0.95% and a removal efficiency of 85%. The wet high-intensity magnetic separation (WHIMS) process reduced the Fe2O3 + TiO2 grade of the concentrate to 1.2%, with 70% removal efficiency. During flotation tests, both pH levels and collector concentration impacted the efficiency and Fe2O3 + TiO2 grade (%) of the concentrate. The lowest Fe2O3 + TiO2 grade of 1.65% was achieved at a pH level of 10 with a collector concentration of 2000 g/t. Flotation concentrates processed with DHIMS achieved a minimum Fe2O3 + TiO2 grade of 0.90%, while those processed with WHIMS exhibited higher Fe2O3 + TiO2 grades (>1.1%) and higher recovery rates (80%). Additionally, studies on flotation applied to WHIMS concentrates showed that collector concentration, pulp density, and conditioning time significantly influenced the Fe2O3 + TiO2 grade of the final concentrate.

1. Introduction

Ceramics, in their broadest definition, are materials derived from the high-temperature firing of inorganic, nonmetallic, predominantly silicate-based raw material blends. This process is designed to induce hardening and enhance mechanical strength, and may or may not involve a glazing step. The proportions of constituent minerals in the ceramic formulation are critical, directly influencing the quality and type of the resultant product [1]. Moreover, the purity and characteristics of the mineral concentrates employed are significant determinants of the final ceramic’s strength and overall quality. Individual minerals impart diverse functionalities to the ceramic body, such as plasticity, mechanical strength, hardness, abrasion resistance, chemical inertness, and the modulation of thermal expansion and melting point (fluxing). Of these, the reduction in the melting temperature is of paramount importance for achieving enhanced vitrification, superior material properties, energy conservation, and cost-efficient manufacturing. This is accomplished through the incorporation of alkali oxides, notably K2O and Na2O, which are primarily supplied by feldspathic minerals or their substitutes, such as nepheline syenites and granites [2,3,4,5].
Diminishing reserves of high-grade ores present escalating challenges to contemporary mineral extraction and processing. Consequently, the mining industry is increasingly compelled to exploit lower-grade deposits, which are inherently burdened with a greater content of gangue minerals. The removal of these non-valuable constituents is essential in order to upgrade the ore for industrial utilization [6,7,8]. Beyond the imperative to process low-grade ores, a rising demand from the ceramics and glass industries for high-purity mineral concentrates also necessitates the recovery of valuable ceramic raw materials from mining and mineral processing wastes, such as tailings and waste rock dumps. These often result from inefficient processing techniques and over-grinding [9,10,11]. Effectively addressing this reclamation is also vital for alleviating the environmental pollution and land degradation stemming from these waste materials.
The methods for processing ceramic raw materials vary according to the mineral composition and the particle size at which these minerals are liberated from their associated gangue. When liberation occurs at coarser sizes, the concentration grade can be effectively improved using straightforward physical separation techniques, such as handpicking, jigging, and magnetic separation [12,13,14,15,16,17]. However, if valuable minerals are liberated from gangue minerals at finer sizes, the application of one or more complex separation techniques, such as flotation, may be required. Thus, the effectiveness of the selected separation process could be highly complex, depending on various factors including ore characteristics, magnetic susceptibility, density, liberation size, and differences in surface properties between valuable and gangue minerals [12,13,14,15,18,19].
Reducing firing temperatures in ceramic and glass manufacturing is crucial for economic efficiency. Even minor reductions in firing temperature can significantly enhance production economics. In pursuit of this objective, materials such as syenite, nepheline syenite, and trachyte have sometimes been employed as fluxing agents in ceramic bodies, acting as alternatives to feldspars [20,21,22]. Phonolite, with its higher alkali and alumina content per unit weight compared to feldspars, offers considerable potential in this regard [23]. The lower fusion point of phonolite provides a distinct advantage over feldspar by enabling a reduction in the melting temperature of ceramic bodies during firing. This leads to benefits such as faster melting, improved productivity, and substantial fuel savings. Furthermore, phonolite enhances the impact, bending, and thermal shock resistance of glass and ceramics and acts as an effective vitrifying agent [24,25,26]. However, the high iron and titanium oxide contents in raw samples are undesirable for industrial applications due to their detrimental impact on the final products during the firing process. This research study, therefore, investigates the potential utilization of abundant phonolite tailings—which constitute approximately 70% of the processed material and present significant storage challenges for mining operations. Our objective is to transform these underutilized waste resources into valuable products through an optimized beneficiation approach combining flotation and magnetic separation techniques, applied both independently and in combination. By selectively removing the high Fe2O3 + TiO2 content from these tailings, we aim to produce a high-alkaline concentrate suitable for ceramic and glass manufacturing applications. Such a concentrate could serve as an effective substitute for traditional feldspar materials in industrial processes, thereby offering a dual benefit: reducing environmental burden through sustainable waste management while creating economic value from previously discarded materials.

2. Materials and Methods

2.1. Materials

The samples utilized in this study were obtained from the tailings of a natural stone quarry and an associated processing plant located in the Direkli region of Isparta in Western Türkiye. The area is characterized by a visible rock reserve estimated at 600,000 m3. Rock extraction is conducted using open-pit mining techniques. The extracted rock typically exhibits a sub-rounded, irregular morphology, often described as ‘potato-shaped’ (Figure 1a). Post-extraction, the rock undergoes cutting and sizing, principally using ST-type machinery, to be used as a building material for wall construction and floor tiling. Irregular portions and offcuts resulting from the cutting process (Figure 1b,c) are directly discarded as tailings. These tailings consist of irregular, cracked, and broken pieces, defective products, and rock fines, which collectively constitute approximately 70% of the processed rock blocks (Figure 2). The particle size of rock fragments within the tailings varies from the millimeter to the micron scale.
Microphotographic analysis (Figure 3) and X-ray diffraction (XRD) data, obtained using an Inel Equinox 1000 diffractometer (INEL, Artenay, France) (Figure 4), indicated that the samples were primarily composed of nepheline, K-feldspar, oligoclase, and amphibole, with minor amounts of pyroxene and biotite. The rock exhibited a porphyritic texture, characterized by relatively coarse-grained feldspathoid and amphibole phenocrysts within a very fine-grained matrix. Magnetite and ilmenite were the predominant opaque minerals. Due to its porphyritic texture, the rock displayed a bimodal particle size distribution, with coarse particles (35%; 0.5 mm to 3 mm) and fine particles (65%; 100 µm to 0.5 mm). The coarse fraction predominantly comprised nepheline, orthoclase, amphibole, pyroxene, and minor magnetite, while the fine fraction consisted mainly of nepheline (approximately 80%). The chemical composition, determined by X-ray fluorescence (XRF) spectrometry using a Spectro XLAB2000 instrument (Spectro, Kleve, Germany), is presented in Table 1. These XRF results are consistent with the mineralogical composition identified by XRD analysis and literature data [27].

2.2. Method

Before experimental studies, initial sample comminution utilized both jaw and roller crushers. Subsequently, a portion of this material was classified into narrow particle size fractions (ranging from 0.833 mm to 0.037 mm) using a Tyler screen series. The degree of liberation for these fractions was assessed by microscopical analysis. Based on the liberation size determined (0.104 mm), the experimental work in this study was conducted on representative samples from the −0.104 to +0.037 mm particle size fraction. The comminution process for the bulk rock sample was achieved by applying a jaw crusher (to <0.5 cm) and a roller crusher. For each beneficiation parameter, the arithmetic mean of three replicate tests was considered the final value.
To remove magnetic minerals from the phonolite sample, laboratory-scale Carpco DHIMS Model MIH (13)111-5 (Carpco, Texarkana, AR, USA) and Carpco WHIMS Model MWL-3465 (Carpco, Texarkana, AR, USA) apparatus were utilized. The variables applied in the magnetic separation processes were determined based on the preliminary studies. In the DHIMS process, roll speeds were varied from 30 to 50 rpm while maintaining a constant feed rate of 2 kg/h. The magnetic field intensity was set to approximately 15,000 gauss, representing the device’s maximum intensity. For the WHIMS process, pulp with a solid content of 15% by weight underwent separation at controlled feed rates ranging from 1.5 to 3 kg/h.
Laboratory-scale flotation tests were performed in a Denver flotation machine (Denver Equipment Company, Denver, CO, USA) utilizing a 1.5 L cell. The initial conditioning phase was conducted at an impeller speed of 1750 rpm and a pulp density of 35% solids (w/w) to facilitate effective collision between mineral particles and collector species. Following this 5 min conditioning period, the pulp was diluted to 15% solids (w/w) with water, and the impeller speed was decreased to 1500 rpm for the flotation stage. Subsequently, air bubbles were introduced to the system and underwent a 1 min defoaming process. The flotation variables utilized in this study are detailed in Table 2. The chemical compositions of the feed, concentrate, and tailings were determined by XRF analysis.

3. Results and Discussion

In ceramic production, the Fe2O3 content in the raw materials can lead to yellow, brown, reddish-brown, and wine-red colors during oxidative firing, and gray-blue and dark gray shades during reduction firing, depending on the concentration used in glazes. TiO2 is also undesirable because of its high melting point and its interaction with Fe2O3, which can cause various color changes [28,29]. Therefore, the effectiveness of the separation process in this study was evaluated based on the efficiency of Fe2O3 + TiO2 removal, the Fe2O3 + TiO2 grade of the concentrate, and recovery (%).

3.1. Magnetic Separation Tests

The primary goal in the beneficiation of ceramic raw materials is to eliminate Fe-Ti-bearing minerals using either a single method or a combination of multiple separation techniques. Magnetic separation is the most common approach due to its quick, simple, and chemical-free nature, as most coloring minerals in industrial mineral deposits are magnetically susceptible [30,31,32]. In general, the intensity of the magnetic field varies based on the objectives of the process; it is generally low when separating magnetic minerals with differing susceptibilities from each other and high when removing magnetic minerals from non-magnetic ceramic raw minerals. This cost-effective method can be conducted in wet or dry conditions to attain pre-concentrates and final concentrates. Consequently, the aim of this part of the study was to derive a ceramic-grade concentrate from low-grade phonolite ore utilizing DHIMS and WHIMS methods. Comparative results from upgrading the phonolite ore using both techniques are presented separately, focusing on the Fe2O3 + TiO2 removal efficiency (%) and recovery (%).

3.1.1. DHIMS Tests

DHIMS is a widely recognized and effective method for eliminating magnetic impurities from raw materials used in ceramics and glass manufacturing. It facilitates pre-concentration at coarser particle sizes, which leads to energy savings by reducing the volume and weight of material requiring subsequent grinding. Additionally, since there is no contact between the magnetic roll and the ore in DHIMS, wear on the roll is minimal, which minimizes contamination. Another advantage of DHIMS is its environmental friendliness; it is not affected by climate variations or water scarcity, eliminating the need for the dewatering of concentrates or solids removal from the circulating water in the plant [33,34]. The efficacy of DHIMS depends on several factors, including magnetic field intensity, feed rate, roll speed, and feed size [31,34,35,36].
As presented in Figure 5a–c, the recovery (%), Fe2O3 + TiO2 grade (%), and Fe2O3 + TiO2 removal efficiency (%) were all dependent on the roll speed. Slower roll speeds increased the duration for which mineral particles were exposed to the magnetic field, thereby promoting the adherence of poorly liberated particles (due to interlocked magnetic phases) to the magnetic roll. These particles subsequently resisted the centrifugal force generated by the rolling motion, leading to the concentrate exhibiting its lowest Fe2O3 + TiO2 grade (<1%) (Figure 5a,b). Furthermore, a direct correlation was observed between recovery rates and the concentrate’s Fe2O3 + TiO2 grade (%). As the roll speed was elevated, the centrifugal force gained prominence, displacing magnetic particles, especially coarser ones, towards the non-magnetic stream. Consequently, at 50 rpm, the Fe2O3 + TiO2 grade reached about 1.5%, and recovery rates progressively increased to about 58%.
The presence of magnetic mineral fines may also contribute to low recovery rates. These fines can coat the surfaces of non-magnetic particles, imparting magnetic properties to them. Consequently, non-magnetic particles might detach from the magnetic roll surface later than expected and report to the magnetic fraction, thereby causing concentrate loss. This mechanism could explain the achievement of a concentrate with slightly over 12% alkali content and 0.95% Fe2O3 + TiO2 at a roll speed of 30 rpm, corresponding to an Fe2O3 + TiO2 removal efficiency of approximately 85%. Although DHIMS effectively removed magnetic minerals from the phonolite sample, as illustrated in Figure 3 and detailed in Section 2, the sample’s very fine-grained matrix contained unliberated iron/titanium minerals (e.g., magnetite, amphibole phenocrysts, amphibole, and biotite). So, the presence of these unliberated phases hindered their complete removal, thereby limiting the Fe2O3 + TiO2 removal efficiency and resulting in a higher final product grade. Increasing the roll speed to 40 rpm led to a decline in Fe2O3 + TiO2 removal efficiency to around 82.5%, yielding a concentrate with 1.11% Fe2O3 + TiO2. A further increase to 50 rpm reduced the Fe2O3 + TiO2 removal efficiency to below 80%, resulting in a concentrate grade of 1.47% Fe2O3 + TiO2 (Figure 5c).

3.1.2. WHIMS Tests

Compared to dry magnetic separation, the wet magnetic separation process reduces dust formation during processing, thereby contributing to a reduction in air pollution. This method can achieve higher separation efficiency and recovery rates than DHIMS, as the presence of water prevents the agglomeration of fine mineral particles. Additionally, WHIMS prevents non-magnetic mineral fines from coating the surfaces of magnetic minerals, effectively enhancing the capture of magnetic minerals [37]. Therefore, in this part of the study, the tailings from the phonolite processing plant were subjected to the WHIMS process. To achieve high separation efficiency during the tests, a pulp with a 15% solid content, determined by preliminary studies, was fed into a wet magnetic separator at controlled feeding rates ranging from 1.5 to 3 kg/h.
The results presented in Figure 6a–c indicate that dispersing minerals in water within the pulp, particularly when considering feed size, facilitated the separation of magnetic from non-magnetic minerals. Consequently, significantly higher recovery rates were obtained with WHIMS compared to DHIMS. The data also demonstrated that the recovery rates (%) increased with increasing feed rates (Figure 6a), ranging from about 65% to 90% as the feed rate varied between 1.5 kg/h and 3 kg/h. Conversely, higher feed rates led to a greater quantity of minerals passing through the magnetic field per unit of time, thereby diminishing the exposure time of the magnetic minerals to the magnetic field. This led to the entrapment of some magnetic minerals within the non-magnetic matrix, thereby producing concentrates with higher Fe2O3 + TiO2 grades (%). Specifically, at a feed rate of 1.5 kg/h, an Fe2O3 + TiO2 removal efficiency of about 82% was achieved, yielding a concentrate with a 1.2% Fe2O3 + TiO2 grade. As the feed rate increased to 2.25 kg/h, the Fe2O3 + TiO2 removal efficiency decreased to about 78%, resulting in a concentrate grade of 1.4% Fe2O3 + TiO2. Further increasing the feed rate to 3 kg/h caused the Fe2O3 + TiO2 removal efficiency to decline to less than 70%, producing a concentrate with an Fe2O3 + TiO2 grade exceeding 1.8%.

3.2. Flotation Tests

One of the primary challenges in beneficiating ores containing Fe2O3 + TiO2-bearing minerals arises when these minerals exhibit low magnetic susceptibility. This characteristic renders their separation by conventional magnetic methods difficult, if not ineffective. Consequently, for minerals that cannot be efficiently removed from the ore by magnetic separation due to their weak magnetic response, flotation emerges as a crucial and often necessary alternative, thereby enabling their removal where magnetic separation alone fails.
Flotation, recognized as the most effective and common technique for removing coloring impurities from industrial minerals, is a separation process that improves concentrate grade by separating valuable minerals from gangues based on differences in their physicochemical surface characteristics [38,39,40]. Conventional industrial mineral beneficiation relies on reverse flotation, where two acidic circuits remove gangues—using cationic long-chain alkyl amines for mica and sulfate/sulfonate for iron/titanium minerals—leaving valuable minerals like feldspars in the pulp [6,13,14]. With the development and application of fatty-acid-based collectors, separation within a single-stage alkaline flotation circuit has been enabled, providing an alternative to conventional processes. This approach not only eliminates the detrimental effects of acid use on the environment and equipment but also offers the potential for enhanced capacity. Oleic acid/oleate, an unsaturated fatty acid, is one of the most used fatty acids, exhibiting greater selectivity compared to its saturated derivatives. Their effectiveness in flotation processes varies depending on their chemical properties in aqueous solutions and the type of minerals present in the ore. Moreover, they are pH-sensitive, existing predominantly as ions (RCOO−) at high pH levels and as ion–molecule complexes ((RCOO)2H−) under moderately neutral to alkaline pH conditions. Their interaction with mineral surfaces can be either physical with oppositely charged particles or chemical with available cationic sites on mineral surfaces [41,42,43,44,45,46,47]. Based on the Fe2O3 + TiO2 removal efficiencies (%) achieved in preliminary flotation tests and data from the literature, subsequent flotation tests investigating collector concentration were carried out at alkaline pHs (pH > 9) [41,42,43,44,45]. The results, presented in Figure 7a–c, are categorized by recoveries (%), Fe2O3 + TiO2 grades (%), and Fe2O3 + TiO2 removal efficiencies (%).
The results indicated that recovery rates were dependent on both pH and collector concentration. A consistent trend was observed across all collector concentrations, with recovery generally peaking at pH 9 and declining at pH values below and above this optimum. Specifically at pH 9, maximum recovery values increased with collector concentration, ranging from approximately 60% (at 1000 g/t) to 80% (at 2000 g/t). In contrast, optimal Fe2O3 + TiO2 removal efficiencies, which corresponded to the lowest Fe2O3 + TiO2 grades in the concentrates, were achieved at pH 10 (Figure 7b,c). Under these conditions (pH 10), Fe2O3 + TiO2 removal efficiencies increased with increasing collector concentration and varied between 65% and 75%, yielding a concentrate with a minimum Fe2O3 + TiO2 grade of about 1.5% (and with the highest alkali content (K2O + Na2O) of 12%) in the presence of 2000 g/t of the collector. For this reason, and to optimize both recovery and concentrate quality, the flotation tests in this study were carried out at pH > 9, with consideration given to both the recovery rate (%) and Fe2O3 + TiO2 grade (%) of the concentrate. The results obtained were determined to be in line with the literature data corresponding to the peak activity of the (RCOO)2H− ion–molecule complex, whose activity decreases towards high, alkaline pH conditions [18,44,45,46,47].
It was also determined that concentrate-forming minerals underwent co-flotation with gangue minerals. This limited selectivity resulted from the interaction between the collectors and industrial minerals (such as feldspars), which involved adsorption onto aluminum (Al) sites on the surfaces of these minerals. Moreover, the co-flotation of concentrate-forming minerals and their tendency to mix with the tailings may be attributed to the activation of their surfaces by cations (e.g., Ca2+, Mg2+) present in the process water, facilitating enhanced collector adsorption [44,48,49]. Therefore, despite high overall flotation recovery at a slightly alkaline pH (pH 9), the poor selectivity resulted in low Fe2O3 + TiO2 removal efficiencies (%) and, consequently, concentrates with an undesirably high Fe2O3 + TiO2 content.

3.3. Magnetic Separation Tests for Flotation Concentrate

In the processing of ceramic raw materials, such as feldspars, magnetic separation is generally used to produce a pre-concentrate for flotation processes or a final concentrate following flotation operations. Thus, considering industrial applications, to obtain higher-grade concentrates, the flotation concentrates obtained with different collector dosages were separately subjected to DHIMS and WHIMS processes. During the tests, pre-determined optimal variables were applied: a magnetic roll speed of 40 rpm and feed rate of 2 kg/h for DHIMS; a pulp density of 15% and feed rate of 2.25 kg/L for WHIMS.

3.3.1. DHIMS Tests for Flotation Concentrate

Flotation concentrates typically contain significant amounts of water, making the subsequent DHIMS application both challenging and costly due to the necessity of complete water removal. Even after drying, fine particles can lead to agglomeration, and the residual flotation reagents may also alter particle surface properties, further complicating dry separation. Nevertheless, at this part of the study, DHIMS was applied, considering not only the low Fe2O3 + TiO2 grade (%) of the concentrate obtained directly from DHIMS but also evaluating whether its application following a flotation stage would similarly lead to a further reduction in Fe2O3 + TiO2 grade (%). In experimental tests, flotation concentrates obtained using 1000 g/t and 2000 g/t collector concentrations at various pH levels were subjected to DHIMS comparatively. The Fe2O3 + TiO2 removal efficiencies (%) given in Figure 8a,b were calculated separately by multiplying the ratio of the concentrate grade obtained through DHIMS by the concentrate grade obtained at each pH value by 100.
The results presented in Figure 8a,b indicate that both the Fe2O3 + TiO2 grade (%) of the concentrates and the Fe2O3 + TiO2 removal efficiencies (%) varied depending on the initial flotation concentrate, which was produced at different pH values. The Fe2O3 + TiO2 removal efficiencies (%) were inversely proportional to the collector concentration, being higher with 1000 g/t of collector compared to 2000 g/t across all pH levels. As detailed in Section 2, the phonolite sample contains various minerals, exhibiting distinct physicochemical surface properties in aqueous solutions. Consequently, after conditioning with the collector at different concentrations and pH levels, these minerals displayed a spectrum of surface hydrophobicities. Minerals that failed to achieve sufficient surface hydrophobicity during conditioning were not effectively recovered by flotation. Therefore, subjecting flotation concentrates produced with lower collector concentrations to subsequent DHIMS yielded higher Fe2O3 + TiO2 removal efficiencies (%).
The application of DHIMS to flotation concentrates revealed that the highest Fe2O3 + TiO2 removal efficiencies (>60%) for both collector concentrations were achieved using concentrates produced at pH ≤ 9, conditions under which the collector was less effective. Conversely, at pH 10, where flotation effectively removed the majority of Fe-Ti-containing minerals due to their acquired hydrophobicity, subsequent DHIMS processing resulted in the lowest Fe2O3 + TiO2 removal efficiencies (ranging from >42% to >55% with increasing collector concentration). Furthermore, the centrifugal force from the roll speed, which preferentially affects coarser particles, may have also influenced DHIMS performance. This factor, related to the feed’s particle size distribution, likely contributed to the observed DHIMS recoveries being below 70%.

3.3.2. WHIMS Tests for Flotation Concentrate

In the processing of industrial minerals like feldspar, WHIMS is often employed as a supplementary stage to both enhance the grade of the flotation concentrate and remove Fe-Ti-containing impurities that may be introduced during concentrate handling and transfer [6]. Accordingly, in this part of the study, flotation concentrates were subjected to WHIMS. The resulting Fe2O3 + TiO2 grades (%) of the concentrates and removal efficiencies (%) for each collector dosage and pH value are presented in Figure 9a,b.
Consistent with the findings in Section 3.3.1, applying WHIMS to flotation concentrates produced with a 1000 g/t collector dosage yielded higher Fe2O3 + TiO2 removal efficiencies (%) compared to those from a 2000 g/t dosage. The maximum Fe2O3 + TiO2 removal efficiencies were approximately 60% and 40% for the lower and higher collector concentrations, respectively. This disparity is attributed to the less complete flotation of Fe-Ti-containing minerals at lower collector concentrations, which results in a higher presence of Fe-Ti-containing minerals remaining in the flotation concentrate. Consequently, except at the optimal flotation pH of 10, these remaining Fe-Ti minerals were more effectively removed by subsequent WHIMS treatment, leading to enhanced Fe2O3 + TiO2 removal efficiencies (%).
In WHIMS, water facilitates the dispersion of mineral particles within the pulp, thereby preventing aggregation and ensuring the optimal exposure of each particle to the magnetic field. This characteristic contributed to WHIMS achieving better recovery results (>85%) compared to DHIMS. However, concerning the Fe2O3 + TiO2 grade (%) of the concentrate, the WHIMS treatment of flotation concentrates yielded higher grades (e.g., >1.10% Fe2O3 + TiO2 at pH 10) than those obtained from the DHIMS treatment of similar flotation concentrates (e.g., <1.05% Fe2O3 + TiO2 at pH 10). The elevated Fe2O3 + TiO2 grades in the WHIMS concentrates were attributed to the higher feed rates employed in this process.

3.4. Flotation Tests for WHIMS Concentrate

Magnetic separation is frequently utilized in the beneficiation of industrial minerals, such as feldspar ores, to generate a pre-concentrate ahead of subsequent flotation stages. This approach facilitates the removal of liberated magnetic minerals from the ore, thereby reducing the volume of material requiring grinding and flotation. Consequently, this pre-treatment step can minimize grinding costs, equipment size requirements, and the consumption of flotation reagents. While flotation offers the primary advantage of greater selectivity, optimal results are often achieved through suitably combined processes, such as implementing magnetic pre-concentration before flotation.
Based on a comparative assessment of WHIMS and DHIMS performance, further flotation experiments were conducted exclusively with WHIMS concentrates due to their markedly greater recoveries (>85%) relative to those from DHIMS (<50%). According to the results of the flotation tests, the concentrate grade (Fe2O3 + TiO2%) was determined to be inversely proportional to the recovery, which was dependent on the feed rate, as given in Figure 6a–c. The minimum recovery corresponded to a high feed rate (3 kg/h), while the minimum grade was achieved at a low feed rate (1.5 kg/h). Therefore, a feed rate of 2.25 kg/h was selected as the optimal condition for WHIMS, which provided a balance between the Fe2O3 + TiO2 grade (%) of the concentrate and recovery (%). This resulted in approximately 80% recovery with an Fe2O3 + TiO2 grade of 1.46%.
The flotation tests were carried out with consideration of the influence of collector concentration, pulp density, and conditioning time. The Fe2O3 + TiO2 removal efficiencies (%) were calculated by multiplying the ratio of the concentrate grade obtained through flotation by the concentrate grade obtained via WHIMS by 100.

3.4.1. Collector-Concentration-Dependent Flotation Tests for WHIMS Concentrate

Fatty-acid-based collectors are among the most commonly employed anionic reagents for the flotation of Fe-Ti-bearing metal oxides (e.g., magnetite, ilmenite, and rutile), rare earth elements (REEs), and soluble salt minerals (e.g., phosphates). While their strong collecting power typically results in poor selectivity, this characteristic was specifically exploited to be advantageous in this research. The goal was to eliminate all Fe-Ti-containing minerals to yield a concentrate with an Fe2O3 + TiO2 grade (%) as low as possible.
In the collector-concentration-dependent flotation tests, which ranged from 1000 to 2000 g/t, the variables were applied based on previously determined and employed conditions at pH 10. The results given in Figure 10a,b indicated that the Fe2O3 + TiO2 grade (%) of the concentrate varied between 1.25% and 1.35%, exhibiting an inverse relationship with collector dosage. Correspondingly, as the collector concentration increased, increasing collector concentrations led to maximum Fe2O3 + TiO2 removal efficiencies (%) of approximately 7.5%, 9.6%, and 15%, respectively.

3.4.2. Pulp-Density-Dependent Flotation Tests for WHIMS Concentrate

The success of flotation processes is governed by several factors, among which pulp density, particle size distribution, and conditioning time are paramount. Pulp density, in particular, is a key determinant of plant throughput, as capacity generally increases with higher solids concentrations in the pulp. However, the percentage of solids also critically impacts concentrate grade and the efficiency of removing undesirable mineral phases. Consequently, lower-density pulps tend to provide greater flotation performance, often leading to higher concentrate grades than those achieved with higher-density pulps. Moreover, elevated pulp densities increase pulp viscosity, which in turn can hinder the effective transfer of hydrophobic minerals via air bubble attachment [50,51,52]. Given these considerations, the aim of this part of the study was to investigate the influence of pulp density on the Fe2O3 + TiO2 grade (%) of the concentrate and the associated Fe2O3 + TiO2 removal efficiency (%). Figure 11a,b present the results from experiments utilizing varying pulp densities, ranging from 15% to 25% solids by weight.
The experimental results indicated a decrease in the Fe2O3 + TiO2 grade (%) of the flotation concentrates as the pulp solids content increased. The initial WHIMS concentrate, the feed for flotation, had an Fe2O3 + TiO2 grade of 1.46%. At the lowest tested pulp density (15%), this grade was reduced to 1.35%, representing an approximate 7.50% decrease in these impurities. Further increases in pulp density to 20% and 25% resulted in Fe2O3 + TiO2 grades of 1.30% and 1.25%, respectively. This trend demonstrated that increasing pulp density, up to a certain point, enhanced the rejection of Fe2O3 + TiO2 from the concentrate. However, raising the pulp density excessively was found to negatively impact critical flotation mechanisms, including the probability of particle–collector collision, the interaction between hydrophobic particles and air bubbles, and an increase in pulp viscosity. Consequently, the maximum pulp density for the flotation tests was set at 25%.

3.4.3. Conditioning-Time-Dependent Flotation Tests for WHIMS Concentrate

Conditioning, the initial agitation of mineral particles with reagents in an aqueous medium, is crucial for effective flotation. This stage ensures reagent dispersion throughout the pulp, facilitating contact between the mineral particles and the added chemicals. Ideally, during an optimal conditioning period, collector species adsorb as a monolayer onto mineral surfaces, imparting hydrophobicity. Inadequate conditioning time results in insufficient collector adsorption, leading to poor flotation recovery. Conversely, prolonged conditioning (over-conditioning) can be detrimental to mineral hydrophobicity due to the potential formation of collector multilayers [53]. Accordingly, the influence of conditioning time on the Fe2O3 + TiO2 grade (%) in the concentrate and the overall Fe2O3 + TiO2 removal efficiency (%) was investigated in this part of the study. To achieve this, flotation tests were conducted on WHIMS concentrates at various conditioning times, ranging from 2.5 to 15 min, while maintaining other flotation variables at constant values.
The primary advantages of obtaining pre-concentrates via magnetic separation include reduced reagent consumption and the more effective utilization of flotation cell capacities during subsequent flotation processes. To facilitate comparison, flotation tests on the magnetic separation pre-concentrates were conducted at pH 10.0 with a collector concentration of 1000 g/t. As presented in Figure 12a,b, a 2.5 min conditioning time resulted in poor Fe2O3 + TiO2 removal efficiency (%), yielding a concentrate with a high Fe2O3 + TiO2 grade (about 1.4%). This result was attributed to insufficient interaction between gangue minerals and collector species due to the short conditioning period. However, longer conditioning times led to increased Fe2O3 + TiO2 removal efficiencies and, consequently, concentrates with lower Fe2O3 + TiO2 grades. Specifically, the removal efficiency increased from about 6% with 2.5 min of conditioning (Fe2O3 + TiO2 grade: 1.37%) to about 10% with 5 min of conditioning (Fe2O3 + TiO2 grade: 1.30%). Extending the conditioning time to 10 min resulted in a sharp increase in removal efficiency to approximately 20%, enabling the production of a concentrate with a significantly lower Fe2O3 + TiO2 grade of 1.18%. Beyond this point, the results obtained with 15 min of conditioning were nearly identical to those observed after 10 min.

4. Conclusions

This experimental study successfully evaluated the potential of upgrading phonolite cutting/sizing plant tailings into high-alkaline, low-iron/titanium raw materials, suitable as feldspar alternatives for the ceramic–glass industry. Various beneficiation strategies, including DHIMS and WHIMS, froth flotation, and their combinations, were employed to remove iron/titanium-bearing minerals selectively.
Direct magnetic separation demonstrated efficacy in impurity removal. DHIMS (at a roll speed of 30 rpm) yielded a concentrate containing 0.95% Fe2O3 + TiO2 with an impurity removal of 85%, although the recovery rate was about 40%. WHIMS (1.5 kg/h) achieved a 1.2% Fe2O3 + TiO2 grade with 70% Fe2O3 + TiO2 removal efficiency and a higher recovery rate of 65%. Conversely, flotation (pH 10, 2000 g/t collector concentration), while capable of ~75% Fe2O3 + TiO2 removal efficiency, resulted in a less favorable concentrate grade of 1.65% Fe2O3 + TiO2 due to the co-flotation of valuable alkali-rich minerals, attributed to Ca/Mg ion activation and collector interaction with Al sites.
Combined processes demonstrated significant advantages. Notably, applying DHIMS to a flotation concentrate produced the lowest overall Fe2O3 + TiO2 grade of 0.90% with a desirable alkaline content exceeding 13.8%, though at 41% recovery. The sequence of flotation followed by WHIMS provided an excellent balance, yielding a concentrate with 1.1% Fe2O3 + TiO2 (K2O + Na2O ~ 13.4%) and about 80% recovery. Furthermore, flotation after WHIMS pre-concentration also effectively reduced impurities, achieving a concentrate grade as low as 1.17% Fe2O3 + TiO2 under optimized conditions (10 min conditioning). These findings underscore the viability of transforming phonolite tailings into valuable industrial raw materials, with combined methods offering promising pathways to achieving optimal product quality and recovery.

Author Contributions

Conceptualization, methodology, formal analysis, data curation, supervision, writing—original draft preparation, writing—review and editing, project administration, funding acquisition, validation, S.O.; investigation, software, S.O. and S.U.; resources, S.U.; visualization, S.O., S.U. and S.Y. All authors have read and agreed to the published version of the manuscript.

Funding

This research was funded by Süleyman Demirel University with grant number SDU BAP #4714-YL1-16.

Data Availability Statement

Data available on request from the authors.

Acknowledgments

The authors thank Y.K. Kadioglu and S. Caran for their technical support during mineral characterization, and Fonolit Barida for providing the phonolite samples.

Conflicts of Interest

The authors declare no conflicts of interest.

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Figure 1. (a) A phonolite rock block; (b,c) the rock tailings formed at the sizing stage of the blocks (photos taken by S. Ulutas).
Figure 1. (a) A phonolite rock block; (b,c) the rock tailings formed at the sizing stage of the blocks (photos taken by S. Ulutas).
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Figure 2. Flowsheet of phonolite cutting/sizing plant.
Figure 2. Flowsheet of phonolite cutting/sizing plant.
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Figure 3. Microphotographs of the rock sample with crossed nicols (Amf: amphibole; Bi: biotite; C: glass; F: feldspar–plagioclase; Fp: feldspathoid; O: oligoclase; P: pyroxene).
Figure 3. Microphotographs of the rock sample with crossed nicols (Amf: amphibole; Bi: biotite; C: glass; F: feldspar–plagioclase; Fp: feldspathoid; O: oligoclase; P: pyroxene).
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Figure 4. XRD pattern of phonolite sample.
Figure 4. XRD pattern of phonolite sample.
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Figure 5. Roll-speed-dependent test results: Assessment of (a) recovery, (b) Fe2O3 + TiO2 grade of concentrate, (c) Fe2O3 + TiO2 removal efficiency.
Figure 5. Roll-speed-dependent test results: Assessment of (a) recovery, (b) Fe2O3 + TiO2 grade of concentrate, (c) Fe2O3 + TiO2 removal efficiency.
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Figure 6. Feed-rate-dependent test results: Assessment of (a) recovery, (b) Fe2O3 + TiO2 grade of concentrate, (c) Fe2O3 + TiO2 removal efficiency.
Figure 6. Feed-rate-dependent test results: Assessment of (a) recovery, (b) Fe2O3 + TiO2 grade of concentrate, (c) Fe2O3 + TiO2 removal efficiency.
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Figure 7. Collector-concentration-dependent test results: Assessment of (a) recovery, (b) Fe2O3 + TiO2 grade of concentrate, (c) Fe2O3 + TiO2 removal efficiency.
Figure 7. Collector-concentration-dependent test results: Assessment of (a) recovery, (b) Fe2O3 + TiO2 grade of concentrate, (c) Fe2O3 + TiO2 removal efficiency.
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Figure 8. Effect of DHIMS application on flotation concentrates with different collector concentrations and pH levels: evaluation of (a) concentrate’s Fe2O3 + TiO2 grade and (b) Fe2O3 + TiO2 removal efficiency.
Figure 8. Effect of DHIMS application on flotation concentrates with different collector concentrations and pH levels: evaluation of (a) concentrate’s Fe2O3 + TiO2 grade and (b) Fe2O3 + TiO2 removal efficiency.
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Figure 9. Effect of HIWMS application on flotation concentrates with different collector concentrations and pH levels: evaluation of (a) concentrate’s Fe2O3 + TiO2 grade and (b) Fe2O3 + TiO2 removal efficiency.
Figure 9. Effect of HIWMS application on flotation concentrates with different collector concentrations and pH levels: evaluation of (a) concentrate’s Fe2O3 + TiO2 grade and (b) Fe2O3 + TiO2 removal efficiency.
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Figure 10. Effect of collector-concentration-dependent flotation tests on WHIMS concentrates: evaluation of (a) Fe2O3 + TiO2 grade of concentrate and (b) Fe2O3 + TiO2 removal efficiency.
Figure 10. Effect of collector-concentration-dependent flotation tests on WHIMS concentrates: evaluation of (a) Fe2O3 + TiO2 grade of concentrate and (b) Fe2O3 + TiO2 removal efficiency.
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Figure 11. Effect of pulp-density-dependent flotation tests on WHIMS concentrates: evaluation of (a) Fe2O3 + TiO2 grade of concentrate and (b) Fe2O3 + TiO2 removal efficiency.
Figure 11. Effect of pulp-density-dependent flotation tests on WHIMS concentrates: evaluation of (a) Fe2O3 + TiO2 grade of concentrate and (b) Fe2O3 + TiO2 removal efficiency.
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Figure 12. Effect of conditioning-time-dependent flotation tests on WHIMS concentrates: evaluation of (a) Fe2O3 + TiO2 grade of concentrate and (b) Fe2O3 + TiO2 removal efficiency.
Figure 12. Effect of conditioning-time-dependent flotation tests on WHIMS concentrates: evaluation of (a) Fe2O3 + TiO2 grade of concentrate and (b) Fe2O3 + TiO2 removal efficiency.
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Table 1. Phonolite sample chemical composition.
Table 1. Phonolite sample chemical composition.
Contentwt%Contentwt%
Na2O5.88CaO + MgO5.50
K2O5.63Fe2O3 + TiO26.99
SiO254.27MnO0.11
Al2O317.65LOI2.57
The Loss on Ignition (LOI) value reported in the XRF results primarily accounts for the mass loss due to the volatilization of structurally bound water from hydrous minerals present in the rock sample, such as clay minerals and micas.
Table 2. Variables used in flotation tests.
Table 2. Variables used in flotation tests.
Process VariablesValues and Units
Pulp density
(Solid % by Wt.)		Conditioning: 35%
Flotation: 15%
Impeller speedConditioning: 1750 rpm
Flotation: 1500 rpm
pH9–10.50
Conditioning time5 min
Flotation time1 min
Collector concentration1000–2000 g/t
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Ozun, S.; Ulutas, S.; Yurdakul, S. Recovery of High-Alkali-Grade Feldspar Substitute from Phonolite Tailings. Processes 2025, 13, 2334. https://doi.org/10.3390/pr13082334

AMA Style

Ozun S, Ulutas S, Yurdakul S. Recovery of High-Alkali-Grade Feldspar Substitute from Phonolite Tailings. Processes. 2025; 13(8):2334. https://doi.org/10.3390/pr13082334

Chicago/Turabian Style

Ozun, Savas, Semsettin Ulutas, and Sema Yurdakul. 2025. "Recovery of High-Alkali-Grade Feldspar Substitute from Phonolite Tailings" Processes 13, no. 8: 2334. https://doi.org/10.3390/pr13082334

APA Style

Ozun, S., Ulutas, S., & Yurdakul, S. (2025). Recovery of High-Alkali-Grade Feldspar Substitute from Phonolite Tailings. Processes, 13(8), 2334. https://doi.org/10.3390/pr13082334

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