Effects of grinding environment on the flotation of Rosh Pinah complex Pb/Zn ore

openUP (July 2007)
Effects of grinding environment on the flotation of Rosh
Pinah complex Pb/Zn ore


Y. Weia and R.F. Sandenbergha

aDepartment of Materials Science and Metallurgical Engineering, University of Pretoria,
Pretoria 0002, South Africa


Abstract
The Rosh Pinah orebody is a complex lead–zinc sulphide system with pyrite gangue and
minor amounts of copper. Laboratory scale milling and flotation testing of ore samples
taken from this operation was performed. Different grinding media and conditions were
used, including ceramic, stainless steel and steel. Flotation tests used a sequential
recovery protocol for selective flotation of first the lead and thereafter the zinc. The
presence of species of oxidation products on the ore after milling was probed using
ethylene diamine tetra-acetic acid (EDTA) leaching. The test data show that, for
comparable grinds and reagent dosages, the choice of grinding media has a marked effect
on the paymetals recoveries and the selectivity. The ceramic mill produced the highest
recoveries but the poorest selectivity. The steel mill produced the converse result. It is
proposed that this is because the ceramic mill produces an oxidizing environment,
allowing sphalerite activation by copper ions, whilst the steel mill produces a reducing
environment, preventing this activation.


Article Outline
1. Introduction
2. Experimental

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openUP (July 2007)
2.1. Ore
2.2. Chemicals and water
2.3. Grinding and flotation
2.4. EDTA leaching, chemical phase analysis and solution analysis
3. Results
3.1. Flotation performance as a function of milling environment
3.2. Flotation kinetics
3.3. Size-by-size recovery distributions for different milling methods
3.4. Pulp chemistry after milling
4. Discussion
5. Conclusions
Acknowledgements
References


1. Introduction
Comminution of complex ores is typically necessary to liberate the sought after minerals
to allow for their selective recovery during recovery processes such as flotation. The
comminution process inevitably involves contact of the ore with the surface of
comminution device and in the case of ball or rod milling also with the grinding medium.
This not only causes wear, but also contamination of the ore with wear debris as well as
with precipitated species on the ore surfaces. If milling is done in an aqueous medium,
corrosion of the ore and the medium can lead to complex interactions, such as galvanic
effects, changes in the Eh of solution and the dissolution and precipitation of species on
the ore surfaces. Galvanic interactions between the steel grinding media and sulphide
ores will typically increase steel consumption and reduce the corrosion of more noble
sulphides, with commensurate higher oxygen consumption, and the possible precipitation
of iron species on the ore particles (Wang and Xie, 1990, Natarajan, 1992 and Natarajan,
1996). Steel media wear during grinding may be reduced in a number of ways, such as
using rubber liners, corrosion resistant media and corrosion inhibitors (Natarajan, 1992,
Natarajan, 1996, Pradip, 1992 and Ayyala et al., 1993).

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openUP (July 2007)
As expected, the interactions between the ore and milling environment affect sulphide
flotation both in recovery and selectivity (Forssberg et al., 1988, Forssberg et al., 1993,
Greet et al., 2005, Kocabag and Smith, 1982 and Wang and Xie, 1990). Rey and
Formanek (1960) first demonstrated the effects of milling on the selective flotation of
Pb/Zn ores. Using synthetic lead–zinc ores (mixing of galena, sphalerite, cerussite and
anglesite) they found that sphalerite showed natural floatability, as well as some extent of
activation by lead ions, which reduced its separation from galena when ground in a
porcelain mill. Grinding this ore in an iron mill reduced the natural floatability of
sphalerite significantly. Many industrial circuits also showed significant effects of milling
on sulphide flotation (Kocabag and Smith, 1982, Forssberg et al., 1988 and Forssberg et
al., 1993). The depression of sulphide flotation after milling in a steel mill, which
typically results in a less oxidising environment, could be due to the unselective coating
of the sulphides by iron oxidation products, or failure to oxidise the collector to its
hydrophobic species. The prevention of corrosion in the milling process may thus reduce
the cost of grinding and improve flotation recovery, but may, unfortunately, reduce the
selectivity of the process. The application of high chromium grinding media was tested
and proved to be a viable measure to improve lead flotation recovery by Greet et al., 2004
and Greet et al., 2005.
In this work, ceramic, stainless and steel mills were used to compare the effect of the
milling environment on the flotation of a complex sulphide ore.

2. Experimental
2.1. Ore
A complex Pb–Zn ore from Rosh Pinah, Namibia was used for all the experiments. The
ore sample taken from the plant mill feed was crushed to below 3.35 mm at the mine and
riffled into 20 kg lots using a spinning riffler. It was further crushed to ?2 mm at the lab
using a gyratory crusher, homogenized and rotatory riffled to 1 kg lots and stored in
plastic bags until final milling just prior to the experiments. The chemical composition of
the sample is shown in Table 1.

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openUP (July 2007)
Table 1.
The chemical composition of the ore (mass %)
Pb Zn Cu Fe S CaO MgO BaO Al2O3 SiO2
2.69 11.91 0.17 6.19 7.48 22.98 10.52 5.22 3.17 25.15

2.2. Chemicals and water
The collector used in the experiments was sodium normal propyl xanthate (SNPX)
provided by Senmin, South Africa and was used as supplied. This collector is currently
used at Rosh Pinah. Industrial frother (Dow 250) was used for all the flotation
experiments. High purity argon gas (>99.999%) was used as protective gas for the EDTA
leach tests and ammonium acetate selective leaching. EDTA and all the other reagents
used in the experiments were of analytical grade quality. Water used in the experiments
was Pretoria tap water.

2.3. Grinding and flotation
Three kinds of mills were used in the experiments namely stainless steel and steel rod
mills of the identical size and a ceramic ball mill. One kilogram of ore was ground wet in
the different mills at a mass concentration of 66.7%. The fineness of ground product,
represented as percent mass passing 75 ?m, was controlled by varying the time of
milling. Conditioning and flotation were conducted in a 2 L flotation cell (Denver 12) at
an unadjusted pH of 7.8–8.2. The pulp was stirred for 3 min before adding SNPX
(50 g/t), followed by conditioning for 5 min, and the addition of frother (Dow frother
250, 12 g/t), followed by 3 min conditioning. Flotation was conducted at a concentration
of 32% solids by mass. The flotation machine was operated at an impeller speed of
1150 rpm. Froth was continuously removed from the cell by an automatic scraper
operating at a speed of 26 revolutions per minute (rpm). The sides of the cell were
washed with tap water during the flotation, and the pulp was kept at a constant level by
adding tap water. The total flotation time was 16 min. Pulp potential (Eh) and dissolved
oxygen (DO) levels were recorded continuously from the beginning of conditioning to
the end of flotation using an Orion composite platinum electrode and a DO probe.

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openUP (July 2007)
Flotation products were analyzed using X-ray fluorescence spectrometry (XRF) and
conventional wet screening. A modified Kelsall first order model (Kelsall, 1961 and
Lynch et al., 1981), with fast and slow-floating distribution, as shown in Eq. (1) was used
to assess the flotation results. The flotation parameters were calculated by fitting the
experimental data through Sigmaplot (version 7.0) according to the regression Eq. (1):

R
t
t
t=Rmax{1-((1-?)e-kf +?e-ks )} (1)
where Rt is the cumulative floatation recovery of mineral at time t, Rmax is the ultimate
flotation recovery, kf, ks are the fast-floating and slow-floating rate constants,
respectively, and ? is the slow floating fraction.

2.4. EDTA leaching, chemical phase analysis and solution analysis
EDTA leaching was used to quantify the precipitates on the ore using the procedure
recommend by Greet and Smart (2002), with a few minor modifications. A 20 mL of
sample of pulp was taken after the 3 min stirring and added to 250.00 g of previously
argon purged (for at least 30 min) 0.1 M EDTA solution of pH 7.5 (adjusted with
potassium hydroxide). The slurry was stirred vigorously under argon for 5 min before
filtering through a 0.22 ?m Millipore filter. A 50 mL sample of the filtrate was analyzed
using ICP. Copper was analysed by ICP plus MS.
The real concentration of the EDTA leachable metal ions in the pulp can be calculated
approximately according to the following formula:

(2)



(3)
where C0 is the measured concentration of the EDTA leachate, C1 is the converted metal
concentration in pulp and C2 is the corresponding equivalent dosage of that metal in pulp.
The solution volumes of total EDTA leaching solution, pulp solution sampled and total
pulp volume, are 0.266, 0.02 and 2.2 L, respectively.

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openUP (July 2007)
The analysis of the oxidised lead was done using the procedure recommended by Young
(1974) as guideline. Two grams of ?0.074 ?m sample was added to 200 mL of argon
purged 60% ammonium acetate solution and stirred under argon for 1.5 h. A 50 mL
sample of the filtrate was analyzed after filtration and dilution, using ICP.

3. Results
3.1. Flotation performance as a function of milling environment
The flotation recovery of various minerals from Rosh Pinah ore, as represented by their
indicator elements, versus grinding fineness, is shown in Fig. 1, Fig. 2 and Fig. 3 for the
different milling environments. These experiments were conducted without any cyanide
addition.


Fig. 1. Flotation recovery of lead (filled symbol) and zinc (hollow symbol) containing
minerals from Rosh Pinah ore as a function of particle size after milling in different
environments.

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openUP (July 2007)

Fig. 2. Flotation recovery of iron (filled symbol) and silicon (hollow symbol) containing
minerals from Rosh Pinah ore as a function of particle size after milling with different
mills.


Fig. 3. Flotation recovery of calcium (filled symbol) and magnesium (hollow symbol)
containing minerals from Rosh Pinah ore as a function of particle size after milling in
different environments.
Milling in stainless steel resulted in the highest recovery of lead over the whole size
range, but unfortunately also gave the highest recovery of zinc, iron, and other non-
sulphide gangue minerals. Milling in the ceramic mill gave similar results, but with lower

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openUP (July 2007)
zinc recovery, while milling in the steel mill resulted in somewhat lower lead recovery,
and also with lower zinc recovery. There was only a slight decrease in zinc recovery as
the grinding fineness increased while the recovery of lead was not affected by the
grinding fineness, indicating that the lead and zinc minerals are liberated from each other
at grinds of more than 75% ?75 ?m. Rougher flotation at a coarser grind followed by
regrinding to liberate the lead and zinc from each other has been suggested by Seke and
Pistorius (2005) as a method to optimise both the grade and recovery of the lead and zinc
concentrates recovered from this ore.
Pyrite (iron) recovery for all the three milling environments was very high and showed
little change as the grinding fineness increased, probably resulting from the good
floatabilities of pyrite with SNPX. The flotation of the non-sulphide gangue minerals
were also significantly affected by the different milling environments with the best
recovery with the stainless steel, and the poorest with the carbon steel.
It can be concluded that milling in a stainless mill increased the recovery for all the
minerals, while milling in a steel mill decreased the flotation of all the minerals, although
to different extents, while milling in a ceramic mill gave intermediate results. However,
the ceramic mill was a ball mill, whereas the stainless and steel mills were rod mills, and
different size distributions might have contributed to the difference in flotation behaviour.

3.2. Flotation kinetics
The kinetics of flotation after comminution in the different milling environments was
characterized using a grind fineness of about 76% ?75 ?m where good liberation between
the lead and zinc minerals should be achieved. The particle size distribution after milling
in the three types of mill is given in Table 2.
Table 2.
The particle size distribution of Rosh Pinah ore after milling in the three kinds of mills to
the same fineness (%)

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openUP (July 2007)

Particle range (?m) Steel mill
Stainless mill Ceramic mill
Con.
Tail.
Con.
Tail.
Con.
Tail.
+106
3.03 11.44 4.17 5.81 4.25 7.87
?106 + 75
2.02
6.54
6.33
7.12
4.26
7.56
?75 + 53
4.27
7.75
7.70
8.39
5.32
8.41
?53 + 45
3.25
4.64
4.55
4.39
2.82
4.80
?45 + 38
2.85
4.62
3.02
3.92
2.67
4.54
?38
19.23 30.36 19.84 24.76 18.45 29.05







Total
34.65 65.35 45.61 54.39 37.77 62.23
Con. = concentrate, Tail. = Tailing.
Fig. 4, Fig. 5 and Fig. 6 show the flotation recoveries of various elements after grinding
Rosh Pinah ore in the three milling environments, while the corresponding modified
Kelsall model parameters for the flotation of the elements are shown in Table 3. The
Kelsall model fitted the experimental data very well, with correlation coefficients of
greater than 0.99 achieved for most cases. Fig. 7 illustrates the relation between lead
grade and recovery for the three mill types. Fig. 8 and Fig. 9 illustrate the flotation
selectivity between Pb–Zn and Pb–Fe, respectively.

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openUP (July 2007)

Fig. 4. Flotation recovery of lead minerals vs. flotation time of Rosh Pinah ore for the
indicated milling methods. The lines illustrate the Kelsall regression fits to the
experimental data points.


Fig. 5. Flotation recovery of zinc minerals vs. flotation time of Rosh Pinah ore for the
milling methods indicated. The lines illustrate the Kelsall regression fits to the
experimental data points.

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