T. Adel W. E. Fore ///(;j E gzén

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

vii

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

I

1 I

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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

3.II

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

7

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

9 .

10

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

13

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

III

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115

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

S6 °°Il

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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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gw

l‘ · •0

Mi

Q CDG.

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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![1; 0.25lb/ton

Q A 0.6 ab/mn20

.. .A Aus?• . • 1.0 2.0 3.0

Figure 6.. Effect of CuSO4 dosage on recovery ofElmwood sphalerite at constant Na2Saddition. _

_——————————————-———————_—————————_”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•—

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0 0 0 D °°0 0 ä

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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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Bates, R.G., Determination Qi BH, Wiley, New York,458-483 (1964).

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Boyce, J.R., Venter, W.J. and Adam, J., "BeneficiationPractice at the Tsumeb Concentrator," In: D.O.Raush and B.C. Mariacher (Editors), World S¥mQ•

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Buckley, A. and Woods, R., "An X-ray PhotoelectronSpectroscopic Investigation of the SurfaceOxidation of Sulfide Minerals," In P.E.Richardson, S. Srinivasan, and R. Woods (Editors)PIOC•_Qi the Int. Symp. gg Electrochem. iB_Mineral Metal Processing, Proceedings vol.84-10, The Electrochemical Society, Pennington,N.J., 286-302 (1984).

Chen K.Y. and Gupta, S.K. "Formation of Polysulfide inAqueous So1ution," Env. Letters, AJ 187 (1973).

Chen, K.Y. and Morris, J.C., "Kinetics of Oxidation ofAqueous Sulfide by O," Enviro. Sci. amd Tech., g_(6), 529-537 (1972).

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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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).

Eadington, P. and Prosser, A.P., "Oxidation of LeadSulfide in Aqueous Suspensions," Trans IMM, JuneC74-C82 (1969).

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Finklestein, N.P., Allison, S.A., Lovell, V.M. amdStewart, B.V., "Natural and InducedHydrophobicity in Sulfide Mineral Systems," In:P. Somasundaran and R. G. Grieves (Editors),Advances lg_Interfacial Applications_;g_FlotationResearch, AIChE, Symp. Ser. No. 150, Z4, 165-175(1975).

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Gaudin, A.M., Miaw, H.L. and Spedden, H.R., "NativeFlotability and Crystal Structure," Proc. 2ndAnnual Congress of Surface Activity, ButterworthSci., London, 3 (1957).

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Glembotskii, V.A., Klassen, V.I., and Plaksin, I.N.,Flotation, Primary Sources, New York (1963).

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II I

I63 I

IJones, M.H. and Woodcock, J,F., "Evaluation of

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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

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72

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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

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—1”E·E1; Z13

Product Meiont E ä Einc Unit Zn Distribution

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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

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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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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

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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

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Samble Carbon—1e Peak Charce Correctxch

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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.