Stage-wise flotation for the removal of colored ...

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ARTICLE IN PRESS Separation and Purification Technology xxx (2010) xxx–xxx

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Stage-wise flotation for the removal of colored minerals from feldspathic slimes using laboratory scale Jameson cell Cengiz Karagüzel ∗ , Güls¸ah Çobano˘glu Dumlupınar University, Faculty of Engineering, Department of Mining Engineering, Kütahya 43270, Turkey

a r t i c l e

i n f o

Article history: Received 9 October 2009 Received in revised form 11 May 2010 Accepted 16 May 2010 Keywords: Fine particle processing Slime Feldspar Flotation

a b s t r a c t Fifteen percent of the total number of ores to be supplied to feldspar flotation facilities is generally comprised of fine particles under 38 ␮m in size. Upgrading of these particles by current commercial beneficiation methods poses an environmental threat besides causing economic losses. In this study, an attempt was made to remove gang minerals containing Fe and Ti from a feldspathic slime sample (−38 ␮m) using laboratory scale Jameson flotation cell in the presence of both anionic BD-15 and cationic G-TAP collectors. The finely sized ore was successfully upgraded to produce Na-feldspar suitable for ceramics industry. A Na-feldspar concentrate assaying 0.18% Fe2 O3 + TiO2 from a slime sample containing 1.06% Fe2 O3 + TiO2 at 50% recovery was obtained. © 2010 Elsevier B.V. All rights reserved.

1. Introduction Feldspars are actually a group of minerals that are aluminasilicates of the alkali metals of potassium, sodium and calcium. These feldspars are rarely found in a pure form and comprise of all the three. The three extreme compositions of feldspars are represented by the minerals orthoclase (KAlSi3 O8 ), albite (NaAlSi3 O8 ) and anorthite (CaAl2 Si2 O8 ). Granite, pegmatites, nepheline syenite, aplites and feldspathic sands constitute the feldspar reserves of the world. The major impurities in feldspar are mica, quartz and other color imparting titanious and ferruginous impurities. Titanious and ferruginous minerals give color to baked ceramic and porcelain materials. For that reason, they should be removed from feldspars [1–8]. The assays in commercially acceptable feldspars are typically 11–13% Na2 O + K2 O, less than 1.5% CaO + MgO, 0.07–0.3% Fe2 O3 + TiO2 and free quartz up to 8–10% [9]. To this end, feldspars are mostly beneficiated by means of magnetic separation, flotation or combination of both methods [5,10,11]. However, the fines are disposed because they cannot be efficiently beneficiated by magnetic separation and they adversely affect flotation recoveries in the form of slime coating [11–14]. Flotation is the most popular method for removing these impurities in the size range of −250 to +38 ␮m. Very fine size particles known as slimes in flotation (for feldspar usually −38 ␮m size) represent large surface areas and result in

∗ Corresponding author. Tel.: +90 274 2652031x4169; fax: +90 274 2652066. E-mail address: [email protected] (C. Karagüzel).

excess collector consumption. Slimes are usually discarded before flotation as they adversely affect flotation by covering valuable mineral surfaces through attractive forces [12,15–17]. Recovery of these minerals from slime which includes a significant amount of valuable minerals will provide economic and environmental gains. Beneficiation of fine particles is technically difficult by physical methods. For that reason, surface based physicochemical methods as flotation, selective flocculation etc. are used. Flotation efficiency of fine particles due to low mass and impact moments and high interfacial energy is very low [18–25]. Bubble-particle collision probability is directly proportional to particle size while it is inversely proportional to bubble diameter. Average bubble size in conventional flotation machines is between 1000 and 2000 ␮m and flotation efficiency is also low due to low collision probability of fine mineral particles [18,19,26–28]. This situation is caused by the inability of bubbles to break the surrounding water layer owing to inadequate kinetic energy of fine particles. It is thus preferred to use flotation systems consisting of small-sized bubbles or systems to increase the kinetic energy of particles. Flotation machines designed to increase bubble-particle collision have recently been introduced in industrial scale. Jameson flotation cell is one of such devices. As of 1990, Jameson cell was used for rougher, scavenger and cleaning flotation of numerous minerals in various sizes in an industrial scale. Today, Jameson cell is used for the beneficiation of Pb/Zn and coal slimes in various mineral processing facilities particularly in Australia, America, India and South Africa [29]. While there are several literature studies regarding its application to industrial mineral samples such as pure quartz [30,31], barite [32] and calcite [32], there are no studies regarding the selective flotation of industrial materials.

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Please cite this article in press as: C. Karagüzel, G. Çobano˘glu, Stage-wise flotation for the removal of colored minerals from feldspathic slimes using laboratory scale Jameson cell, Separ. Purif. Technol. (2010), doi:10.1016/j.seppur.2010.05.012

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ARTICLE IN PRESS C. Karagüzel, G. Çobano˘glu / Separation and Purification Technology xxx (2010) xxx–xxx

Table 1 Chemical analysis of slime sample. Component

Al2 O3

SiO2

Na2 O

MgO

CaO

K2 O

TiO2

Fe2 O3

LF

Oxide (%)

19.75

66.38

9.58

1.02

0.94

0.78

0.35

0.71

0.49

Jameson cell generates 400–600 ␮m bubbles in size [33]. Since lack of collision is the biggest problem in fine particles, flotation efficiency of particles is higher compared to other methods, particularly mechanical cells [34,35]. Studies carried out on Jameson cell on the effect of particle size showed that fine particles were able to float at higher efficiencies compared to large particles [30,31]. In this study, flotation of albite to remove colored impurities from feldspathic slime was investigated in a Jameson flotation cell. The aim was to obtain concentrates of different characteristics by stage-wise operation of Jameson flotation cell to satisfy ceramics industry. 2. Experimental materials and methods Slime sample was obtained from the flotation facility of Kaltun Mining Inc. operating in the Aegean Region of Turkey where large reserves of feldspar ore exist. The mine meets its need of feldspar ore from Menderes massive located in the same region. The main colored impurities in these ores are iron and titanium bearing minerals such as anatase, rutile, sphene, biotite and ilmenite [1]. The slime sample used in this study consisted of albite, quartz and microcline and colored minerals. Feldspathic slime sample was physically and chemically analyzed by means of a Rigaku Miniflex powder X-ray diffractometer (XRD), a Spectra X-LAB 200 X-ray fluorescence (XRF) and a Fritsch brand particle sizing device. Analysis results are respectively given in Table 1 and Figs. 1 and 2. Characterization studies revealed that the complete feldspathic slime sample was under 38 ␮m in size (Fig. 2) and contained 1.06% Fe2 O3 + TiO2 and 9.58% Na2 O (Table 1). XRD studies also showed that albite was the major mineral while quartz and mica were minor minerals (Fig. 1). Microscopic studies further identified muscovite and biotite type mica minerals. Minerals such as sphene, anatase, ilmenite, hematite, titanious and ferruginous representing typical of Menderes massive remained below the analysis limit of XRD. In flotation experiments, anionic type commercial BD-15 reagent, which consists of 48% oleic acid, 36% linoleic acid, max. 3% palmitic, stearic acids and max. 3% linoleic acid and a cationic type reagent G-TAP (fatty alkyl propylene diamine) were used as the collectors. While sodium hexametaphosphate was used as dispersant, H2 SO4 and NaOH was used as pH modifiers. Aerofroth 65 (AF-65), where its successful use in Jameson cell is reported was used as Frother [36].

Fig. 1. XRD pattern of slime. A, Albite; M, Mica; Q, Quartz; Fe-Min, colored minerals.

The zeta potential measurements and microflotation studies were carried out using concentrated feldspar and colored minerals. Colored minerals (17.7% Fe2 O3 , 1.67% TiO2 and 0.57% Na2 O), a group of altered granite sample, 1.0 mm × 0.3 mm in size was a product of Kaltun Co.’s Permroll type high-intensity magnetic separator. The concentrated magnetic products were then ground in the ring mill (−38 ␮m) for the flotation experiments. The concentrated feldspar sample (0.067% Fe2 O3 , 0.061% TiO2 and 13.24% Na2 O), obtained from flotation plant of Kaltun Co. was cleaned with alcohol in order to remove the collector residues from the surfaces. The sample was the rinsed with pure water. Finally, the sample was ground in agate mortar in order to prepare the sample for flotation experiments (−38 ␮m). The zeta potential measurements were carried out using Zetasizer Nano-ZS (Malvern Inst, UK) using solid–liquid suspensions. The suspension used for the measurements was prepared by placing 0.1 g material into 100 mL bottle and mixing it for 10 min using a shaker. Then, the suspension was kept for 2 min in order to allow the coarse particles settle down. Finally, 2 mL of suspension was taken and added to the measurement cell, and the zeta potential measurements were carried out. Microflotation tests were carried out in a 150-mL column cell (25 mm × 220 mm) with a 15 ␮m frit and magnetic stirrer. The sample of 1 g was conditioned in 150 mL of solution containing the desired collector for 10 min and then floated for 1 min with nitrogen gas at a flow rate of 50 cm3 min−1 . The float and unfloat fractions were dried and weighed to calculate the percent floated. Flotation experiments were carried out in Jameson flotation cell which is schematically demonstrated in Fig. 3. As seen in Fig. 3, Jameson cell consists of two main parts, which are the downcomer and separation tank. Whereas the downcomer (2 cm in diameter, 180 cm in length) is the part where flotation micro-events take place and a rapid collection is enabled, separation tank (20 cm in diameter, 90 cm in length) is the part encapsulating the froth zone and enabling the separation of hydrophobic particle-bubble aggregates from hydrophilic particles. Pulp feeding, waste discharge and air flow rate in the process can be measured with digital flowmeters as well.

Fig. 2. Particle size distribution of the slime.



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Fig. 3. Schematic view of experimental set-up.

Experimental studies were carried out by means of single-stage, stage-wise method and the combination of both methods. Conditioning prior to flotation was performed in a 100-L conditioning tank for 10 min in the presence of reagents. Flotation pulp was fed into Jameson flotation cell at a pressure of 140–170 kPa. As flotation efficiency was high in pilot and industrial trials of Jameson cell, the percent solids (PS) by wt. was kept in the range of 2.5–5% and used at 2.5% thereafter. Other fixed and variable operation parameters are demonstrated in Table 2. While colored minerals (titanious and ferruginous impurities) were floated by reverse flotation techniques using BD-15 collector, feldspar was floated by the conventional flotation techniques with cationic G-TAP in single-stage experiments. Titanious and ferruginous impurities were floated using BD-15 and feldspar was floated using G-TAP. Thereafter, products comprising of tailings and concentrate were obtained following the single-stage flotation process carried out under operating conditions demonstrated in Table 2. Middling products were obtained by gradually increasing the concentration of anionic and cationic reagents in stage flotation experiments. All experiments were carried out using sodium hegza meta phosphate as dispersant. The parts floated for 2 min in each stage of the flotation process were collected as concentrate while unfloated part of each stage were fed to the conditioning tank for 10 min reconditioning by closing the tailing valve shown in Fig. 3. At the end of the experiment, the last unfloated portion was taken as tailings (Fig. 4). Thus four concentrates having different mineral contents and one tailing were collected. In other words, while the stage-wise system performs well for scavenging in the presence of G-TAP, it does well for cleaning in the presence of BD-15. The final experiments of the present study were carried out using the flowsheet in Fig. 4, which is the combination of singlestage and stage-wise flotation techniques and utilizes rougher and final flotation stages. The technique used in the rougher flotation

Fig. 4. Flowsheet including rougher and stage-wise flotation circuits.

was the single-stage method while the stage-wise flotation was the technique for the final flotation stage. As seen in Fig. 4, a rougher concentrate was fed to stage-wise flotation. Five products were obtained in the final flotation stage following the removal of colored minerals by anionic reagent at pH 9 in rougher flotation stage. In the final flotation stage, significant operating parameters for fine particle flotation such as superficial gas velocity, feed velocity, and bias factor were studied in detail. Flotation results were presented as tables and figures. In addition, mineralogical analysis of the product obtained from the flotation experiments were carried out using the OPTECH VF10X model microscope. The decantation method was also applied to determine the effect of very fine particle removal on flotation. A sample of 500 g was put into a 10-L measuring cylinder of 30 cm in height and mixed with water. Then, samples were taken from the different heights of the cylinder (25–22.5–20–15 cm) by siphoning after 2 min of settling. The remaining settled sample was utilized in the single-stage Jameson flotation experiment. Color tests on baked samples were carried out in Kütahya Ceramic Inc. For the color tests, the flotation concentrates or slime samples obtained from Jameson cell were pressed to a circular shape of 1.5 cm diameter by 0.5 cm height using 5 g samples and with the help of a mould. The samples were then baked for 2 h at 1250 ◦ C within a period of 2 h using a Protherm furnace. Color determination of baked products was performed with a Minolta Color Measurement Device which provides color values for L, a and b; in the triaxial color scale, L shows the whiteness; + value of “a” shows red and − value of “a” shows green, while + value of “b” shows yellow and − value shows blue. The baked colors of slimes produced out of flotation products were found to be 65.16, −0.15 and +16.77 for L, a and b, respectively. 3. Results and discussions

Table 2 Constant and variable parameters applied in flotation experiments.

3.1. Single-stage flotation experiments

Operating parameters Particle size (dp ) pH Frother (AF 65) amount (Qfr ) Na-Hegza meta phosphate amount (Qd ) BD-15 amount (Qc ) G-TAP amount (Qc ) Conditioning time (t) Flotation time (t) Superficial gas velocity (Jg ) Bias factor (BF) Jet velocity (Jh ) Wash water flow (Qww )

−38 ␮m 8.5–9 (BD-15 usage) 2.5–3 (G-TAP usage) 20 ppm 100–1500 g t−1 (for single-stage flot) 100–2000 g t−1 (for single-stage flot) 300–2000 g t−1 (for stage-wise flot) 100–1000 g t−1 (for single-stage flot) 100–1000 g t−1 (for stage-wise flot) 10 min (for each stage) 2 min (for each stage) 0.48, 0.65, 0.79 cm s−1 0.3, 0.5, 0.8 10.6, 13.3, 15.9 m s−1 1.2 L min−1

It is known that feldspars are hydrophilic due to the hydrated layer on their surface at basic pH levels. While colored minerals typically oxides are floatable by anionic reagents, silicate minerals, particularly feldspars, are floated by cationic reagents [11,12,37–39]. G-TAP which is used as cationic collector in this study is reported to be used efficiently in the flotation of feldspar minerals [40]. However, anionic BD-15 type collector has been used in the flotation of colored minerals in Kaltun Company’s Mineral Processing Plant. Microflotation studies and electrokinetic measurements were carried out to find out the effect of these collectors on flotation recovery and selectivity. According to electrokinetic measurements the iep of albite and colored minerals are 2.5 and 5.5, respectively (Fig. 5). Fig. 5 illustrates that the recovery of the colored minerals in the




Fig. 7. Effect of decantation amount on Fe2 O3 + TiO2 % grade of feldspar concentrate. Fig. 5. Effects of pH on zeta potential and microflotation recovery.

microflotation cell using BD-15 flotation is 78% under alkali conditions (pH 9–9.5), while G-TAP flotation recovery for albite is 85% in acidic conditions (pH 2.5–3). The experiments carried out using G-TAP reveals that the increase in pH leads to a dramatic decrease in recoveries due to the collector precipitation, while flotation performed with BD-15 exhibits slight changes in recovery. Aggregation of fine particles in aqueous suspensions is governed by attractive forces and will inevitably affect the success of selective separation. Dispersants are generally used to prevent aggregation in fine particle flotation. In this study, 500 g t−1 sodium hexametaphosphate was used as a dispersant to provide the dispersion of clay minerals in the pulp. Surface charge of colored minerals was changed to negative sign because of the phosphate ions, while surface charge of silicate minerals became more negative. Therefore, dispersion condition occurred due to the prevention of hetero- and homo-coagulation owing to electrostatic repulsion. Because of this, selectivity was supplied resulting from the stable flotation condition. Fig. 6 shows the effect of dispersant dosage on the selectivity of flotation. Decantation was also used to remove the ultrafine por-

tion of the slimes. Fig. 7 shows that decantation of ultrafine particle on flotation recoveries was marginal. In this study, both anionic and cationic reagents were tested on a slime rich feldspar sample and the flotation results in terms of feldspar grades and the sum of colored minerals are presented as a function of reagent dosages in Figs. 8 and 9. As seen in Fig. 8, at pH 9 and 1000 g t−1 BD-15 addition, the Fe2 O3 + TiO2 value decreased to 0.49% in the slime sample with a total alkali content (Na2 O + K2 O) of 10.93%. Moreover, the yield of Na-feldspar was found to be 67% at 1000 g t−1 BD-15 dosage. The Fe2 O3 + TiO2 content in the slime sample could not be reduced below 0.3% with the anionic collector due to the existence of mica and particularly biotite type mica. However, Aksay et al. [11] reported that Fe and Ti minerals were removed from a feldspar ore in the presence of less number of mica minerals at pH 9 in a singlestage by means of fatty acid based anionic collector (Aero 704). Fig. 9 indicates that the total alkali content (Na2 O + K2 O) increased by increasing G-TAP dosage. In addition, Fe2 O3 + TiO2 content is relatively high at the presence of low G-TAP concentration due to the natural hydrophobicity of biotite. Rath and Subramanian [41]

Fig. 6. Effect of dispersant dosage (Na-hegza meta phosphate) on Fe2 O3 + TiO2 % grade of feldspar concentrate.

Fig. 8. Effect of collector dosage (BD-15) on Fe2 O3 + TiO2 % and Na2 O + K2 O% grades of feldspar concentrate (Jg = 0.79 cm s−1 ; BF = 0.8; Qf = 9.8 L min−1 ; SR = 2.5%, Qfr = 20 ppm).


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Fig. 9. Effect of collector dosage (G-TAP) on Fe2 O3 + TiO2 % and Na2 O + K2 O% grades of feldspar concentrate (Jg = 0.79 cm s−1 ; BF = 0.8; Qf = 9.8 L min−1 ; SR = 2.5%, Qfr = 20 ppm).

pointed out to a flotation study carried out with the Hallimond tube where biotite mica floated at 85–90% efficiency without the use of any kind of collector in the presence of 10−2 M KNO3. In this study, a decrease in Fe2 O3 + TiO2 content with increasing G-TAP concentration was observed. A concentrate with 0.49% Fe2 O3 + TiO2 at 500 g t−1 collector was obtained at 48% Na2 O recovery. However, it was noted that a number of coloring minerals excluding biotite could not be collected by means of cationic reagent, and thus it was deemed unnecessary to carry out the color analysis of samples since there was no adequate whiteness in compressed and baked samples.

5

BD-15 quantity was increased from 300 to 2000 g t−1 while G-TAP quantity was increased from 100 to 1000 g t−1 . Results of the stagewise flotation experiments in which both reagents were tested are separately presented in Tables 3 and 4. Table 3 indicated that colored minerals were removed in every stage in which BD-15 was increased. But the increase in the number of stages and reactive dosages can decrease feldspar recoveries. The unfloated product with BD-15 as collector assayed 0.48% Fe2 O3 + TiO2 and 11.01% Na2 O + K2 O. Contrary to this, based on microscopic observations, little amount of colored minerals, i.e. biotite type of mica having Fe-silicate was removed in the first stage of G-TAP flotation (Table 4). In other stages, combining the second, third and fourth stages, sufficient whiteness could not be achieved according to firing tests. Thus studies showed that application of an improved tree technique developed by Karaguzel et al. [12] can enhance the results. The suggested technique is shown in Fig. 4. Fig. 4 presents the removal of colored minerals in the rougher flotation utilizing BD-15 collector, removal of biotite in the first stage of G-TAP flotation carried out after BD-15 flotation and flotation of feldspar in the later stages. Table 5 indicates the stage-wise flotation results performed after the rougher flotation. The results of stage-wise flotation test are shown in Table 5. When the sample produced in the anionic rougher flotation stage shown in Fig. 8 was transferred to the final flotation consisting of stage-wise flotation process given in Table 5, different products were obtained in different stages. It is evident that biotites were separated in the presence of low dosage (100 g t−1 ) cationic reagent in the first stage. In this stage, Fe2 O3 content was determined as 0.24%. When the products obtained in the intermediate stages were combined, the Fe2 O3 + TiO2 content and Na recovery were determined as 0.31 and 64%, respectively. These products can be utilized either by combined or separate concentrates with respect to their technological field of use. Particularly the product obtained in the second stage can be of ceramic quality depending on its 0.11% Fe2 O3 content and L: 78.57, a: −0.97 and b: +9.8 color values.

3.2. Stage-wise flotation experiment Owing to the fact that colored minerals containing iron and titanium had different physicochemical characteristics, the stage-wise flotation system proposed by Burat et al. [1] was adapted for the Jameson cell study. In the stage-wise flotation, the reagent amount was increased proportional in each stage. The tailings obtained in each stage was re-fed and conditioned for subsequent stages.

3.2.1. The effect of operation parameter on selectivity 3.2.1.1. Effect of superficial gas velocity (Jg ) on selectivity. One of the most critical parameters affecting recovery and selectivity in Jameson cell is the superficial gas velocity as it controls the bubble diameter, gas hold-up and jet length in Jameson cell. Increasing the air flow rate within the stable flow rate results in bubble diameter and gas hold-up increase [42]. The bubble diameter

Table 3 Effect of BD-15 quantity on selectivity of stage-wise flotation. Parameter −1

BD-15 (g t

Fe2 O3 + TiO2 )

300 500 1000 2000 Feed

Na2 O + K2 O

Product

Weight (%)

G (%)

R (%)

G (%)

R (%)

Stage I float Stage II float Stage III float Stage IV float Sink Total

21.57 11.76 19.95 14.24 32.48 100.00

1.05 0.62 0.58 0.51 0.48 1.06

21.35 6.85 11.00 6.87 14.76 100.00

9.81 10.24 10.81 10.94 11.01 10.36

20.42 11.63 20.82 15.04 34.52 100.00

Table 4 Effect of G-TAP quantity on selectivity of stage-wise flotation. Parameter −1

G-TAP (g t 100 250 500 1000 Total

Fe2 O3 + TiO2 )

Na2 O + K2 O

Product

Weight (%)

G (%)

R (%)

G (%)

R (%)

Stage I float Stage II float Stage III float Stage IV float Sink

5.67 6.21 32.31 27.17 28.64 100.00

0.69 0.48 0.37 0.61 2.03 1.06

3.68 2.80 62.10 15.58 55.04 100.00

10.76 11.09 11.13 11.00 9.13 10.36

5.88 6.64 34.71 28.85 25.24 100.00


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Table 5 Application of stage-wise flotation using G-TAP in rougher flotation product at 1000 g t−1 BD-15. Parameter −1

G-TAP (g t 100 250 500 1000 Feed

Fe + Ti )

Na + K

Product

Weight (%)

G (%)

R (%)

G (%)

R (%)

Stage I float Stage II float Stage III float Stage IV float Sink Total

9.15 20.37 32.34 16.44 22.76 100.00

0.46 0.23 0.32 0.39 0.88 0.49

8.43 9.48 21.15 12.98 40.54 100.00

11.34 11.00 11.31 10.70 8.93 10.93

9.49 20.50 33.47 16.09 18.60 100.00

is particularly significant for fine particle flotation. At the same time, fine air bubbles are required for inducing the attachment of hydrophobic particles to the air bubbles. Air bubbles at a diameter of 400–600 ␮m generated in Jameson cell provide a suitable environment for the flotation of particles minus 100 ␮m in size [30,33]. Superficial gas velocity (Jg ) is the upward superficial velocity of air within the flotation cell; Jg is determined by dividing air flow rate let in the column (cm3 s−1 ) by the cell cross-sectional area (cm2 ). Mohanty and Honaker [43] developed a relationship between the production rate (velocity) and superficial gas velocity. In general, low values of Jg (0.4–0.8 cm s−1 ) are used for cleaning application, while high values (1–2 cm s−1 ) are used for scavenging and roughing flotation [42]. Particle size is also important for choosing the Jg as it affects froth stability. And low values of Jg (0.4–1.0 cm s−1 ) are effective for particles finer than 100 ␮m [42]. In order to determine the optimum superficial gas velocity, experiments at superficial gas velocities of 0.48, 0.65, and 0.79 cm s−1 were carried out as shown in Fig. 10. Examination of Fig. 10 reveals that Fe2 O3 + TiO2 and Na2 O + K2 grade are respectively 0.31 and 11.01% at superficial air velocity of 0.65 cm s−1 . When the superficial gas velocity is lower than 0.65 cm s−1 , adequate selectivity cannot be obtained owing to the failure to create sufficient kinetic energy of particles within the downcomer mixing zone. Excess increase in jet length within the downcomer due to the excess increase of jet velocity results in a decrease in the area where micro-events take place. An increase in Fe2 O3 + TiO2 was observed due to the fact that probability of particle-bubble encounter decreased due to the decrease in gas hold-up resulting from bubble coalescence [44]. Another reason for affecting selectivity and increased superficial gas velocity is the formation of fluctuation created by the turbulence within the separation and froth zones. Sahbaz et al. [33] reported on flotation

Fig. 10. Effect of superficial gas velocity on Fe2 O3 + TiO2 % and Na2 O + K2 O% grades of feldspar concentrate (BF = 0.8; Qf = 9.8 L min−1 ; SR = 2.5%, Qfr = 20 ppm).

Fig. 11. Effect of jet velocity on Fe2 O3 + TiO2 % and Na2 O + K2 O% grades of feldspar concentrate (Jg = 0.65 cm s−1 ; BF = 0.8; SR = 2.5%, Qfr = 20 ppm).

of unburned carbonaceous matter (UCM) 90% of which was under 100 ␮m; the selectivity decreased due to Jg values of more than 0.9 cm s−1 , resulting in bubble coalescence and particularly fluctuation created in the froth zone. While Mohanty and Honaker [43] pointed out that 1 cm s−1 Jg superficial gas velocity in coal flotation is the optimum value. 3.2.1.2. Effect of feed flow rate (jet velocity) on selectivity. Jet velocity (Jv ) is the feed flow rate into downcomer and is formulated as

Fig. 12. Effect of bias factor on Fe2 O3 + TiO2 % and Na2 O + K2 O% grades of feldspar concentrate (Jg = 0.65 cm s−1 ; Jv = 10.6 m s−1 ; SR = 2.5%, Qfr = 20 ppm).



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Fig. 13. Pictures of slime and baked products with secondary stage-wise and combined concentrate XRD patterns.

follows: Jv =

Qf An

where Qf is the feed flow rate and An represents the nozzle cross-section area. In this study, the nozzle cross-section area was measured at 12.56 mm2 and the jet velocity is studied by means of changing the feed flow rate. Experimental results regarding the jet velocity are given in Fig. 11. An increase in jet velocity causes gas hold-up to increase as well. Gas hold-up is closely related to the probability of bubble-particle encounter. This relationship was formulated as Pk = 1.209ε3/2 by Oteyaka [26]. While Pk represents the encounter probability, ε represents gas hold-up in this formulation. The gas hold-up rate is between 10 and 20% in normal cells whereas it is between 40 and 70% in the Jameson cell downcomer and this situation leads to an increase in the encounter probability. As seen in Fig. 11, Fe2 O3 + TiO2 content is 0.26% at low feed velocity and increases to 0.46% at a feed velocity of 16 m s−1 . The cause for the change in selectivity at increased feed velocity was attributed to the fluctuation created in the froth zone of the separation tank. However, the cause for the same change within the downcomer was due to the inadequate quantities of air bubbles and insufficient retention time of particles in the downcomer. The relationship between jet velocity (cm s−1 ) and air entrainment rate (L m−1 ) for various jet lengths, nozzle diameters and downcomer diameters was discussed in coal flotation by Tas¸demir et al. [31] who stated that gas hold-up and air entrainment rate increased as the jet velocity increased. It was pointed out that an increase in the total combustible recovery obtained about 7.8% using the low jet velocity (11.3 m s−1 ) and normal operating range (Vj ; 18.0 cm s−1 ) [29]. It was observed in our study that there was no significant difference in Na recovery and that selectivity decreased due to the increased gas hold-up and air entrainment. 3.2.2. Effect of bias on selectivity Bias is the calculation of net water flow rate within the whole system. One of the variables that contribute to the attainment of optimum performance by the Jameson Cell is the bias factor, which is defined by the fraction of the wash water flowing downward and reporting to tailings stream. It is possible to determine the total bias

factor within the whole system by means of the following equation (bias is the factor determining the existence of froth zone). Bias factor =

QT − Qf QWW

where QT is the flow rate of tailings, Qf is flow rate of feed and QWW is the flow rate of wash water. In general, a positive bias factor is maintained to limit the entrainment of fine hydrophilic particles into the froth product. The effect of bias factor on selectivity is demonstrated in Fig. 12. As seen in Fig. 12, while the Fe2 O3 + TiO2 content of the concentrate decreased with increasing the bias, Na2 O grade was escalated. It is seen that selective flotation is realized by preventing entrainment of fine particles resulting from hydrodynamic forces by using the wash water. Mohanty and Honaker [43] pointed out in a study carried out using fine coal sample that positive bias over 0.6 and wash water flow rate at the value of 2.5 L min−1 are required in order to beneficiate clean coal from high ash coal. The same results were obtained in this study and Fe2 O3 + TiO2 content at 0.8 bias factor reached to 0.2%. Optimum results for superficial gas velocity (Jg ), jet velocity (Jv ) and bias factor are respectively 0.65 cm s−1 , 10.6 m s−1 and 0.8. When the quality of concentrates are analyzed separately and in combination, baking color of the second stage product was L: 79.12, a: −0.91 and b: +13.02. While the products beneficiated in the secondary stages can be utilized as the first quality product, combined products are utilized as the second quality products. These samples can be used in ceramics sector in the range of L: 70–80, a: ±1, and b: +10 and +20. Fig. 13 presents the pictures of secondary stage-wise and combined concentrate XRD patterns and baked products of the same materials. 4. Conclusions Increasing demand for raw materials in industry increases the slime-sized waste. These wastes cannot be utilized using the available methods such as classical flotation owing to technological difficulties. Jameson cell was used as an alternative method for the utilization of feldspathic slime sample for the first time. The conclusions obtained from this study are as follows:




(i) It was observed that beneficiation of the slime sample containing Fe and Ti minerals in a single stage was not sufficient and the flotation flowsheet consisting of single stage-wise rougher and stage-wise (final) flotation with cationic and anionic reagents were recommended. (ii) Various products were obtained in the final flotation stage and a product containing 0.11% Fe2 O3 and 0.07% TiO2 with 70% recovery was obtained as a result of combination of intermediate stages. (iii) The color tests show that the product representing L: 76.15, a: −1.01 and b: +11.15 values are in line with the standards. (iv) The operation parameters (superficial gas velocity, jet velocity, bias) were systematically studied in order to increase selectivity in Jameson cell. The optimum superficial gas velocity, jet velocity and bias were determined to be 0.65 cm s−1 , 10.6 m s−1 and 0.8, respectively. (v) The kinetic results obtained from the stage-wise flotation system employed in this study are extremely significant for cell selection and modeling in an industrial scale. Acknowledgements The authors are thankful to Dr. B. Öteyaka, Dr. M.S. Çelik and Dr. O. S¸ahbaz for their scientific support. The authors also would like to acknowledge helpful suggestions provided by Dr. G. Jameson during the design stage of the Jameson cell. In addition the authors would like to acknowledge to Kutahya Ceramic Inc. for color tests and Kaltun Mining Inc. for supplying sample. References [1] F. Burat, O. Kokkilic, O. Kangal, V. Gurkan, M.S. Celik, Miner. Metall. Process. 24 (2004) 75–80. [2] I. Bayraktar, S. Ersayin, O.Y. Gulsoy, Proceedings of the VII. International Mineral Processing Symposium—Balkema, Istanbul, Turkey, 1998, pp. 315–318. [3] I. Bayraktar, S. Ersayin, O.Y. Gulsoy, Z. Ekmekci, M. Can, Proceedings of 3rd. Industrial Raw Material Symposium, Izmir, Turkey, 1999, pp. 22–33. [4] C. Karaguzel, Investigation of useability of Simav region feldspar in ceramic industry, Dumlupinar University, M.S. Thesis, Kutahya, Turkey, 2000. [5] M.S. Celik, B. Pehlivanoglu, A. Aslanbas, R. Asmatulu, Miner. Metall. Process. 18 (2) (2001) 101–105. [6] www.epa.gov. [7] www.usgs.gov. [8] Turkish Standarts (TSE), Feldspar—Used for Ceramic Industry, 1987, Ankara, Turkey. [9] V.V Klyachin, Translated from Steklo i Keramika, 9 (1972), 30–31.

[10] I. Dogu, A.I. Arol, Powder Technol. 139 (2004) 258–263. [11] K.E. Aksay, A. Akar, E. Kaya, I. Cocen, Asian J. Chem. 21 (3) (2009) 2263–2269. [12] C Karaguzel, I. Gulgonul, C. Demir, M. Cinar, M.S. Celik, Int. J. Miner. Process. 81 (2006) 122–132. [13] I. Bayraktar, S. Ersayin, O.Y. Gulsoy, Miner. Eng. 1 (12) (1997) 1363–1374. [14] E.C. Orhan, I. Bayraktar, Miner. Eng. 19 (2006) 48–55. [15] I. Gulgonul, The mechanism of slime effect on the boron minerals flotation, Master Thesis, Istanbul Technical University, Istanbul, Turkey, 1995. [16] P. Bandini, C.A. Prestidge, J. Ralston, Miner. Eng. 14 (5) (2001) 487–497. [17] Z. Xu, J. Liu, J.W. Choung, Z. Zhou, Int. J. Miner. Process. 68 (2003) 183–196. [18] R.H. Yoon, G.H. Luterell, Miner. Process. Extractive Metall. Rev. 5 (1989) 101–122. [19] R. Sivomohan, Int. J. Miner. Process. 28 (1990) 247–288. [20] A.G. Lange, W.M. Skinner, R.St.C. Smart, Miner. Eng. 10 (7) (1997) 681–693. [21] Z. Dai, D. Fornasiero, J. Ralston, Adv. Colloid Interface Sci. 85 (2000) 231– 256. [22] T. Yalçın, A. Byers, K. Ughadpaga, Miner. Process. Extractive Metall. Rev. 23 (2002) 181–197. [23] J. Rubio, F. Capponi, E. Matiolo, D. Nunes, C.P. Guerrero, G. Berkowitz, Proceedings XXII International Mineral Processing Congress, Cape-Town, South Africa (2003) 1002–1014. [24] D. Tao, Sep. Sci. Technol. 39 (4) (2004) 741–760. [25] R.T. Rodrigues, J. Rubio, Int. J. Miner. Process. 82 (2007) 1–13. [26] B. Oteyaka, Modelisation D’une Colonne De Flottation Sans Zone D’ecume Pour La Separation Des Particules Grossieres, PhD Thesis, Universite Laval, Quebec, Canada, 1993. [27] A.V. Nguyen, H.J. Schulze, Colloidal Science of Flotation, Marcel Dekker, New York, 2004. [28] B.P. Binks, T.S. Horozov (Eds.), Colloidal Particles at Liquid Interfaces, Cambridge University Press, Cambridge, 2006. [29] http://www.jamesoncell.com/index.cfm?action=dsp content&contentID=19. [30] M. Cinar, O. Sahbaz, F. Cinar, S. Kelebek, B. Oteyaka, Miner. Eng. 20 (2007) 1391–1396. [31] A. Tasdemir, T. Tasdemir, B. Oteyaka, Miner. Eng. 20 (2007) 1331–1336. [32] E. Mozaffari, A study of coarse particle recovery by the froth flotation in the Jameson cell, PhD Thesis, University of Notingham, England, 1998. [33] O. Sahbaz, B. Oteyaka, S. Kelebek, A. Ucar, U. Demir, Sep. Purif. Technol. 62 (2008) 103–109. [34] G. Harbort, S. De Bono, D. Carr, V. Lawson, Miner. Eng. 16 (2003) 1091–1101. [35] M. Uçurum, Powder Technol. 191 (2009) 240–246. [36] O. S¸ahbaz, Removal of unburned carbon from bottom ashes of Tuncbilek Power Plant by the use of Jameson cell, Dumlupinar University Institute of Science, Master Thesis, Kutahya, Turkey, 2005. [37] M.S. El-Salmawy, Y. Nakahiro, T. Wakamatsu, Proceedings of the IV. International Mineral Processing Symposium, Antalya, Turkey, 1992. [38] M.S. El-Salmawy, Y. Nakahiro, T. Wakamatsu, Miner. Eng. 6 (12) (1993) 1231–1243. [39] I. Bayraktar, O.Y. Gulsoy, M. Can, E.C. Orhan, Proceedings of the 4th. Industrial Raw Material Symposium, Izmir, Turkey, 2001. [40] C. Demir, A.A. Abramov, M.S. Çelik, Miner. Eng. 14 (7) (2001) 733–740. [41] R.K. Rath, S. Subramanian, Miner. Eng. 10 (12) (1997) 1405–1420. [42] G.M. Evans, B. Atkinson, G.J. Jameson, The Jameson cell, in: K.A. Matis (Ed.), Flotation Science and Engineering, Marcel Dekker Inc., 1995. [43] M.K. Mohanty, R.Q. Honaker, Miner. Eng. 12 (4) (1999) 367–381. [44] G.J. Harbort, E.V. Manlapig, S.K. De Bono, Trans. IMM Section C 111 (2002) 1–10.
