desulphurization of tailing

.Int. J. Miner. Process. 60 2000 5774www.elsevier.nlrlocaterijminpro

Environmental desulphurization of four Canadianmine tailings using froth flotation

M. Benzaazoua a,), B. Bussiere a,1, M. Kongolo b,2, J. McLaughlin c,3,`P. Marion b,4

a Uniersite du Quebec en Abitibi Temiscamingue, Unite de Recherche et de Serice en Technologie Minerale, 445 Boul. de lUniersite, Rouyn-Noranda, Proince Quebec, Canada J9X 5E4

b CNRS Laboratoire Enironnement et Mineralurgie, B.P. 40, F-54504 Vandoeure-Les-Nancy Cedex, France `c Noranda Technology Centre, 240, Boul. Hymus, Pointe-Claire, Quebec, Canada H9R 1G5

Received 5 August 1999; accepted 27 December 1999

Abstract

Environmental desulphurization is an attractive alternative for management of acid generatingtailings. This process placed at the end of the primary process treatment circuit will reduce a largeamount of the problematic tailings by concentrating the sulphide fraction. To produce desulphur-ized tailings, non-selective froth flotation is the most adapted method. The desulphurization level

.is fixed by the sulphide content of the tailings and their neutralization potential NP . The final .residue should have enough NP to safely compensate for its acidity potential AP .

In this paper, the authors present the results of a battery of tests conducted in Denver cells tostudy the sulphide flotation kinetics of four different mine tailings samples which contain 2.9 S%,3.4 S%, 16.2 S% and 24.2 S%, respectively. Tailings P, M, and G are cyanide free and can befloated at pH values of less than 10 by using amyl xanthate as the collector agent. However,tailings D, which come from a gold cyanidation process, did not provide good sulphide recovery

because of pyrite depression even after cyanide elimination followed by a sulphide surface.activation and a pH decrease . To successfully overcome this problem, amine acetate was used.

) Corresponding author. Fax: q1-819-797-6672. .E-mail addresses: [email protected] M. Benzaazoua ,

. [email protected] B. Bussiere , [email protected] M. Kongolo ,` . [email protected] J. McLaughlin , [email protected] P. Marion .

1 Fax: q1-819-797-6672.2 Fax: q1-333-83-59-62-55.3 Fax: q1-517-630-9379.4 Fax: q1-333-83-59-62-55.

0301-7516r00r$ - see front matter q2000 Elsevier Science B.V. All rights reserved. .PII: S0301-7516 00 00006-5

( )M. Benzaazoua et al.r Int. J. Miner. Process. 60 2000 577458

This collector allows easy flotation without a pretreatment stage. Dosage of the two type ofcollector were optimized for the tailings studied. The results of all kinetic tests and collectordosage optimization were combined to establish a model which will estimate the cost ofdesulphurization. A preliminary analysis shows that the expense of the implementation of thisdesulphurization technology is comparable to other rehabilitation methods. q 2000 ElsevierScience B.V. All rights reserved.

Keywords: acid mine drainage; desulphurization; flotation kinetics; pyrite; modeling; acid generating potential

1. Introduction

Throughout the world, many mining operations deal with sulphide ores to extract . .various valuable minerals such as chalcopyrite CuFeS2 for copper, sphalerite ZnS for

.zinc and galena PbS for lead. Some gold and silver ores are also sulphide rich. Themining process generates substantial tonnages of tailings that contain various amounts ofsulphides, of which the major component is pyrite. These tailings have to be properlymanaged to avoid the pollution problems caused by natural weathering of sulphides. Infact, pyrite and other sulphides lost during mineral processing will oxidize in aqueousmedia under the effect of oxygen. Thiobacilli bacteria appreciably accelerate thesulphide oxidation under certain geochemical conditions. The result of this pollution is

.called acid mine drainage AMD . A detailed description of the process involved and theproblems associated with AMD have been discussed by many authors Kleinmann et al.,

.1981; Lowson, 1982; Evangelou, 1995; Gray, 1997; etc. . While the methods used toprevent AMD are diversified, they are usually quite expensive. A relatively newalternative has been proposed recently to reduce rehabilitation costs. This methodseparates the sulphidic fraction that can later be managed more easily due to the reduced

volume. Moreover, the desulphurized fraction has the requisite properties essentially.non-reactive chemically and with good geotechnical characteristics to be used as the

fine material in a cover which has the capillary barrier effect demonstrated by columntests conducted for over a 1-year period Bussiere et al., 1997a,1997b; Bussiere et al.,` `

.1998; Benzaazoua et al., 1998b .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 andror silver hydrometallurgy

and some for waste management strategy. Regarding the available literature, we are able . .to mention the work of McLaughlin and Stuparyk 1994 , Stuparyk et al. 1995 who

evaluated the production of low sulphur tailings at INCOs Clarabelle concentrator, the .work of Balderama 1995 on various tailings impoundments in the United States about

.controlling AMD, and the flotation test series of Leppinen et al. 1997 who focused onrecovering residual sulphide minerals from the tailings of the Pyhasalami CuZn mine

.in Finland. Others who have performed studies on this topic are Bussiere et al. 1995 ,` . . .Humber 1995 , Luszczkiewicz and Sztaba 1995 and Benzaazoua et al. 1998a .

For non-selective flotation of sulphide mineral, the most common and most investi-gated reagents are the xanthate-based collectors, which are characterized by their ability

( )M. Benzaazoua et al.r Int. J. Miner. Process. 60 2000 5774 59

to collect for sulphide mineral in general. The length of their radical chain is the cause .of their selectivity Crozier, 1992 . Xanthate of the amyl type are more intended for

.non-selective flotation of sulphides including pyrite because of their collection power.In some cases, pyrite flotation may be inhibited. The main factors giving rise to thisphenomenon are: surface state of the grains can be affected by natural oxidation or

.oxidation by dissolved cyanides as demonstrated by Wet et al. 1997 , and pH of thepulp, 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 .

Other collectors can also be used successfully for sulphide non-selective flotation. .This was demonstrated by the works of Bradshaw and OConnor 1994 on thiocarba-

.mates, OConnor and Dunne 1991 on mercaptobenzothiazoles and Hodgkinson et. .al., 1994 on amines. However, these collectors are not recommended for desulphuriza-

tion processes due to their high costs and relatively slow flotation kinetic compared toxanthates.

The work described in this paper concerns froth flotation of sulphide tailings to .produce a low sulphur fraction non-acid generating and a sulphide concentrate fraction.

Tailings from four different mines of the Abitibi region in Quebec, Canada were studiedto optimize flotation conditions in terms of collector dosage and flotation time. Thesulphur content of these tailings ranged from 2.9 to 24.2 wt.%. One of the tailings pulp .tailings D contains cyanide and was studied in order to investigate the inhibiting roleof cyanide in sulphide flotation. The modified Sobek standard static test was used todetermine the flotation level required to produce non-reactive end result tailings. Theresults were compiled and modeled to make useful simulation tools for the economicissues related to the desulphurization technology for control of AMD.

2. Methods and materials studied

2.1. Reagents

Flotation requires different types of reagents to produce the proper surface tension forthe desired mineral, collection and adequately make the proper chemical characteristicsfor the pulp. The choice of collectors and frothing reagents was based on the preliminarytests conducted by the authors as well as extensive literature review. The technicalspecifications of the reagents used to achieve the flotation experiments are the follow-ing: Collectors: KAX-51: Potassium amyl xanthate; purity of 80 wt.%; molecular weight202.3 g; from Prospec Chemicals; ARMAC-C: Cocoalkylamine acetate; purity of 95%to 100%; from Akzo Nobel Chemicals; Frother: D-200: Polypropylene glycol methylether; Dow Chemical; and Actiator: Copper sulphate; industrial product diluted at 10%

.solution unknown purity .

2.2. Tailings

Four tailings from four important Canadian mines were chosen for this study becauseof their diverse characteristics which are representative of existing sulphide tailings in


Table 1Mine descriptions

Type of Valuable Sulphide Silicate Carbonate Ore feeddeposit metals minerals minerals minerals rate

Mine P Sulphide Gold, copper Pyrite, Chlorite, quartz Siderite, Closed minedeposit pyrrhotite, and sericite ankerite,

chalcopyrite calciteMine D Sulphide Gold, silver Pyrite, Quartz, micas, Calcite, 3300 trday

deposit chalcopyrite, chlorite, dolomitesphalerite plagioclases

Mine M Massive Zinc, copper, Pyrite, Chlorite, talc, Small 2500 trdaysulphide gold, silver pyrrhotite, quartz amountdeposit chalcopyrite,

sphaleriteMine G Massive Zinc, copper, Pyrite, Undetermined Undetermined 2500 trday

sulphide gold, silver pyrrhotite, silicates carbonatesdeposit chalcopyrite,

sphalerite

Canada. A brief description of these mines is presented in Table 1. The tailings weresampled from the outlet of the processing plant as a slurry with approximately 50 wt.%solid and were then stored with minimal air contact.

The contents of sulphur, zinc, and copper were determined by ICP analysis. The goldgrade was determined by fire assay. The chemical composition of the different tailingssamples and the calculated sulphide composition are presented in Table 2. Tailingssamples P and D have low sulphur contents. The gold contents of the two tailings is 0.56and 0.23 ppm, respectively. One can observe in Table 2 that the main sulphide mineralin all of the tailings studied is pyrite. Sphalerite is a significant component in tailings G

.and M 1.24 and 1.74 wt.%, respectively , whereas chalcopyrite does not exceed 0.2wt.% in any of the tailings samples. Pyrrhotite, when present, is indistinguishable frompyrite in this process and is of negligible content. Also, total cyanide from a freshfiltered solution from pulp D was analyzed and gave a value of 240 mgrl CNy.

Table 2Chemical and mineralogical composition of the four tailings studied

Pulp P Pulp D Pulp M Pulp G .S wt.% 2.89 3.4 16.2 24.2 .Zn ppm wt. 244 116 11700 8300 .Cu ppm wt. 355 533 566 715 .Au ppm wt. 0.56 0.23 0.09 0.04

.Pyrite % 5.31 6.24 29.10 44.34 .Sphalerite % 0.04 0.02 1.74 1.24

.Chalcopyrite % 0.10 0.15 0.16 0.21


Table 3ABA static tests for the tailings studied

Pulp P Pulp D Pulp M Pulp G .S total wt.% 2.9 3.4 16.2 24.2

2y .SO4 wt.% 0.14 0.1 0.40 0.20AP kg CaCO rt 90.3 106.3 506.3 756.33NP kg CaCO rt 130.0 25.0 120.0 105.03

.NNP total S kg CaCO rt 39.7 y81.3 y386.3 y651.33 .Net NNP sulphide S kg CaCO rt 44.1 y79.0 y373.8 y645.03

NPrAP 1.4 0.2 0.2 0.1

Another important characteristic of the tailings is the acid generating potential or net.neutralization potential, NNP . From the various methods used to evaluate this parame-

ter, the modified Acid Base Accounting test was chosen because of its relative reliability .Sobek et al., 1978; Lawrence and Wang, 1997 . The NNP is calculated as the difference

. .between the neutralization potential NP and the acidity potential AP . AP is estimatedfrom the sulphide sulphur content by chemical analysis and the NP is determined by

.volumetric titration using an NaOH 0.1 M solution for the pulp mixed in excess of a .HCl 0.1 M solution 2 g solidr20 ml solution . The results are summarized in Table 3

and indicate tailings M, G, and D are acid generating. Results for tailings P show the NPof these tailings is greater than their AP indicating tailings P should be in the uncertainzone of acid generating potential.

Particle size analyses were done on the studied tailings because of the importance ofthis factor for both the flotation process and sulphide oxidation. The analyses were done

.with a laser-based instrument the Malvern Matsersizer . One can see in Table 4 andFig. 1 that particle size distributions corresponding to the four tailings are quite close toeach other except for tailings P, which is slightly a larger size. This led us to considerparticle size a not too significant parameter for this study. Table 4 presents the relative

.density as determined by a Helium pycnometer Micromeritics which is indicative ofthe sulphide contents of the materials.

.The initial pH of the cyanide free pulps tailings P, M, and G was between 7 and 9. .An alkaline solution NaOH was added to fix the pH between 9 and 10 which is the

.optimum for pyrite flotation using the xanthate collector Duc, 1992 . The pulp .containing cyanides tailings D was very alkaline because of the lime added during the

Table 4Main results of the grain size analysis

Tailings P Tailings D Tailings M Tailings G .D80 mm 105 58 65 56 .D50 mm 37 23 28 30 .Mean mm 108.6 63.8 88.8 70.1

3 .D grcm 3.04 2.85 3.04 3.34r


.Fig. 1. Size diameter distribution of the four tailings studied histogram right and cumulative left .

.process pH)11.5 . For tests with xanthate, the pulp D was slightly acidified, after twosuccessive flushes with fresh water and activation with copper sulphate, to avoid thedepression of pyrite using xanthates. However, when this pulp was floated usingARMAC-C, no pH modification was required.

2.3. Conditioning and flotation

Experimental procedure is schematized in Fig. 2. High density slurries were dilutedby using process water from the processing plant. This preserves all the initial

.physicochemical characteristics of the pulp residual reagents, pH, Eh, etc. . The targetsolid was 30%. Time of conditioning was 10 min after simultaneous collector KAX or

. .ARMAC-C in various concentrations and frother additions 16 mlrkg tailings . Allflotation tests were carried out in a Denver D-12 lab flotation machine. The cell volumewas 2.5 l. Speed of the rotor-stator was adjusted to 1500 rpm and airflow was fixed at2.25 lrmin. To obtain consistent results, the froths were manually removed by the sameoperator with a spatula for all of the flotation tests. The pH was measured and adjustedby adding a diluted H SO solution for acidification or a diluted NaOH solution for pH2 4increase.

Fig. 2. Schematic representation of the experimental procedure for the flotation tests.


3. Results and analyses

3.1. Flotation kinetics

The time and the collector dosage required for the flotation of a given pulp can bedetermined by the production of a series of successive concentrates. To reach theobjectives, as far as environmental desulphurization is concerned, it is necessary to studythe flotation kinetics of sulphide minerals for each tailings pulp and collector concentra-tion studied. Therefore, modeling work on the desulphurization process and the associ-ated flotation cost estimation are studied concurrently to achieve this purpose.

3.1.1. Amyl xanthateExperiments were carried out to look at the flotation kinetics for the four tailings

under the conditions cited above using an amyl xanthate collector. The results arepresented in the form of sulphur recovery % vs. time and residual sulphur % vs. time.

3.1.1.1. Pulp P. Because of the low sulphur content of tailings P, low collector dosageswere tested. The experiments show that the collector concentration has an influence only

.at xanthate concentrations below 20 grt see Fig. 3 . However, the slightly largerparticle diameter for this tailing material could be partially responsible for the fastflotation of the sulphide minerals. In fact, residual sulphur levels were reached below 0.3

. .wt.% at 4 min flotation Fig. 3b with a sulphur recovery of 90% Fig. 3a and weightconcentrate of 9%.

3.1.1.2. Pulp D. To float the sulphides in these tailings with a xanthate collector, freecyanide must be eliminated by a double wash and the sulphides must be activated. Theactivator used is copper sulphate at 300 grt dosage. CuSO was added to the pulp after4two rinses with fresh water. As shown in Fig. 4, increasing the collector dosage from 40to 80 grt does not have any significant influence on the sulphide recovery; the finalsulphur recovery reaches a plateau at approximately 80%. This corresponds to a contentof 0.76 wt.% residual sulphur and a concentrate of 12% of the total weight tailings.

. .Fig. 3. Flotation kinetic as a function of collector dosage for tailings P. a Sulphur recovery vs. time %. bResidual sulphur % vs. time.


. .Fig. 4. Flotation kinetic as a function of collector dosage for tailings D. a Sulphur recovery vs. time %. bResidual sulphur % vs. time.

3.1.1.3. Pulp M. The first battery of tests for tailings M was done without sulphideactivation. Fig. 5a shows sulphur recovery was improved by increasing xanthate dosage

up to 120 grt. Doubling this concentration had a little effect on the total recovery see. .Fig. 5a . However, the residual sulphur was still high at 2.4 wt.% S see Fig. 5b . A

second series of tests was undertaken using 120 grt xanthate concentration after addingcopper sulphate to activate the sulphides. The desulphurized fraction of these tests

.contains 1.8 wt.% S corresponding to about 95% sulphur recovery when the activator .was added during conditioning Activ t0 . No significant effect was observed when

.the activator was added at the 8th min floating Activ t8 . The weight percentage ofthe concentrate obtained is about 48 for the two cases of this test series.

3.1.1.4. Pulp G. Among the tailings studied, tailings G was the easiest to desulphurizedespite its high sulphur content. Fig. 6a shows that the sulphur recovery reaches 96%with xanthate concentrations between 120 and 240 grt. The desulphurized final tailings

.obtained contain 1.4 wt.% S Fig. 6b and the concentrate obtained represent 50% of thetotal initial weight.

3.1.2. Amine acetateThe amine acetate collector was not tested for all of the tailings because of its

.relatively high price compared to xanthate Can$4.5rkg and Can$2.1rkg, respectively .

. .Fig. 5. Flotation kinetic as a function of collector dosage for tailings M. a Sulphur recovery vs. time %. bResidual sulphur % vs. time.


. .Fig. 6. Flotation kinetic as a function of collector dosage for tailings G. a Sulphur recovery vs. time %. bResidual sulphur % vs. time.

However, the amine acetate collector was necessary for pulp D cyanides containing.pulp because of the poor recoveries obtained using xanthate with or without an

activation stage. It can be noted here that for sulphide flotation, amine acetate is notaffected by the presence of cyanide. Sulphur recovery has been enhanced up to 95%with a concentration of 100 grt amine and the % S reached in the final tailings was 0.22

. .wt.% Fig. 7 . Even at lower concentration 50 and 75 grt , the performance of this .collector is still very good 85% and 90% S recovery, respectively at 10 min. The

amine acetate is confirmed for its immunity from the cyanide presence allowing forgood simultaneous non-selective sulphide flotation. The concentrate obtained in this caseis more diluted, possibly due to a relatively higher gangue minerals entrainment ratecaused by this collector compared to the amyl xanthate. Thus, the final concentraterepresents 20% of the total initial weight of the tailings.

3.2. Collector optimization

Another challenge for achieving good desulphurization performances is to properlyoptimize the collector addition and to minimize its consequent consumption during the

.flotation process. As shown in the kinetic curves for each tailings flotation Figs. 36 ,once beyond a certain collector concentration level, an addition of collector had onlylittle effect on sulphide recovery. Fig. 8 clearly shows this phenomenon and the optimalcollector concentration for each flotation,. Graphically, the optimum location is the

.inflexion point of the curves start of the plateau . For tailings M and G, 140 grt ofxanthate and 12-min flotation are required. A 20 grt xanthate concentration and a flash

. .Fig. 7. Flotation kinetic as a function of collector dosage for tailings D. a Sulphur recovery vs. time %. bResidual sulphur % vs. time.


Fig. 8. Collector dosages vs. residual sulphur in the tailings after various time flotation corresponding to thefour pulps studied.

flotation is sufficient for tailings P. Finally, optimum desulphurization for pulp D isabout 50 grt of amine and 10-min flotation. Note that in order to facilitate thepresentation of data in Fig. 8, dashed curves are used. Therefore, the lines have no realsignificance.

4. Interpretations and discussions

4.1. Kinetics flotation modeling

Among the flotation kinetic models found in the literature, the most suitable and mostwidely used to represent the recovery kinetics is the one proposed by Lynch et al.


. .1981 . This model was used by other authors e.g. Yuan et al., 1996; Klimpel, 1997 , ..and is a first-order model with rectangular distribution of the flotabilities see Eq. 1 .

.The experimental kinetic curves Figs. 3a7a were fitted using this mathematicalequation in order to determine the two parameters of the model: the final recovery Rfand the flotation rate constant k. The fitting was achieved at a correlation coefficientbetween 0.9 and 1.

1RsRf 1y 1yexp ykt 1 . . 5kt

.where R is the sulphur recovery % .Each point of Fig. 9a,b and Fig. 10 correspond to a flotation test that was presented in

Figs. 4a7a. Fig. 9a and b represents the sulphur recovery evolution vs. the modelparameters k and Rf, respectively. Fig. 10 represents the sulphide recovery as afunction of the collector dosage for tailings M, G, and D. Tailings P do not figurebecause of its very low sulphide content. It is important to note that the curve obtainedin Fig. 10 has the same shape as the one obtained for pure pyrite flotation using xanthate .Kongolo, 1991 .

As observed in Figs. 9a and 10, the shape of the relationships Rk and Rd is .similar to the one of the mathematical model Eq. 1 , which is used to fit these results.

Thus, the best fit equations obtained are the following:1

Rs105 1y 1.5yexp y5k 2 . . 55k1

Rs105 1y 5yexp y0.45d 3 . . 50.45dwhere R is the acceptable sulphide recovery %, k the flotation rate constant andd the optimal collector concentration for any given tailings.

.Moreover, the relationship between R and Rf is linear Fig. 9b and is .represented by Eq. 4 where Rf is the final recovery:

Rs0.996Rfy4.25. 4 .These three equations allow the development of a universal model that represents the

desulphurization process. From a known sulphide recovery R, one can evaluate the

Fig. 9. Kinetic model parameters as functions of cumulative sulphur recovery.


Fig. 10. Collector dosage as functions of cumulative sulphur recovery.

. .parameter k with Eq. 2 and the parameter Rf with Eq. 4 . These two parameters . .inserted in Eq. 1 lead to the estimation of the flotation time using Eq. 1 . It is also

.possible to estimate the collector dosage by placing R in Eq. 3 .

4.2. Acceptable desulphurization leel

As mentioned in the previous section, it is possible to predict the collector dosage andthe flotation time needed to reach a certain sulphide recovery. However, what is anacceptable sulphide recovery level for the environment? The main objective of thedesulphurization is to produce a sulphide concentrate and a non-acid generating fractionwhich must have enough NP to neutralize the acid produced by the residual sulphide. Todetermine if the tailings are acid or non-acid generating, different criteria have been

.proposed in the literature e.g. SRK, 1989; Morin and Hutt, 1997 . The static testcriterion used in this study is the NPrAP ratio which is represented in Fig. 11 with theprojection of the NNP of each tailings studied. This criterion considers a material as acidgenerating if the ratio is less than 1, in the uncertain zone if the ratio NPrAP is between1 and 2 and non-acid generating if the ratio is greater than 2. In this study, it wassupposed that the NP is constant during the flotation tests. This assumption is based on

Fig.11. NPrAP criterion for classifying tailings samples on the basis of static tests.


.preliminary results and the study of Catalan et al. 1999 . Moreover, removing the .sulphides in the concentrate particularly when the entrainment is low leads to a higher

NP in the final tailings due to the relative enrichment of the carbonate minerals.The main challenge of the desulphurization consists of optimizing the two most

important parameters of the global desulphurization costs: the flotation time and thecollector dosage. To achieve this, the environmental sulphur recovery needed R mustbe calculated. R corresponds to the sulphur to recover for moving the initial NPrAP

.point onto the line NPrAPs2 see Fig. 11 . For tailings M, R must be equal orgreater than 84.2%; R must be equal or greater than 90.9% for tailings G; and Rmust be equal or greater than 91.8% for tailings D. Because of its low acid generatingpotential, tailings P need a sulphur recovery of only 31.1%. Environmental flotation timecan be defined as the flotation time needed to obtain desulphurized tailings with the

.desired environmental characteristics i.e. without acid generating risk . Flotation timewas optimized using the environmental sulphur recovery of the tailings evaluated with

.NPrAP criterion and the kinetic model for which the final recovery Rf and time . .constant k have been first estimated with Eqs. 2 and 3 . As can be seen in Table 5,

the laboratory environmental flotation time is between 9 and 13 min for the cyanide free .tailings tailings M, G, and P and about 9 min for tailings D which contain cyanides

.flotation with AMAC-C . It can also be observed that it is not possible to desulphurizetailings D with KAX and reach the environmental sulphur recovery. Table 5 summarizesall of the desulphurization data including optimal collector concentration for eachmaterial studied.

4.3. Economic ealuation and adantages of desulphurizationThe previous sections showed that it is possible to remove most of the sulphide

minerals in acid-generating tailings by flotation and to produce two distinct fractions: asulphide concentrate and a desulphurized fraction non-acid generating. It was alsodemonstrated the flotation time and the collector dosage could be estimated from thecalculated environmental sulphur recovery. However, this new environmental approachwill be applied only if the costs associated with its implementation are comparable to theone associated to other existing techniques.

Table 5Optimization characteristics of the desulphurization process

Tailings M Tailings M Tailings G Tailings P Tailings D Tailings DqKAX qCuSO qKAX qKAX ycyanidation qcyanidation4

qKAX qCuSO qKAX qARMAC-C4S % tails for NPr 1.92 1.92 1.68 2.08 0.40 0.40APs2Environmental 84.20 84.20 90.91 31.1 91.76 91.76sulphur recovery %Environmental 12.8 12.8 9.2 flash impossible 8.7

.flotation time min 60 80 10 90w xOptimal collector

.grtm for 12 min


The total desulphurization costs are a combination of capital costs and operatingcosts. The capital costs are composed of the purchase costs of flotation equipment .flotation cells, pumps, tubing and installation costs. It is assumed in this study that theexisting buildings have the necessary space to integrate the new facilities. The selectionof flotation equipment is related to the production rate of tailings and flotation time. As

.mentioned before, the flotation time is estimated from Eq. 1 where the parameter k . .and Rf are evaluated with Eqs. 2 and 3 , respectively. The scale factor used for the

flotation time, to transpose laboratory data into industrial data, was 2. The installationcosts were considered to be equivalent to mill flotation equipment costs. Table 6

presents a description of the desulphurization costs estimated from the Mining Source-. .book, 1997 for three of the mines studied Mine D, G, and M . Mine P was not

.investigated because of its very low acid generating potential see Fig. 11 . As shown inTable 6, the operating costs of sulphur flotation are less than Can$0.35rt for cyanide

.free tailings Mine G and M . The operating costs for producing a desulphurized fractionof Mine D tailings, which contain cyanides, are about Can$0.55. The capital costs of thedesulphurization are estimated, for the three mines studied, to be around Can$1 000 000.These capital costs are for new equipment and could be reduced by using old flotationcells.

Even if the desulphurization cost is not negligible, this alternative could be in manycases an economic solution from the environmental point of view. For example, inCanada, many mines are presently using paste backfill which contain a certain amount

.of pyrite from 5 to 70 wt.%; e.g. Nantel, 1998; Ouellet et al., 1998 . The use of pastebackfill reduces the amount of tailings at the surface by a factor of about 60% Hassani

.and Archibald, 1998 . By combining the desulphurization process and the paste backfilloperation, it would be possible to practically eliminate the storage of harmful materials .sulphidic tailings at the surface and thus, reduce the rehabilitation costs of the tailingspond. These costs are usually considered between Can$100 000 to 200 000rhectare for

sulphidic materials Geocon, 1995; McMullen et al., 1997; Ricard et al., 1997; Bussiere`

Table 6Economical evaluation of the desulphurization for mines D, G, and M

Mine D Mine G Mine M .Production rate of tailings trday 3300 2135 2300

.Environmental sulphur recovery % 91.8 90.9 84.2Type of collector ARMAC-C KAX KAX

.Collector costs Can$rt 0.40 0.18 0.12Other operating costs including power 0.15 0.15 0.15

.and maintenance Can$rt .Total operating costs Can$ 0.55 0.33 0.27

.Flotation time min 17.5 18.5 25.73 .Flotation cells volume m 16 8 16

Number of cells required 8 11 8 .Cell costs Can$ 576 k 528 k 576 k

Other capital costs including installation, 576 k 528 k 576 k .pump, and tubing Can$ .Total capital costs Can$ 1152 k 1056 k 1152 k


.et al., 1998 . Another alternative for existing mines is to desulphurize the tailings, pump .the two fractions desulphurized tailings and sulphur concentrate at different locations

in the tailings pond and use the desulphurized fraction as cover material for the sulphidicfraction. Other studies have shown that tailings have suitable geotechnical characteristicsfor use as the moisture retaining layer in a multi-layered cover with the capillary barrier

.effect Aubertin et al., 1995; Bussiere et al., 1997b; Benzaazoua et al., 1998b . By using`the desulphurized fraction as cover material, the rehabilitation costs could be reduced bya factor between 10% and 35% due to a reduction of transportation costs. More detailsabout the economical advantages of using desulphurization can be found in Bussiere et`

. .al. 1997a and Bussiere et al. 1998 .`

5. Conclusions and perspectives

This study showed that the environmental desulphurization is technically feasible bynon-selective flotation. Cyanide-free pulps can be desulphurized by froth flotation withamyl xanthates. However, cyanided pulps must be floated with a more expansivecollector such as amine acetate without a pretreatment. The process generates both a

.sulphide concentrate and non-reactive desulphurized tailings i.e. non-acid generating .Through the tests realized, a model governed by four equations has been established.

This model will allow for the estimation of the optimal collector dosage and the flotation .time needed to obtain final tailings with an acceptable of NPrAP ratio )2 . Environ-

mental sulphur recovery is dependent on the NP. The standard or the modified ABA test .Sobek et al., 1978; Lawrence 1990 usually estimates this parameter. However, recentworks show the NP determination can be improved by including the mineralogical

.aspect carbonates vs. silicates mineral varieties as part of the effective NP evaluationLawrence and Scheske, 1997; Lawrence and Wang, 1997; Li, 1997; Skousen et al.,

.1997; Kwong and Ferguson, 1997 .The global costs involved with the desulphurization are economically attractive

compared to other tailings management methods. For the four cases examined, operat-ing costs are between Can$0.1 and 0.4rt of tailings. These costs include the capital,operating, and maintenance costs corresponding to the flotation circuit and the extensionof the existing process flow sheet. Desulphurization of mine tailings must be evaluatedin its overall context as an attractive alternative to the other techniques existing fortailings 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 permanentmonitoring of liquid effluent quality. Another worthwhile technique may be consideredat the same time as desulphurization is the use of paste fill technology to place thesulphidic fraction backs underground. The costs of surface rehabilitation could belimited, by this way, to the expense of desulphurization, disposal, and revegetation.

Acknowledgements

This work was financed through the Canadian MENDrNEDEM program, whichincludes a contribution from the Ministere des ressources naturelles du Quebec and`


CANMET. Our acknowledgements also go to the CRDAT who was a financialcontributor of the project. We would also like to thank Nil Gaudet and Jean Lelievre for`their technical works and Denis Bois, Peter Radziszewski, and Robert Houot for theirinteresting review of the paper. Finally, the authors would like to thank all our minepartners.

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