t. adel w. e. fore ///(;j e gzén
TRANSCRIPT
THE\COLLECTORLESS FLOTATION OF SPHALERITEL,
bvJohn Raymond Craynon
Thesis submitted to the Faculty of the
Virginia Polytechnic Institute and State University
in partial fulfillment of the requirements for the degree of
MASTER OF SCIENCEin
Mining and Minerals Engineering
APPROVED:
R. H. éééé, Chairman
6G.T. Adel W. E. Fore ///(;jE gzén
July, 1985Blacksburg, Virginia
·
THE OOLLECTORLESS FLOTATION OF SPHALERITE
,s, by .&‘ JOHN RAYMOND CRAYNON
(ABSTRAOT)
The flotation of sphalerite has been demonstrated without the use
of collectors. The effect of redox potential, pH, and copper-activation
have been investigated in tests using samples of pure mineral. It has
been found that in general, collectorless flotation of sphalerite can be1
accomplished at potentials greater than -2OO mV, SHE, and is more
readily carried out in acidic solutions. It has also been shown that
although copper-activation was necessary to achieve flotation recoveries
above 35%, an excessive addition of cupric ions may result in a decrease
in floatability.1
Batch flotation experiments conducted using Elmwood Mine sphalerite
ore have shown that in addition to copper—activation, the addition of
sodium sulfide was required to obtain high grades and recoveries. If
the ratio of the addition of these reagents is maintained such that the
atomic ratio of cupric ions to sulfide ions is O.31, good flotation is
observed over a range of reagent dosages.
X-ray photoelectron spectroscopy (XPS) was conducted on pure
mineral samples after microflotation testing. Based on the sulfur
species identified on highly flotable samples, possible mechanisms for
collectorless flotation of sphalerite have been suggested. These
include: i) elemental sulfur formed under oxidizing conditions is 1
responsible for collectorless flotation; ii) polysulfides or (
1 1
b
metal—deficient sulfides formed as a result of mineral oxidation are
responsible for collectorless flotation; and iii) removal of HS- ions,
which may render the surface hydrophilic, under oxidizing conditions.
The third mechanism is based on the assumption that clean, unoxidized
sphalerite surfaces are naturally hydrophobic. Evidence has been
presented to suggest that the first mechanism may be responsible for
collectorless flotation in acidic solutions, while the second mechanism
may be of greater importance in nearly neutral or basic solutions where
elemental sulfur is thermodynamically less stable.
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TABLE OF CONTENTS
PageABSTRACT........................................... iiACKNOWLEDGEMENTS................................... ivLIST OF FIGURES.................................... viiLIST OF TABLES..................................„..viii
INTRODUCTION....................................... 1General.................................. ..... 1Literature Review............................. 3Scope of Work................................. 8
EXPERIMENTAL....................................... 9Materials..................................... 9
Ore Samples.............................. 9Pure Minerals............................ 9Reagents................................. 1O
Equipment..................................... 13Electrodes............................... 13
Potential Electrodes................ 13Dissolved Oxygen Probe.............. 13Sulfide Ion Electrode............... 14pH Electrode........................ 14
Microflotation Apparatus................. 15Batch Flotation Apparatus................ 15 ‘
X—ray Photoelectron Spectroscopy......... 15Procedure..................................... 18
Measurement of Operating Variables.,..... 18Potential........................... 18Dissolved Oxygen.................... 18Sulfide Ion......................... 18pH.................................. 19
Microflotation........................... 19Batch Flotation.......................... 2OX—ray Photoelectron Spectroscopy......... 21
EXPERIMENTAL RESULTS............................... 23Microflotation................................ 23
Effect of Potential...................... 23Effect of Copper-Activation.............. 25Effect of pH............................. 27
v
PageBatch Flotation............................... 3O
Effect of Sodium Sulfide Dosage.......... 30Effect of Cupric Sulfate Dosage.......... 3OEffect of Cupric Ion/Sulfide Ion Ratio... 32Effect of Dissolved Oxygen............... 32Effect of Permanganate................... 36Effect of Potential...................... 36
X—ray Photoelectron Spectroscopy.............. 37
DISCUSSION......................................... 43Flotation..................................... A3Mechanisms.................................... A6
Induced Hydrophobicity by Oxidation...... A6Elemental Sulfur.................... L6Polysulfides or Metal-DeficientSulfides............................ 48
Inherent Hydrophobicity.................. 52
SUMMARY AND CONCLUSIONS............................ 55
INDUSTRIAL APPLICATION............................. 57
RECOMMENDATIONS FOR FURTHER WORK................... 58
LITERATURE CITED................................... 59
Appendix A. Batch Flotation Conditions andMetallurgical Balance Sheets.......... 67
Appendix B. Microflotation Conditions............. 8OAppendix C. XPS Spectra and Calculations.......... 88Appendix D. Records of Operating Electrodes
in Batch Flotation....................156Appendix E. Assay Procedure.......................183
VITA...............................................186
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1
1LIST OF FIGURES
PageFigure 1. Sohematic Diagram of the Micro-
flotation Apparatus.................... 16
Figure 2. Schematic Diagram of the BatchFlotation Apparatus.................... 17
Figure 3. Potential versus Per Cent Floatabilityfor Unactivated Sphalerite............. 24 —
Figure 4. Potential versus Per Cent Floatabilityfor Sphalerite Activated with VariousConcentrations of Cupric Sulfate....... 26
Figure 5. Potential versus Per Cent Floatabilityfor Unactivated and Copper-ActivatedSphalerite in Buffer Solutions of pH
1OII•OOIOOIOIOOOOIOOOIOICOUOOFigure
6. Recovery versus Cupric Sulfate Additionfor Elmwood Mine Sphalerite at Constant
Additj•OI1l•IIOOOIOOIOlClOFigure
7. Recovery and Grade versus Cupric Ion/Sulfide Ion Atomic Ratio for ElmwoodMine Sphalerite at pH 6.75............. 33
Figure 8. Record of Operating Electrode Responsefor a Batch Flotation Test Where High
· Recovery and Grade Were Obtained....... 34I
Figure 9. Unresolved Sulfur 2-p Spectra for HighlyFlotable Copper-Activated Sphalerite... 38
Figure 10. Curve-Resolved Sulfur 2-p Spectra forHighly Flotable Copper-ActivatedSphalerite............................. 39
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i
LIST OF TABLES
Page”
Table 1. Summary of Oomposition of Elmwood MineSphalerite as Determined by ElectronMicroprobe Analysis...................... 11
Table 2. Summary of Preparation Conditions ofSamples for X-ray Photoelectron
Table 3. Effect of Dissolved Oxygen Levels onFlotation in Selected Tests.............. 35
Table L. Identification of Sulfur Species Presenton the Mineral Surface as Determined from
Table 5. Relative Abundance of Various SulfurSpecies on the Surface of XPS Samples.... L2
viii
INTRODUCTION
GENERAL
The process of froth flotation is of extreme importance to the
mineral processing industry. Over 2 x 109 tons of ore are processed
using this technique annually (Leja, 1982).
This process relies on interfacial phenomena. The chemical
environment is therefore extremely important in determining the
interaction between the solid, liquid, and gas phases involved. The
common method calls for particular surfactants, called collectors, to be
adsorbed onto the solid surface, rendering it sufficiently hydrophobic
to adhere to a rising stream of air bubbles.
As most minerals are naturally hydrophilic, the use of these
collectors, as well as other reagents to promote collector adsorption is
required. However, some minerals, such as graphite, sulfur and talc, do
not require collector for their flotation (Gaudin, Miaw and Spedden,
1957). Most common sulfide minerals have long been thought to be
hydrophilic and, have been treated as such in flotation practice.
However, some sulfides have been shown to float well without the
addition of collector, suggesting that perhaps they are naturally
hydrophobic. Another possible explanation for the observed
collectorless flotation is that the chemical environment favors the
formation of hydrophobic surface species which lead to good flotability. I
This study attempts to examine the effect of pulp chemistry on the
collectorless flotation of sphalerite, a zinc sulfide. The posssible
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2
flotation mechanisms for collectorless flotation have been examined with
a focus on identifying the hydrophobic species formed during flotation.
The collectorless flotation process studied in the present work might
lead to a savings in reagent costs and to increased efficiency.
T
LITERATURE REVIEW
There have long been questions about the specific character of
sulfide minerals. Very early in the development of the science of
mineral processing, it was noted that sulfide minerals had less affinity
for water than the associated gangue minerals. The bulk oil flotation
technique exploited this difference by inducing the sulfide minerals
into an oil phase in the presence of water (Haynes,1860 and
Everson,1886). This technique was soon abandoned due to the high
consumption of oil.
Another early process that made use of the wettability
characteristics of sulfides was known as the skin flotation technique
(Bradford,1885). In this method of separation fine ore powder was
slowly introduced onto the surface of a quiescent water bath. The
sulfide minerals floated on the water and the gangue, which was easily
wetted, sank. These early processes became obsolete when the use of a
rising stream of air bubbles to buoy oil—coated sulfide minerals to the -
pulp surface was developed (Ballot, Sulman, and Picard, 1905).
Strong evidence has been presented both for and against the natural
flotability of sulfide minerals. Sulman (1930) reported that galena and
chalcopyrite were flotable without collectors. Other investigators,
such as Ravitz and Porter (1933), Ravitz (1940), and Herd and Ure (1941)
specifically suggested that galena was naturally flotable, if cleaned of
all oxidation products. Ravitz and Porter (1933) further suggested that
the role of xanthate, the most common sulfide mineral collector, was toI
chemically clean the mineral surface, thus revealing its inherent IIII
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4
flotability.
Contradictory evidence was presented, however, by Hagihara (1952)
who used the electron diffraction technique to show that xanthate did
indeed adsorb on galena surfaces. Sutherland and Wark (1955)
demonstrated that different xanthates produced different contact angles
on the same sulfide mineral surfaces. These facts seem to
overwhelmingly refute Ravitz and Porter's conjecture.
The natural flotability of sulfides was challenged in work done by
Knoll and Baker (1941). They showed that clean galena did not float
without collector, nor did it adsorb xanthate. Another study where
sodium sulfide (Na2S) was used in an attempt to produce a clean galena
surface indicated a need for collector (Wark,1938). In addition the
depressant effect of Na,S was noted. Gaudin (1932) discovered that this
effect was true for both the reducing agent, Na,S, and the oxidizing
agent, potassium permanganate (KMnO4) in a pure chalcopyrite system.
This controversy was temporarily laid to rest by Taggart, del Giuduce,
and Ziehl (1934) who speculated that the samples had been contaminated
by oily substances or that the frothers had collecting properties
themselves. Others (Plaksin, 1959; Glembotskii, Klassen, and Plaksin,
1963) have suggested that natural hydrophobicity is a result of the
adsorption of molecular oxygen. The speculation was that this
adsorption led to dehydration of the mineral surface allowing air to
displace water more readily at the surface.
Oollectorless flotation has gained renewed interest in recent
years. Boyce, Venter, and Adam (1970) reported that it was being
5 I1
applied for galena and sphalerite recovery at the Tsumeb concentrator in
South Africa. Lepetic (1974) showed that chalcopyrite was flotable
using only a frother after dry, autogenous grinding. Pyrrhotite has
also been reported to be naturally flotable (Hodgson and Agar, 1984).
In addition, Rey and Formanek (1960), Mino (1957), and Plaksin,
Khazinskaya, and Tyurnikova (1955) have observed flotability of
sphalerite without collector.
Heyes and Trahar (1977) discovered that the flotability of
chalcopyrite depended on the potential of the pulp. They also took
precautions to eliminate possible contamination by organics. They
showed, by using alkaline salts to promote frothing, that organic
frothers were not responsible for the phenomenon of collectorless
flotation. Gardner and Woods (1979) confirmed Heyes and Trahar's
results using electrochemical techniques. Using potential sweep
voltammetry, they identified the presence of elemental sulfur as one of
the oxidation products in alkaline pH. Since the mineral floated only
under oxidizing conditions, where it can be oxidized to form elemental
sulfur on its surface, these investigators considered the elemental
sulfur to be the hydrophobic entity responsible for the collectorless
flotation. One thing to note however, is that in addition to the
elemental sulfur, they also identified the presence of iron hydroxide
which is hydrophilic.
Finklestein, Allison, Lovell and Stewart (1975) differed with this
view, finding no correlation between elemental sulfur on the mineral I
surface and natural flotability. Other work in this area has been done1
ü_„„1____l
» IE6 I
by Clifford, Purdy, and Miller (1974), Trahar (1983), and others, who
found evidence demonstrating a correlation between elemental Sulfur and
flotation. Heyes and Trahar (1984) have recently reported the
importance of elemental Sulfur in the collectorless flotation of pyrite
and pyrrhotite.
Research done by Furstenau and Sabacky (1981) demonstrated that
most Sulfide minerals, including sphalerite, were naturally flotable
when ground in an essentially oxygen—free environment. Their work
supported that of other researchers (Rao, 1969; Gaudin, 1957; Rogers,
1962; Gaudin et al, 1959; Sutherland and Wark, 1955; Yonezawa, 1960)
demonstrating that copper activation improved the flotation of
sphalerite. Recent work (Perry, Tsao, and Taylor, 1984) has shown that
copper activation produces a copper-Sulfide compound on the mineral
Surface.
Other researchers have given alternative explanations for
collectorless flotation. Yoon (1981) postulated that the use of sodium
sulfide produced a Sulfur-enriched mineral surface which was able to be
floated. This theory has been supported by research done by Luttrell
and Yoon (1982), Luttrell (1982), and Luttrell and Yoon (1984a) on
chalcopyrite and sphalerite ores. In addition, Luttrell (1982) pointed
out that there is an optimum ratio of copper activation to sodium ,
Sulfide addition yielding maximum recoveries for sphaleriteflotation,verifying
an earlier suggestion by Yoon (1981). He also reiterated theI
importance of potential in chalcopyrite flotation. :
Luttrell and Yoon (1983b, 1984 a and b) indicate the presence of IIII
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polysulfides on the surface of highly flotable chalchopyrite. Their
results indicate the presence of an intermediate oxidation state between
sulfide ion, S2- and elemental sulfur, SO. This species, which can be
inferred from their X-ray photoelectron spectroscopy (XPS) data, had
previously been identified as the sulfur peak from covellite, Cu2S, by
Buckley and Woods (1981). Since the formation of polysulfides occurs at
oxidizing potentials, the theory that they are responsible for flotation
is consistent with results obtained earlier by Heyes and Trahar (1977),
Gardner and Woods (1979) and, Trahar (1983). Hamilton and Woods (1983)
have examined the mechanism for formation of the polysulfides. The
assertion that the oxidation of the mineral to form polysulfide
responsible for flotation has been recently supported by Hodgson and
Asar (1984).Hamilton and Woods (1984) and Buckley and Woods (1984) have offered
a seemingly different explanation for this oxidation product. Instead
of polysulfide being repsponsible for the collectorless flotation, they
considered the metal-deficient mineral surface to be responsible for the
flotation. They have suggested that this species, though
thermodynamically unstable, would have a sulfur lattice structure very
similar to that of the unoxidized mineral. By using electrochemical
techniques, Hamilton and Woods (1984) were able to identify the
potentials involved in the production of these metal-deficient sulfides
and relate these potentials directly to the flotation behavior of
several sulfide minerals.
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SCOPE OF WORK
This investigation has included microflotation of pure sphalerite
for the purpose of establishing the effect of redox potential, pH, and
copper activation on its flotation behavior without collector. Batch
flotation has been carried out on samples of Elmwood Mine sphalerite ore
to determine the role of dissolved oxygen, redox potential, and sodium
sulfide and copper sulfate additions. Further, an attempt has been made
to determine if there exists an optimum ratio between Cu and S
additions. In an effort to explain the mechanisms involved in the
collectorless flotation of sphalerite, X-ray photoelectron spectroscopy
(XPS) was used to examine the samples obtained from microflotation
tests. From this data, possible mechanisms for the collectorless
flotation of sphalerite have been suggested.
I (Y
EXPERIMENTAL
MATERIALS
0re Samples
The ore samples used for this study were obtained from the Jersey
Miniere Elmwood mine in Carthage, Tennessee. The company shipped coarse
run-of-mine ore which was crushed upon receipt to -28 mesh and split
into 1000 gram lots. These were stored at or below -20 C to minimize
surface oxidation. Prior to flotation testing, a bag of sample was
removed from the freezer and split into two 500-gram lots, and ground in
a ceramic ball mill. The average analysis of the feed ore used in this
work was 15.2% Zinc.
Pure Minerals
The pure sphalerite sample used for microflotation tests was also
from the Elmwood mine as obtained from Ward Scientific Company. -
Specimens were visually checked for contamination and impure samples
were discarded. The pure specimens were then crushed with a mortar and
pestle, and the -65 + 100 mesh fraction was obtained by hand screening.
The crushed samples were then further examined visually for foreign
materials and obvious surface discoloration. All particles that were of
a questionable nature were removed with tweezers and discarded.
_The -65+100 mesh sample, thus prepared, was placed in a Vacuum
dessicator and stored under Vacuum until use. The pure mineral assayed
65.9% zinc by weight. The complete composition of the sphalerite as
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determined by electron microprobe analysis is in table 1.
Reagents
Sodium sulfide flakes from Fisher Scientific were used as a surface
cleaning agent in some batch flotation tests. These flakes were
composed of 60 to 62% Na2S, more than 35% water of crystallization, 1.5%
sodium chloride and 2% unspecified sodium salts. For microflotation
tests, where higher quality sodium sulfide was required, reagent grade
sodium sulfide crystals (Na2S'9H20), also from Fisher Scientific, were
employed. These were carefully scraped and washed with doubly distilled
water as recommended by Chen and Morris (1972) to remove surface
oxidation products before weighing and dissolution. This high-grade
sodium sulfide solution was used to make the pulp potential more
reducing. Fresh solution was prepared daily. The certified A.C.S.
grade potassium permanganate obtained from Fisher Scientific Company was
used to adjust the pulp to more oxidizing conditions in some batch
experiments.
When sphalerite activation was desired, certified A.C.S. grade
cupric sulfate (CuS04'5H20) from Fisher Scientific was used.
Dowfroth 250 from Dow Chemical Company was used as frother in all
batch flotation experiments. A 1% solution of this frother was prepared
to more accurately control the dosage. Lime, sodium hydroxide and
hydrochloric acid were used to adjust pH. For some microflotation
experiments, buffer solutions were prepared. The pH A buffer was made
from 0.5M potassium biphalate. The pH 7 buffer made from sodium 11
1 1
11
Table 1. Summary of the composition of pure Elmwood sphaleriteas determined by electron microprobe analysis.
ELEMENT PER CENT BY WEIGHT
Zinc(Zn) 65.86
Sulfur(S) 32.62
Gallium(Ga) 0.61
Germanium(Ge) 0.49
Iron(Fe) 0.24
Cadmium(Cd) 0.18
Total 100.0
W
12
biphosphate (KHZPOA). A solution of 0.05M sodium borate, sodium
carbonate, and sodium hydroxide was used to make a pH 10buffersolution.
These chemicals were obtained from various suppliers.
Compressed ultra—pure nitrogen from Airco Industrial Gases was used
as the carrier gas for microflotation experiments.
All water used in preparing solutions and in microflotation tests
was doubly distilled. A Corning Megapure system was used for
seoond—stage distillation. Batch flotation experiments were conducted
in tap water drawn from the Virginia Tech-Blacksburg, Virginia, water
system.
1
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EQUIPMENT
Electrodes and Meters
Potential Electrodes ——The potential across a bright
platinum-saturated calomel electrode pair was monitored during flotation
experiments. A standard porous calomel electrode was used for all
microflotation work. For batch flotation, a reverse—sleeve calomel
reference electrode was used to provide improved response time and to
lessen the possibility of clogging. The electrodes were refilled weekly
with saturated potassium chloride solution.
A pin-type platinum electrode was used for the microflotation
potential measurements. In batch flotation, a disc—type platinum
electrode was used to increase the surface area presented for slurry
contact and to facilitate proper cleaning. These electrodes were
developed by Orion Research and supplied by Fisher Scientific.
The platinum—calomel pair was connected, along with the sulfide ion
electrodes and the pH electrode, to a Fisher Accumet model 75O digital
pH/millivolt meter through a Fisher model 753 electrode switching box.
The output from this meter was recorded as a function of time using a
Pedersen model 27MR strip—chart recorder.
Oxygen Probe —-Dissolved oxygen was monitored in the batch
flotation work using a Lazar Research Labs model DO-166 dissolved oxygen Iprobe. This was connected to an Altex model 35OO digital
millivoltmeter. The dissolved oxygen levels were recorded as a function
of time using a Hewlett—Packard strip—chart recorder.I
I
I14 I
IThe membrane and reference solution of the oxygen probe were
changed weekly and replaced with manufacturers standard equipment
supplied by Cole-Parmer Equipment Company. The probe was zeroed in a 5%
NaSO3 solution and standardized to atmospheric conditions at the ambient
temperature and pressure.
Sulfide Ion Electrode -—In selected batch flotation tests, the
sulfide ion conoentration was monitored using an Orion Research
silver/sulfide specific ion electrode coupled with an Orion Research
double-junction reference electrode. The outer—chamber electrolyte of
the reference electrode was changed daily and the inner—chamber
electrolyte was changed weekly as suggested by the manufacturer.
This electrode pair was two-point calibrated using doubly distilled
water as a blank and a sodium sulfide solution as a known concentration.
The sulfide standard was prepared weekly and prevented from oxidizing
using the SAOB II buffer described in the electrode instructions.
pH Electrode --Hydrogen ion conoentration was monitored by use of
an epoxy body, glass membrane pH electrode from Fisher Scientific. The
electrode was calibrated in standard pH buffers obtained from American
Scientific Products and Fisher Scientific. The electrode was cleaned I
between tests using O.1 N HC1 and rinsed thoroughly with I
doubly—distilled water. I
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Microflotation Apparatus
Microflotation was carried out in a set—up indentical to that used
by Luttrell (1982). The diagram in figure 1 shows the arrangement. The
pure mineral samples were floated in a Partridge and Smith type cell
(1971) by sparging ultra-pure nitrogen through a medium porosity glass
frit at the base of the cell. Gas flow was controlled and measured
using a Gilmont micrometer capillary flowmeter. The particle suspension
was gently stirred using a Teflon coated magnetic stir bar and a Sybron
Nuova II magnetic stirrer.
Batch Flotation Apparatus
The automated laboratory flotation cell developed by Luttrell and
Yoon (1983) was used for all batch flotation experiments. A diagram is
included as figure 2. This apparatus required no modifications for this
work.
X-ray Photoelectron Spectroscopy
The x—ray photoelectron spectroscopy (XPS) was done using an XSAM
model 600 spectrophotometer manufactured by Kratos. The resolution
limit of this machine was 0.46 eV. Additional information regarding
this equipment may be obtained by contacting the manufacturer or the
Polymer Lab, Department of Materials Engineering, Virginia Tech.
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18
PROCEDURE
Measurement of Operating Variables
Potential -—Potentials were measured by immersing a bright
platinum-saturated calomel electrode pair into the suspension and
reading a millivolt value from the voltmeter. This value was converted
to the hydrogen scale by assuming the saturated calomel electrode had a
potential of +O.242 volts, as determined by Bates (1964). This
so—ca1led "potential" is not intended to represent the reversible Nernst
potential (Heyes and Trahar, 1977; Gardner and Woods, 1979). However,
in general, positive potentials can be considered to be oxidizing and
negative potentials, reducing.
The performance of the electrode pair was regularly checked using a
ZoBell solution (Garrels and Christ, 1965). In addition, the surface of
the platinum was mechanically cleaned with fine emery paper to ensure
that any surface poisoning would be removed.
Dissolved Oxygen ——The dissolved oxygen level was monitored by
submerging the probe in the flotation pulp. The millivolt reading of
the voltmeter was converted to dissolved oxygen content by a
relationship where O.1 volts equals 1 ppm oxygen. The probe was
calibrated to atmospheric conditions and checked after each test.
Sulfide Ion --When sulfide ion concentration was measured, the
specific ion- reference electrode pair was placed into the pulp
19suspension.The meter was standardized to give the S level directly6
in ppm. This value was measured at the beginning of the experiment, at
the onset of conditioning, at the beginning of flotation, and at the
conclusion of the experiment in batch flotation tests.
pH -—The pH was measured in an identical manner as were sulfide ion
concentrations in batch flotation. In microflotation tests, pH was
measured coincident with potential during conditioning,
Microflotation
The microflotation tests were carried out on pure sphalerite using
the apparatus described earlier. In each test, 1.0 g of -65 +100 mesh
particles was placed in an 150-ml beaker with approximately 70 ml of
doubly distilled water. In the experiments conducted at constant pH,
the buffers described earlier were used in place of doubly distilled
water.
The potential measuring electrodes were placed in the suspension
and the potential was adjusted to the desired level by reagent
additions. 0nce the potential steadied at this level, the suspension
was conditioned for 5 minutes. Except for those tests in which buffer
solutions were used, no attempt was made to adjust the pH from the
"natural pH" that resulted after reagent additions. After conditioning, 1
the mixture was transferred to the flotation cell and flotation
~immediately begun. A one-minute flotation time was used. Forty ml/min
ultra-pure nitrogen was passed through a glass frit for bubble
II
20 :I
generation.
For tests where the sphalerite was desired to be copper-activated,
the mineral sample was initially placed in a 250-ml Erlenmeyer flask
with 50 ml CuS04 solution. This mixture was conditioned for 15 minutesutilizing a wrist—action shaker. After this time, the CuSO4 was
decanted off and the mineral was rinsed once with doubly—distilled H20•
Then the sample was transferred to an 150-ml beaker, conditioned and
floated as above. After flotation, both the floating and non—floating
fractions were filtered, dried and weighed.
Batch Flotation
For batch flotation, approximately 500 g samples of the Elmwood
ore, prepared as described earlier, were ground in a 10—inch ceramic
ball mill with a 50% by volume (5 kg) charge of steel balls and 300 ml
tap water. This pulp was transferred to the 2-liter plexiglas cell of
the automated flotation machine. The necessary reagents were added,
conditions adjusted, and conditioning begun. One minute prior to the
starting of flotation, 1.5 ml of a 1% DF 250 solution (0.06 lb/ton) was
added to the pulp.
At the appropriate time, the paddles of the machine were activated
and the froth product collected for three minutes. This productwasmixed
with tap water to make up the volume of the cell, and floated to
exhaustion as a cleaner stage. I
The various pulp parameters were continuously monitored during I
conditioning and subsequent rougher flotation. In some tests, the level
IIIIII
21
of dissolved oxygen in the pulp was used to determine the starting point
for flotation. In most tests, the pulp was conditioned six minutes
after the addition of Na2S and/or CuS04 dosages.For the series of tests which were designed to verify the ratio of
Cu2+ to S2- necessary for good flotation, the procedure described by
Luttrell (1982), was used. The sodium sulfide was added and the pulp
conditioned for 10 minutes. The copper sulfate was then added and the
pulp conditioned an additional 5 minutes before floating to exhaustion.
In this series of tests, the pH was controlled at 6.75 j 0.1 by
adding HC1 as required. However, in the majority of batch tests, the pH
was monitored but not controlled . A summary of reagent additions, pH,
and other pertinent information for each test is included in Appendix A.
X—ray Photoelectron Spectroscopy
The samples for XPS analysis were prepared as for microflotation.
More specifically, they were stirred in a beaker and reagents were added
to bring potential to the desired level. Copper—activated specimens
were in contact with 5 X 10-5 M CuS04for 15 minutes, then mixed with
doubly—distilled water prior to potential adjustment. The treatment of
these samples is summarized in table 2.
The samples were then dried in a vacuum dessicator and stored under
vacuum until just prior to running the spectra. The powdered sample was
then mounted on double-stick tape and placed on the probe for insertion
into the spectrophotometer. A wide—scan spectrum was then collected,
followed by narrow scans for elements of interest, such as sulfur.
22
Table 2. Summary of preparation conditions for samples for
XPS analysis.
Sample No. Weight Product Flotability Potential
1C 0.12g Concentrate 11.65 +476mV
2T 1.03 Tailing 0.00l
+829
3T 1.05 Tailing 0.00 -276
4T* 0.97 Tailing 4.90 +829
5C* 1.08 Concentrate 98.18 +517
6T* 1.04 Tailing 0.00 -458
*Copper—activated
EXPERIMENTAL RESULTS
MICROFLOTATION
The results of microflotation experiments were interpreted using
flotability as the dependent variable. The outcomes were examined to
determine the effect of potential, Cu2+ activation and pH. Flotability
was defined as the percentage of the original 1-g sample of pure
sphalerite collected in the head of the cell at the end of flotation. A
summary of all experiments is included in Appendix B.
Effect of Potential
It was shown in this work that the flotation behavior of sphalerite
is dependent on the potential of the solution prior to flotation. This
agrees with the results obtained by other researchers for other sulfide
mineral systems, such as Yoon and Luttrell (1984a) and Gardner and Woods
(1979) for the chalcopyrite system.
Figure 3 shows these results for unactivated Elmwood sphalerite.
This curve shows that flotation is negligible at potentials less than
+200 mV, versus the standard hydrogen electrode (mv, SHE), and reaches a
maximum of 35% at +432 mV. It drops off considerably at potentials
greater than +600 mV. These results indicate that under certain
potential conditions the surface of untreated sphalerite is at least
partially hydrophobic.
II
II
24 I
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Effect of Copper ActivationU
In view of the observations of Furstenau and Sabacky (1981) and
others, an attempt was made to improve the flotability of the sphalerite
by activating the surface with cupric sulfate (CuS04'5H20). As was
discussed in the procedure section, the mineral sample was treated with
various concentrations of CuS04 solutions to provide this activation.The samples thus treated were conditioned at numerous potentials and
then floated. The results in figure 4 show that the behavior of
activated sphalerite is also dependent on potential. The concentrations
of copper used in the activation step determined the exact flotation
characteristics of the sphalerite.
At a copper sulfate concentration of 5 X 10-5 M, the range and
amount of flotation was drastically improved from the non—activated
case. Recovery of 33% of the sphalerite was obtained at potentials as
low as 0 mV, SHE. The flotability was above 80% between +200 and +600
mV, reaching a maximum of 91% between +250 and +425 mV.
Increasing the copper sulfate concentration by one-hundred fold to
5 X10_3
M increased the range of good flotability greatly. More than
40% of the sphalerite floated at potentials down to -300 mV. The
flotability was greater than 80% in the potential range of -100 to +450
mV with a maximum of 97% around +150 mV. The recovery decreased rapidly
for potentials greater than +450 mV, becoming less than 2% at +800 mV.
Further increase in copper concentration did not improve the
flotation, however. At a copper sulfate concentration of 5 X 10-2 M,
flotation behavior became complex. After an initial rise in flotability
26
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27
to a maximum of 88% at +50 mV, the recovery fell to 30% at +275 mV
before rising to a second maximum of 67% at +535 mV. This bimodal
response indicates that excess copper ions gave rise to two distinct
flotation mechanisms, each having its own potential-dependent behavior.
Effect of pH
In the above tests, no attempt was made to control the solution pH.
To examine the effect of this parameter on the hydrophobicity of pure
sphalerite, microflotation experiments were carried out in pH 4, 7, and
10 buffer solutions. Both activated and non-activated series were
conducted. The results of these experiments are shown in figure 5.
In both the activated and non-activated cases, the best results
were obtained at pH A. The non-activated sphalerite had a maximum
flotability of 67% at approximately +200 mV at pH A, compared with a 20%
maximum for pH 7 and 7% for pH 10. Good flotability was exhibited for
all potentials greater than 0 mV for pHA, with 25% of the sphalerite
still floating at +1000 mV.
In the activated case, the flotability demonstrated even more
variation with pH. At pH A, the recovery was greater than 80% for all
potentials between -100 and +900 mV, approaching 100% at potentials
greater than 0 mV. Those tests conducted at pH 7 also showed
flotabilities greater than 90% fo potentials between 0 and +700 mV. The
recovery dropped drastically at potentials less than 0 mV and greater
than +700 mV.
The experiments conducted at pH 10 showed a much narrower range of
28
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l‘ · •0
Mi
Q CDG.
_I z
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, '‘ 88
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29
lgood flotability. They exhibited a slow increase in recovery from about
7% at -350 mV to a maximum of about 92% at +350 mV. The flotability
decreased rapidly for potentials above the maximum point.
The limit for reducing potentials obtainable in pH A and 7 buffers
was due to limitations in the chemical system. It was found that
addition of Na2S, in an effort to reduce the potential below the minimum
values shown in figure 5, resulted in a marked increase in pH.
IIsoI
BATCH FLOTATION
Batch flotation experiments were conducted on 500-g lots of Elmwood
mine sphalerite ore. Many pulp parameters, e.g. potential, dissolved
oxygen, pH, and sulfide ion concentration, were monitored. Further, the
effect of various reagent additions were examined. A complete summary
of all batch experiments is in Appendix A.
Effect of Sodium Sulfide Dosage
Figure 6 illustrates the influence of Na2S and CuS04 addition onflotation recovery of zinc from the Elmwood ore. It can be seen that
any addition of Na2S at CuS04 levels of less than 1 lb/ton, depressedthe sphalerite flotation compared to tests where no sodium sulfide was
added. At cupric sulfate dosages greater than 1 lb/ton however, the
recovery increased if 0.25 lb/ton Na2S were added. The grade of the
flotation concentrate was high for all of these tests. Increasing the
Na S dosage to 0.50 lb/ton brought a decrease in flotability.
Additional tests conducted at a sodium sulfide addition of 1.5 lb/ton
produced zinc recoveries of less than 2.5% at CuS04 dosages even up to
3.0 lb/ton, demonstrating that too much sodium sulfide is detrimental to
flotation.
Effect of Cupric Sulfate Dosage
The effect of cupric sulfate activation on sphalerite recovery is
also shown in figure 6. Flotation is enhanced greatly from nil at 0.0
lb/ton addition, to a maximum of 66 to 85% at dosages between 0.5 and
31 (
100
A Ö ¢.80 “If__
I60 ^>—11gl sNLU
. tonAO El Ocä Ib
_——————————————-———————_—————————_”T777777777777777ET'———————————_———————““““’““““““’“"’“1
32
2.0 lb/ton. This general behavior is observed for 0.0, 0.25, and 0.5
lb/ton additions of Na2S. In the case of 0.5 lb/ton sodium sulfide
addition, the recovery reaches a maximum at a 1.5 lb/ton dosage of CuS04»
then falls markedly at 2.0 lb/ton. ·
Effect of Cupric Ion/Sulfide Ion Ratio
As the results in figure 6 indicate, there is a relationship
between both copper sulfate and sodium sulfide additions and zinc
recovery in the Elmwood sphalerite system. This relationship is more
clearly demonstrated in figure 7. Using an identical technique as
Luttrell (1982), results showing the most favorable ratio of added
copper ions to added sulfide ions for optimum recovery were obtained.
The maximum recovery was obtained at atomic ratios of 0.31. The
recovery curves were identical for Na S additions of both 0.25 and 0.50
lb/ton. The grade curves showed a slight difference, however.
Effect of Dissolved Oxygen
To examine the importance of dissolved oxygen in collectorless
flotation, several experiments were conducted in which various methods
were used to bring the pulp oxygen concentration to a level of 9 ppm.
This level was chosen by examining the dissolved oxygen level at
flotation for a test in which good recovery was obtained for a
high—grade sphalerite product, as can be seen in figure 8. A summary of
the most significant of these tests conducted at 9 ppm is included in
table 3. The recoveries and grades for cleaner flotation show little
u
ä100 __EEzä_
Ü 0.25lb/tonA O.5 lb/ton
B0A
E60 60Q: 1 -— .2* XE IK1.1.1
40 E:LI.JE — „ 5
. <20 20 11am
O 0.25 0.50 0.75 0002* /52* 1=111T10
Figure 7. Effect of Cu2+/S2- ratio om recovery amdgrade of Elmwood sphalerite at pH 6.75.
11
. l
31+
_zg LuddO~
*¤+2asG)1 ·ärcs—+¤¤
LZtx ¤¤LEowc:PS
-+-+:>«4J$3“° **6 *’
LuE<r
E w.¤roI 2s:U
N E3+ä
OLp.} •c>:>~+
+-Z3 88•—
(DCDn1“‘“
0 0 0 D °°0 0 ä
O O an~o <r ~ Q} g
I II(/\‘“)‘·I3
I I Igw
O~ äz Y\ ~Q L;)Lu
Qwdd
cw O oo ~O ·<r CO
35
Table 3. Recovery and grade for flotation experiments conductedusing 0.5 lb/ton Na2S and 1.0 lb/ton CuS0a underdifferent conditions.
Test Conditions Product Grade Recovery
11 Conditioned without air Rougher 37.46 75.66for 5 minutes Cleaner 54.09 69.20
Conditioned without air Rougher 38.37 73.7930 until dissolved oxygen Cleaner 57.83 67.96
reached 9 ppm (12 min.)
Conditioned with air Rougher 49.18 73.5631 until dissolved oxygen Cleaner 55.06 67.76
reached 9 ppm (3 min.) ”
Conditioned without air32 10 minutes, KMnO4 added Rougher 42.29 90.57
to bring dissolved Cleaner 54.77 68.46oxygen to 9 ppm. (0.07g1<M¤0A)
1
V36
V
variation.
In test 32, the pulp was conditioned for 10 minutes and then 0.07 g
of KMn0 added. This produced an immediate increase in dissolved
oxygen. The rougher concentrate in this test had a recovery of 90.57%
at a grade of 42.29% Zn. This greatly exceeded the rougher recoveries
obtained in the other tests and indicates the importance of dissolved
oxygen, which is important in determining the pulp potential.
Effect of Permanganate Addition
In the series of tests conducted to examine the effect of potassium
permanganate, the results indicate that there is an optimum level of
permanganate addition. At a level of 0.2 lb/ton, the grade of the
concentrate was 55.88% Zn at a recovery of 29.58%. When the dosage of
KMn0 was increased to 0.8 lb/ton, the grade decreased to 43.84% Zn, but
the recovery was increased to 39.74%. Further increases in permanganate
addition to 1.4 and 2.0 lb/ton gave very inferior results, with grades
and recoveries indicating sphalerite depression.
Effect of Potential
As figure 8 and those in appendix D show, the pulp potential was
monitored for all of the previously discussed batch flotation tests.
The flotation data can be correlated to potential as well as the factors
already examined by using this data. In general, the flotation of the
ore was similar to that of the pure mineral observed in the
microflotation tests. Specifically, the best flotation was obtained
37
when the pulp potential at the beginning of concentrate recovery was in
the range of 0 to +600 mV SHE. Pulp potentials outside of this range
greatly depressed the flotation of the sphalerite, although many other
factors also affected the flotation behavior in a given test.
X-RAY PHOTOELECTRON SPECTROSCOPY
X-ray photoelectron spectroscopic (XPS) analysis was conducted on
pure sphalerite samples after they had undergone microflotation testing.
This surface analysis was done to identify the species that might have
been responsible for flotation or the lack of flotation. In the cases
where good flotability was observed, the microflotation concentrate was
subjected to the analysis. In the other cases, the unfloated material
was used. A standard of pure sulfur was also examined.
The raw spectra were deconvoluted using software developed by
Kratos for use on their system. This provided detailed information
regarding the oxidation states of sulfur present on the surface.
Complete spectra and information for each sample is contained in
appendix C. An example of raw and curve resolved Sulfur 2p spectra for
sample 5C is included as figures 9 and 10.
Using accepted charge correction techniques, the binding energy of
each peak was determined. This involved determining the binding energy
of the carbon 1s peak and comparing it to the standard value of 284.9
eV. This difference was assumed to represent the shift in binding
energies of all other peaks. When this shift was used to adjust the4
peak location of the curve—resolved S—2p peaks, the peaks could be
g 138
“
6000
E: 4000U7ZuJI—·Z
2000
·168 163 158 153
BI NDING ENERGY (eV)Figure 9. Raw Sulfur—2p Spectra for Sample 5C of pure
I
Sphalerite.
II
39 III
IO0
A
IB**50ZI)
0I00
B
U7b- I°
‘
.1 •„• s‘.
'I
. • léxI, ‘· \4/ • °·< X ‘~ 'O _ - •,•° • 0 •'(‘ ‘~
I65 I63 I6] I5?BINDING ENERGY (eV)
Figure 10. Su1fur—2p spectra fer sample 5C ofpure sphalerite. (A)Curve smoothed(B) Peak reselved. ····g¤ ____S 2“
-___ S2-. I X I
40
compared to literature values for identification of the species present.
Table A gives those peak identifications.
In addition, the ratio of peak areas of the different sulfur
species gives a measure of the relative amounts of each on that sample
surface. These relative abundances are included in table 5. As it can
be seen, the samples that were treated at high oxidizing potentials (2T
and LT) have a lot of oxidized surface species such as SO and other more
oxidized hydrophillic species, such as 32032- and 3042-, while samples
3T and 6T, which were treated at extremely reducing potentials, show a
higher percentage of sulfur species in lower oxidation states. In
addition, sample 50, which alone showed good flotabilty, was the only
sample to show the presence of 3- or SXZ- on the surface. This seems to
indicate a strong correlation of this species to flotability.
2 41
Table 4. Identification of surface sulfur species present onElmwood sphalerite after various treatments asdetermined by XPS analysis.
Sample Corrected Peaks Identification ReferenceS—2p1/2 S-2p3/2
Sulfur 163.3 162.1 S° 1,2
5C 163.3 162.1 S°2_ 1161.6 160.4 SX160Ä4 159Ä2 S2’(w/ Cu) 3
1C 163.5 162.3 S°162.4 161.2 S2-(w/ Zn) ä
2-2T 164.2 163.0 SO3 5163.3 162.1 S° _162.0 160.8
S2—(w/Zn)
2-4T 164.7 163.5 SO3 ·163.3 162.1 S°160.9 159.7
S2_(w/Cu)
3T 163.4 162.2 S°162.4 161.2 S2-
6T 163.5 162.3 S°162.4 161.2 S2-
References1. Luttrell (1982) 3. Langer, Helmer, and Weichert (1969)2. Nordling (1972) 4. Vesely and Langer (1971)
5. Hercules (1970)
IIII
42 IIII
Table 5. Relative surface abundance of various Sulfur Speciespresent on Elmwood sphalerite after microflotationas determined by XPS analysis.
Abundance of SpeciesCopper (Z of Total S on Surface)
Sample Activ. S2- S 2- So SO 2- Pot. Flot.X 3
5C yes 38.52 49.53 11.95 ----— +517 98.18
1C no 74.81 --—-- 25.19 -—--- +476 11.65
2T no 18.10 -—-—- 54.60 27.30 +829 0.00
_ 4T yes 4.56 ----- 67.92 27.52 +829 4.90
3T no 71.61 --—-- 28.39 ----— -276 0.00
6T yes 72.14 --—-- 27.86 ----- -458 0.00
DISCUSSION
FLOTATION
The results of microflotation experiments presented in previous
sections demonstrate that pure sphalerite can be floated without use of
collector, primarily at potentials above 0 mV, SHE. In order for this
flotation behaviour to be significant, however, the sphalerite must be
copper-activated. This activation, which can be represented by:
ZnS(s) + Cu2+(aq) --§> cus(S) + Zn2+(aq)
(Sutherland and Wark, 1955), changes the potential range where
collectorless flotation is possible as well. The critical potential for
flotation onset can be lowered to below 0 mV by increasing the cupric
sulfate dosage added.
The pH is also a factor in determining the potential range where
collectorless flotation is possible (see figure 5). In addition, the pH
greatly influences the maximum flotability obtained. A wider range of
good flotability was obtained at pH A where nearly 100% flotability was
observed at potentials between -100 and +900 mV for activated
sphalerite. The potential range for maximum flotation at pH 10, on the
other hand, was extremely narrow, with the point of maximum recovery of
about 92% occuring at +350 mV. This behavior seems to agree with
Luttrell and Yoon's (1984a) reasoning that pH may be important in
forming hydrophobic surface species such as elemental sulfur or
polysulfides, and is consistent with the findings of Clifford, Purdy and V
Miiier (1974). i4 3
44
Similar trends in behavior were noted for batch flotation of
Elmwood sphalerite ore. The ore also required copper—activation to
float without collector. The ore also required the addition of Na2S as
well to obtain maximum recovery and grade. This upholds observations by
Yoon (1981) and Luttrell and Yoon (1984).
The effect of Cu2+/S2' ratio was shown in figure 7. The ratio,
previously noted by Luttrell (1982) as producing the best results at a
value of 0.17, was found in this case to have a different point yielding
optimum recovery, 0.31. This difference may be attributed to the
difference in the characteristics of the ores used in these two studies.
It is important to note that just as Luttrell found, the flotation
behavior was independent of reagent dosages if this ratio were
maintained. The findings of this work generally concur with those of
Luttrell‘s work, and thus, his argument relating the cupric ion/sulfide
ion ratio to the semi—conducting properties and the resulting
flotability of the sphalerite surface is reinforced. That is to say
that, because of the insulating nature of sphalerite (band gap 3.67
eV)(Teichman, 1964), copper-activation may be necessary to create a
semi—conducting surface which can become involved in electrochemical
oxidation reactions. Ralston, Alabaster, and Healy‘s (1980) work
indicating that surface formation of elemental sulfur is closely related
to the semi—conducting properties of the mineral seems to support this
contention.
The results of the present study also support the findings of
Gaudin (1932) indicating that in general sodium sulfide and potassium
45
permanganate act as depressing agents in flotation. The results
presented here provide further support for the evidence presented by
others that this is due to the effect these reagents have on pulp
potential.l
V46
MECHANISMS
Induced Hydrophobicity by Oxidation
Elemental Sulfur —-Metal sulfides have long been thought to be
thermodynamically unstable in the presence of oxygen. Because of this
instability, their surfaces are readily oxidized under the prevailing
conditions of flotation circuits. Sphalerite behaves slightly
differently since, unlike most metal sulfides, it is an insulator rather
than a semi-conductor. This electronic property is changed, however,
when the surface of the ZnS is copper—activated. Because it is believed
that the copper ions replace the zinc in the sphalerite lattice during
activation, the surface essentially becomes CuS. The surface, which
would now be semi—conducting, would be more easily oxdized to form
elemental sulfur.
The XPS spectra show the presence of SO on all samples, regardless
of the treatment conditions used. The amount of elemental sulfur on the
surface (based on peak intensities and distribution of each species)
varied from sample to sample. The samples treated at high oxidizing
potentials showed the most sulfur as expected. Unexpectedly, however,
the samples treated under reducing potentials had the next highest
amounts. This was probably due to the oxidation of sulfide ions present
in a film of sodium sulfide solution adhering to the mineral surface
during the drying of the sample. The samples which exhibited the best
flotability appeared to have the least amount of SP on the surface.
The possibility that some sample oxidation took place during the
drying prior to XPS analysis, raises doubts about the validity of the
47 I
interpretation of the data to explain flotation behavior. The samples
which floated well without Na2S added may provide more reliable
information. Further, the presence and relative amounts of other
sulfur-containing species agrees well with expected behavior at the
treatment potentials.
The apparent lack of correlation between elemental sulfur on the
surface and flotation behavior would seen to negate the importance of
elemental sulfur as a flotation inducer. However, since the large
quantity of SO on the surface of the reduced samples (3T and 6T) can be
explained by the oxidation of extraneous S2- ions upon drying, and the
lack of flotability of the oxidized samples (2T and 4T) by the presence
of higher—oxidation-state, strongly hydrophillic sulfur species such as52032-, the assertion that the observed flotability of samples 5C and 1C
is due to the presence of elemental sulfur, is a valid one. It is also
consistent with the proposal of Gardner and Woods (1979) that elemental
sulfur is responsible for collectorless flotation in the chalcopyrite
system. Recent work by Heyes and Trahar (1984) has demonstrated that
the collectorless flotation of pyrite and pyrrhotite is related to the
production of elemental sulfur on the surface, lending further support
to this proposal.
In addition, the results indicating that sphalerite floats better
at acidic pH's is consistent with this theory. Many researchers have
documented that the oxidation of sulfide minerals in acidic solutions
results in the formation of elemental sulfur (Vizsolyi, Veltman, and
Forward, 1963; Sato, 1966; Majima and Peters, 1966; Eadington and
I48 I
Prosser, 1969; Bjorling, 1973). Thermodynamics also favors the
increased stability of this elemental sulfur formed at pH's less.than
7.5 (Garrels and Christ, 1965).
Polysulfides or Metal—Deficient Sulfides —-As it can be seen in
figure 10 and tables 3 and A, the surface of sample 5C has a
considerable amount of a sulfur species with a binding energy and
oxidation state intermediate to S2- and SO. This species is best
identifiable as a polysufide. Polysulfides can form as a result of the
interaction of sulfur with an aqueous solution of sulfide (Chen and
Morris, 1972) or by the aging or oxidation of sulfides or hydrosulfides
in solution (Karchmer, 1970). Chen and Gupta (1973) have also
demonstrated that when sulfur is produced in the presence of sulfide,
polysulfides are immediately formed in some cases.
The pH of a system is of extreme importance in determining the
stability of polysulfides. Chen and Morris (1972) and Chen and Gupta
(1973) discovered that the concentration of polysulfides in a pH 8
solution is several times greater than that of elemental sulfur. The
situation is reversed at pH 6 however. The concentration of
polysulfides should increase with increasing alkalinity, according to
the mass balance calculations done by Chen and Morris.It has been demonstrated (Allen and Hickling, 1957) that ‘
polysulfides can adsorb on a metal surface through the following IsX2’ + M ---> M===sXwhere
SX2—represents the polysulfide and M the metal. Chen and Morris II
II
49I
(1972) showed evidence that the oxygenation of mildly alkaline sulfide
solutions is 1OO times greater in the presence of transition metals. It
is very possible that transition metal sulfides could behave in much the
same way. This metal—polysulfide complex would probably be hydrophobic
since the bonding of the sulfur atoms in the polysulfide chain is very
much like that in elemental sulfur.
Thus polysulfides may also play important roles in the oxidation
and flotation of copper—activated sphalerite as well as other sulfides.
They may also be important components of electrochemical reaction
systems involving sulfide minerals. Similar electrochemical behavior
has been suggested for both production of elemental sulfur and
polysulfides on galena surfaces (Ho and Conway, 1978). This would seem
to indicate that it is difficult to differentiate between the two
processes electrochemically.
The sample 5C was prepared in distilled water with a pH presumably
of around 7. This corresponds to the optimum pH for formation of
polysulfides from a sulfide solution (Chen and Morris, 1972). Since
sphalerite is one of the most soluble sulfides, in the absence of
oxygen, (Yoon, 1981) it is concievable that enough S2- ions are present
to form these polysulfides.
Additional support is provided by the XPS data presented earlier
(Table 3) which shows a peak at approximately 161.6 eV. This
corresponds roughly to the peak assigned previously by Buckley and Woods
(1981) and Luttrell (1982) as being Cu2S. But as Luttrell points out,
this peak could also be considered as an indication of the surface
50 'n
presence of metal polysulfides. He suggests these may form by the )
reaction2Cu+ + SX2' ----9 Cu-SX-Cu
where CuY represents the copper ions in the surface lattice and x
represents the number of sulfur atoms in the polysulfide chain (usually
2-5). This is entirely possible since the formation of polysulfides
from the CuS present on the surface would probably produce Cu+ ions on
the surface by the reaction
XS2'+YCu2+ ----9 SX2”+YCu++(X-Y)e·
where X and Y are probably equal to provide charge balance. Since the
oxidation state of the sulfur in the middle of the polysulfide complex
is nearly O and that on the end sulfur atoms is approximately -1, the
XPS spectra may show these two peaks (Luttrell, 1982). The peaks at
161.6 and 163.3 eV for sample 5C correspond very closely to those
indicated by Luttrell.
More recently, Hamilton and Woods (1983, 1984) and Buckley and
Woods (1984) have attributed these intermediate peaks to the presence of
a metal-deficient sulfide which they consider important to collectorless
flotation. Their XPS data shows that the copper contained in this
surface compound is present as copper (I), implying a formal oxidation
state for the sulfur of -1/2, the same as is found in the polysulfide,
E22“_Fromthis data, an electronic structure of Cu2S4 was theorized for
the metal-deficient sulfide found on the chalcopyrite. They also i
assumed that all of the sulfur atoms would have the -1/2 formal valence
I
51I
state.
Other recent work (Perry, Tsao, and Taylor, 1984) suggests that the
surface compound formed on copper-activated sphalerite is a copper (I)
sulfide based on XPS and Auger parameter data. This compound, though
identified as possibly being chalcocite (Cu2S) could actually be a
metal-deficient sulfide. Thus, the copper—activation of sphalerite at
appropriate oxidizing potentials could produce a metal-deficient sulfide
which would cause flotation behavior such as was found for sample SC.
Since, in general, polysulfide and metal-deficient sulfide
formation is enhanced at higher pH, it is possible that these species
are responsible for collectorless flotation in alkaline or nearly
neutral solutions. The flotation behavior observed in acidic solutions
may be caused by elemental sulfur, on the other hand.
III
521Inherent Hydrophobicity
The results of the microflotation tests conducted on pure and
copper-activated sphalerite indicate that flotability decreases with
increasing additions of sodium sulfide and thus with decreasing
potential. To better understand the role that sodium sulfide additionplays in this system, it is important to consider the proportion of
total sulfide present as S2-, HS-, or HZS in aqueous solutions as a
function of pH. Jones and Woodcock (1978) demonstrated that HS- ions
make up the greatest portion of the total sulfide at pH values between 7and 13. Therefore, it might be considered that hydrosulfide ions are
depressing the natural flotation of sphalerite and copper activated
sphalerite under the reducing conditions.
HS- ions have long been identified as depressants in xanthate ·flotation because of the competition of the hydrosulfide and xanthate
ions for the mineral surface (Gaudin, 1957). Gardner and Woods (1979)
and Trahar (1983) olaimed that HS- ions produce a reducing environment
that prohibits the formation of elemental sulfur in collectorless
flotation. Luttrell (1982) suggests alternatively that the adsorption
of hydrosulfide ions on the mineral surface might render it hydrophilic.
The S—H group can act as a proton donor and thus easily form hydrogen
bonds (Vinogradov and Linnel, 1971) and, if the water molecules
surrounding the mineral are proton acceptors, the adsorbed HS"would
give the mineral an hydrophilic character.
This argument implies that elemental sulfur may not be required forcollectorless flotation. Failure to find a strong correlation between
53
elemental sulfur present on the mineral surface and flotability has been
the case in much previous work (Finklestein et al, 1975; Heyes and
Trahar, 1977; Pritzker, Yoon, and Dwight, 1980; Furstenau and Sabacky,
1981; and Luttrell, 1982). As was previously discussed, the present
results cannot be easily interpreted using the elemental sulfur theory
alone.
It has previously been indicated that the flotation response for
both activated and non-activated sphalerite is best in acidic pH. If HS
is responsible for hydrophilic depression of the mineral, this improved
flotability may be due to the removal of HS° from the system as HZS gas.
0f course, acidic conditions also favor the formation of elemental
sulfur.
Under oxidizing conditions, the mineral may be depressed in a
different way. At high oxidizing potentials, XPS shows large amounts of
elemental sulfur, yet neither of these samples was flotable. This lack
of flotability can be explained; at high oxidizing potentials,
hydrophilic species such as CuO, CuS203, and Mn02 are readily
formed as shown by XPS spectra. These species adversely affect the
hydrophobic nature of the sphalerite surface.
The assertion that some species (e.g. HS-, CuS04) is rendering the
mineral surface hydrophilic is based on the assumption that clean
sulfides are inherently hydrophobic. The increased flotability of
copper—activated sphalerite could be explained on the basis that the CuS
formed is less soluble than the ZnS (Yoon, 1981). The inherent
flotability of sulfides is due to the inability of sulfide ions to form
54
H-bonds with the surrounding water molecules (Finkelstein et al, 1975;
Furstenau and Sabacky, 1981). The weakness of this natural flotability
theory is that sulfide minerals are extremely succeptible to oxidation.
At even 1 x 10'lo M levels of dissolved oxygen, oxidation is likely to
occur (Gaudin, 1972).
The present study has verified that collectorless flotation of
sphalerite is related to the oxidation mechanism of the sphalerite or
the CuS on the surface. Since oxidation of sulfide proceeds in an
orderly progression fromS2- -9
SX2_-9 S- -9 SO -9 S2032- -9 3032- -9 soaz',
the flotability of a mineral will be most influenced by the degree of
oxidation of the mineral at the time of flotation. The results of the
flotation experiments and the XPS experiments indicate that the chemical
enviromment can cause this oxidative progression to be limited or to
progress fully, producing the desired flotation behavior. In addition,
these results favor the conclusion that polysulfides or metal-deficient
sulfides are the species responsible for collectorless flotation of
sphalerite in nearly neutral and slightly alkaline solutions. Elemental
sulfur is probably responsible for flotation in acidic solutions,
however.
SUMMARY AND CONCLUSIONS
The results of the present investigation may be summarized as
follows:
1. Microflotation experiments conducted on non—activated and
copper-activated samples of pure sphalerite in the absence of collector
demonstrated that flotation was related to the potential of the chemical
system. In general, better flotability was obtained in the range of
potentials from 0 to +600 mV versus SHE.
2. The use of copper sulfate was essential in obtaining good
flotation in both the microflotation and batch flotation tests. The
concentration of the cupric sulfate used for activation was extremely
important in determining the exact potential range where good
flotability occurred. -
3. In batch flotation tests conducted with Elmwood mine sphalerite
ore, the use of both cupric sulfate and sodium sulfide was required to
obtain maximum grade and recovery in collectorless flotation practice.
Good flotability was obtained for a wide range of reagent additions as
long as the atomic ratio of Cu2+yS2_ was maintained at 0.31.
A. The atomic ratio of Cu2+yS2— required for optimum flotability
of the Elmwood sphalerite differed from that obtained by earlier
research on a different ore. This seems to indicate that the ratio may
be a function of each particular ore.
5. ·The sphalerite recovery in batch flotation tests was found to
improve in neutral to acidic solutions. The increased flotability was
55
W56
postulated to be caused by the increased stability of SO in acidic
solution or to the removal of hydrophillic HS- by protonization to HIP
gas.
6. The X—ray photoelectron spectroscopic analyses of the flotation
products indicated the possibility of polysulfide ions or metal-
deficient sulfides on the surface of a highly-flotable copper-activated
sample. The chemical reactions necessary to produce this species could
have taken place under the conditions of this test. This supports
Luttrell and Yoon's work (198Ab) as to the role of polysulfides in
collectorless flotation.
7. Sodium sulfide and potassium permanganate were found in general
to depress flotation. This may support the theory that sulfide surfaces
are inherently hydrophobic. The depressant effect of NaZS may be caused
by the adsorption of HS-_which may hydrogen bond to surrounding water
molecules. KMnOZ‘may cause the formation of hydrophillic oxidation
products, such as manganese dioxide, on the mineral surface.
8. Collectorless flotation of sphalerite may be viewed as being
the result of superficial oxidation of the mineral surface.
Insufficient or over oxidation can result in inferior flotation
behavior. The use of specific reagents, such as potassium permanganate
or sodium sulfide, can speed up or limit the oxidation.
III
INDUSTRIAL APPLICATION
The results of the present study have demonstrated that
collectorless flotation of sphalerite is possible using only a frother
after treatment with copper sulfate and sodium sulfide. Since the cost
of sodium sulfide is substantially less that that of collectors used in
industry today, the collectorless flotation of sphalerite may result in
savings in operation costs. In addition, the need for pH regulators
such as lime is eliminated in collectorless flotation.
The current economic crisis in the mining industry makes it
essential that cost cutting measures be implemented. With the price of
zinc at only 34.5 cents per pound, many mining companies have found it
impossible to make a reasonable profit and have ceased to operate. More
efficient and cost—effective methods of mining and beneficiation are
required to rejeuvenate the industry.
In addition, the results of this work underscore the importance of ·
the potential of the flotation pulp as a major variable in achieving
good flotation behavior, as has been pointed out by other investigators
(Luttrell and Yoon, 1984a; Heyes and Trahar, 1977; Walker, Stout, and
Richardson, 1982). Industrial utilization of potential electrodes in
flotation control may lead to more efficient metal processing and
recovery.:IIIII
I57 I I
II
REOOMMENDATIONS FOR FURTHER WORK
Based on the results of the present investigation, further study is
recommended in the following areas:
1. The physical and chemical significance of the cupric
ion/sulfide ion ratio should be investigated. Special attention should
be paid to the crystal structure and electronic properties of the
sphalerite that are important in determining this ratio.
2. Photo—conductivity and other semi—conducting properties of
copper-activated sphalerite should be investigated and quantified.
3. Additional XPS studies should be conducted to determine the
surface species present on sphalerite after treatment at various pH's
and cupric sulfate concentrations.
L. Research to determine the collectorless flotation
characteristics of covellite and chalcocite should be conducted and the
results compared to those obtained for copper-activated sphalerite.
5. Analysis of the surface of highly-flotable sphalerite should be
conducted whicle the mineral is still in the flotation pulp at the
prevailing test conditions using a technique such as Fourier Transform
Infrared Spectroscopy (FTIR). This would prevent the problems
associated with the possible oxidation of the mineral surface as a
result of the drying required prior to using analytical techniques such
as XPS.
58
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59
I60
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Clifford, R.K. and Miller, J.D., "Formation andDetection of Elemental Sulfur on the Surface ofSphalerite in Aqueous Sustems," AIME Annual
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Meeting, Feb. 24-28, Dallas, Texas (1974).
Clifford, R.K., Purdy, K.L. and Miller, J.D.,"Characterization of Sulfide Mineral Surfaces inFroth Flotation Systems Using ElectronSpectroscopy for Chemical Analysis," AIChE Symp.Series No. 150, Z4, 138-147 (1975).
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II I
I63 I
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IIIII
- .. _.1..1.1_.„..__...._...._..........„.................................I
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66
Walker, G.W., Stout III, J.V. and Richardson, P.E.,"Electrochemical Flotation of SulfideszReactions of Chalcocite in Aqueous Solution,"56th Colloid and Surf. Sci. Symp., Virginia Tech,June 13-16 (1982).
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APPENDIX A: BATCH FLOTATION CONDITIONS AND
METALLURGICAL BALANCE SHEETS
67
68
This appendix contains a summary of the conditions
under which each batch flotation test was conducted, xas well as the metallurgical balance sheets from these
tests. The charts of the continuously monitored opera-
ting parameters, such as potential, are contained in
appendix D. The procedure for determining the assays
reported in the metallurgical balance sheets is given
in appendix E.
69
I
II70 I
71
6Predort Ue1ont E E 21nc Un1t Zn D1:tr1ü0t1on
üonc„ 1.25.„ X1 . 65 25 . 7 ·]· 222 .. 2-5 1 , 415
16111no 50.07 15.43 135;.75 54.25Pemd 100.00 16.45 1645.06 10U„00
1eet 7
Product Heiont E E Zinc Unit In Distribution
Lonc. 1.72 17.13 25.46 1.7651.1611 10.65 14.35 153.26 5.13Ta111n¤ 57.60 17.07 1455.33 5;.11Feed 100.00 16.7% 1675.05 100.00
Teet 5
Product Me1ont Z E 21nc Unit Zn D1stribution
Conc. 1.02 5.54 10.04 0.55C1.T611 4.64 10.57 45.04 2.5516111nu 54.34 17.36 1637.74 56.52Feed 100.00 16.57 1656.52 100.00
Test
t L~1et·1 €.’11"'11l T. 1 n c Un 1 t. n C11 tri b u 1; 1 on
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15111no 52.55 14.55 1357.07 54.4;
Feed 100.00 14.65 1467.57 100.00
72
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Luna. Ü.:1 7.7; 4.75 0.3ÜÜ}„1&11 $.5E ?.qS $5.73
1 .1 1 4 $15 . 51311
L 1 1
Fvmduct M210nL E Z 21nc Unzt in D15tF1ÜUt1üM
CGN:.E.71 :.44
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14
Fvmüuct N&1¤hL E Z 21n: Un1L Zn Ü1EÄF1ÜUtlÜÜ
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73 ·1
Test 15
Product Meiont E L Ein: Unit Zn Distribution
ton:. 1.15 53.34 :1.83 5.0735.%} @7.39 5.0w
13111no @:.12 11.04 10:1.57 3:.83
lest 15
Produc1 Meioht E Z lin; Unit Zn Distribution ·
üonc. 21.27 53.13 1134.77 34.38C1.Tai1 3.14 2.13 5.83 0.48Teilind 75.53 2.83 214.41 15.14FEeE·d 1 D1} .. 110 1 #1 . 1 é, 1 /1 1 é, . Ü:) 1 @..10 . €Qi·CT·
Test 17
Product Meiont E E Exnc Unit Zn Distribution
Eon:. 4.54 47.11 213.38 15.44E1.1e11 4.37 13.30 ?ü.@5 :.57Teilino 30.43 11.34 1080.45 77.88Feed 1w0.ü0 13.35 1385.28 100.00
—1”E·E1; Z13
Product Meiont E ä Einc Unit Zn Distribution
Kon:. 5.58 43.33 234.28 20.32C1.Tai1 :.23 12.33 77.*4 5.52Ta111nm 3?.32 11.85 1041.00 73.:5Feed 105.üu 14.13 1413.23 100.mü
74
Teet 1%
P Hmhack. 14e1kv1t Z; 3.
1.3% 0.3e3.31 10.35 35.%0 2.22
le111no %e.43 1;.37 1573.45 %?.7eFeed 100.00 15.14 1e14.71 100.00
Teet 40
1DVCuUJ§3 lnexcnwt 3. T1 Znwwc Urd
tLonc.0.7e %.e3 7.32 0.50C1.Tei1 5.17 11.20 57.%3 3.%%
Teiiino %4.07 14.78 1388.07 %5.51Feed 100.00 14.53 1453.32 100.00
Teak 21
Product NEIÜÜÜ E E Z1n: Unit Zn D1atr1b0t1on
Con:. 15.17 55.51 1008.82 71.23C1.Tei} 5.88 14.22 125.00 5.%0
1ei1ino 72.%7 3.85 281.27 1%.87Feed 100.00 14.15 1415.5% 100.00
Teet 22
Product Me1onk.3. T1 iinc Unit Zr« Diatributfon
Eon:. 1%.17 4%.77 %54.13 75.757.4% 21.5e 1:1.50 12.82
1e111nd 73.34 1.85 143.%% 11.43
Feed 100.00 12.50 125%.el 100.00
I
7 5
ieat 23
1%*0601:1- bkaicdwt E T; Zirwc Urrit. KV1 Di.;tr‘1tu1t1.¤s.
Luna . 1. . .1 5 C1 . 1 1 1. 1 . 50 *33 . .:.12
0.1* 51.31 3.57F &·6e<;1 1 00 . 00 1 . . 21 1 00 . ·.Q>·..¤
-1Prmduct Heiont Z 2 Zinc Unit in 0i;tr1but1on
C].Tai1 10.07 15.2* 153.91 *.37Teilino 32.1* 12.1% 1003.*3 13.1*Feed 100.00 15.71 1571.2e 100.00
Test 21
Product Meiont 2 E Zinc Unit Zn Üiatrimution
Eon:. 20.52 öl.35 1220.5* 61.00C1.Tai1 1.25 27.&1 3%.76 2.33Tailind 77.Eü 2.23 173.éT 11.éTFeed 100.00 1%5*.02 100.00
Teat 27
Product Meiont E L Ein: Unit Zn Diatribution
Conc. 20.15 55.©5 1121.35 E5.15
1.** 155.5% 11.51Feed 100.00 13.13 1311.*3 100.00
1
76
1 e- e 1 .~:1¥E1
Product Meiont L 1 Zinc Unit in Dxakributeon
01.1e11 1.78 28.51 51.03 3.15
Te111nd 75.02 5.37 418.87 25.87
Feed 100.00 1:.20 1:1%.81 100.00
Teat EP
Prcwhnüt Meicüwt 7.
T.Con:.18.:7 53.70 1002.58 :;.34
C1.Te11 5.83 3:.83 214.:: 13.35
Ta111nd 75.50 5.18 3%1.0% 24.31
Feed 100.00 1:.08 1:08.33 100.00
Teak 30
Product WEIÜÜÜ E E Ein: UNIÄ En Diatribution
tion .. 1. : . 87 57 . 1 . 87 . 8*:
C1„Te11 10.80 7.7% 84.13 5.83
Te111nd 72.23 5.24 378.4% 2:.48
Feed 100.00 14.44 1444.00 100.00
Teat 31
Product weicht E E Zine Unit En üzatribution
Eon:. 18.0: 55.0: :7.7:
_ C1.Te11 3.8% 21.88 85.11 5.80
T8t11ÜU 78.05 4.87 387.82 2:.44
Feed 100.00 14.:7 14:7.41 100.00
77
1e;1 3B
Unit ir¤ D1etr1but10n
Län:. 15.54 54.77 1015.43 :5.4:51.1611 13.23 24.75 327.57 22.114A€kE 1 1 TIC} c>E . I3 E- L2;. üvié 1 IE5 . i3‘.‘·‘?·.·1Äß
Feed 10ü.00 14.53 1453.27 100.00
Tee 1: 1'L'1 +1
Fnoduct·‘‘‘ Nülüüf L E lan: Unit In Dietrimutiün
Bon:. 5.74 :1.51 355.3: 24.32C1.1e11 7.11 45.53 352.1: 24.101ai1ind 57.15 5.:5 753.55 51.55F eed 1 00 .. 00 1 4 . :1 1 4 :1 . 37 Z1 00 . 00
Teet 515
Product Me1ont E L Zine Unit Zn ÜÄEÜVIÜUÄICÜ
Con:. 7.44 :3.05 4:5.35 27.41C1.1ei1 4.44 45.37 201.40 11.7:Teilino 55.12 11.52 1041.55 :0.53Feed 100.00 17.12 1712.34 100.00
Teet R15
Product Meidht L E Zinc Un1t Zn Ü1et:1but1¤n·‘‘‘
Con:. 1:.20 :1.57 1075.22 72.52C1.Ta1] 15.53 12.43 230.33 1:.33Tailind :4.77 2.34 151.5: 10.75Feed 100.00 14.10 1410.11 100.00
7813Et H11
Fw"ncu»nl. lunicnat 7; T; 2111;. Lhwlt ir: KH atv 1b1=t1cne
unna. 15.5: 30.24 555.41 53.2551.1311 27.Cm= 5.358 231.13l
E E7.]?
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7 Q? 1-
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Len;. 3.13 33.37 413.47 23.3%C].Ta1}1nn 10.5d 33.52 353.53 23.57Tallinn 51.33 10.22 531.15 51.04Feed 100.00 13-23 1323.43 133.33
1 1; F;213
?Vndnnt M@lnnt L E Zinn Un1t Zn Diatrlüntlnn
Lnnn. 7.35 55.17 427.55 27.34Cl.7ai] 5.24 32.51 170.55 11.05
Tallinn 57.41 10.55 ?45.40 ö1.31Füäd 100.00 15.47 1543.53 100.00
Täät R25
Prndnnt Hülnnt 1 E Ein: Unlt Zn Dlatrlnntlnn
inn:. 22.50 51.14 1171.11 73.33
C1.Tel1 3.23 27.33 223.23 14.13Tglllnn 35.55 2.52 154.15 12.21
1* 100 . 00 1 1 1550 . 50 1 00 . 1:0
79Ä
21nc Unit Zn Üietvieutien
Lena. 17.34 52.40 #05.52 el.01
ü1.1e11 2'°'·°.47 5.55 237.52 15.iE
Te11wnd 55.15 5.75 315.00 21.77
Feed 100.00 14.55 1435.24 100.00
Tee? F1
PVCEFAC+ wexcürt E UL Zim: LWn‘t Zn 1;1etritn1t1¤n
Lena. 15.5Ü.'·· 43.54 725.43 35.74
C1.Te11 5.04 10.54 53.53 Z.?3
1e111nm 75.35 13.37 1045.07 57.33
Feed 100.00 15.25 1525.13 100.00
Teak PE
Fnedmct·‘‘‘ weicht E E Eine Unit Zn Dietributien
ten;. 4.25 17.54 75.50 5.15
Te11irm1 55.72 14.5f1 1357.54 54.54
Feed 100.00 14.53 1453.44 100.00
Teat F3
Pveduct weicht R E Ein: Unit En D1etFib0t1¤m
Gene. 10.50 44.22 454.32 30.27
Teilind 55.45 11.55 1055.41 55.73
Feed 100.00 15.34 1533.73 100.00
Teet P4
Fvmüwct wejmht E E Ein: Unnt In Ül£tV1DUÄ1üü
Umm:. Z,75 11.11 30.55 2.EZ
Teiline 57.22 14.01 1352.05 57.75
Feed 100.00 13.53 1352.54 100.00
II
IAPPENDIX B: MICROFLOTATION CONDITIONSIIIIII
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APPENDIX C: XPS SPECTRA AND CALCULATIONS
III
Since most specimens exhibited complex carbon-
ls peaks, the assignment of peak location for
calculation of sample charging was not trivial.
The assignment of particular peak locations was based
on an analysis of the geometry of the carbon-ls peaks,
an assumption that sample charging would be relatively
consistent between samples, and a practical examination y
of the identity of the various sulfur species indicated
using different charge corection factors. Peak
location and the differnece in binding energy between
peaks in the Sulfur-2p region as given in the liter-
ature were used to correlate the data obtained in the
present work and to define the charge correction
factor for each sample. The assignments of the binding
energy for the carbon-ls peak thus calculated result
in reasonable identities for the sulfur species found
on each sample.
89
90
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Samble Carbon—1e Peak Charce Correctxch
Sultwr etc. 287.7 ev 2.5 ev1C 291.P 7.02T 291.1 6.2’ 3T 292.0 7.14T 290.8 ° 5.5SC 291.3 · 6.46T 290.0 5.1
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APPENDIX E: ASSAY PROCEDURE
183
ASSAY PROCEDURE
An amount of flotation product, between one-
half and two grams, depending on the expected con-
centration, was accurately weighed into a 125-ml
Erlenmeyer flask. Twenty ml of aqua regia (50% HC1-
50% HNO3) was added and the flask placed on a hotplate
and heated to boiling. A reflux boiling funnel was
placed in the mouth of the flask to prevent excessive
splashing of the solution and to wash the sides of”
the flask. The sample was slowly boiled until the
solution had nearly all evaporated. The sides of the
flask were then rinsed with enough water to dissolve
any metal ions.
The samples, thus digested, were allowed to cool
and filtered through Whatman number 2 filter paper to
remove any undissolved solids. The filtrate was then
transferred to a volumetric flask and diluted with
doubly distilled water. Further dilution was often
necessary to bring the zinc concentration in the
unknown sample to within the range of the calibration
standards.N
The zinc content of each sample was then deter-
mlhéd using a Spectraspan IV Plasma Emission Spectro-
184
185
meter, set to detect zinc in the wavelength region “
around 2025.51 Angstroms. All glassware used in
these procedures was cleaned in Micro cleaning solution,
rinsed with tap water, soaked for a minimum of 2
hours in 20% HNO3, then rinsed with distilled anddoubly distilled water. This procedure was carried
out to ensure that no extraneous metal ions would
interfere with the analysis.
