Matching laboratory and plant performance––a case study of the Elura lead circuit, Pasminco Australia Limited

Minerals Engineering 17 (2004) 811–824 This article is also available online at: www.elsevier.com/locate/mineng

Matching laboratory and plant performance––a case study of the Elura lead circuit, Pasminco Australia Limited M.C. Pietrobon a, S.R. Grano

b,*

, C. Greet

a

a

b

Amdel Ltd., P.O. Box 338, Torrensville Plaza, Adelaide, SA 5031, Australia Ian Wark Research Institute, University of South Australia (The ARC Special Research Centre for Particle and Material Interfaces), Mawson Lakes Campus, Adelaide, SA 5095, Australia Received 1 September 2003; accepted 3 February 2004

Abstract A common problem in assessing the applicability of laboratory flotation data is determining how closely conditions during laboratory testing of the ore matches those that occur in a plant. Typical laboratory based investigations do not consider factors such as water quality, pulp temperature and the pulp chemical environment during both grinding and conditioning on laboratory flotation performance. Disregarding these important factors may seriously compromise the validity of conclusions based solely on laboratory experimental results. Furthermore, considering these factors during the flowsheet design stage in the development of new ore resources may reduce financial risk associated with building a new plant. A methodology is described here which embraces these factors in an attempt to bridge the gap between laboratory and plant. The methodology should be considered a vital first step in all laboratory based experimental programs and has implications for fundamental studies on the flotation behaviour of single minerals. This methodology increases the confidence with which findings of laboratory based studies can be transferred to the plant and highlights the effect of variables which may be accessible to manipulation in the plant. Pulp chemical parameters, such as pulp dissolved oxygen demand, pH, Eh, mineral oxidation and solution composition were manipulated in the laboratory by adjusting the duration of gas purging after grinding, the process water input and pulp temperature. A key finding was that matching of the laboratory pulp chemistry to the plant pulp chemistry was dependent upon oxidation after grinding. In turn, mineral oxidation was dependent upon the pH in grinding, water composition and the duration of oxygen purging after grinding. The methodology closely matched the dissolved oxygen demand and surface chemistry of laboratory and plant pulps.
2004 Elsevier Ltd. All rights reserved. Keywords: Sulphide ores; Froth flotation; Grinding; pH control

1. Introduction The Elura ore body (Pasminco Australia Ltd.) is a huge vertical plug of massive sulphide ore rich in zinc, lead and silver. There are four main ore types: (1) pyrrhotitic ore consisting of
15% coarse grained (40– 50 lm) pyrrhotite,
45% pyrite and fine grained (15 lm) galena/sphalerite, (2) pyritic ore consisting of
55–60% pyrite,
0–5% pyrrhotite and fine grained (15 lm) galena/sphalerite, (3) siliceous–pyrite ore consisting of
55–60% pyrite and pyrrhotite substituted by silica and * Corresponding author. Tel.: +61-8-8302-3106; fax: +61-8-83023683. E-mail addresses: [email protected] (M.C. Pietrobon), [email protected] (S.R. Grano), [email protected] (C. Greet).

0892-6875/$ - see front matter
2004 Elsevier Ltd. All rights reserved. doi:10.1016/j.mineng.2004.02.003

very fine grained (8 lm) galena/sphalerite, and (4) siliceous–pyrrhotite ore consisting of
15% pyrrhotite with pyrite substituted by silica and very fine grained (8 lm) galena/sphalerite (Allen et al., 1993). The ore dealt with here is the pyritic ore. The high iron sulphide (i.e., pyrite plus pyrrhotite) content of Elura ores affects the processing of these ores in flotation. Specifically, Elura ores exhibit a high dissolved oxygen demand and acid generating capacity due to iron sulphide oxidation. The effect of dissolved oxygen demand is dealt with in more detail below. The fine grained nature of the ore necessitates both fine primary grinding and regrinding. The concentrator comprises sequential flotation circuits recovering galena followed by sphalerite. The Elura concentrator flowsheet and the chemical conditions employed in processing are discussed further below. In this study, the focus was on matching the pulp and

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surface chemistry in both the grinding and conditioning stages of laboratory experiments to that in the plant lead circuit at the equivalent, key process points (viz, mill discharge and flotation feed). The ultimate objective was to match the flotation behaviour in laboratory experiments to plant flotation behaviour. This study formed part of a larger study in the AMIRA P260C (ÔMineral Flotation’) project, a subproject of which was on assimilating plant and laboratory flotation behaviour. The adopted approach to match flotation behaviour was based on the understanding that the flotation rate constant is dependent upon both chemical factors (i.e., surface properties of the exposed minerals in the conditioned feed to flotation) and physical factors during flotation itself (i.e., bubble size, particle size, gas flow rate, hydrodynamics). The flotation rate constant in the collection zone is given by Eq. (1), (Ralston, 1994): k ¼E

3Gfr h 2db  Vr

ditioned feed, differences in the physical variables (i.e., Gfr and db ) will give rise to differences in the flotation rate constant, assuming that the capture efficiency is the same for the two cells. This is the basis for scale up in other studies which have focused on bubble surface area flux ðSb Þ (Gorain et al., 1997). The effect of these physical variables on the correlation of laboratory cells to plant scale cells is the subject of a different, but related, investigation (Pietrobon, unpublished work). In this study, emphasis was placed on matching the flotation behaviour of plant and laboratory slurries when using the same physical variables in flotation. This is achieved by using the same laboratory flotation cell and physical conditions in the laboratory cell for tests on both plant streams and laboratory samples. 1.1. Research methodology To successfully assimilate plant and laboratory flotation data it is useful to consider the chemical and physical variables separately. In the case of the chemical variables, these are manifested through the values of Ea and Ed which are dependent upon the surface properties of the minerals in the conditioned feed, while the physical variables are manifested through the values of db , Gf and Ec . A two step approach to the problem of assimilating plant and laboratory flotation data was adopted in the project overall:

ð1Þ

where E is the capture efficiency, Gfr is the gas flow rate, db is the bubble size, Vr is the reference volume and h is the height of the collection zone. The capture efficiency is given by Eq. (2), (Derjaguin and Dukhin, 1960–1961): E ¼ Ec  Ea  Ed

ð2Þ

where Ec is the collision efficiency, Ea is the attachment efficiency and Ed is the detachment efficiency. The collision efficiency ðEc Þ is controlled by particle and bubble size, as well as the flotation cell hydrodynamics and is not considered further here. Rather, emphasis is placed on matching the parameters that are influenced by particle hydrophobicity (i.e., Ea and Ed ). In these cases, pulp and surface chemistry control the particle hydrophobicity and the efficiency of particle-bubble attachment and detachment. In the case of two cells of different volume (e.g., plant and laboratory cells) treating the same con-

(a) Matching the collection efficiency (E) by Ôpulp chemical matching’ of the laboratory and plant conditioned feeds (Fig. 1). The flotation behaviour of the plant conditioned feed and laboratory conditioned feed should be identical if the same physical variables are used in flotation and the same mineral particle size distributions are treated. This corresponds to the case of testing the plant conditioned Ore

Process Water

Reagents Laboratory Grinding

Media

Reagents Plant Grinding Circuit Plant Grinding Circuit Product Reagents Plant Conditioning

Reagents Laboratory Conditioning b

c

Plant Conditioned Feed d a

Key Points for Matching

Laboratory Flotation Cells

Plant Flotation Cells

Laboratory Flotation Cells

‘Pulp Chemical Matching’

‘Physical Variable Scale-Up’

Fig. 1. Overall methodology for correlating the laboratory flotation performance to the plant flotation performance.

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813

The work described here focused on pulp chemical matching of laboratory pulps to plant streams in order to develop methods of simulating flotation behaviour of plant pulps using laboratory generated pulps. The work also highlights the significant number of parameters which can influence the ultimate flotation performance of a sulphide mineral system. The case of the Elura lead circuit of Pasminco Australia Limited was chosen for this study. In the study a survey of the plant was conducted in which both pulp and surface chemical information was gathered. A specifically chosen sample of slurry from the plant was also tested in a laboratory scale flotation cell for correlation to the laboratory based studies on plant mill feed. Furthermore, a sample of mill feed was collected and tested at laboratory scale. In these latter experiments, selected variables of relevance to the Elura lead circuit were manipulated, and their effect on pulp chemistry and flotation determined. It was anticipated that through careful matching of the various significant parameters that a close correlation between the plant and laboratory flotation behaviour may be found.

feed slurry with the same cell and with the same physical conditions (i.e., c and d in Fig. 1) and the Ôchemically matched’ laboratory conditioned feed (i.e., b in Fig. 1). Most accurate matching would be achieved by matching the capture efficiency (E) on a size-by-size basis. This aspect of the matching has not been discussed in this paper. The size-by-size flotation behaviour is important as detachment efficiency, at least, is dependent upon particle size, as well as particle hydrophobicity. However, in this study very similar particle size distributions for both the plant and laboratory feeds are used allowing the effect of particle size to be disregarded in this instance. (b) Providing appropriate scale up of the physical variables of flotation using db , Gf and Ec . For two different cells treating the same chemically matched feed, differences in flotation rate will be caused by differences in these variables. Thus, modifying the measured capture efficiency by correction with these physical variables should result in the same values of the rate constant for two different cells. This corresponds to the case of testing the plant conditioned feed in the plant flotation cells (i.e., a in Fig. 1), the chemically matched laboratory feed in the laboratory cell (i.e., b in Fig. 1), and the plant conditioned feed in the same laboratory cell (i.e., c and d in Fig. 1). Another important factor to consider in the scale up criteria is the froth recovery as distinct from collection zone recovery (Savassi et al., 1997). The results from this work are described elsewhere (Pietrobon, unpublished work).

2. Experimental 2.1. Elura grinding and lead circuits The Elura circuit processes approximately 140 tonnes per hour of lead/zinc ore through a SAG/ball mill circuit (Fig. 2). The primary cyclone pack classifies the SAG mill discharge, with the cyclone overflow stream reporting directly to two conditioning tanks (1 and 2)

d80 53 µm SAG Feed

pH10 SAG Mill

Conditioning Tank 1

O2

Ball Mill

Process Water Lime

Ball Mill Discharge pH 7.5

Recycle Streams Conditioning Tank 2

O2 60g/t 3418A Rougher 1

Key Points for Matching from Plant to Laboratory

10g/t 3418A

Scavenger Concentrate

Rougher 2

pH 9 Scavenger

Cleaner Conditioner

Scavenger Tailing

70g/t NaCN

Cleaner Block (including regrind to d80 20 µm)

1st Cleaner Tailing

Pb Final Concentrate

Fig. 2. Elura lead circuit (June 1998) showing reagent addition points, d80 and pH values. Key points for chemical matching are the ball mill discharge and Tank 1 product streams. Laboratory flotation tests were conducted on Tank 1 product.

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that are purged with oxygen gas. These tanks are in series and purge the pulp with a combined total of
300–350 m3 h1 of oxygen gas (or 2.1–2.5 dm3 kg1 of ore). The discharge from the second tank is lead rougher feed. The cyclone underflow is reground in a ball mill which is closed-circuited with a cyclone pack. Experimental work focused on the grinding and conditioning stages of the Elura lead circuit, with key points for matching being the plant ball mill discharge and Tank 1 product (Fig. 2). Other studies have shown that the grinding chemical environment (viz, Eh, pH, media type, etc.), amongst other factors, controls to a large extent fine galena maximum recovery (Cullinan et al., 1999). Hence, it is vital to match the pulp chemistry of the ball mill discharge stream, in addition to and distinctly from the flotation feed stream. Further research is proceeding using a specially designed laboratory mill that will allow closer control of the chemical conditions during grinding (Grano et al., 2001). Tank 1 product was also chosen for matching as this stream is at flotation feed particle size, includes a stage of oxygenation, but excludes recycle streams which feed Tank 2 (Fig. 2). Hence, the extent of post-grinding oxygenation became an important variable to match the pulp chemistry and flotation behaviour from plant to laboratory. The collector used was di-isobutyl dithiophosphinate (Cytec 3418A) which other studies have shown to be a powerful galena collector under the chemical conditions relevant to galena flotation at Elura (Grano et al., 1991). Specifically, these studies have shown that this collector exhibits higher galena recovery than the equivalent xanthate collector at Eh values less than 100 mV (SHE). This suggests that this collector type may be beneficial to galena recovery from pulps which show high dissolved oxygen demand such as those in the Elura concentrator. The interaction of collector type and oxygenation is the subject of another study. The collector, 3418A, is added to Tank 2 product prior to rougher flotation, as well as stage added to scavenger feed (Fig. 2). Hence, in laboratory flotation experiments on Tank 1 product, 3418A was added in conditioning to induce galena flotation. 2.2. Plant survey A block metallurgical survey of the lead flotation circuit was carried out and a SAG mill belt feed sample (
400 kg) collected for laboratory experiments. The SAG feed sample was crushed, riffled and sub-sampled into 1 kg lots using standard methods. During the survey period, laboratory flotation tests were conducted on Tank 1 product stream for the purpose of referencing to laboratory experiments on SAG mill feed after laboratory grinding. Furthermore, pulp samples were collected for solution cation and anion analysis throughout the circuit. Conditioning of selected pulp samples with EDTA was conducted to give a qualitative analysis of

the extent of mineral oxidation. Samples for surface analysis were also collected during the survey period. A large quantity of process water was collected for use in laboratory testing. The various methods and techniques are described further below. 2.3. Process water quality A sample of process water was collected during the survey and used in some flotation tests. The sample was also analysed for major chemical species as part of the matching study. The solution data obtained from the analysis of the process water and other water samples obtained from the pulps were chemically balanced using a chemical speciation program called MinteqA2/Prodefa2 (version 3.0, 1991, US Environmental Protection Agency). This program is a geochemical assessment model for environmental systems and serves as a quantitative tool for predicting the equilibrium behaviour of metals under different chemical environments. This was used to predict species most likely to precipitate from the process water and pulp solutions. The method of determining the contributions to the dissolved and precipitated species in the process pulps will now be described. The grinding circuit can be considered to be a chemical reactor in which the feeds to the grinding circuit (viz, SAG feed ore, process water, reagents such as lime in this case) (Fig. 2), contribute to the dissolved and precipitated species in the pulp of the grinding circuit product. Mathematically combining these inputs (using unit time as the basis) gives rise to a simulated pulp composition prior to precipitation. This method assumes that dissolution of the feed ore minerals is similar to that which occurs in the presence of reagents and conditions used in the plant process. It is therefore important to measure the dissolution of the ore feed sample at laboratory scale under conditions similar to the plant. This is especially important in the case of pH which controls dissolution of ore components. In experiments on SAG feed which aimed at measuring ore dissolution during grinding, the grind pH was controlled to the cyclone overflow value (pH 9.1). The simulated pulp composition is then allowed to come to equilibrium, with precipitation and exchange with the atmosphere allowed, through the MinteqA2 program. The equilibrated solution composition is then compared to the actual process water composition at the selected sample point. Furthermore, the amount and type of species precipitated are predicted by the MinteqA2 program and may be compared to EDTA extraction and surface analysis data. 2.4. Mineralogy Detailed mineralogy of the SAG feed sample was obtained using quantitative X-ray diffraction. The

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Table 1 Mineralogy of the SAG feed sample (wt.%) Mineral phase Galena 9.2

Sphalerite 11.9

Chalcopyrite 1.3

Pyrite 47.1

Pyrrhotite 2.1

Siderite 12.2

Quartz 6.6

Ankerite 4.4

Calcite 5.0

analysis was conducted by Sietronics using X-ray diffraction and quantified using Siroquant (Table 1). The mineralogy results indicate that the SAG feed sample had high levels of iron sulphides (approximately 50% by weight), mainly as pyrite.

Gas

5 3

2.5. EDTA metal ion extractions A 3% w/w solution of AR grade ethylenediaminetetra acetic acid disodium salt (EDTA) was made in deionised water and adjusted to pH 7.5 with sodium hydroxide. The EDTA stock solution was stirred while being continuously nitrogen purged. A pulp sample was syringed into the EDTA solution and conditioned for 5 min. The ratio of EDTA solution to pulp was maintained at 20:1. The slurry was then filtered (at 0.45 lm) to remove solids and colloidal matter. The solutes were assayed using ICP, previously calibrated with liquor samples containing EDTA at the same concentration. The solids were retained after leaching to obtain the dry weight (solids g) and assayed. Calculations follow that shown below (Rumball and Richmond, 1996): EDTA extractable metal as a %of total metal ¼ 20 

ðtotal mg in leach liquorÞ ðsolids g  metal ion assayÞ

ð3Þ

4

1

2

Fig. 3. Specialised grinding mill (1) contains a ball charge (2) and a divider plate (3) that contains openings covered with 300 lm aperture screens. The pulp is removed from the sampling chamber (4) and pumped to a conditioning vessel (5) that contains pH, Eh, dissolved oxygen and temperature probes. The slurry circulates back to the mill from the conditioning vessel.

be monitored and adjusted continuously during grinding (Fig. 3). The importance of pH control during grinding is commented upon further below. The grind pH was 9.1, adjusted with either lime or sodium hydroxide. 2.7. Flotation test procedure

2.6. Laboratory grinding Laboratory grinding used 1 kg of crushed SAG feed sample, ground for an appropriate time in a specialised mild steel ball mill (Fig. 3), with 9.0 kg of white iron, conically shaped cylpeb balls until a d80 of 53 lm was obtained. This type of grinding media is used in the plant ball mill. Particular care was taken to closely match the particle size distribution of the laboratory produced pulp to that of the plant pulp (i.e., Tank 1 discharge) measured during the survey period. Results of the particle size matching are commented upon further below. Particle sizes were measured using a combination of laser sizing and screens. The samples were ground at 62.5% solids in either process or demineralised water. The specialised mill allows the control of chemical conditions within the mill during grinding (Cases et al., 1990). While similar in design to traditional laboratory mills, the design incorporates a peristaltic pump that circulates part of the slurry during grinding through a conditioning vessel. In this way, the pH and other pulp chemical parameters can

In the case of laboratory ground SAG feed samples, the slurry was transferred to a 3 dm3 Agitair flotation cell and diluted with water (process or demineralised water) to approximately 28% solids (w/w). An oxygen purging period of specific time and gas flow rate was then employed, followed by conditioning with 3418A collector (60 g/t) for 2 min. The effect of oxygenation time, pulp temperature and water quality were investigated. In the case of plant samples (Tank 1 product), a precise volume of pulp (and therefore solids) was collected and tested undiluted at 28% solids (w/w). The pulp was then oxygen purged for varying times at a fixed gas flow rate, adjusting the pH to 9–9.2 with lime as required. The pulp was then conditioned with 3418A collector (60 g/t) for 2 min. In the case of both laboratory ground SAG feed and plant samples, four separate lead concentrates were collected at individual times of 0.5, 1.5, 2.0 and 4.0 min. The entire surface of the flotation cell was scraped at a fixed rate to ensure reproducibility. Frother addition was not necessary when using process water.

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2.8. Dissolved oxygen demand

the initial analysis, surface etching of the examined area involves sputtering the sample in the chamber with an Arþ ion beam for a specific time. This removes the overlying layer (
50 nm), and the underlying surface reanalysed.

The dissolved oxygen demand of the pulp was measured using a method adapted from Spira and Rosenblum (1974). Pulp was transferred to a laboratory flotation cell and agitated. The dissolved oxygen concentration, Eh, pH and temperature were all recorded. The pulp was then purged with oxygen gas at a rate of 10 dm3 min1 until the dissolved oxygen value in the pulp reached
20 mg dm3 . The oxygen gas flow was then ceased, and the dissolved oxygen concentration measured as a function of time. Once the dissolved oxygen reached a steady state value (changing at
0.1– 0.2 mg dm3 min1 ) the sequence was repeated at least 3 more times.

3. Results 3.1. Plant survey 3.1.1. Metallurgical survey Attention will turn firstly to the metallurgical survey of the plant. While the plant metallurgical performance itself is not a principal focus of this current discussion it is pertinent to make some comments at this stage. The mass balanced cyclone overflow (and Tank 1 product) assays were 6.0% Pb, 9.0% Zn and 34.6% Fe (Fig. 4). These streams constitute the fresh feed to the lead circuit. Lead rougher flotation feed is made up of fresh feed (100% mass) and the recycle streams, making a direct comparison of plant and laboratory flotation results difficult. There is a large circulating load derived from the scavenger concentrate (22% mass) and first cleaner tailing (i.e., cleaner block tailing) (15% mass). The cumulative concentrate grades and recoveries for the rougher and scavenger concentrates with respect to lead rougher flotation feed (i.e., fresh feed plus recycle streams) are shown in Table 2. In increasing the lead recovery by 18.6% in the plant scavengers there was a 11.0% increase in zinc recovery, suggesting poor galena/sphalerite selectivity in this part of the circuit.

2.9. X-ray photoelectron spectroscopy X-ray photoelectron spectroscopy (XPS) is a surface sensitive technique which analyses a surface region to 10 nm in depth (i.e., the first few atomic layers of a surface). The photoelectron peak position and intensity provide quantitative information, giving the surface composition and chemical states of detected elements. An area of approximately 0.5 cm2 was analysed and therefore the analysis represents an average for this area. The samples were placed into the preparation chamber and out-gassed under vacuum. The background pressure during analysis was typically 1 · 108 mbar. Samples examined by XPS were deslimed using a solution replacement treatment which does not incorporate ultrasonics. This method uses pH adjusted water (in this case pH 9.1) to remove colloidal and excessive solution species. After

Cyclone Overflow 100.0 100.0 100.0 6.0 9.0 34.6

SAG Feed

SAG Mill

Ball Mill

Ball Mill Discharge

Recycle Streams

Scavenger Concentrate 26.6 7.2

Rougher Feed 142.1 123.2 143.6 6.25 8.1 36.2

Rougher 1

Rougher 2

Rghr 1 Conc. 72.8 26.9

8.9 4.9

18.7 15.3

25.7 40.1

Scavenger Tail

Rghr 2 Conc.

13.9 29.5

13.5 5.5

4.1 5.1

Cleaner Conditioner

Scavenger 24.2

1.60

7.6 35.8

96.7 9.6

96.4 36.4

Cleaner Block Tail 15.5 6.2

Key

9.7 5.8

17.9 41.0

Pb Rec% Zn Rec% Fe Rec% Pb %

Zn%

Fe%

Pb Final Concentrate 75.8 53.4

3.3 3.5

3.6 14.6

Fig. 4. Elura lead circuit (June 1998) mass balance during the survey period showing the mass balanced lead, zinc and iron grades and recoveries.

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Table 2 Plant lead rougher and scavenger concentrate grades and recoveries with respect to flotation feed Stream

Pb

Zn

Cumulative conc. grade (%) Rougher conc. 1 Rougher conc. 1 + 2 Rougher conc. 1 + 2 plus scav. conc.

Cumulative conc. rec. (%)

26.9 23.3 15.5

Rougher flotation feed

6.25

Cumulative conc. rec. (%)

51.3 64.5 83.1

7.2 10.5 21.5

100.0

100.0

3.1.2. Pulp chemical survey Key pulp chemical parameter values during the survey are shown in Table 4. The pH in the SAG mill discharge is at 10.0 due to lime addition to SAG feed. Lime addition, process water, and ore dissolution inputs all contribute to the dissolved and precipitated metal ion composition of the grinding circuit product. The proportion of each contribution is discussed further below. However, it is clear that the process water is a major contributor to the grinding circuit product composition, at least in the case of magnesium, sodium, chloride and sulphate. The pH of the ball mill discharge is lower due to oxidation of sulphide minerals and grinding media during ball mill grinding for which there is no separate lime addition point (Fig. 2). At this sample point, the Eh is at its lowest value in the circuit. The Eh increases in passing the pulp to the cyclone overflow and after oxygenation in Tank 1. Clearly, it is important to match

These plant results may be compared with the results for laboratory flotation tests on Tank 1 product conducted during the survey period (Table 3). In the latter case, the head grade to the flotation test (Tank 1 product) is lower because of the absence of the higher grade recycle streams. Factors which contribute to the difference between the plant cells and laboratory cells flotation behaviour are not the focus of this paper, but include: (1) differences in the inherent floatability of the recycle streams relative to fresh feed (Runge et al., 1997), (2) differences in the capture efficiency in the collection zone (Pietrobon, unpublished work), and (3) differences in froth zone recovery of value and gangue minerals by genuine flotation and/or entrainment mechanisms (Savassi et al., 1997). Differences in the ultimate recovery of lead and zinc are most likely due to the lack of stage addition of 3418A in the final stages of laboratory flotation of Tank 1 product.

Table 3 Average concentrate grades and recoveries for flotation tests on Tank 1 product during the survey period Cumulative flotation time (min)

Pb

Zn

Cumulative conc. grade (%) 0.5 2.0 4.0 8.0 Tank 1 head

Cumulative conc. rec. (%)

Cumulative conc. rec. (%)

32.0 26.0 23.5 18.5

38.0 63.0 70.5 79.5

4.7 7.5 11.0 15.0

6.0

100.0

100.0

Table 4 Pulp chemical measurements of sample points during the survey period Sample Point

pH

Eh (mV) SHE

Temperature (C)

Dissolved ions (mg dm3 ) Ca

SAG discharge Ball mill discharge Cyclone overflow Tank 1 discharge Process water

Mg

Pb

Zn

EDTA extractable metal ions (%) Na

Cl

SO2 4

HCO 3

Pb

Zn

Fe

10.0

+50

35

1350

95

0.5

0.4

1300

1250

1200

55

0.1

0.0

0.0

7.5

–90

35

1300

165

0.5

0.5

1400

1250

1450

150

0.1

0.0

0.0

9.1

+110

32

1150

135

0.4

0.3

1300

1150

1000

150

4.4

0.2

0.5

8.8

+245

30

1250

145

1.7

1.4

1300

1250

1350

38

9.7

0.4

0.7

6.3

–

30

905

145

4.6

25.0

1340

1150

1200

38

–

–

–

818

M.C. Pietrobon et al. / Minerals Engineering 17 (2004) 811–824

pulp chemical conditions during both grinding itself and post-grinding during the oxygenation stage. The ore is not oxidised in the SAG mill discharge as indicated by the very low EDTA extractable metal ions at this sample point. The minerals then oxidise during oxygenation in Tank 1 as indicated by the increase in EDTA extractable metals for this sample point. Specifically, galena oxidises more rapidly than either sphalerite or iron sulphide minerals. It is usually accepted that galena oxidation is detrimental to its recovery in flotation. This was evidenced by the increase in EDTA extractable lead for the flotation tailings streams relative to the concentrate streams (not shown). However, it is considered likely that the high oxygen demand of the pulp prior to oxygenation in Tanks 1 and 2 plays a determining role. Here, collector adsorption on unoxidised galena may be inhibited by the low dissolved oxygen content of the pulps due its rapid removal by sulphide mineral and media oxidation. Fig. 5 shows the dissolved oxygen content of sampled pulps as a function of time (i.e., dissolved oxygen demand). This shows clearly that the dissolved oxygen demand of the fresh pulp after grinding (i.e., cyclone overflow) is much greater than that after oxygenation in Tanks 1 and 2. This diagram demonstrates the importance of measuring the dissolved oxygen demand, as well as the instantaneous dissolved oxygen level itself which is a balance between oxygen inputs and outputs. Another important variable is temperature, elevated values (>50 C) of which can adversely affect galena flotation through a number of different mechanisms. The results of surface analysis of the plant samples are discussed below in relation to its comparison with

20

Dissolved Oxygen (mg dm-3)

18

Cyclone Overflow

16

Tank 1 Product 14

Tank 2 Product 12 10 8

the surface analysis of the laboratory ground SAG feed sample. 3.2. Laboratory investigations 3.2.1. Head grade and particle size distribution Initial focus was on ensuring that the sub-samples of SAG feed were representative of the feed during the survey period, and that these samples could be ground using the specialised ball mill to a particle distribution that was very similar to the plant Tank 1 product stream. It was possible to achieve the appropriate grind particle size distribution by adjusting the mass of media charge in the mill and the time of grinding (Table 5). It was also shown that the sample was representative of the plant flotation feed at the time of the survey (Table 5), with similar head grades apparent for the plant survey sample and the average head grade of flotation tests on the SAG feed sample. 3.2.2. Laboratory grinding conditions Whilst it was relatively straightforward to achieve close agreement for the head grades and particle size distributions using standard procedures, achieving a similar agreement for the various pulp chemical parameters was more problematic. A particular issue was matching of the chemical conditions during grinding (i.e., pH, Eh and other dependent parameters in laboratory grinding). This was further complicated by the fact that the plant grinding circuit has two size reduction units in which two different chemical conditions exist (viz, SAG mill and ball mill). It was decided to match the pH in the laboratory experiments to the pH of the plant cyclone overflow (pH 9.1), as a compromise position between the SAG and ball mill discharge pH values (10.0 and 7.5, respectively). It is recognised that the grind pH controls to a great extent the oxidising/ reducing conditions during grinding, which in turn controls recovery and selectivity in flotation. It is in the ball mill where most size reduction occurs and the increase in mineral surface area the greatest, so most emphasis should be placed on matching chemical conditions to the plant ball mill. Table 5 Head assay and particle size distributions of plant Tank 1 product and laboratory ground sample of SAG feed

6 4

Parameter

Tank 1 product from the plant survey

Laboratory ground sample of SAG feed

Head grade (%)

Pb Zn Fe

5.2 8.9 36.0

5.7 9.1 34.7

Particle size (lm)

d90 d50 d10

46.2 14.6 2.2

43.3 12.4 1.5

2 0 0

1

2

3

4

5

6

7

8

9

10

Time (Minutes) Fig. 5. Dissolved oxygen concentration as a function of time (dissolved oxygen demand) for the plant cyclone overflow, Tank 1 product and Tank 2 product streams after the discontinuation of oxygen gas purging.

M.C. Pietrobon et al. / Minerals Engineering 17 (2004) 811–824 20

pH Control ; Lime 1500g/t ; Eh -30mV

10

18

Dissolved Oxygen (mg dm-3 )

Pulp pH During Grinding

10.2

819

No pH Control ; Lime 1200g/t ; Eh -60mV

9.8 9.6 9.4 9.2 9 8.8 8.6 0

5

10

15

20

Grinding Time (Minutes)

16 14

0 min. Oxygenation Time 20 min. Oxygenation Time 60 min. Oxygenation Time

12 10 8 6 4 2 0 0

Fig. 6. Pulp pH as a function of grinding time in the specialised grinding mill for batch addition of lime (no pH control) and continuous lime addition (pH control). The target grind pH in both cases was 9.1 at the completion of grinding.

In the current experimental work, use was made of the specialised grinding mill (Fig. 3) to control the pH of the laboratory grinding mill pulp to a value of 9.1 (Fig. 6). Two separate experiments were conducted. In one case, the grind pH was controlled by use of a single lime addition to the mill feed in a batch wise manner to a value of 9.1 at the completion of grinding. The lime addition necessary to achieve this grind pH, determined in preliminary experiments, was 1200 g/t. It was found that in the initial stages of grinding that the pH was 10.2, a value far away from the plant ball mill feed value (pH 8.8). In contrast, continuous control of the grinding pH using the specialised mill allowed the grinding pH to be closely controlled to 9.1. It is noted that in both modes of lime addition, the mill discharge pH was 9.1. However, in the case of pH control during grinding, the lime addition required to achieve pH 9.1 was increased to 1500 g/t, while the Eh value in the mill discharge was also higher (Fig. 6). The plant ball mill Eh was lower ()90 mV) and lime addition to the SAG mill feed was lower (1000 g/t). It is difficult with the design of the specialised grinding mill described here to independently control pH and Eh. Further work is proceeding with a new mill (Magotteaux Mill), which will allow both pH control during grinding and also gas sparging into the mill. With this mill a closer match of both the pH and Eh during grinding should be possible. 3.2.3. Dissolved oxygen demand Experiments were conducted in which the SAG feed sample was ground in the specialised laboratory grinding mill (controlled pH 9.1), and the slurry purged with oxygen gas in a laboratory flotation cell for different periods of time and at a fixed gas flow rate. This stage was to simulate the oxygenation which took place in Tank 1 (and Tank 2) of the plant. With increasing oxygenation the laboratory ground pulps exhibit re-

1

2

3

4

5

6

7

8

9

10

Time (Minutes) Fig. 7. Dissolved oxygen concentration as a function of time (dissolved oxygen demand) for the SAG mill feed sample ground in the specialised laboratory mill (grind pH 9.1) using circuit water after the discontinuation of oxygen gas purging. The oxygenation times were 0, 20 and 60 min prior to the dissolved oxygen demand test.

duced oxygen demand (Fig. 7). In the absence of oxygenation, the mill discharge shows very high dissolved oxygen demand which was very similar to the dissolved oxygen demand of plant cyclone overflow (i.e., prior to Tank 1) (Fig. 5). With 20 min oxygenation (Fig. 7), the dissolved oxygen demand of the laboratory ground sample approached that of Tank 1 product (Fig. 5). It is clear that 20 min of oxygen purging, under these conditions, closely matched the dissolved oxygen demand of Tank 1 product. This oxygen purging time was subsequently used in flotation experiments on the SAG feed sample which had the objective of matching the flotation performance of the laboratory flotation test on Tank 1 product. With 60 min oxygenation (Fig. 7), it is clear that the minerals are very heavily oxidised and do not consume oxygen significantly. Reasons for the beneficial effect of moderate oxygenation on galena recovery from Elura ore is the subject of another study. 3.2.4. Dissolved and precipitated species Tests conducted at 20 min oxygenation time and with demineralised water were analysed to determine the dissolved species at pH 9.1 while using sodium hydroxide as the pH regulator (Table 6, Stream 2). Sodium hydroxide was used to control the pH as lime could not be used to exclude calcium, while sodium was more abundant in the process water than potassium. Thus, addition of sodium hydroxide did not disturb the solution composition markedly. This is in effect the solution composition of the pulp without using process water and may be compared to the actual solution composition of Tank A discharge solution (Table 6, Stream 6). There is a relatively minor concentration of ions in solution suggesting only modest dissolution of

820

M.C. Pietrobon et al. / Minerals Engineering 17 (2004) 811–824

Table 6 Dissolved and precipitated species in the pulp during the plant survey (Tank A solution), process water, after laboratory grinding in demineralised water (ore dissolution) and in circuit water (simulated pulp) pH Stream 1. Process water 2. Ore dissolution 3. Reagents (lime) 4. Sum (1 + 2 + 3) 5. MinteqA2 in solution

6.8 9.1 – 4.9

Precipitates 6. Tank A solution 7. MinteqA2 for Tank A

Ca

Mg

Pb

Zn

Na

Cl

905 35 710 1650 1412

145 8 – 153 153

4.6 0.2 – 4.8 4.8

25.0 0.1 – 25.1 25.1

1340 110 – 1450 1450

1150 25 – 1175 1175

145 145

1.7 1.7

1.4 1.4

1300 1300

8 120 33

3.1 0.4 4.4

23.7 0.1 25

150 1400 50

SO2 4 1200 115

HCO 3

1315 745

38 36 – 74 74

1250 1250

1350 1130

38 38

)75 2300 1125

)605 2100 )1355

36 18 56

–

Gypsum 8.4 4.5

Precipitates 8. (5–6) ¼ 9. Simulated pulp 10. (5–9) ¼

Dissolved ions (mg dm3 )

1250 1045 Gypsum

9.1

162 1000 412

the ore components under these conditions. It is clear that process water and lime addition are required to adequately simulate the plant solution composition. This is further highlighted by mathematical addition of the process water (Table 6, Stream 1), ore dissolution (Table 6, Stream 2), and lime (Table 6, Stream 3) inputs to the grinding stage to produce a simulated solution composition prior to equilibration (Table 6, Stream 4). The addition is based on the relative flows of each input, assuming also that the water in the combined product is derived only from the process water input. Clearly, most of the dissolved magnesium, sodium, chloride and sulphate in this simulated grinding circuit product are derived from the circuit water. In the case of calcium, significant quantities emanate from lime addition to the grinding circuit, as well as the circuit water. Chemical equilibration of this simulated solution shows that gypsum will precipitate, and that the pH will decrease, with time (Table 6, Stream 5). Comparison of this equilibrated, simulated solution to the actual composition of Tank A discharge (in Table 6, Stream 8) shows that significant (on a relative basis) additional calcium, lead and zinc have been removed during grinding beyond that predicted by simple solution equilibration. In the case of lead and zinc, these will have precipitated, and possibly adsorbed, at the higher pH value of the former plant stream. In the case of calcium, adsorption onto minerals seems likely. In the case of sulphate, Tank A discharge shows a higher concentration than that expected after solution equilibration, being similar to that obtained from the simulated solution without equilibration. This suggests that while gypsum precipitation is expected under solution equilibrium conditions, equilibrium has not been obtained in the grinding circuit pulps. Furthermore, calcium is removed from solution,

most likely via adsorption, without complete precipitation of gypsum. Chemical equilibration of Tank A discharge solution also shows that gypsum may precipitate, and that the pH will decrease, with time (Table 6, Stream 7). Attention will now turn to the actual composition of the SAG feed sample after laboratory grinding at pH 9.1 in circuit water, with lime addition for pH control, and with 20 min of post-grinding oxygenation (Table 6, Stream 9). As for Tank A discharge, the SAG feed sample under these conditions shows less calcium, lead and zinc in solution relative to that predicted on the basis of the mass balance of inputs and equilibration (Table 6, Stream 10). It is likely that in the case of this particular laboratory condition, calcium, lead and zinc are being removed from the process water input due to precipitation and adsorption. In the case of sulphate, the SAG feed sample under these conditions shows a higher concentration, again suggesting that calcium is removed from solution predominantly not as precipitated gypsum, but most likely via adsorption onto minerals. Finally, comparison of Tank A discharge (Table 6, Stream 6) with the laboratory value under similar conditions (Table 6, Stream 9) shows reasonable agreement except for chloride and sulphate. The flotation results, described further below, which show a close correlation between the plant and laboratory demonstrate that this difference may not be significant. 3.2.5. EDTA extractable metal In the same experiments as those measuring dissolved oxygen demand, in which the SAG feed sample was ground in the specialised laboratory grinding mill (controlled pH 9.1) and the slurry purged with oxygen gas in a laboratory flotation cell for different periods of

M.C. Pietrobon et al. / Minerals Engineering 17 (2004) 811–824

30

EDTA Extractable Lead (%)

Demineralised H2O + NaOH Demineralised H2O + Lime

25

Process H 2O + NaOH Process H 2O + Lime

20

15

10

5

0 0

10

20

30

40

50

60

Time (Minutes) Fig. 8. EDTA extractable lead as a function of oxygenation time for the SAG mill feed sample ground in the specialised laboratory mill (grind pH 9.1) using either circuit or demineralised water, with either lime or sodium hydroxide as the pH regulator.

20

Dissolved Oxygen (mg dm-3)

time and at a fixed gas flow rate, the amount of EDTA extractable metal was also measured (Fig. 8). The extent of galena oxidation is a function of oxygenation time, pH regulator and process water type. Both process water and the use of lime as the pH regulator decreased galena oxidation after grinding relative to that obtained with demineralised water and sodium hydroxide. It seems probable that calcium in lime inhibits galena oxidation in an indirect mechanism which may involve other sulphide minerals, notably pyrite. To further investigate this possibility the dissolved oxygen demand of these specific tests, after 20 min oxygenation, are shown in Fig. 9. For the condition which gave the highest galena oxidation (demineralised water and sodium hydroxide), the dissolved oxygen demand was also the highest. Conversely, for the condition which gave the lowest galena oxidation (process water and lime), the dissolved oxygen demand was also the lowest. The similar effects attributed to lime and process water suggests that calcium has a specific effect in decreasing galena oxidation and oxidation of sulphide minerals in general. It is not the purpose here to explore this mechanism. In any case, for the laboratory condition which most closely approximates the plant condition (i.e., process water and lime, 20 min of oxygen purging) similar EDTA extractable lead as the plant Tank 1 discharge is apparent (Table 7). For the other laboratory conditions investigated, galena oxidation was apparently much higher during the post-grinding oxygenation period. In the case of EDTA extractable zinc and iron, similar levels were also found in the laboratory and plant (Table 7). It is worthwhile noting that in the case of EDTA extractable metals, that this parameter is a variable dependent upon grinding conditions, pH and pH regu-

821

18

Demineralised H2O + NaOH

16

Demineralised H2O + Lime Process H2O + NaOH

14

Process H2O + Lime

12 10 8 6 4 2 0 0

1

2

3

4

5

6

7

8

9

10

Time (Minutes) Fig. 9. Dissolved oxygen concentration as a function of time (dissolved oxygen demand) for the SAG mill feed sample ground in the specialised laboratory mill (grind pH 9.1) using either circuit or demineralised water, with either lime or sodium hydroxide as the pH regulator, after the discontinuation of oxygen gas purging. The oxygenation time was 20 min prior to the dissolved oxygen demand test in all cases. Table 7 EDTA extractable metals for plant and laboratory samples EDTA extractable metal ions (%) Plant Tank 1 discharge Laboratory SAG feed (pH 9.1, process water, lime pH regulator, 20 min O2 purging)

Pb

Zn

Fe

9.7 9.5

0.40 0.32

0.78 0.70

lator type, water type and oxygenation after grinding, amongst other factors. 3.2.6. Pulp temperature The temperature of the plant flotation pulps during the survey period were measured to be between 32–35 C. It was noted that in laboratory tests on plant pulps discussed above, that the pulp temperature may decrease by as much as 10 C due to addition of dilution water. Elevated pulp temperatures can affect galena recovery and selectivity (Lin, 1989; O’Connor and Mills, 1990) and therefore the effect of temperature was further investigated. Flotation tests with process water were performed at temperatures of 30, 45 and 60 C. It was observed that lead recovery was unaffected by pulp temperatures to 45 C, and therefore it was concluded that temperature did not influence galena recovery in the current work. It is worth noting that at 60 C the lead flotation rate was severely retarded. In terms of matching the pulp chemistry from plant to laboratory, the pulp temperature is an important parameter to consider and requires further investigation.

M.C. Pietrobon et al. / Minerals Engineering 17 (2004) 811–824

3.2.8. Flotation behaviour After having established a close correlation in pulp and surface chemistry between the plant and laboratory at key points in the circuit (viz, in grinding and in postgrinding conditioning), laboratory flotation tests were carried under conditions which replicated the pulp chemistry in the plant. These conditions were: (a) grind pH 9.1 continuously controlled with lime as the pH regulator, (b) process water, (c) 20 min O2 purging after grinding also controlled at pH 9.1. Furthermore, 60 g/t of 3418A was used in the laboratory experiments with the same physical conditions in the laboratory cell (i.e., Gfr , db , impellor RPM) as used in the flotation tests on Tank 1 discharge during the survey period. A very close

correlation in the lead grade/recovery relationship, on a unsized basis, was found for this laboratory test and that on Tank 1 discharge (Fig. 10a). There was less correlation in the lead recovery/zinc recovery selectivity relationship (Fig. 10b), for flotation times greater than 2 min, for reasons which are presently unclear. Further experimental work would be required to determine the sensitivity of zinc recovery to the various parameters outlined here as a step towards understanding the reasons for this discrepancy. Key parameters which would control selectivity against zinc would be, amongst others, (1) grind pH, (2) grind Eh, (3) grind media type, (4) post-grinding oxygenation, and (5) collector concentration. It is likely that more work would be required to more closely match the grind pH and Eh and to determine the sensitivity of zinc recovery to these parameters. The new Magotteaux Mill, allowing independent control of pH and Eh during grinding, will be used for this purpose. Cumulative Concentrate Pb Grade ( %)

3.2.7. Surface analysis using X-ray photoelectron spectroscopy Samples of Tank A discharge from the plant and laboratory SAG feed under conditions which replicates the plant in most of the common pulp chemical parameters were examined by XPS. This was carried out as a final check that close chemical matching had indeed been obtained. It is recognised that surface atomic concentrations determined by XPS (Table 8), will in themselves be dependent upon grinding conditions, pH and pH regulator type, water type and oxygenation after grinding, amongst other factors. The results showed that there was very similar abundance of surface atoms for these two samples, both with and without etching (Table 8). A feature was that the sulphur 2p regions in both samples indicated that the major sulphur species present was sulphide and that sulphate was almost undetectable. The ratio of sulphide:sulphate was
10:1. This supports the argument that calcium sulphate had not precipitated, either in solution or on mineral surfaces during grinding in the plant or laboratory. The extent of oxidation, as measured by the atomic concentration of oxygen, was very similar between the two samples. The surface analysis results indicated that, provided sufficient care was taken to prepare the samples in the laboratory under conditions which replicated the plant, a very close match of the pulps could be produced.

Zn Recovery (%)

822

40 35

(a)

30 25 20 15

Laboratory SAG Feed Plant Tank 1 Discharge

10 5

Pb Recovery (%)

0 20

(b)

15 10 5 0 0

Pb Recovery (%) 10

20

30

40

50

60

70

80

90 100

Fig. 10. Lead grade/lead recovery relationship (a) and lead recovery/ zinc recovery relationship (selectivity) (b) for flotation tests (three tests in total; average values shown) on the plant Tank A product stream and the SAG mill feed sample ground in the specialised laboratory mill (grind pH 9.1) using circuit water, with lime as the pH regulator and 20 min oxygenation time. In both cases, 60 g/t of 3418A was added.

Table 8 Atomic concentrations for major elements determined by XPS, for the initial surface and after an ion beam etch, for the plant Tank 1 discharge and a laboratory test on SAG feed Elements (%)

Ratio of peak areas

C

O

Fe

Pb

Zn

S

Ca

Mg

Al

Si

þ

HC/ CO2 3

S2 /SO2 4

Tank A discharge

Initial Etched

20.0 5.7

39.3 28.0

6.1 19.2

2.7 1.9

1.4 3.3

9.8 16.8

1.9 2.1

12.1 16.8

1.4 0.7

4.9 5.3

3:1 3:1

10:1 >10:1

Laboratory SAG feeda

Initial Etched

23.8 6.4

38.0 29.2

5.6 20.8

2.5 1.6

1.1 3.0

10.6 18.1

1.6 1.8

10.9 13.2

1.3 0.6

4.2 4.9

3:1 3:1

10:1 >10:1

a

Conditions of the Laboratory SAG feed are the same as those shown in Table 7. þ HC is carbon in hydrocarbon.

M.C. Pietrobon et al. / Minerals Engineering 17 (2004) 811–824

4. Discussion With careful control of the key parameters in both grinding and post-grinding conditioning it is possible to closely match the pulp and surface chemistry of laboratory pulps to the plant pulps. A list of the key parameters and how to manipulate them at laboratory scale is shown in Table 9. In the current study, most difficulty was found with matching the laboratory grinding chemistry to the plant grinding chemistry. This was because the design of the specialised mill did not allow independent manipulation of the grind pH and gas sparging during grinding. It is recognised that most plant mills are open to the atmosphere with continuous ingress of oxygen into the mills exacerbating oxidation of sulphide minerals and particularly grinding media. With the new Magotteaux Mill a closer match to the plant should be possible and the sensitivity of the process to grind pH and Eh more closely determined. It is important to consider conditions during both grinding and post-grinding conditioning. While the latter was a specific focus of the current investigation due to the oxygenation stages in the plant, the former is equally important. For example, in the evaluation of different grinding media types at laboratory scale it is important to replicate the plant grinding conditions to increase the confidence in which the effect of the media type may be translated to plant performance. A key parameter in the current study, was the extent of oxygenation after grinding. This affected the oxidation state of the minerals as reflected in the dissolved oxygen demand of the pulp and the amount of EDTA extractable metals. While the exact role of oxygenation

823

is still presently unclear, the surface passivation (oxidation) of iron sulphide minerals would appear to be an important factor. However, it is not clear whether oxygenation is required to simply raise the instantaneous dissolved oxygen level at the point of collector addition to allow its adsorption onto galena, or that extensive, prior oxidation of the iron sulphide minerals is required. It is possible that only a modest dissolved oxygen addition is required to raise the dissolved oxygen level at the point of collector addition only, a possibility which would have implications for oxygen consumption and its control in the plant. The extent of oxygenation required will also be dependent on the collector type. Other key parameters were the use of circuit water and lime as the pH regulator. Not only were their use essential to match the dissolved species, these parameters also significantly affected sulphide mineral oxidation as evidenced by changes in the dissolved oxygen demand and EDTA extractable lead with changes in water and pH regulator type. A possible mechanism to explain this was adsorption of calcium onto sulphide minerals reducing their oxidation rate and oxygen consumption. Evidence for calcium adsorption during grinding in both plant and laboratory was found by comparison of the mill input and output streams. Further, calcium did not apparently precipitate as gypsum during grinding suggesting that the noted decrease in calcium concentration during grinding was due to its adsorption. In general, flotation in demineralised water resulted in worse selectivity of lead against zinc and iron compared to that achieved with circuit water. Specifically, flotation in demineralised water increased the recovery of zinc and iron compared to circuit water. When calcium

Table 9 Key parameters for matching Mill inputs

Parameter

Controlled by:

Ore

The same ore sample used and correct sampling procedures to ensure head grades are identical The same media sample used in laboratory and plant The same addition used in laboratory and plant. Preferably the same reagent sample and purity The same water quality used, either a simulated water or sampled water

Grinding media type Reagent addition Process water Mill outputs

Post-grinding conditioning

Particle size pH Eh Temperature Dissolved oxygen demand Dissolved species EDTA extractable metal ions Particle surfaces

Grinding media and grind time pH regulator to control the pH continuously during grinding Grinding media, gas purging to the mill, also dependent on grind pH Water source/grinding media/type of mill Grinding media/gas purging to the mill, also dependent on grind pH Process water chemical species/grinding media, also dependent on grind pH and pH regulator Dependent on Eh/pH/dissolved oxygen/process water/grinding media Dependent on Eh/pH/dissolved oxygen/process water/grinding media

Conditioning stages pH Eh Dissolved oxygen demand Temperature

Conditioning time/gas flow rate pH regulator to control the pH continuously Aeration/Oxygenation after grinding, also dependent on grind and conditioning pH Aeration/oxygenation after grinding, also dependent on grind and conditioning pH Heating/% solids dilution/conditioning mechanism

824

M.C. Pietrobon et al. / Minerals Engineering 17 (2004) 811–824

was added to the demineralised water (in the form of Ca(NO3 )2 ), to achieve the same concentration which occurs in the plant pulps (
1000 ppm), selectivity against zinc and iron were partially restored at the same pH (9.1). It appears likely that calcium has a specific effect in depressing sphalerite and pyrite, changing their oxidation rate and oxygen demand behaviour.

5. Conclusions Key parameters in the successful duplication of plant slurry flotation behaviour by the use of SAG mill feed samples in laboratory scale experiments were identified. A key finding was that matching of the laboratory pulp chemistry to the plant pulp chemistry was dependent upon oxidation after grinding. In turn, mineral oxidation was dependent upon the pH in grinding, water composition and the duration of oxygen purging after grinding. The methodology closely matched the dissolved oxygen demand and surface chemistry of laboratory and plant pulps.

Acknowledgements The authors gratefully acknowledge the financial support of AMIRA International and Pasminco Australia Limited. This paper was presented at the 34th Annual Meeting of the Canadian Mineral Processors 2002 from whom permission has been obtained to publish in Minerals Engineering.

References Allen, P., Jackman, R., Peck, R., 1993. Lead–zinc concentrator practice by Pasminco Limited, Elura mine, Cobar, NSW. In: The Sir Maurice Mawby Memorial Volume, second ed. Australasian Institute of Mining and Metallurgy, Parkville, Vict, pp. 516–518.

Cases, J.M., Kongolo, M., Donato, P., Michot, L., Erre, R., 1990. Interaction between finely ground galena and pyrite with potassium amyl xanthate in flotation. I. Influence of alkaline grinding. Int. J. Min. Process. 28, 313–337. Cullinan, V., Grano, S.R., Greet, C., Johnson, N.W., Ralston, J., 1999. Investigating fine galena recovery problems in the lead circuit of Mount Isa Mines lead/zinc concentrator. Part I: grinding media effects. Min. Eng. 12, 655–669. Derjaguin, B.V., Dukhin, S.S., 1960–1961. Theory of flotation of small and medium sized particles. Trans. Inst. Min. Metall. 70, 221–246. Gorain, B.K., Burgess, F., Franzidis, J.P., Manlapig, E.V., 1997. Bubble surface flux––a new criterion for flotation scale-up. In: Sixth Mill Operators Conference, Madang, Papua New Guinea. Australasian Institute of Mining and Metallurgy Publisher, Parkville, pp. 141–148. Grano, S.R., Lauder, D.W., Johnson, N.W., Smart, R.St.C., Ralston, J., 1991. Comparison of ethyl xanthate and diisobutyldithiophosphinate collectors for the lead roughing of the Hilton ore of Mount Isa Mines Ltd.. In: Fifth Extractive Metallurgy Conference. Australasian Institute of Mining and Metallurgy Publisher, Parkville, pp. 203–210. Grano, S.R., Ralston, J., Fornasiero, D., 2001. P260D Mineral Flotation. Project proposal to AMIRA International. Lin, I.J., 1989. The effect of seasonal variation in temperature on the performance of mineral processing plants. Min. Eng. 2 (1), 47–54. O’Connor, C.T., Mills, P.J., 1990. The effect of temperature on the pulp and froth phases in the flotation of pyrite. Min. Eng. 3 (6), 615–624. Ralston, J., 1994. A unified approach to flotation. In: Fifth Mill Operators Conference, Roxby Downs. Australasian Institute of Mining and Metallurgy Publisher, Parkville, pp. 15–27. Rumball, J.A., Richmond, G.D., 1996. Measurement of oxidation in a base metal flotation circuit by selective leaching with EDTA. Int. J. Min. Process. 48, 1–20. Runge, K.C., Harris, M.C., Frew, J.A., Manlapig, E.V., 1997. Floatability of streams around the Cominco Red Dog lead cleaning circuit. In: Six Mill Operators Conference, Madang, Papua New Guinea. Australasian Institute of Mining and Metallurgy Publisher, Parkville, pp. 157–164. Savassi, O.N., Alexander, D.J., Johnson, N.W., Franzidis, J.P., Manlapig, E.V., 1997. Measurement of froth recovery of attached particles in industrial flotation cells. In: Six Mill Operators Conference, Madang, Papua New Guinea. Australasian Institute of Mining and Metallurgy Publisher, Parkville, pp. 149–155. Spira, P., Rosenblum, F., 1974. The oxygen demand of flotation pulps. In: 6th Annual Canadian Mineral Processors Operators Conference, pp. 73–106.