Extraction of gold from an Indian low-grade refractory gold ore through physical beneficiation and t

Extraction of gold from an Indian low-grade refractory gold ore

through physical beneficiation and thiourea leaching

D.S.R.Murthy *,Vinod Kumar,K.V .Rao

National Metallurgical Laboratory,CSIR,Jamshedpur 831007,India

Received 26July 2002;received in revised form 8November 2002;accepted 8November 2002

Abstract

The work carried out on the processing of a low-grade and refractory gold ore using thiourea is presented.Direct leaching of the ‘‘as-received’’ore with thiourea has yielded only 16%gold extraction.Starting with ore containing 1.4g/t of gold,it has been possible to generate a sulphide concentrate with 6.1g/t of gold by physical beneficiation of the ore through froth flotation.Thiourea leaching of this flotation concentrate after pressure,pre-oxidation has resulted in gold extraction of 98%,with an overall gold recovery of 74%.Pressure oxidation was found to be slightly superior to roasting for the pre-oxidation of the refractory gold ore/concentrate.

D 2003Elsevier Science B.V .All rights reserved.

Keywords:Low-grade gold ores;Leaching;Refractory ores;Floatation

1.Introduction

With the fast depletion of high-grade gold ores,the dependence on low-grade ores may increase globally in the near future.When a gold ore is below the cut-off level (1.75g/t),it will not be economical to use such ore directly.However,depending on the nature of the ore,it may be possible to generate a gold concentrate with enhanced gold content by suitable physical beneficiation techniques.Once such a con-centrate with a higher assay value of gold is produced,it may be possible to extract the gold with an advantage,because only lesser quantity of the material will be used in the subsequent down-stream leaching

operation.A low-grade gold ore from the southern region of India has been used in this study.

Although cyanide has been the primary lixiviant for the extraction of gold for over a century now,it is associated with certain shortcomings such as toxicity,long leaching times and considerable interaction with other nonferrous metals present in the gold ores.Because of this,attempts are being made by various investigators to search for alternative nontoxic lixi-viants (V on Michaelis,1987;Tran et al.,1992;Swa-minathan et al.,1993)for gold.One such leachant,thiourea (Brent Hiskey,1984;Groenewald,1976;Hisshion and Waller,1984)is known to give much faster dissolution of gold.As opposed to alkaline base leachants,thiourea forms gold cationic ions.Because of this,a variety of options such as activated car-bon adsorption,cationic ion exchange,solvent extrac-tion,electrowinning,cementation,etc.,are available

0304-386X/03/$-see front matter D 2003Elsevier Science B.V .All rights reserved.PII:S 0304-386X (02)00197-4

*Corresponding author.

E-mail address:dsrmurthy _nml@https://www.360docs.net/doc/1f11311258.html, (D.S.R.Murthy).

https://www.360docs.net/doc/1f11311258.html,/locate/hydromet

Hydrometallurgy 68(2003)125

–130

for the recovery of gold from thiourea leach liquors (Deschenes,1986).The dissolution of gold in acidic thiourea solutions in the presence of ferric ion may be represented as:

AutFe3tt2C SeNH2T2

?Au?CSeNH2T2 t2tFe2te1TGold ores that are not amenable to direct leaching by cyanide or any other leachant are known as refractory ores.The primary reason for this poor response is mineralogical(Robinson,1983),with the gold being finely disseminated in the sulphide miner-als such as pyrite,pyrrhotite,chalcopyrite,arsenopyr-ite,etc.In a broad classification of the gold ores(La Brooy et al.,1994)based on their response to cyani-dation,gold ores with greater than90%recovery were considered nonrefractory and those with less than 50%gold recovery were said to be highly refractory. In the case of such refractory ores,a pre-oxidation step often becomes necessary to oxidise the base metal sulphides and expose the gold particles for subsequent leaching.The main methods of pre-oxida-tion recommended for the refractory gold ores are roasting(McQuiston and Shoemaker,1975;La Brooy et al.,1994;Angelidis and Kydros,1995),biological oxidation(Lawrence and Bruynesteyn,1983;Deng and Liao,2002)and pressure oxidation(Mason,1990; Thomas,1991;Murthy et al.,1992).

The reactions involved in the roasting of the sulphides may be given as:

2FeS2t5:5O2?Fe2O3t4SO2e2T2CuFeS2t6:5O2?2CuOtFe2O3t4SO2e3TThe oxidation of the sulphide minerals during pressure oxidation may be represented as:

2FeS2t7:5O2tH2O?Fe2eSO4T3tH2SO4e4T

4FeS2t15O2t8H2O?2Fe2O3t8H2SO4e5T

2CuFeS2t8:5O2t2H2O

?2CuSO4tFe2O3t2H2SO4e6T

In the present work,the low-grade refractory gold ore was beneficiated by froth flotation and the gold concentrate generated was pre-oxidised and then leached with thiourea to extract the gold.

2.Experimental

2.1.Reagents and materials

All the reagents used in this work were of analyt-ical grade.A low-grade refractory gold ore from the Southern region of India was used in this investiga-tion.The chemical analysis of the‘‘as-received’’ore, which contained1.4g/t of gold,is shown in Table1.

2.2.Procedures

X-ray diffraction studies of the‘‘as-received’’ore as well as the flotation concentrate were carried out using a Seifert XRD3003PTS unit.The initial1kg flotation experiments and the bulk flotation experi-ments were carried out using standard flotation equip-ment.A series of experiments were conducted to arrive at the optimum values of grind size,collector and frother dosages required for the flotation studies. All the flotation experiments were carried out in duplicate and the average results are presented.

Leaching experiments were conducted using a lab-oratory size stirrer in a4-L glass beaker after adding all the required reagents and500g of the‘‘as-received’’ore or the pre-oxidised ore/flotation concentrate.Anal-ysis of gold from the leach liquor was carried out using a Shimadzu inductively coupled plasma spectropho-tometer.Gold analyses in all the solid samples were done by fire assay.The analyses of all other constitu-ents were done by conventional methods.Pressure oxidation of the‘‘as-received’’ore/concentrate was carried out in500g batches under mild acidic pH Table1

Analysis of the as-received ore

Constituent Amount present Au 1.4g/t

S0.880wt.% Cu0.021wt.%

Ni0.025wt.% Fe11.80wt.%

D.S.R.Murthy et al./Hydrometallurgy68(2003)125–130 126

conditions in a 7.0-L capacity Parr autoclave under oxygen pressure at 150j C,10atm pressure for 2h.Roasting of the ‘‘as-received’’ore/concentrate was conducted in a muffle furnace at 750j for 2h,with frequent rabbling.All the experiments were conducted in duplicate,whereas leaching tests were carried out in triplicate and the average results are presented.

3.Results and discussion 3.1.Material characterisation

Mineralogical as well as XRD studies were carried out on the ‘‘as-received’’ore,whereas only XRD studies were conducted on the flotation concentrate.From the megascopic/hand specimen studies,it is inferred that the ‘‘as-received’’ore is a quartz–biotite schist.Microscopic studies of the polished sections indicated the presence of pyrite as the main sulphide mineral with traces of chalcopyrite.No other sulphide mineral or any native gold was observed.The pyrite in the sample was present along the grain boundaries and inter-granular spaces of the silicates.Rarely they were present as inclusions within the silicates.The grain size of these pyrites is between 5and 25A m and most probably,these pyrites may contain some gold in them.The liberation size of these pyrites is below 150mesh (100A m).

The X-ray diffraction patterns of the head sample as well as the flotation concentrate,which are shown in Figs.1and 2,indicated quartz as the major mineral with subordinate amounts of biotite.The presence of pyrite or chalcopyrite could not be detected in the XRD pattern of the head sample because of their low con-centrations.However,in the flotation concentrate,sulphide mineral phases such as pyrite and chalcopyrite could be identified along with quartz and biotite.Since no gold phase could be identified,it was concluded that the gold present in the sample was of invisible type.3.2.Physical beneficiation

3.2.1.Froth flotation

Since the head sample contained around 0.88%of sulphur,it was decided to use froth flotation to collect the sulphide minerals and thereby generate a sulphide concentrate with enhanced gold assay values.One hundred kilograms of the head sample was drawn from the bulk sample by the coning and quartering method and was used for the further studies.3.3.Grind variation studies

A series of grind variation studies were conducted using 1kg batches of wet ground samples.Grinding was done in a Denver rod mill for different lengths of time i.e.5,8and 11min and the fineness in terms of %–200mesh (Tyler)(74A m)was noted.The

ground

Fig.1.XRD pattern of the ‘‘as-received’’

ore.

Fig.2.XRD pattern of the flotation concentrate.

D.S.R.Murthy et al./Hydrometallurgy 68(2003)125–130127

material was subjected to froth flotation in a Wemco (Fagergren)flotation cell under conditions mentioned in Table2.Sodium silicate was used to depress the silicate minerals.Sodium isopropyl xanthate(SIPX)

as collector and pine oil as frother were used.The flotation experiments were conducted both in acidic and alkaline pH conditions.Three floats(concen-trates)and one tail were collected for each grind. All three concentrates were mixed and analysed for sulphur.These results are presented in Table3,from where,it can be seen that the concentrates obtained at 65.5%–200mesh size(74A m)analysed2.30%S and the yield was12.4%by weight.Though there was up-grading in the sulphur content,the yield was not encouraging and this may be due to coarse grinding. In the experiments conducted with77.8%–200mesh (74A m),there was marked improvement both in the sulphur content and yield.Further experiments con-ducted with still finer size,i.e.86.1%–200mesh (74A m)resulted in poor yield of sulphur.Hence,a grind of77.8%–200mesh(74A m)was chosen for further flotation experiments.

3.4.Optimisation of collector dosage

The dosage of the collector was varied between0.06 and0.18kg/t and that of the pine oil was maintained at 0.05kg/t.These results are presented in Table4which shows that the flotation concentrate obtained with0.06 kg/t of collector dosage was not encouraging since its sulphur assay was only1.79%though the yield was appreciable.When the collector dosage was increased to0.12kg/t,an increase in the sulphur assay of the floats coupled with reasonable yield of sulphur was

Table2

Flotation conditions used

Reagent Quantity

(kg/t)Conditioning

time(min)

Flotation

time(min)

Product

Sodium

silicate

1.0002––SIPX0.0402–

Pine oil0.0171/22Float I SIPX0.0402–

Pine oil0.0171/22Float II SIPX0.0402–

Pine oil0.0171/22Float III

Table3

Results obtained on grind variation studies a

Time(min)/grind size %–200mesh(74A m)Product Wt.(%)Assay

(%)S

Dist.

(%)S

5/65.5Floats12.4 2.3032.0

Tails87.60.6968.0

Head

(Calc.)

100.00.89100.0

8/77.8Floats19.8 2.5356.0

Tails80.20.4644.0

Head

(Calc.)

100.00.89100.0 11/86.1Floats11.8 2.4032.4

Tails88.20.6767.6

Head

(Calc.)

100.00.87100.0

a Average values of two tests.Table4

Results obtained with variation in collector dosages a

Collector

dosage(kg/t)

Product Wt.(%)Assay

(%)S

Dist.

(%)S 0.06Floats21.3 1.7943.9

Tails78.70.6256.1

Head

(Calc.)

100.00.87100.0

0.12Floats19.8 2.5356.0

Tails80.20.4644.0

Head

(Calc.)

100.00.89100.0

0.18Floats26.4 2.6148.7

Tails83.60.5451.3

Head

(Calc.)

100.00.87100.0

a Average values of two tests.

Table5

Results with various frother dosages a

Frother

dosage(kg/t)

Product Wt.(%)Assay

(%)S

Dist.

(%)S 0.025Floats22.4 2.2757.7

Tails77.60.4842.3

Head

(Calc.)

100.00.88100.0

0.050Floats19.8 2.5356.0

Tails80.20.4644.0

Head

(Calc.)

100.00.89100.0 0.075Floats15.9 2.4346.6

Tails84.10.5755.4

Head

(Calc.)

100.00.87100.0

a Average values of two tests.

D.S.R.Murthy et al./Hydrometallurgy68(2003)125–130 128

found.Further experiments with0.18kg/t of collector dosage were not encouraging.Therefore,0.12kg/t of

collector was considered to be suitable.

3.5.Optimisation of frother dosage

In order to establish the ideal frother dosage, flotation experiments were conducted with optimum grind and collector dosage,while the frother,pine oil was varied in the range of0.025–0.075kg/t.These results are presented in Table5based on which,it was established that the suitable frother dosage was 0.05kg/t.

3.6.Bulk production of the concentrate

Bulk sulphide flotation experiments were con-ducted in a Denver unit cell under the optimum conditions established earlier(grind size77.8%–200 mesh(74A m),collector0.12kg/t and frother0.05 kg/t).The bulk rougher concentrate was further up-graded by adopting one cleaning stage without using any reagents.The products were analysed for both sulphur as well as gold contents.The average values of the results obtained in these experiments are presented in Table6from where,it can be seen that the sulphide concentrate obtained analysed4.30%S,and6.10g/t Au with sulphur and gold yields of86.5%and75.6%, respectively.The concentrate thus obtained was used in the subsequent leaching experiments.The chemical analysis of this concentrate is shown in Table7.

3.7.Leaching of gold

Leaching of gold was carried out in acidic thiourea solutions in the presence of ferric ion.Leaching conditions were maintained identical in all the experi-ments.The results obtained on the leaching of the‘‘as-received’’ore and the pre-oxidised ore using roasting/ pressure oxidation are shown in Table8from where it can be seen that the roasted ore has resulted in63.4% gold extraction,whereas the pressure oxidised ore has yielded67.5%gold extraction,as against16.4%gold extraction obtained in the direct leaching of the‘‘as-received’’ore.

The results obtained on the leaching of the flotation concentrate after pre-oxidation using roasting/pressure oxidation are presented in Table9from where it can be observed that with calcination,the gold extraction

Table6

Results of bulk flotation with one cleaning a

Grind used Product Wt.Assay Dist.(%)

%–200mesh (74A m)(%)S

(%)

Au

(g/t)

S Au

77.8Conc.17.8 4.30 6.1086.575.6

Middlings27.80.170.80 5.315.5

Tailings51.40.140.258.28.9

Head

(Calc.)

100.00.88 1.44100100

a Average values of two tests.

Table7

Analysis of the flotation concentrate

Constituent Amount present Au 6.1g/t

S 4.30wt.%

Cu0.15wt.%

Ni0.18wt.%

Fe21.30wt.%Table8

Thiourea leaching of the‘‘as-received’’gold ore a

Nature of

the ore used

Assay of

gold in the

starting material

Assay of

gold in the

leach residue

%Gold

extracted b

As-received 1.4 1.1716.4 Calcined 1.40.5163.4 Pressure

oxidised

1.40.4567.5

Amount of ore taken=500g,solid/liquid ratio=1:4,concentration of thiourea=0.15M,concentration of ferric sulphate=0.012M. Temperature=25j C.

a Average of three separate leaching experiments.

b Based on the fire assay of the leach residue.

Table9

Thiourea leaching of the gold concentrate a

Nature of

the concentrate

Assay of gold

in the starting

material(g/t)

Assay of

gold in the

leach(g/t)

%Gold

extracted b

Roasted 6.10.2396.2 Pressure oxidised 6.10.1298.0 Amount of concentrate taken=500g,solid/liquid ratio=1:4, concentration of thiourea=0.15M,concentration of ferric sulphate=0.012M.Temperature=25j C.

a Average of three separate leaching experiments.

b Based on the fire assay of the leach residue.

D.S.R.Murthy et al./Hydrometallurgy68(2003)125–130129

was96.2%,whereas the pressure oxidised concen-trates has given a gold extraction of98.0%.Taking into account that the percent distribution of gold recorded into the flotation concentrate was75.6%, the overall gold extractions obtained in the thiourea leaching of the pre-oxidised concentrate using roast-ing and pressure oxidation are found to be72.7%and 74.1%,respectively.

4.Conclusions

It has been possible to physically beneficiate a low-grade refractory gold ore with 1.4g/t of gold to generate a sulphide concentrate containing6.1g/t of gold by froth flotation technique.Direct thiourea leaching of the‘‘as-received’’ore has given a gold extraction of only16.4%while pre-oxidised ore has yielded gold recoveries of63.4%and67.5%,corre-sponding to roasting and pressure oxidation,respec-tively.Leaching of the flotation concentrate after pre-oxidation resulted in gold recoveries of72.7%and 74.1%,with respect to roasting and pressure oxida-tion.Physical beneficiation of the ore followed by subsequent leaching of the gold concentrate was found to be effective because of the dual advantages of handling a much smaller amount of material during leaching coupled with higher recoveries of gold than the direct use of the ore itself.As a pre-oxidation technique,pressure oxidation was found to be slightly more effective than roasting. Acknowledgements

The authors are thankful to Prof.S.P.Mehrotra, Director,National Metallurgical Laboratory,Jamshed-pur for his keen interest in this work and permission to publish this paper.References

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