Journal of Minerals & Materials Characterization & Engineering, Vol. 5, No.2, pp 103-117, 2006
jmmce.org Printed in the USA. All rights reserved
103
Enrichment of Molybdenum and Fluorite by Flotation of Fluorite Ore
Containing Molybdenum
Abuzer Akgün, İbrahim Teğin, Recep Ziyadanoğulları*
Department of Chemistry, Faculty of Science and Art, Dicle University,
21280 Diyarbakir, TURKEY
*For correspondence: Fax: [phone] Tel : [phone]/3159 E-mail: [e-mail]
ABSTRACT
The molybdenum and fluorite were obtained in different phases with the flotation of fluorite
ore containing molybdenum obtained from Karaamağaradere-Keban-Elazığ district in Turkey. It
has been found that the original ore contains 1.08 % Mo, 0.05 %Cu, 1.20 % Pb, 1.12 % Zn and
12.0 % F. It was determined that molybdenum and fluorite could not be enriched in different
phases with satisfactory yields by flotation of original ore at the defined size. Thus, the ore was
sulfurized before the flotation. Concentrates containing molybdenum and the other concentrates
containing fluorite were collected with fairly high yields in different phases by the flotation of the
sulfurized ore. In the optimum flotation conditions, the Mo, Pb, Cu and Zn were obtained over 95
% yields and approximately 6 % fluorite passed into concentrate phase. Mo and Cu containing
phase were enriched with high yield by selective flotation of concentrate obtained from the
concentrate. On the other hand, the most of Pb and Zn remained in tailing. In order to gain
molybdenum, the appropriate concentrate was obtained by this process with hydrometallurgical
method. It was also determined that the trace amounts of Ta, Nb and V in ore were collected in the
concentrate phase.
In the first step of flotation, the fluorite left in the tailing contained 29.6 % of CaF2. After
flotation of this tailing, concentrate grade of CaF2 raised to 96 %. 0.5 % of sulfur was found in
this concentrate. It was concluded that this concentrate contained fluorite with suitable particle
size (-100 mesh) and at sufficient concentration for metallurgical processes.
Key Words: Molybdenum, fluorite, sulfurization, flotation

104 Akgün, Teğin, and Ziyadanoğulları Vol.5, No.2
1. INTRODUCTION
Fluorite (CaF2) is an important fluorine mineral, which is mostly used for the production of
hydrofluoric acid and as a flux in steel making. Other uses are the manufacture of glass, fiberglass,
pottery and enamel, etc.
Recently, there have been numerous reports on the modification of the chemical scheme in
fluorite flotation, in order to increase the separation efficiency and lower the operation costs.
It was [1] reported that methanoic acid could selectively adsorb onto fluorite surfaces in
preference to calcite surfaces at ambient temperature. This means that fluorite flotation to remove
calcite might be achieved at a much lower temperature if methanoic acid is used as collector,
resulting in a large reduction in energy consumption. It was [2, 3] found that sodium N-dodecanoyl
sarcosine (anionic collector) in combination with dodecylammonium chloride (cationic collector)
as co-collectors could greatly enhance the floatability of fluorite, compared with the use of the
anionic collector alone. Also, fluorite flotation could be improved by using effective depressants
when fatty acid is used as collector. For example, acidized sodium silicate was reported to depress
calcite strongly in preference of fluorite at ambient temperature, leading the fluorite flotation to be
achieved at a lower temperature [4].
Molybdenum and fluorite are important materials for modern industry. Therefore,
production of fluorite and molybdenum from fluorite ore containing molybdenum are significant.
The most important minerals of molybdenum are molybdenite (MoS2), wulpherite (PbMoO4), and
molybdite (MoO3). However, for molybdenum production from ores containing molybdenum in
the first step, enrichment is required. Therefore, the flotation method was applied. Molybdenum
usually occurs as molybdenite in nature and enrichment of molybdenum is much easier in
comparison to other ores with high yield. Molybdenum has different constituent in minerals.
Therefore, it was determined that flotation yields are low. Moreover, recovery of concentrate of
molybdenum and fluorite separately reduce flotation yield [5].
Flotation tests on fluorite ores with various carbonate content (CaCO3 contents from 4.43%
to 74.40%) were carried out using acidized sodium silicate as modifier and oleic acid as collector.
A high purity fluorite concentrate (99% CaF2) was obtained. It was demonstrated that acidized
sodium silicate has an activating effect on fluorite and a selective depressing effect on calcite in
rougher flotation, at the same time effectively removing the floated calcite from the rougher
concentrate [6].
In this study, we aimed to collect molybdenum and fluorite in separate phases and to
increase flotation yield. Therefore, before the flotation, the sulfurization process was made in order

Vol.5, No.2 Enrichment of Molybdenum and Fluorite 105
to change the ore structure and surface properties [7, 8]. Sulfurized sample were subjected to
flotation and results were evaluated [9].
2. MATERIALS AND METHOD
2.1. Materials
The fluorite ore containing molybdenum used in this study was obtained from
Karamağaradere district of Keban-Elazığ, which is in eastern part of Turkey. About 50 kg of mine
sample was taken. Analyses showed that the sample contained 1.08 % Mo, 0.05 % Cu, 1.20 % Pb,
1.12 % Zn, 12.0 % F, 0.016 % Ta, 0.20 % Nb and 0.14 % V
98% (w/v) H2SO4, 37% (w/v) HCl, 65% (w/v) HNO3 and KClO3 were purchased from
Merck. The K-Amyl xanthate and Aeroflot 65 used in flotation were provided from Cyanamid Co.
in the USA.
A flame atomic absorption spectrometer (Unicam 929 Model AAS) was used for
determination of Mo, Cu, Pb, Zn, Ta, Nb and V concentrations in the solution. The fluor analysis
was carried out titrimetrically [10]. Denver Mark and 890 Model pH meter were used for flotation
and determination of pH of samples, respectively.
The sulfurization process was conducted in an autoclave of 1.3 liter internal volume,
resistant to 250 atm pressure and to 350 oC
2.2. Method
The mine sample was first crushed and then ground to -100 mesh size, sieved and dried at
110 oC.
For sulfurization, 1000 g of sample was sulfurized with gas containing different amounts
H2S + H2O in an autoclave for 1 hour at 120 oC. Then, samples obtained from autoclave were
floated with potassium amyl xanthate.
3. RESULTS AND DISCUSSION
3.1. Separation of Molybdenum from Fluorite Ore:
3.1.1. Flotation of Original Ore

106 Akgün, Teğin, and Ziyadanoğulları Vol.5, No.2
After sieved to -100 mesh size of the original ore, flotation process was conducted under
the conditions indicated blow.
Flotation condition:
Particle size : - 100 mesh
Solid / Liquid Rate : 100 g / L
Collector : 0. 2 g Z5 (potassium amyl xanthate), 3 minute mix
Frother : 0, 5 mL Aeroflot 65 (1%) 2 minute mix
pH : 6.5-9.0
Mix Speed : 900 rpm
At the end of flotation process, results aimed were not attained, and flotation yield was low.
Approximately 12 % of fluorite passed into the concentrate phase. Since an efficient separation
could not be carried out, the experimental results were not given in detail. Thus, original ore was
sulfurized in an autoclave before flotation. Then, these samples were floated.
3.1.2. Flotation of Sulfurized Samples
Firstly, samples of original fluorite ore containing molybdenum were ground, sieved to -
100 mesh size, dried at 110 oC and then reacted with gas mixtures containing different amounts of
H2S + H2O for 1 hour at 120 oC. For this purpose, the six samples each of which was 1000 g, were
reacted with mixtures of H2S and H2O separately as seen in Table 1.
Table 1. The used amount of gas mixtures of H2S and H2O.
Experimental No Amount of H2S (g) Amount of H2O (g)
1
2
3
4
5
6
7.66
11.49
12.44
13.40
15.32
19.44
100.0
60.0
65.0
70.0
80.0
100. 0

Vol.5, No.2 Enrichment of Molybdenum and Fluorite 107
After sulfurization, the samples were floated under conditions stated below.
Flotation condition:
Particle size : - 100 mesh Solid / Liquid Rate : 100 g / L
Collector : 0. 2 g Z5 (potassium amyl xanthate) 3 Minute mix
Frother : 0, 5 mL Aeroflot 65 (1%) 2 Minute mix
Mix Speed : 900 rpm
The sixth different sulfurized samples were floated under the same conditions. The results
are given in Table 2-7.
Table 2: The values obtained by flotation of the first sulfurized sample
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g)
Tailing
(g)
Mo Pb Cu Mo Pb Cu Mo Pb Cu
3.0
2.0
2.3
2.1
1.5
1.3
6.50
7.05
7.51
8.01
8.53
8.97
33.60
22.19
27.95
23.48
22.33
20.15
65.50
77.40
71.55
75.48
77.16
79.96
96.0
91.7
95.0
92.5
95.2
94.3
93.6
79.6
83.2
74.2
81.0
69.0
53.0
63.2
69.2
85.0
77.0
75.0
3.05
4.45
3.65
4.24
2.60
5.03
3.30
4.29
3.57
3.77
4.33
4.10
0.08
0.14
0.12
0.18
0.17
0.18
0.06
0.11
0.08
0.11
0.07
0.07
0.11
0.16
0.29
0.41
0.30
0.46
0.035
0.030
0.020
0.009
0.014
0.015
Table 3: The values obtained by flotation of the second sulfurized sample
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g) Tailing
(g)
Mo Pb Cu Mo Pb Cu Mo Pb Cu
1.50
1.50
1.15
1.04
1.10
1.10
6.54
6.97
7.54
8.00
8.52
9.02
29.76
19.83
26.59
20.25
20.83
18.49
69.56
79.88
72.93
79.30
78.38
81.34
94.9
91.2
92.3
95.4
91.7
92.2
88.6
68.0
78.0
78.0
69.4
52.0
67.4
74.0
68.5
75.0
65.6
66.0
3.42
4.95
3.72
5.06
4.73
5.38
3.56
4.10
3.50
4.61
3.98
3.36
0.11
0.18
0.12
0.18
0.15
0.18
0.08
0.12
0.11
0.06
0.11
0.10
0.19
0.49
0.36
0.34
0.47
0.71
0.025
0.016
0.022
0.047
0.022
0.021

108 Akgün, Teğin, and Ziyadanoğulları Vol.5, No.2
Table 4: The values obtained by flotation of the third sulfurized sample
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g)
Tailing
(g)
Mo Pb Cu Zn Mo Pb Cu Zn Mo Pb Cu Zn
1.20
1.30
1.30
1.30
1.45
1..30
6.51
7.00
7.50
8.00
8.50
9.01
26.27
27.19
26.74
24.80
24.63
20.49
73.16
71.29
72.53
74.49
75.54
79.23
100
98.2
100
88.7
96.3
92.0
87.0
88.4
95.9
79.0
86.0
66.2
87.0
92.3
95.0
88.9
91.5
87.6
52.0
79.0
85.8
43.0
52.0
37.0
4.10
3.90
4.03
3.86
4.21
4.89
3.96
3.90
4.29
3.81
4.17
3.47
0.16
0.17
0.18
0.18
0.19
0.21
2.22
3.24
3.58
1.93
2.35
2.02
-
0.03
-
0.16
0.05
0.11
0.21
0.20
0.06
0.34
0.23
0.66
0.008
0.005
0.003
0.007
0.006
0.007
0.80
0.17
0.22
0.86
0.71
0.88
Table 5: The values obtained by flotation of the fourth sulfurized sample
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g)
Tailing
(g)
Mo Pb Cu Zn Mo Pb Cu Zn Mo Pb Cu Zn
1.15
1.20
1.15
1..20
2.00
2.05
6.53
6.97
7.55
8.00
8.55
9.02
28.31
24.84
29.79
27.79
31.74
30.77
71.12
75.24
70.63
71.72
67.70
69.50
95.0
95.3
96.0
99.0
94.4
100
78.0
79.0
80.0
90.0
81.0
95.0
86.5
88.0
90.
88.5
86,0
95.0
52.5
76.5
59.0
65.0
48.0
52.3
3.61
4.13
3.47
3.83
3.20
3.50
3.31
3.80
3.21
3.87
3.05
3.69
0.15
0.18
0.15
0.16
0.14
0.15
2.06
3.44
3.38
2.60
1.69
1.90
0.08
0.07
0.06
0.02
0.09
-
0.38
0.35
0.34
0.17
0.34
0.01
0.009
0.007
0.007
0.008
0.010
0.003
0.75
0.35
0.64
0.55
0.86
0.77
Table 6: The values obtained by flotation of the fifth sulfurized sample
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g)
Tailing
(g)
Mo Pb Cu Mo Pb Cu Mo Pb Cu
3.0
2.0
2.10
1.5
2.1
2.36
6.51
7.01
7.51
8.00
8.49
9.0
12.04
12.23
12.45
13.33
12.21
13.69
86.45
87.44
87.11
86.78
87.31
86.28
88.5
91.0
90.7
94.4
84.0
89.8
72.0
74.0
74.4
76.0
78.0
81.0
71.0
83.0
81.6
79.6
72.0
74.8
7.92
8.02
7.85
7.63
7.41
6.79
7.16
7.25
7.16
6.83
7.65
7.09
0.28
0.33
0.32
0.29
0.30
0.26
0.15
0.12
0.10
0.07
0.20
0.13
0.39
0.36
0.35
0.33
0.31
0.25
0.017
0.010
0.011
0.012
0.016
0.014

Vol.5, No.2 Enrichment of Molybdenum and Fluorite 109
Table 7: The values obtained by flotation of the sixth sulfurized sample
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g)
Tailing
(g)
Mo Pb Cu Mo Pb Cu Mo Pb Cu
2.0
4.0
5.0
7.2
7.00
7.02
7.00
7.01
29.21
33.33
70.40
90.66
69.69
215.34
228.43
257.22
89.3
91.4
89.8
90.6
73,0
82.1
91.7
95.0
80.0
88.4
83.5
85.7
3.29
7.40
4.13
3.78
2.98
7.37
4.68
4.40
0.13
0.33
0.17
0.16
0.16
0.10
0.15
0.14
0.46
0.25
0.13
0.08
0.014
0.007
0.011
0.010
As seen in tables, the best result was obtained from third sulfurized sample by flotation. In
order to understand these results better, the results were plotted in Figure 1 with respect to different
pH values.
0
20
40
60
80
100
6789
pH
Flotation yield (%)
Mo
Pb
Cu
Zn
Figure 1: The effect of pH on the flotation yield
As seen in Figure 1, the best of flotation yield was at pH 7.5. Thus, six different sulfurized
samples were floated at this pH and the results are given in Figure 2.

110 Akgün, Teğin, and Ziyadanoğulları Vol.5, No.2
0
20
40
60
80
100
7,510 12,5 15 17,5 20
H2S (g)
Flotation yield (%)
Mo
Pb
Cu
Figure 2: The effect of sulfurization at pH 7.5 on the flotation yield of sulfurized samples
When Figure 2 is examined, it can be seen that the highest flotation yield was achieved by
flotation of third sulfurized samples.
In the next step, flotation process was conducted to test the effect of pulp density and
results are given in Table 8.
Table 8: The effect of pulp density on the floatability of results from third sulfurized sample
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g)
Tailing
(g)
Mo Pb Cu Zn Mo Pb Cu Zn Mo Pb Cu Zn
2.40
3.50
3.60
3.75
7.50
7.51
7.52
7.50
25.73
61.84
67.04
83.33
73.52
188.73
232.23
266.20
97.4
96.4
97.6
94.6
94.6
93.0
94.0
92.0
94.0
94.0
.95.4
85.6
86.0
85.5
89.0
82.0
4.08
4.20
4.71
4.29
4.41
4.50
5.03
4.62
0.18
0.19
0.21
0.17
3.74
3.86
4.45
3.86
0.04
0.05
0.03
0.08
0.06
0.11
0.09
0.13
0.004
0.003
0.004
0.003
0.21
0.22
0.16
0.27
When Table 8 is examined, it can be seen that there is not a significant difference in
flotation yield as pulp density changes. However, the results are better when solid/liquid ratio was
done at 300 g ore / 1 L
3.1.3. The effect of the activator and the depressant on the flotation yield
3.1.3.1. The effect of the activator:

Vol.5, No.2 Enrichment of Molybdenum and Fluorite 111
The results of the flotation in which CuSO4 was used as activator are given in Table 9.
Table 9: The values obtained from flotation of third sulfurized sample using activator (CuSO4)
As seen in Table 9, floated amount of sample was too much. Moreover, in order to
investigate whether flotation yield could be increase more, a series of experiments were carried out
by using depressant (Na2SiO3) as well as activator.
3.1.3.2. The effect of the activator and depressant together:
In these series of experiments, flotation studies were conducted by using CuSO4 and
Na2SiO3 and results obtained are given in Table 10.
Table 10: The values obtained from flotation of third sulfurized sample using the activator and the
depressant together at pH 7.5
Note: 1 mL solution contains 5.10-3 g Na2SiO3.
* Conditions under which the best results were obtained for Mo, Cu, Pb and Zn recovery
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
CuSO4
(mg)
Time
Min
pH Froth
(g)
Tail
(g)
MoPb Cu Zn Mo Pb Cu Zn Mo Pb Cu Zn
5.0
10.0
10.0
1.50
1.20
1.20
7.50
7.52
8.50
33.00
38.83
31.83
66.32
59.92
67.98
100
100
100
87.0
84.0
90.0
80.0
99.0
95.0
60.0
92.0
72.0
3.27
2.78
3.05
3.16
2.59
3.38
0.12
0.12
0.15
2.03
2.65
2.52
-
-
-
0.24
0.33
0.18
0.015
0.001
0.004
0.68
0.15
0.46
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
CuSO4
(mg)
Na2Si
O3
(mL)
Time
(Min)
Froth
(g)
Tail
(g)
Mo Pb Cu Zn Mo Pb Cu Zn Mo Pb Cu Zn
5.0
15.0
20.0
20.0
20.0
25.0
30.0*
35.0
0.5
1.0
1.50
1.0
0.5
1.0
1.0
1.0
1.30
2.10
2.20
2.30
2.20
1.50
1.40
1.30
29.85
19.75
19.56
19.94
22.74
24.90
23.47
23.85
70.29
79.66
79.91
79.43
78.27
74.64
75.92
75.49
100
95.0
91.0
96.0
89.0
96.3
97.3
96.3
87.0
80.0
66.0
67.0
73.0
92.0
93.0
86.0
94.0
84.0
55.0
76.0
72.0
95.6
98.0
100
66.0
78.0
67.0
69.0
65.0
89.0
95.2
66.2
3.61
5.18
5.01
5.18
4.21
4.16
4.46
4.35
3.48
4.85
4.03
4.02
3.84
4.42
4.74
4.31
0.15
0.21
0.14
0.19
0.15
0.19
0.21
0.21
2.46
4.41
3.82
3.86
3.20
4.00
4.53
3.10
-
0.07
0.13
0.05
0.15
0.05
0.04
0.05
0.23
0.31
0.52
0.50
0.42
0.13
0.11
0.22
0.004
0.010
0.028
0.015
0.018
0.003
0.001
-
0.55
0.31
0.52
0.44
0.51
0.17
0.07
0.50

112 Akgün, Teğin, and Ziyadanoğulları Vol.5, No.2
As seen in Table 10, although molybdenum yield decreased a bit, yields of other elements
increased considerably. Particularly, the optimum results were obtained when 30.0 mg CuSO4 and
1.0 mL Na2SiO3 were used in the flotation. In addition, amount of sample floated by depressant
fairly decreased and 94 % of fluorite was depressed (not shown in Table). The results obtained
from flotation of the samples at up to 30 % solid/liquid ratio as the pulp density changed were
similar (not shown in Table). Additionally, this concentrate phase (flotation carried out with 30 mg
CuSO4) contained 0.05 % Ta, 0.68 % Nb and 0.41 % V and was produced with recoveries of 76 %
Ta, 84.4 % Nb and 76.2 % V by flotation.
Separation of Mo and Cu from Pb and Zn in the Concentrate
The concentrate phase (flotation carried out with 30 mg CuSO4) containing 4.46 % Mo,
4.74 % Pb, 0.21 % Cu, 4.53 % Zn, 0.05 % Ta, 0.68 % Nb and 0.41 %V was used in this process. In
order to separate Mo and Cu from Pb and Zn, the experiments were conducted with respect to pH.
Thus, these experiments were performed with -100 mesh size and -160 mesh sizes, and pH 11 to
pH 13. Results obtained are given in Table 11.
Table 11: The values obtained by flotation of the collective concentrate
-100 mesh size
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g)
Tail
(g)
Mo Pb Cu Zn Mo Pb Cu Zn Mo Pb Cu Zn
2.50
2.30
2.50
2.50
2.50
3.00
11.0
11.50
12.00
12.50
12.75
13.00
78.20
62.34
50.79
45.26
38.59
31.31
21.55
37.71
48.97
53.43
61.04
69.87
98.0
95.5
90.0
98.0
93.0
97.3
97.6
94.0
77.6
57.0
47.5
31.6
98.0
96.7
93.9
65.8
74.0
84.0
87.7
75.8
60.5
55.0
25.4
30.0
5.58
6.82
7.90
9.65
10.74
13.85
5.90
7.14
7.23
5.96
5.82
4.77
0.26
0.32
0.38
0.30
0.40
0.56
5.07
5.50
5.39
5.49
2.98
4.33
0.42
0.53
0.91
0.17
0.51
0.17
0.53
0.75
3.14
3.82
4.08
4.64
0.019
0.018
0.026
0.134
0.089
0.048
2.59
2.90
3.65
3.82
5.54
4.54
-160 mesh size
3.00 13.0 33.46 66.98 84.6 31.368.434.411.27 4.410.434.65 1.03 4.86 0.1244.97
As seen in Table 11, the best separation was performed at pH 13 and with -100 mesh
particle size. However, in order to obtain better flotation results, -100 mesh size samples were
sulfurized again. For resulfurization, two samples were prepared in two different autoclave
mediums: Sample A containing 0.96 g H2S + 5 g H2O steam for 500 g of sample and Sample B
containing 1.92 g H2S+ 10 g H2O for 500 g of sample. Additionally, -200 mesh size samples were

Vol.5, No.2 Enrichment of Molybdenum and Fluorite 113
sulfurized under the same conditions. The best results were obtained from the flotation of the
samples of -100 mesh size sulfurized in a medium containing 0.96 g H2S + 5 g H2O steam for 500
g of sample and are given in Table 12.
Table 12: The values obtained by flotation of resulfurized collective concentrate
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
pH Froth
(g) Tailing
(g)
Mo Pb Cu Zn Mo Pb Cu Zn Mo Pb Cu Zn
3.0
2.5
2.30
12.00
12.50
13.00
74.06
36.05
22.45
25.88
64.68
77.69
97.6
96.0
98.0
94.0
64.0
12.6
97.0
88.0
94.0
83.0
44.0
22.5
5.87
11.87
19.46
6.01
8.40
2.65
0.27
0.51
0.87
5.06
5.51
4.54
0.41
0.28
0.11
1.10
2.64
5.33
0.024
0.038
0.016
2.98
3.92
4.52
As seen in Table 12, most of Mo and Cu were floated at pH 13, while most of Pb and Zn
were depressed. Then, Mo, Cu, Ta, Nb and V were recovered from this concentrate. For this
process, these elements were separated by a hydrometallurgical method (9).
Moreover, the flotation yields of Ta, Nb and V, which were not given in Table 12, were
found as 86 %, 91 % and 89 %, respectively and their concentrate grades of Ta, Nb and V were
ascertained as 0.16 % Ta, 2.43 % Nb and 1.49 % V.
In the next step, the sample was floated to test the effects of pulp density. As results given
in Table 12, the best of yield was obtained with 10 % solid/liquid ratio.
Flotation of Fluorite in the Tailing
Tailing of first step flotation (collective flotation) contained 14.4 % F (29.6 % CaF2). It was
tried to enrich fluorite by flotation of this sample. For this aim, the effect of some parameters such
as pH, amount of the collector, the activator and the depressant were examined. Results obtained
under flotation conditions stated below are given in Table 13.
Flotation condition:
Particle size : - 100 mesh
Solid / Liquid Rate : 100 g / L
Collector : Na-Oleate (5.10-2 g/mL) 3 minute mix

114 Akgün, Teğin, and Ziyadanoğulları Vol.5, No.2
Activator : Al2(SO4)3 (5.10-2 g/mL)
Suppressor : Na2SiO3 (5.10-3 g /mL)
Mix Speed : 900 rpm
Table 13: The values obtained from the flotation performed for enrichment of fluorite
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Time
(Min)
Al2(SO4)3
(mL)
Na2SiO3
(mL)
Na
oleat
(mL)
pH Froth
(g)
Tailing
(g)
F F CaF2 F CaF2
1.0
1.0
1.10
1.20
1.0
1.0
1.0
1.0
1.0
1.0
1.0
1.0
1.0
1.0
1.0
1.0
7.5
8.0
8.5
9.0
24.77
16.81
19.20
24.39
74.49
82.48
80.53
75.03
78.0
44.4
53.2
71.3
45.35
38.03
39.90
42.10
93.21
78.18
82.02
86.53
4.2
9.7
7.3
5.5
8.7
19.9
15.7
11.3
*1.0 mL Al2(SO4)3 = 5.10-2 g , 1.0 mL Na2SiO3 =5.10-3 g, 1.0 mL Na-Oleate = 5.10-2 g
As seen in Table 13, the highest flotation yield was at pH 7.5. However, since these results
were not satisfactory, the sample was floated with respect to the amounts of the activator, the
depressant and the collector and the results obtained are given in Table 14.
Table 14: The values obtained from the flotation performed with respect to the activator,
depressant and collector
Recovery
(%)
Concentrate
Assay (%)
Unfloated
Assay (%)
Al2(SO4)3
Na2SiO3
Na oleate
(mL)
Time
(Min)
pH Froth
(g)
Tailing
(g)
F F CaF2 F CaF2
-
1.0
1.0
1.50 7.5 29.03 70.40 94.0 46.63 95.85 1.2 2.5
1.0
1.0
2.0
2.0 7.50 32.39 66.48 95.0 42.24 86.82 1.1 2.2
1.0
1.0
1.5
1.5 8.0 15.77 83.57 48.0 43.83 90.10 8.9 18.4
1.0
1.0
2.0
2.0 8.0 35.61 63.76 94.5 38.21 78.55 1.2 2.6
1.0
0.5
1.0
1.45 7.5 23.51 76.90 73.4 44.96 92.41 5.0 10.2
1.0
0.5
1.0
2.0 8.0 30.23 68.96 94.0 44.78 92.04 1.3 2.6

Vol.5, No.2 Enrichment of Molybdenum and Fluorite 115
As seen in Table 14, the highest flotation yield and in concentrate grade of CaF2 were
obtained by the flotation conducted in a medium containing 1.0 mL Na2SiO3 and 1.0 mL Na-
oleate. Thus, these experiments were repeated three times and the similar results were obtained.
Then, the flotation studies were performed by changing the amounts of Na2SiO3 and Na- oleate.
The best result was obtained from the flotation in which 1.0 mL Na2SiO3 and 1.0 mL Na- oleate
was used at pH 7.5 and it was determined that grade of sulfur was 0.5 %.
CONCLUSIONS
The finding of processing of fluorite ore containing molybdenum are as follows.
• Collecting molybdenum and fluorite in different phases by direct flotation was not
possible, and the flotation yield was low
• Since molybdenum and other minerals bound to gang minerals, phase separation
was not possible. Liberation with respect to particle size was unsatisfactory and this
resulted in low flotation yield.
• In order to achieve an efficient separation by flotation, fluorite ore containing
molybdenum was sulfurized to change its structure and surface. Thus, the sample
was sulfurized in a medium containing different amounts of H2S and H2O steam.
12.44 g H2S + 65 g H2O gas mixture was enough for sulfurization one kg of sieved
ore. The sulfurization was finished in 1 hour in the autoclave for at 120 oC.
• The minerals containing fluorite and molybdenum were separated with high yields
by flotation of sulfurized sample. At the end of procedure, 94 % of fluorite was
remained at the tailing. Beside of Molybdenum, Pb, Zn, Cu, Ta, Nb and V also
passed into concentrate phase with high yields. After the collective flotation,
concentrate phase contained 4.46 % Mo, 4.74 % Pb, 4.53 % Zn, 0.21 % Cu, 0.05 %
Ta, 0.68 % Nb and 0.41 % V. In the flotation performed with respect to pulp
density, the flotation yield was remained at the same levels with the samples
consisted of up to 30 % solid/liquid ratio. Additionally, the flotation yield was
increased when Na2SiO3 was used as depressant and CuSO4 was used as activator.
• In order to increase the grade of Mo, Cu, Ta, Nb and V from Zn and Pb, the flotation
process was repeated. Flotation was performed at higher pH and optimum pH was
found as 13. However, it was decided to repeat sulfurization of concentrate since
flotation yield was not satisfactory.

116 Akgün, Teğin, and Ziyadanoğulları Vol.5, No.2
• 0.96 g H2S + 5 g H2O gas mixtures were enough for the sulfurization of 500 g
sample. After the flotation, the concentrate grades of Mo, Pb, Cu, Zn, Ta, Nb and V
were found as 19.46 %, 2.65 %, 0.87 %, 4.54 %, 0.16 %, 2.43 %, and 1.49 % and
flotation yields were determined as 98 %, 12.6 %, 94 %, 22.5 %, 86 %, 94 % and
82.2 %, respectively. So, while Mo, Cu, Ta, Nb and V was collected in the
concentrate phase, Pb and Zn passed into the tailing. The concentrate was enriched
approximately 10 times with respect to the original ore.
• It was ascertained that the tailing of collective flotation contained 14.4 % F (29.6%
CaF2). This sample was refloated for enrichment. The effect of the parameters such
as pH, activator, depressant and collector were investigated. The highest flotation
yield was at pH 7.5. Using Al2(SO4)3 as activator did not increased the flotation
yield significantly. On the contrary, in the flotation in which Na2SiO3 and Na oleate
was used, concentrate grade of CaF2 raised to 96 % and flotation yield was 94 %.
Additionally, it was determined that the concentrate contained 0.5 % S.
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Vol.5, No.2 Enrichment of Molybdenum and Fluorite 117
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