International Journal of Mineral Processing Accepted 8 ... · Version préliminaire – non publiée power. In some cases, pyrite flotation may be inhibited. The main factors giving

Version préliminaire – non publiée

International Journal of Mineral Processing Accepted 8th July2002

PHYSICO-CHEMICAL PROPERTIES OF TAILING SLURRIES DURING

ENVIRONMENTAL DESULPHURIZATION BY FROTH FLOTATION

M. Benzaazoua1(URSTM, University of Quebec in A-T, Canada)

and M. Kongolo2 (LEM, École Nationale Supérieure de Géologie, France)

1 Université du Québec en Abitibi Témiscamingue, Unité de Recherche et de Service en Technologie

Minérale, 445 Boul. de l’Université, Rouyn-Noranda, Province Québec, Canada J9X 5E4.

FAX : 1 (819) 797-6672, E-mail : [email protected].

2 CNRS Laboratoire Environnement et Minéralurgie, B.P. 40, F-54504 Vandoeuvre-Lès-Nancy Cedex,

France.

FAX : (33) 3 83 59 62 55, E-mail : [email protected],

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mailto:[email protected]

mailto:[email protected]


_____________________________________________________________________________________

Abstract

Environmental desulphurization has been established as an alternative solution to control acid mine

drainage due to the reactivity of sulphide tailings when exposed to open air conditions. In fact, this process

placed at the end of the primary treatment circuit greatly reduces the amount of problem tailings by

concentrating the sulphide fraction. An acceptable target for sulphide content can be estimated from the

acid potential (AP), neutralization potential (NP) and the net neutralization potential (NNP) of the mill

tailings. To produce desulphurized tailings, non-selective froth flotation is the most commonly method

used in previous work. In this paper, the authors have focused the physicochemical properties of the pulp

as the main parameters affecting the non-selective sulphide flotation. The pyrite depression due to lime

addition during the former process represents the main problem. Several laboratory tests were conducted

using a Denver cell to choose the best and most economic collector. Other tests were done to select the

best frothing agent. The pH and redox potentials were investigated as parameters of great importance in

flotation performance. For studying the sulphide flotation kinetics, two mine tailings are chosen which are

characterized by a weak neutralization potential (under 37 kg CaCO3/t). Tailings S and L are cyanide free

and contain respectively 5.27 and 10 Wt. % sulphur. Collector dosage was optimized for these tailings as

well as the flotation time and the results show that Tailing L needed more collector than Tailings S.

Desulphurization costs were estimated to 0.35 $ per ton (dry tailings) which is very comparable and

competitive to existing method for tailings management.

Keywords : Desulphurization, tailings management, froth flotation, sulphides, collection.

___________

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1. Introduction

Throughout the world, many mining operations concentrate various valuable metals such as copper, zinc,

lead, gold and silver, etc., by treating sulphide ores. The mining process generates substantial tonnages of

tailings that contain various amounts of sulphides (mainly pyrite). These tailings are managed to avoid the

pollution problems caused by natural weathering of sulphides that lead to acid mine drainage. While the

methods used to prevent AMD are diversified (under water disposal using dams, dry natural or synthetic

covers, paste backfills, etc.), they are usually quite expensive. Over the last few years, froth flotation has

been proposed for tailings desulphurization as a new management technique with a view to reduce

rehabilitation costs. This method separates the sulphidic fraction so that it can be managed more easily

later due to the reduced volume. It can be made into paste backfill or simply disposed in a local area which

can be rehabilitated later. Moreover, the desulphurized fraction has the requisite properties for later use as

a mine cover. This have been previously demonstrated by column tests conducted for a one year period

(Bussière et al., 1997a-b; Bussière et al., 1998; Benzaazoua et al. 1998a). The environmental sulphur

recovery, which corresponds the sulphur proportion to be floated to produce a non-acid generating final

tailings, depends mainly on the intrinsic neutralizing potential of the tailings and the physico-chemical

properties of the corresponding pulp. Pyrite is usually depressed during the polymetallic ore treatment by

increasing the pH to approximately 11 by using lime.

There is extensive literature on sulphide flotation, especially on pyrite. Many authors have worked on

sulphide concentration by non-selective flotation for mineral processing purposes (preparation of

concentrates intended for gold and/or silver hydrometallurgy) and some for waste management strategy.

Regarding the available literature, we are able to mention the work of McLaughlin and Stuparyk (1994)

who evaluated the production of low sulphur tailings at INCO’s Clarabelle concentrator (also in Stuparyk

et al., 1995), the work of Balderama (1995) on various tailings impoundments in the United States about

controlling acid mine drainage, and the flotation test series of Leppinen et al. (1997) who focused on

recovering residual sulphide minerals from the tailings of the Pyhasalami Cu-Zn mine in Finland. Others

who have performed studies on this topic are Luszczkiewicz and Sztaba (1995), Humber (1995), Bussière

et al. (1995), Benzaazoua et al. (1998b), (1999), (2000a) and Benzaazoua and Bussière (2000b).

For non-selective flotation of sulphide mineral, the most common and most investigated reagents are the

xanthate-based collectors, which are characterized by their ability to collect for sulphide minerals. The

length of their radical chain is the cause of their selectivity (Crozier, 1992). Xanthates of the amyl type are

commonly used for non-selective flotation of sulphides (including pyrite) because of their collection

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power. In some cases, pyrite flotation may be inhibited. The main factors giving rise to this phenomenon

are (i) the surface state of the grains can be affected by natural oxidation or oxidation by dissolved

cyanides as demonstrated by Wet et al. (1997) and the pH of the pulp, especially when the xanthate

concentration is low, pH above 10 causes depression of the mineral (Duc, 1992). At higher xanthate

concentration, this effect disappears (Kongolo, 1991, Benzaazoua et al., 2000a). Fornasiero et Ralston

demonstrated the effect of iron speciation on the amyl xanthate adsorption within pyrite. Now is well

known that in addition to dixanthogen, iron xanthate and iron-hydroxide xanthate complexes contributes

to the flotation of pyrite (de Donato et al., 1989a,b; Kongolo, 1991).

In this paper, a number of tests were conducted using a Denver flotation Cell on mine tailings. The main

objective is to optimize the most important parameters that influence the non-selective sulphide flotation

performance placed at the end-circuit of a typical metallic ore processing. Attention was paid to the

physicochemical properties of the pulp from the time of its sampling and during the process of aging,

conditioning and flotation. To achieve this, some tailings were sampled and submitted to a series tests

consisting of optimizing the types and concentrations of the collector, type of froth, pH and oxido-

reduction potential and solid percentage. Consideration was given to lowered zinc and copper recoveries.

Finally, the paper will give an approximate estimation concerning operating costs related to the

desulphurization process regarding the two tailings studied.

2. Experimental section

2.1. Samples

Two tailings from Canadian mines were chosen for this study because they are considered representative

of typical sulphide tailings from hard rock Canadian mines. The tailings were sampled from the outlet of

the processing plant as a slurry with approximately 25 solid percentage and were stored with minimal air

contact to preserve as much as possible their physicochemical properties.

The contents of sulphur, zinc and copper were determined by ICP analysis. The chemical composition of

the different tailings samples and the calculated sulphide composition are presented in Table 1. Tailings

sample S has low sulphur contents compared to tailings sample L. The main sulphide mineral in the

studied tailings is pyrite (Table 2). Some pyrrhotite occurs in tailings L. Sphalerite and chalcopyrite are

accessory components (around 0.2 % sphalérite and les than 0.1 % chalcopyrite). The tailings studied are

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all cyanide free. Acid generating potential (also called Net Neutralization Potential NNP) is another

important characteristic of the tailings. Among the various methods used to evaluate this parameter, the

modified Acid Base Accounting test was chosen because of its simplicity and reliability (Lawrence and

Wang, 1997, Morin, 1997). The NNP is calculated as the balance between the Neutralization Potential

(NP) and the Acidity Potential (AP). AP is estimated from the sulphide sulphur content by chemical

analysis and the NP is determined by volumetric titration (using an NaOH 0.1M solution) of the pulp

mixed with an excess of HCl 0.1M solution The results are summarized in Table 3 and show that the two

tailings studied are acid generating and have a relatively low NP. Tailing L is more acidic than tailings S

due to their different sulphide contents.

Grain size analyses were done on the tailings because of the importance of this factor both in flotation

processes and in sulphide oxidation. The analyses were done with a laser based instrument (Malvern

Matsersizer). One can see in Table 1 that grain size distributions are very close for the two tailings

samples studied and lead us to consider them to be of negligible importance in this study. Table 1 also

shows the relative density as determined by an Helium pycnometer (Micromeritics) which indicates the

sulphide contents of the materials.

2.2. Reagents

Flotation requires different types of reagents to condition the superficial tension for the desired minerals

(all sulphides in our case) and to assure the collection mechanisms and the proper chemical condition for

the pulp. The technical specifications of the reagents tested for flotation experiments are the following :

Collectors:

- KAX-41: Potassium propyl xanthate, from Prospec Chemicals Inc

- KAX-51: Potassium amyl xanthate, from Prospec Chemicals Inc.

- FLEX 31 : xanthate derivative, from Prospec Chemicals Inc

- SPRI 105 Phosphorodithioate Salt Dithicarbamate, from Prospec Chemicals Inc.

- SPRI 206 Phosphorodithioate Salt Dithicarbamate, from Prospec Chemicals Inc.

- AERO 3477 Dithiophophates, from Cytech Canada

- AERO 7279 Dithiophosphates + dithiocarbamate, from Cytech Canada

- S 7151 Dithiophosphates + dithiocarbamate, from Cytech Canada

Frothing Agents:

- D-200: Polypropylene glycol methyl ether; Dow Chemical.

- Sasfroth Sc39, 161, Sc26, Sascol 105, from Prospec Chemicals Inc.

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pH modifiers:

- Diluted solution of H2SO4

- Diluted solution of NaOH

2.3. Flotation tests

Slurries were sampled at the mine concentrators in a way that all the initial physicochemical

characteristics of the pulp (residual reagents, pH, Eh, etc.) were preserved the more possible. The target

solid percentage was 30 solid % for all test. Time of conditioning was 10 minutes after simultaneous

collector (at various concentrations) and frother additions (16 µl/kg tailings). All flotation tests were

carried out in a Denver D-12 lab flotation machine. The used cell volume was 2.5 liters. Speed of the

rotor-stator was adjusted to 1500 rpm and airflow was fixed at 2.25 liters per minute. To obtain consistent

results, the same operator manually removed the froths with a spatula for all of the flotation tests. The pH

was measured and adjusted by adding a diluted H2SO4 solution for acidification or a diluted NaOH

solution for pH increase depending on tests.

3 Results and discussions

The optimization began by fixing the more efficient reagents to obtain the best desulphurization results.

Several types of collector and frothers were verified in the study. To achieve this objective, only tailings L

were chosen.

3.1 Choice of collector

The flotation experiments on tailings L were done at three different pH levels (7.5, 9 and 10.5) to test the

various collectors cited in the experimental section. The frother was D-200, which is commonly used for

sulphide flotation and the target in term of collector dosage was 150 g/t (purity of each product was not

taken into account). The feed of all experiments varied between 9.2 and 9.8 wt% suggesting relative

constant head grade. The results are summarized in Table 2 which contains information about total

sulphur , zinc and copper recoveries, the concentrate weight percentage and its sulphur grade. In term of

non selective sulphide recovery, the best results are obtained with an alkaline pH using xanthate collectors

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(KAX 51, 41 and FLEX 31). The choice of optimal collector was done by considering the following

criteria (in decreasing importance) :

- Price (must not exceed 2.5 $/kg)

- The sulphur recovery (97 % is considered as a sufficient recovery

- The zinc and copper recoveries (maximum)

- The weight and sulphur grade of the concentrate (lesser weight and higher sulphur grade)

The best results were obtained using xanthates as the collector under alkaline pH conditions. It reached 93

% sulphur recovery and approximately 60 and 90 % recoveries for copper and zinc respectively. The

concentrate represents 28 % of the total tailings weight and has a sulphur grade of 33 %. In comparison

with other expensive collector, xanthate gives the best price/efficiency ratio. Thus, KAX 41, which is a

potassium propyl xanthate, was chosen for the rest of experiments in this study.

3.2 Choice of the frother

In the same manner as the collectors, several frothers were tested using the same performances criteria. A

concentration of 150 g/t amyl xanthate (78 % purity) was used for these tests and the pH target was 10.5.

The final choice was Sasfroth Sc 39 which offered the best efficiency in terms of sulphur recovery and

concentrate quality. The results are summarized in Table 3.

3.3 Pulp aging

Pulp aging was investigated as a natural parameter leading to pyrite activation. Pulp was freshly sampled

at a mine concentrator "L" than submitted to chemical analyses of the pulp solution in one hand and to

flotation testing at regular time-intervals in the other hand. The pulp geochemistry (Fig.1) shows that

sulphide oxidation occurs within the first 10 days which leads to a decrease in pH and an increase in Eh.

Metal species are also released as shown by conductivity and iron analyses. Calcium as well as oxidized

sulphur species (probably in the form of sulfates) seems to precipitate progressively. Two types of

flotation tests were conducted at regular increased time-intervals. The first consisted of floating the pulp

without any pH regulation and the second test was to neutralize the pulp prior to maintaining the pH at 11

using soda ash. The results (Fig.2) demonstrated that the pyrite is depressed due to lime addition during

the former process. Acid was added to clean up aged surfaces. A pH of 7 was sufficient prior to raising the

pH to 11 using Caustic soda. The aging leads to the same results as the surface cleaning. The flotation

became good after 14 days of aging.

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3.4. pH and redox effect

At the end-circuit of the processing plants, slurries are generally alkaline as it necessary for copper, lead

and zinc mineral flotation as well gold ore cyanidation. The pH is usually higher than 10.5, as this

condition is necessary to assure the depression of barren sulphide minerals (mainly pyrite, arsenopyrite

and pyrrhotite initially contained in the ore) when using various specific and selective collectors. To reach

this goal, the pH is set by adding an alkaline reagent like lime during the different steps of processing.

The redox potential of the pulp at end-circuit is relatively low (around – 200 mV) immediately after

sampling tailings L. It tends to increase with natural aging (as demonstrated in the above section) or

during conditioning and flotation steps due to agitation and air bubble diffusion. In this work, pH and Eh

were investigated as the main parameters controlling the collector adsorption onto sulphide minerals. In

fact, this two parameters control the chemistry of the pulp as well as the surface mineral composition.

Some flotation tests using Denver cell at a KAX 41 concentration of 100 g/t shows that the pH has a

strong effect on the sulphide recovery as shown in Figure 3. The best recoveries were obtained at pH

between 6 and 11. Moreover, the results obtained from the other tests shows that there is no effect of Eh

on the sulphur recovery at pH approximately 6. However, recovery is very sensitive with respect to the Eh

variation at a pH of 11 as shown in Figure 4. The Eh increases as the sulphur recovery decreases. The pH

and Eh control the soluble species within the pulp and consequently the superficial phases within the

surface of the mineral. This fact, demonstrated elsewhere (e.g. Fornasiero and Ralston, 1991), explains

that the adsorption of the collector depends on these two parameters without forgetting the role of the

acidic condition in the clean up of the pyrite surfaces from depressant species as calcium hydroxides.

3.5. Collection kinetics

Time and collector dosage required for the flotation of a given pulp can be determined by the production

of successive concentrates at various collector dosages. To find the optimal conditions for environmental

desulphurization , it is necessary to study the non selective flotation kinetics of the different sulphide

minerals for each tailings sample studied (tailings S and L) and for different collector dosages.

Experiments were carried out to investigate the flotation kinetic of the two non-cyanided tailings under the

conditions cited above using an amyl xanthate collector. The results are presented in Figures 5 and 6

where the sulphur recovery % versus time and residual sulphur % versus time are presented for tailings

sample S and L respectively.

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For Tailings S, experiments showed that the collector dosage has an influence only at xanthate

concentrations under 20 g/t (see Figure 5). The residual sulphur content reached was below 0.4 Wt % after

8 minutes which corresponds to a recovery of approximately 95 % and a weight percentage of the

concentrate of 28 %. Concerning copper and zinc, in the optimum condition, recovery reached 70 % for

both metals.

Tailings L has more sulphides than the first one studied (see Table 1). The residual sulphur content was

about 0.4 % and was reached with a collector concentration of 100 g/t (Fig.6). Increasing KAX dosage

does not improve the sulphur recovery which stabilized at 97 %. Moreover, concentration of KAX below

40 g/t seems to have no effect on the sulphide flotation. In the best cases, the weight concentrate

percentage was around 45 %. This indicates an important entrainment of gangue minerals.

4. Cost estimation

Despite the fact that the main objective of this work was not to extensively develop the economical aspect

of desulphurization process, a tentative of cost estimation was done to give at least a general idea. The

desulphurization tend to produce a sulphide concentrate and a non-acid generating fraction which must

have enough neutralization potential to neutralize the acid produced by the residual sulphide. The

classification criterion used here is based on the net NP/AP (SRK, 1989, Morin and Hutt, 1997). This

criterion considers a material as acid generating if the ratio is less than 1. The NPs of the tailings after

desulphurization remain stable or increase slightly due to the relative enrichment of the carbonates mineral

which happens after removing the sulphides in the concentrate (particularly when the entrainment is low).

The environmental sulphur recovery needed called R can be calculated as corresponding to the sulphur

proportion to recover for decreasing the initial NP/AP ratio to a value equal or greater to 1 (i.e. with an

acidy potential equal to the neutralization potential). For Tailings S, R must be equal or greater than 89.8

% and R must be equal or greater than 90.8 % for Tailings L. These high recovery values are due to the

low neutralizing potential of the studied tailings.

The main challenge of the desulphurization consists of optimizing the two most important parameters of

the global desulphurization costs, i.e. the flotation time and the collector dosage needed to reach the

acceptable sulphide recovery. This two parameters can be estimated by using some mathematical models

established in a previous works (Benzaazoua et al. 2000a). These models are described in equation 1, 2, 3

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and 4 and represent respectively the general kinetic model for flotation (eq.1), the acceptable sulphur

recovery R as a function of the model parameters k (eq.2), Rf (eq.3) and collector dosage d (eq.4).

[

−−−= )exp(111 kt

ktRfr ] (1)

where r is the sulphur recovery (%),Rf the final recovery and k the flotation rate constant.

( )[

−−−= k

kR 5exp5.1

511105 ] (2)

R = 0.996Rf – 4.25 (3)

([

−−−= d

dR 45.0exp5

45.011105 )] (4)

where R is the acceptable sulphide recovery % and d the optimal collector concentration for any given

tailings.

These models are based on numerous desulphurization kinetic tests conducted on four different mine

tailings. The test, where all flotation condition were similar, consisted of varying the collector

concentrations (Benzaazoua et al., 2000).

The environmental flotation time can be defined as the flotation time needed to obtain desulphurized

tailings with the desired environmental characteristics (i.e. without acid generating risk). Flotation time

was optimized using the acceptable environmental sulphur recovery of the tailings (evaluated with NP/AP

criterion) and the kinetic model for which the final recovery "Rf" and time constant "k" have been first

estimated with equations 2 and 3. As it can be seen in Table 4, the laboratory environmental flotation time

is around 3 minutes for the studied tailings. Table 4 summarizes all of the desulphurization data including

optimal collector concentration for each material studied.

Thus, the operating costs of desulphurization are 0.35 CND $/t approximately for the two tailings studied.

The capital costs of the desulphurization are estimated, for the two mines studied, to be around 1 000 000

CND $. These capital costs are for new equipment and could be reduced by using old flotation cells. Even

if the desulphurization cost is not negligible, this alternative could be in many cases an economic solution

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from the environmental point of view particularly when the mines use paste backfill (Bussière et al.

1997b; 1998).

5. Summary and conclusions

This study demonstrates that environmental desulphurization is possible for the cases studied herein even

if they were characterized by low neutralizing potential. Free-cyanide pulps can be desulphurized easily

by froth flotation with amyl xanthates. The process generates a sulphide concentrate and non acid-

generating desulphurized tailings. The results allow to choose the adequate collector and frother to reach

this objective. Moreover, it have been demonstrated that the pH and redox parameters must be set prior to

the flotation because of their crucial role in the collector adsorption on the sulphide surfaces. Natural pulp

aging acts as an activating process for pyrite flotation. Through these tests, the relationships between

desulphurization parameters have been established. These relationships can be used to estimate the

optimal collector dosage and the flotation time needed to obtain a final tailings with acceptable NP/AP.

Due to the low neutralizing potential of the two tailings studied, the desulphurization process needs

around 20 g/t and less than 100 g/t collector for tailings S and L respectively. Concerning the required

times, they are about 8 minutes for both tailings. In the investigation of an optimal waste management

strategy, this study confirms that desulphurization of mine tailings must be regarded as an alternative to

the other existing techniques (Bussière et al., 1997b; 1998). Desulphurization of mine tailings must be

evaluated in its overall context as an attractive alternative to the other techniques existing for tailings

management. In many cases, it leads to a major reduction in costs related to the supply and transportation

of natural materials (such as clay and gravel) or the permanent monitoring of liquid effluent quality.

Another worthwhile technique may be considered at the same time as desulphurization; the use of paste

fill technology to place the sulphidic fraction backs underground. The costs of surface rehabilitation could

be limited, by this way, to the expense of desulphurization (around 0.35 CND $ per ton), disposal, and

revegetation.

Acknowledgments

This work was financed through an industry-university program. All the actors in this proms are

acknowledged particularly Denis Bois. We would like to thank also Nil Gaudet for the technical

contribution. Finally, the authors would like to thank all the mine partners who participate in this project.

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List of tables

Table 1 Chemical, mineralogical, static ABA test and physical characteristics of the two tailings studied

Table 2 Results of tailings L desulphurization using various types of collectors

Table 3 Results of tailings L desulphurization using various types of frothing agents

Table 4 Optimization characteristics of the environmental desulphurization

List of figures

Fig.1 Evolution of the grade of some soluble species released in the pulp solution

Fig.2 Sulphur recovery (plot on the left) and residual sulphur (plot on the right) evolution during the

desulphurization tests

Fig.3 Effect of the pH on the sulphur recovery during desulphurization tests

Fig.4 Effect of the Eh of the pulp on the flotation kinetic at pH 11 (plot on the left) and pH 6 (plot on the

right).

Fig.5 Flotation kinetic of Tailings S for 8 KAX dosages (in g/t). % Sulphur recovery vs. time (plot on the

left) and Residual sulphur % vs. time (plot on right)

Fig.6 Flotation kinetic of Tailings L for 8 KAX dosages (in g/t). % Sulphur recovery vs. time (plot on the

left) and Residual sulphur % vs. time (plot on right)

14


Tables

Table 1 Chemical, mineralogical, static ABA test and physical characteristics of the two tailings studied

Tailings S Tailings L

S Wt % 5.27 10

Zn ppm Wt 0.15 0.12

Cu ppm Wt 0.03 0.07

Pyrite 9.2% 17.4%

Sphalerite 0.22% 0.19%

Chalcopyrite 0.04% 0,10%

S (sulphate) Wt % 0.27 0.62

S (sulphide) Wt % 5 9.38

AP kg CaCO3/t 165 313

NP kg CaCO3/t 16 27

Net NNP (sulphide S) -140 -266

Net NP/AP 0.1 0.09

D90 (µm) 100.6 76.3

D50 (µm) 14.25 17.85

D10 (µm) 2.43 2.09

Gs 2.85 3.11

15


Table 2 Results of tailings L desulphurization using various types of collectors

% recovery

Collector Price CND $ Final pH Wt% Conc. %S Conc. S Cu Zn

7.5 26.9% 31.8% 92.5 36.3 66.7

KAX 51 2.1 9.0 21.1% 40.5% 89.8 46.0 94.2

10.5 27.7% 32.7% 93.4 36.1 66.4

7.5 23.6% 37.0% 91.9 34.6 76.1

KAX 41 2.05 9.0 22.2% 37.5% 89.0 45.8 94.2

10.5 28.5% 33.1% 95.9 60.5 88.1

7.5 17.7% 37.1% 69.0 26.6 80.5

Aero 407 4 9.0 20.9% 36.6% 85.3 43.1 93.8

10.5 23.4% 30.5% 75.5 34.4 78.8

7.5 27.4% 29.7% 83.2 49.0 77.3

Aero 3477 5 9.0 21.7% 44.3% 89.7 46.0 94.1

10.5 28.2% 30.0% 92.2 49.1 79.3

7.5 29.9% 28.5% 88.0 48.0 77.3

Aero 7279 6 9.0 23.0% 37.3% 88.8 44.6 93.9

10.5 28.3% 28.4% 91.8 50.6 81.4

7.5 26.5% 32.4% 89.9 51.5 80.8

SPRI 105 5 9.0 21.4% 40.6% 91.3 42.7 94.5

10.5 28.4% 29.9% 91.7 52.0 80.3

7.5 27.1% 29.5% 83.2 50.4 80.7

SPRI 206 5.45 9.0 28.4% 28.3% 88.2 56.4 87.8

10.5 25.8% 25.1% 67.0 56.4 87.1

7.5 29.8% 30.6% 94.2 45.3 71.7

FLEX 31 3.25 9.0 23.0% 38.1% 92.5 46.0 94.1

10.5 27.7% 31.0% 90.4% 46.8 80.0

7.5 25.3% 33.0% 86.2 47.1 77.2

S 7151 6 9.0 21.5% 39.5% 89.4 41.1 94.3

10.5 28.0% 30.3% 90.3 55.7 88.0

16


Table 3 Results of tailings L desulphurization using various types of frothing agents

Récupération %

Frother Wt% Conc. %S Conc S Cu Zn

Sasfroth SC39 28.0% 32.0% 94.5% 35.0% 63.5%

Sascol 105 30.9% 23.3% 80.0% 36.5% 63.1%

Sasfroth 161 28.6% 30.6% 88.9% 41.4% 74.3%

Sasfroth SC26 29.0% 29.3% 87.3% 41.0% 73.7%

D-200 29.8% 30.5% 93.0% 54.7% 84.4%

Table 4 Optimization characteristics of the environmental desulphurization

Tailings S Tailings L %S tailings for NP/AP = 1 0.51 0.86

Environmental sulphur recovery % 89.8 90.8

Optimal [collector] g/tm 25 90

Environmental flotation time, min. 3 2

17


Figures

Ca mg/l

300

350

400

450

500

0 10 20 30 40

days

Fe mg/l

0

0,05

0,1

0,15

0,2

0,25

0 10 20 30 40

days

S mg/l

0500

10001500

20002500

3000

0 10 20 30 40

days

pH

8

8,5

9

9,5

10

10,5

11

0 10 20 30 40

days

Conductivity µmohs

2000

2100

2200

2300

2400

2500

2600

0 10 20 30 40

days

Eh (SHE) mV

250

300

350

400

450

500

0 10 20 30 40

days

Fig.1 Evolution of the grade of some soluble species released in the pulp solution

50556065707580859095

100

0 10 20 30 40

Days

Sulp

hur

reco

very

(%)

0

1

2

3

4

5

6

7

8

0 10 20 30 4

Days

Res

idua

l sul

phur

(%)

0

Natural

H2SO4 pH 7 than NaOH pH 11

Fig.2 Sulphur recovery (plot on the left) and residual sulphur (plot on the right) evolution during the

desulphurization tests of tailings L

18


80%82%84%86%88%90%92%94%96%98%

100%

4 5 6 7 8 9 10 11 12

pH

Sulp

hur

reco

very

(%)

Fig.3 Effect of the pH on the sulphur recovery during desulphurization tests on tailings L

pH 6

0%

20%

40%

60%

80%

100%

0 5 10 15Time (min.)

Sulp

hur r

ecov

ery

(%)

140 mV -200 mV

pH 11

0%

20%

40%

60%

80%

100%

0 5 10 15Time (min.)

Sulp

hur r

ecov

ery

(%)

200 mV 150 mV -15 mV -140 mV -230 mV

Fig.4 Effect of the Eh of the pulp on the flotation kinetic at pH 11 (plot on the left)

and pH 6 (plot on the right) of tailings L.

0%

20%

40%

60%

80%

100%

0 2 4 6 8 10Time (min.)

Sulp

hur

reco

very

(%)

5 10 24 3444 50 77 99

0%

1%

2%

3%

4%

5%

0 2 4 6 8Time (min.)

Res

idua

l sul

phur

(%)

10

5 10 24 3444 50 77 99

Fig.5 Flotation kinetic of Tailings S for 8 KAX dosages (in g/t)

% Sulphur recovery vs. time (plot on the left) and Residual sulphur % vs. time (plot on right)

19


0%

20%

40%

60%

80%

100%

0 5 10 15Time (min.)

Sulp

hur

reco

very

(%)

10 19 38 5778 98 116 145

0%

2%

4%

6%

8%

10%

0 5 10 15Time (min.)

Res

idua

l sul

phur

(%)

10 19 38 5778 98 116 145

Fig.6 Flotation kinetic of Tailings L for 8 KAX dosages (in g/t)

% Sulphur recovery vs. time (plot on the left) and Residual sulphur % vs. time (plot on right)

20

Documents

International Journal of Mineral Processing Accepted 8 ... · Version préliminaire – non publiée power. In some cases, pyrite flotation may be inhibited. The main factors giving