metallurgical testwork on kevitsa ore in pilot scale

Eastern Finland Office 21 June 2007 C/MT/2007/19 Outokumpu

Metallurgical Testwork on Kevitsa Ore in Pilot Scale

Research Report C/MT/2007/19

TABLE OF CONTENTS

Documentation page 1 SUMMARY 2 GENERAL 3 ORE SAMPLE 4 CRUSHING AND GRINDING 4.1 Crushing 4.2 Grinding 5 FLOTATION CIRCUIT 5.1 Pilot Phases 1 and 2 (Autumn 2006) 5.2 Pilot Phase 3 (Jan-Feb 2007) 5.3 Equipment 5.4 Flotation Chemicals 6 PILOT PLANT OPERATION 6.1 Operating Parameters 6.2 Process Water 6.3 Chemical Analyses 7 RESULTS 7.1 Selective Flotation 7.1.1 Copper Flotation 7.1.2 Nickel Flotation 7.2 Bulk Flotation 7.3 Floatable and Nonfloatable Nickel 7.4 Material Balances 7.5 Recovery Correlations 7.6 Flotation Times 7.7 Concentrates Produced 7.8 Samples Outside 7.9 Ion Concentrations in Tailings and Concentrates 8 ESTIMATE FOR AVERAGE ORE 9 DEWATERING TESTS ON CONCENTRATES 10 CONCLUSIONS

Research Report C/MT/2007/19

GEOLOGICAL SURVEY OF FINLAND

DOCUMENTATION PAGE

Date: 21 June 2007Type of report

Research Report Authors

Reijo Kalapudas Markku Klemetti Tapio Knuutinen

Commissioned by

Krister Söderholm, Peter Walker and John Pedersen / Kevitsa Mining Oy / Scandinavian Minerals Ltd.

Title of report

Metallurgical Testwork on Kevitsa Ore in Pilot Scale

Abstract Metallurgical test work was carried out for 575 tons of Kevitsa main ore at the pilot plant of Geological Survey of Finland. Both selective flotation for producing separate copper and nickel concentrates, and bulk flotation were tested. In selective flotation a high grade copper concentrate with about 30 % Cu was produced with copper recovery of 77-80 %. About 25-40 % of gold also reported to copper concentrate. The nickel content in this product could be decreased to level 0,5-0,6 %. The nickel concentrate produced by selective flow sheet contained 6-18 % Ni depending on the process conditions and chemicals used. The nickel recovery changed from 69 to 64 % when concentrate grade increased from 7,2 to 16,7 % Ni. In phase 3 of pilot campaign the concentrate contained 12,2 % Ni as an average with a recovery of 65,3 %, and 44,3 % Fe and only 2,7 % MgO yielding a high Fe/MgO ratio; 16,3. These nickel concentrates with a good quality could be produced without the use of special chemical, TETA. The grades and recoveries of precious metals in nickel concentrate in phase 3 were: platinum 10,7 g/t, recovery 58 %; palladium 8,0 g/t, recovery 43 %; gold 1,56 g/t, recovery 15 %. The nickel recovery in selective flotation remained below the target, which was set to about 70 %. The main rea-son for this was the high content of nonfloatable nickel occurring in silicates and partly in goethite in the feed of pilot run. The level of nonfloatable nickel was 460-600 ppm whereas the average Kevitsa main ore contains 275 ppm only. It was estimated that for an average Kevitsa main ore, containing 0,30 % Ni of which 275 ppm in silicates, the recoveries with the grades of 10 -12 - 16 % Ni would be 79,5 – 78,9 – 77,7 % for sulphide Ni, and 72,0 – 71,5 – 70,3 % for total Ni. The sulphide nickel recoveries are those achieved in pilot plant and total nickel recoveries about 6 % higher due to lower silicate nickel in the average main ore. In bulk flotation at alkalic conditions high metal recoveries were reached: Cu 94,9 %; Ni 84,4 %; Pt 78 %; Pd 90 %; and Au 77 %. The bulk concentrate grades were then 6,06 % Cu and 3,65 % Ni. Keywords

Kevitsa, PGM ore, copper flotation, nickel flotation, bulk flotation Geographical area

Sodankylä, Finland Map sheet

Other information

Report serial

Research Report Archive code

C/MT/2007/19 Total pages

55+127 Language

English Price

Confidentiality

Confidential Unit and section

Geoservices, 651 Project code

Signature/name

Kauko Ingerttilä, Laboratory Manager

Signature/name

Reijo Kalapudas, Senior Metallurgist

Research Report C/MT/2007/19 2(53)

1 SUMMARY Metallurgical test work was carried out for 575 tons of Kevitsa main ore type at the pilot plant of Geological Survey of Finland. Both selective flotation for producing separate copper and nickel concentrates, and bulk flotation was tested. The feed of pilot plant contained about 0,52 % Cu and 0,40 % Ni in pilot phases 1 and 2 in the autumn 2006. The feed grades during pilot phase 3 in 2007 were 0,45 % Cu and 0,36 % Ni. In selective flotation a high grade copper concentrate with about 30 % Cu was produced with copper recovery 77-80 %. About 25 to 40 % of gold also reported to copper concentrate. The nickel content in this product could be decreased to level of 0,5-0,6 %. The nickel concentrate produced by selective flow sheet contained 6-18 % Ni depending on the process conditions and chemicals used. As average figures the nickel recovery changed from 69 to 64 % when concentrate grade increased from 7,2 to 16,7 % Ni. The best quality of nickel concentrate was achieved in phase 3 when the product contained 12,2 % Ni as an average with a recovery of 65,3 %, and in addition, 44,3 % Fe and only 2,7 % MgO yielding a high Fe/MgO ratio, 16,3. The grades and recoveries of precious metals in nickel concentrate in phase 3 were: platinum 10,7 g/t, recovery 58 %; palladium 8,0 g/t, recovery 43 %; and gold 1,56 g/t, recovery 15 %. In bulk flotation at alkalic conditions high metal recoveries were reached: Cu 94,9 %; Ni 84,4 %; Pt 78 %; Pd 90 %; and Au 77 %. The bulk concentrate grades were then 6,06 % Cu and 3,65 % Ni. It was seen in phase 3 of pilot campaign that nickel concentrates with grade up to 18 % Ni and a high Fe/MgO ratio could be obtained without the use of special chemical, TETA. At the same time the reagent consumptions could be decreased also otherwise as much less collectors and depressants were used than earlier and, in addition, sulphuric acid and copper sulphate were not used at all. The main challenge in selective flotation was the improvement of nickel recovery. It was, how-ever, pointed out by mineralogical studies and phase analysis of nickel that the feed of pilot plant contained significant amount of nickel in nonfloatable form. The analysed content of nickel in silicates was 460 to 605 ppm (12-17 % of total nickel) in the pilot feed whereas the average level for Kevitsa ore is about 275 ppm (9,2 % of total nickel). In addition to silicates, some nickel was also found in goethite indicating some degree of oxidation in the ore sample received to pilot tests. The average recovery of sulphide nickel in pilot phase 3 was 78,8 % when the grade of concen-trate was 12,2 % Ni. By using the sulphide nickel recoveries of pilot tests it was estimated that for an average Kevitsa main ore, containing 0,30 % Ni of which 275 ppm occur in silicates, the recoveries with the grades of 10 -12 - 16 % Ni would be 79,5 – 78,9 – 77,7 % for sulphide Ni, and 72,0 – 71,5 – 70,3 % for total Ni. The recoveries of total Ni are about 6 % higher than what were achieved in pilot phase 3 and are based on the estimation (supported by analyses) that in the average main ore the share of sulphide nickel is clearly higher and the amount of silicate nickel lower than in the feed of pilot plant.


2 GENERAL About 575 tons of Kevitsa main ore was run in the pilot plant of GTK in three phases between September 7, 2006 and February 16, 2007. For the test work the ore was crushed below 6 mm and homogenised in a stockpile. The feed rate to grinding/flotation circuit was either 1 t/h or 0,85 t/h. The pilot plant was run totally 50 days, most of time by selective flotation to produce separate copper and nickel concentrates. Bulk flotation was run four days from Nov 28 to Dec 1, 2006. Two concentrates, copper and nickel, were produced by direct selective flotation flow sheet. The target was first to float most of copper to its own concentrate so that Ni-content in this product would remain below 0,8 % and preferably below 0,4 %. The original target for nickel concentrate was: grade about 12 % Ni with the recovery of nickel close to 70 % or higher if achievable. During different test periods the grade of concentrate varied from 6 to 18 % Ni. The content of Fe and MgO in nickel concentrate was also paid attention to, especially in the last phase of pilot run. It was important to have adequate amount of Fe and low enough MgO to achieve ratio Fe/MgO more than 5. Of the precious metals, gold should report predominantly to copper concentrate, and Pt and Pd to nickel concentrate in order to be payable. In bulk flotation all valuables were floated to bulk concentrate when aiming at high recoveries. Concentrates were thickened and filtered and stored in plastic drums. Samples of concentrates were also shipped for downstream metallurgical tests. Water in the pilot run was pumped from the tailings pond of GTK which contains ions typical for sulphide tailings. During two days of pilot tests (12-13 October) water of the Kitinen river close to the Kevitsa site was used as process water. The daily operation of pilot plant was normally in two shifts whereby the operation time (when the feed was on) was 10-12 hrs. Longer periods of 20 to 36 hrs in three shifts were also run. The average nickel grade of feed ore was 0,40 % in pilot phases 1 and 2, and 0,36 % in phase 3 which are higher than the target grade of average main ore; 0,30 % Ni. The assays of ore feed kept very stable during the pilot phases pointing out that the ore sample was homogeneous. 3 ORE SAMPLE An ore sample in excess of 600 tons was blasted from the Kevitsa main ore in May 2006. The first sample, 333 tons, of this ore was transported from Kevitsa to GTK, Outokumpu, in June 2006. The second shipment, 121 tons, took place in October and the third one, also 121 tons, in December. So, totally 575 tons were transported to Outokumpu.


300 tons of ore sample 1 was run in pilot phase 1 (Sept 7 – Oct 13), the rest of it (33 tons) and ore sample 2 in phase 2 (Nov 13 – Dec 1), and ore sample 3 in phase 3 (Jan 24 – Feb 16, 2007). The chemical compositions of pilot feed samples are given in Table 1. The assays represent average analyses of feed samples which were collected several times during each day. The day by day assays are presented in Appendix 1. Table 1. Analyses of pilot plant feed Cu Ni Fe S Pt Pd Au MgO % % % % g/t g/t g/t % Ore sample 1 (333 t) 0,512 0,398 5,33 2,14 0,390 0,401 0,205 23,9 Ore sample 2 (121 t) 0,547 0,402 5,70 2,24 0,356 0,335 0,163 24,9 Ore sample 3 (121 t) 0,448 0,360 5,39 1,85 0,329 0,336 0,186 24,9

It is seen that the nickel content was quite close to 0,400 % in ore samples 1 and 2. The grade kept continuously steady during the whole phases 1 and 2. Both copper and nickel were lower in pilot phase 3; 0,448 % Cu and 0,360 % Ni. The mineralogical compositions of feed samples from three days of pilot run are presented in Table 2. The total amount of sulphides in these samples is 5,9-6,6 wt-%. The main minerals of them are pyrrhotite, chalcopyrite and pentlandite. According to numbers in the table the con-tent of pyrrhotite increased during the last phase of pilot run (samples from 1st and 15th of Feb-ruary 2007) compared to sample from the first phase (12th of October 2006). The content of pentlandite was lower in pilot phase 3 (2007) than in phases 1 and 2 (2006). (The increase of amount of chalcopyrite was not supported by chemical analyses). There was also 0,9-1,1 % goethite in the ore and according to numbers in Table 2 the content increased slightly at the end of pilot tests. The occurrence of goethite indicates that the feed sample of pilot test, originating close to the surface, contained a small amount of oxidized ore. Mineralogical study of flotation tailings in phase 3 showed that part of nickel occurred in goe-thite (in which Ni had been found to occur also in separately picked oxidized ore samples). Higher than expected amount of nickel was analysed to occur in nonfloatable form in silicates and goethite in the ore samples received. The level of nonfloatable nickel was 460 ppm (12 % of total Ni) in samples 1 and 2, and 605 ppm (17 % of total) in sample 3. According to mineralogical analysis the PGM´s in the feed of pilot tests were found to occur in the following minerals (PGM mineral distribution): Tellurides (moncheite, merenskyite, melonite) 95,8 % Arsenides (sperrylite) 3,6 % Bismuthides (michenerite) 0,6 %


Table 2. The mineral composition of pilot plant feed (samples of cyclone overflow)

Oct 12, 2006 Feb 1, 2007 Feb 15, 2007wt-% wt-% wt-%

Diopside 37,17 33,84 34,78Tremolite 11,74 16,31 15,18Hornblende 4,49 5,67 5,57Olivine 14,87 13,68 13,71Plagioclase 2,82 1,34 1,35Serpentine 7,79 6,62 6,00Biotite 1,53 1,72 1,92Talc 5,72 7,25 7,70Chlorite 2,99 3,93 3,63Quartz 0,57 0,07 0,11Carbonate 0,49 0,38 0,33Ilmenite 0,08 0,03 0,06Magnetite 1,79 1,71 1,55Apatite 0,03 0,02 0,06Pyrite 0,12 0,12 0,06Pyrrhotite 2,81 3,16 3,74Sphalerite 0,00 0,00 0,00Chalcopyrite 1,75 1,88 1,87Pentlandite 1,20 0,95 0,90Goethite 0,92 0,92 1,11Process metal 0,40 0,10 0,10Other 0,07 0,02 0,02Unclassified 0,67 0,29 0,26TOTAL 100,02 100,00 100,00

4 CRUSHING AND GRINDING 4.1 Crushing Prior to grinding, the ROM ore sample was crushed down to -6 mm particle size and homoge-nized. The over sized rocks (>500 mm) were broken with hydraulic hammer on a steel plate. Crushing was done in three stages. The crushing circuit was fed with a vibrating feeder that was loaded with a frond end loader from the ROM ore stockpile. The primary crusher was a LOKOMO C63B jaw crusher, from which the crushed product was fed with a belt conveyor to the first screening stage, where the ore was screened into +16 mm and -16 mm fractions. A belt scale was installed in the first belt conveyor for measuring the amount of ore feed to the crushing circuit. From primary screening, the +16 mm fraction was fed to the secondary cone crusher, LOKOMO G128, and the secondary crusher product was combined to the screen under size. The combined product was fed to the secondary screening, where 6 mm screen cloth was used. The under size of the secondary screening was the final crushed product, and it was fed to the homogenization stockpile to be further conveyed to the pilot plant grinding circuit. The over size of the secondary screen continued to the tertiary crusher, LOKOMO G108, and the crusher product was returned to the secondary screening stage.


A schematic illustration of the crushing circuit is presented in Figure 1.

Double deckvibrating screen

Single deckvibrating screen #6 mm

Jaw crusher( primary )

Gyratory crusher( tertiary )

Feed

Gyratory crusher( secondary )

Homogenization Stockpile

Moving Scalper

Stockpile unloadingdevice

Vibrating Feeder

Figure 1. Kevitsa pilot plant crushing circuit. 4.2 Grinding Grinding circuit configuration In Kevitsa pilot plant run, a conventional rod mill - ball mill grinding circuit was used. The cir-cuit type was chosen for the ease of operation and for producing a stable particle size distribu-tion for the downstream flotation. Prior to grinding, the ore was crushed in three-stage crushing circuit down to -6 mm particle size and stored in the homogenization stockpile. The crushed ore was taken into the feed bin with an automatic discharger working according to feed bin weight. The feed bin was lying on the weight cells and it was equipped with an inverter-controlled belt feeder and an impact scale to monitor and control the feed rate to the grinding circuit. The schematic grinding circuit flow sheet is presented in Figure 2. The grinding circuit consisted of a rod mill in open circuit and a ball mill in closed circuit with a vibrating screen and a hydrocyclone. Ore and water were fed to the rod mill. The ore feed rate was most of the time 1 t/h during pilot phases 1 and 2 (in 2006), and 0.85 t/h in phase 3 (2007). The rod mill had an inner diameter of 0.84 m and an inner length of 1.75 m, which makes 1.13 BMU ( Base Mill Units ). The rod charge consisted of Ø 75 mm and Ø 50 mm rods with a power draw of 7.4 kW. The mill revolution speed was 32 rpm.


The rod mill product was pumped to a 1.5 mm vibrating screen. The main purpose of the screen was to prevent blockages in the hydrocyclone inlet. The screen underflow was pumped to a 75 mm AKW hydrocyclone, whose overflow reported through a flow meter and a particle size analyzer to flotation circuit. The screen overflow and the cyclone underflow gravitated to the ball mill. The ball mill had an inner diameter of 0.84 m, inner length of 1.35 m (0.87 BMU), and a steel ball charge (Ø 25-40 mm) giving 10.8 kW power draw.

Combined milldischarge pump

FeedBin

Homogenization stockpile 0 - 8 mm

Belt FeederTorsion scale

Rod Mill 3021000 x 1800 mm15 kW

Ball Mill 3011000 x 1500 mm15 kW

Automatic Reclaiming according to the Bin weight

Cyclone feed pump

Feed rate0.85-1 t/h

Vibratingscreen# 1.5 mm

Slurry flow

Water flow

To flotation

Flow meterParticle size analyzerPulp density measurement

Cyclone D 80 mmVortex 16-20 mmApex 15 mm

Pumpsump level control

Rod millwater

Ball millwater

Feed pressurecontrol

Figure 2. Kevitsa pilot plant grinding circuit Grinding circuit balances During the pilot plant operation, feed samples of the crushed ore were collected daily. Figure 3 shows the average particle size distribution of the plant feed during the run. The plant feed had the following particle size distribution: 82.7 % -5.6mm 18.1 % -0.500 mm 7.1 % -0.075 mm. The D80 of the plant feed was 5.35 mm. During the run, three different grinding finenesses were tested having the D80-values of 66 µm (fine), 77 µm (medium) and 84 µm (coarse). In the autumn 2006 the grinding fineness was 75-80 % -75 µm most of the time. During pilot phase 3 (in 2007) the fineness was 82-88 % -75 µm. Figure 4 shows the average particle size distributions of different grinding finenesses.


0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

10 100 1000 10000

Particle Size (µm)

Pass

ing

(%)

Figure 3. Kevitsa pilot run, plant feed

0.0

10.0

20.0

30.0

40.0

50.0

60.0

70.0

80.0

90.0

100.0

10 100 1000

Particle Size (µm)

Pas

sing

(%)

FineMediumCoarse

Figure 4. Kevitsa pilot run, flotation feed The detailed grinding balances and particle size distributions are presented in Appendices 2 and 3. Table 3 summarizes the main grinding parameters with different grinding products.


Table 3. The main grinding characteristics with different finenesses

Particle Sizes Energy Consumption

Grinding Feed Rod Mill Cyclone Overflow Work Rod Ball Total

d 80 Discharge % Index Mill Mill Mills

mm d 80 mm d 80 mm - 75 µm kWh/t kWh/t kWh/t kWh/t

Coarse 5,35 0,42 0,084 73,8 16,5 6,4 (42 %)

8,8 (58 %)

15,2

Medium 5,35 0,42 0,077 79,9 15,5 6,5 (42 %)

9,0 (58 %)

15,5

Fine 5,35 0,42 0,066 86,3 17,0 7,8 (42 %)

10,8 (58 %)

18,6

Pilot Grinding vs. Bench Scale (Mergan) The Mergan ball mill test on bench scale was done for the feed ore of pilot plant to determine the grinding energy and Work Index with three different finenesses. The feed was an ore sam-ple of 5 kg crushed to -1 mm. The grinding times were 60, 80 and 100 min to achieve different finenesses. The determination of energy consumption is based on grinding time, mass of sam-ple and direct measurement of mill power. The results obtained with Mergan mill are presented in Table 4 together with grinding charac-teristics of pilot plant. The energy consumptions measured with Mergan mill for the feed of -1 mm were also converted to energies for coarser feed, -6 mm, corresponding to the feed of pilot rod mill. The grinding energy vs. fineness is shown in Figure 5. It is seen that the Mergan energies con-verted to feed size -6 mm are very close to grinding energies in pilot plant. With the high grinding fineness used in pilot phase 3, 86 % -75 µm (d 80 = 66 µm), the energy consumption was 18,6 kWh/t. The same energy was obtained when Mergan result was converted to feed size -6 mm. With a fineness of 80 % -75 µm the grinding energy would be about 15,5 kWh/t. The size distributions of Mergan grinding and pilot plant cyclone overflow are presented in Appendix 3/1. Table 4. Grinding characteristics in Mergan test and pilot plant

d 80 Energy WI d 80 Energy d 80 Energy WIµm % -75µm kWh/t kWh/t µm kWh/t µm % -75µm kWh/t kWh/t

Feed 640,3 14,8 5350 5350

Product 84 73,8 15,2 16,5Product 80,9 76,5 10,56 14,7 80,9 14,4 77 79,9 15,5 15,5Product 68,3 85,3 14,05 17,3 68,3 18,5 66 86,3 18,6 17,0Product 58,4 91,7 17,51 19,2 58,4 22,5

Mergan test for feed ore -1 mmEnergy calculated

for pilot feed -6 mmon the basis of Mergan Measured in pilot plant


Kevitsa Grinding Energy

5

10

15

20

25

50 60 70 80 90d 80 (µm)

Ener

gy (k

Wh/

t)

Mergan

Mergan calculated for pilot feed size

Pilot plant

Figure 5. Grinding energy in pilot plant and in Mergan test Regrinding Sala SAM 7.5 tower mill was used for regrinding in nickel flotation circuit. The grinding me-dia consisted of 4-8 mm mild steel balls. During phases 1 and 2 (autumn 2006) of pilot tests the regrinding was only used in the begin-ning of phase 1 and in the end of phase 2. It was found that regrinding was not necessary when special chemicals (MBS and TETA) were used for depression of pyrrhotite. At the end of phase 2 the regrinding was taken into use when special chemicals were not used. The feed of regrinding consisted of first cleaner tailings and scavenger concentrate. In phase 3 (2007) of pilot run the regrinding was used regularly. The special chemicals were not used and it was seen that the depression of pyrrhotite succeeded better when regrinding was used. It was concluded that the meaning of regrinding was to create chemically reducing pulp conditions for pyrrhotite depression in the final nickel cleanings. The feed of regrinding mill was the second cleaner concentrate which after regrinding continued to cleaners 3-5. The size distributions of regrind mill feed and product on Jan 25, 2007 are shown in Table 5. The mass of NiCC2 was 45,1 kg/h and the mill power 4,5 kW which gives the energy con-sumption of 99,8 kWh/t per feed of regrinding and 5,3 kWh/t per primary feed (850 kg/h) in regrind mill. It should be noted that the fineness and energy input in regrinding were not optimized in the pilot run. The energy consumption and fineness were rather high as the presented numbers in-dicate.


Table 5. Size distributions of regrind mill feed and product

Screen Mill feed Mill product µm % passing % passing

90 99,6 75 98,9 100,0 32 75,3 91,7

5 FLOTATION CIRCUIT 5.1 Pilot Phases 1 and 2 (Autumn 2006) The flotation circuit changed during the test work from period to period, but the basic configu-ration was as follows: The flotation flow sheet consisted of two flotation circuits, one for copper and the other for nickel flotation, separated by a Lamella thickener between the circuits for water change and pulp density control. Copper flotation consisted of a rougher flotation followed by a four-step cleaning in closed cir-cuit where each cleaner tailing was returned to previous cleaner step. The first cleaner tailings were returned to rougher flotation. The final copper concentrate was collected to a thickener. The nickel flotation circuit did undergo a lot of modifications during the pilot plant run. The last circuit configuration of phase 2 had a three-stage rougher and two-stage scavenging. The scavenger tailings were fed to tailings flotation, where high and low sulphur tailings were split. The scavenger concentrate was combined with the first cleaner tailings and returned to rougher flotation through a regrind mill. The rougher concentrate from three rougher flotation cells was cleaned three times. Each cleaner tailing was returned to previous cleaner flotation, the first cleaner tailing to rougher flotation via regrinding. The basic flow sheet was developed from the laboratory test work done before and during the pilot plant operation. The last flotation flow sheet in autumn 2006 for selective copper-nickel flotation is described in Figure 6. Besides selective flotation the bulk flotation was studied during the last days of the pilot phase 2 (Novemver 28 – December 1, 2006). The copper flotation and the intermediate thickening were by-passed and bulk flotation was done in the previous nickel flotation circuit. The last circuit configuration of bulk-flotation is presented in Figure 7. The bulk flotation circuit consisted of three-stage roughing followed by a three-stage scavenger flotation. Rougher concentrate was cleaned twice in a closed cleaner circuit, and the first cleaner tailing was combined with the scavenger concentrate, thickened and returned back to rougher flotation through the regrinding mill. Totally 23 different flow sheet modifications were tested during the pilot phases 1 and 2. The circuit modifications from day to day are described in detail in Appendix 7.


Cy O/F

CuCC4 to thickener

NiCC3 To Thickener

Cu Rougher FlotationNi Rougher Flotation

Ni Cleaner Tailings 1/2

2 x 150 l 2 x 150 l

2 x 500 l 2 x 500 l 2 x 500 l

2 x 50 l

2 x 50 l

2 x 150 lNi Cleaner Flotation 1/1

Ni Cleaner Flotation 2


Cu CleanerFlotation 1


CuRC

CuCC1CuCT1

CuCT2

NiRC1

NiCC1/1

NiCT2NiCT3

NiCC2

CuRT

CuCT3

CuCC2

1 x 50 l



CuCC3

CuCT4

1 x 50 l

Lamellathickener

LCy O/F to Cu flotation through grinding circuit

DewateringCyclone

LCy U/FLamella U/F

2 x 500 l

2 x 150 l

CombinedTailings toTailingsPond

1 x 150 l

1 x 150 l

2 x 150 l

NiT

Ni Cleaner Flotation 1/2

High SulphurTailing

Low SulphurTailing

2 x 150 l

NiRghF1 NiRghF2 NiRghF3 NiRghF4NiRghF5

NiRC2 NiRC3NiRC4 NiRC5

Re-GrindingMill

NiCC1/2

Figure 6. Selective copper-nickel flotation at the end of pilot phase 2 (autumn 2006)

Bulk Rougher Flotation

Cleaner Tailings 1/3

2 x 500 l 2 x 500 l 2 x 500 l

2 x 150 lCleaner Flotation 1/1

Cleaner Flotation 2

RC1

CC1/1

CT2 CC2 To Thickener

2 x 500 l

2 x 150 l

1 x 150 l

2 x 150 l

Cleaner Flotation 1/2

Low SulphurTailing

2 x 150 l

RghF1 RghF2 RghF3 RghF4RghF5

RC2 RC3 RC4 RC5

Re-GrindingMill

CC1/2

Cy O/F From GrindingCircuit

Bulk Scavenger Flotation

BulkTails

2 x 150 l

CC1/3

Cleaner Flotation 1/3

ThickenerO/F to Tailings Pond

LamellaThickener

RC6

RghF6

Figure 7. Bulk flotation flow sheet 5.2 Pilot Phase 3 (Jan-Feb 2007) The third lot of ore sample, 121 t, was received from Kevitsa in December 2006. The sample was prepared for the pilot plant as previously, and the phase 3 of pilot tests was run during Jan 24 – Feb 16, 2007.


The purpose of pilot phase 3 was to study the flotation of separate Cu- and Ni-concentrates without using the special reagents, TETA and MBS. TETA was left totally out and sodium me-tabisulphite (Na2S2O5) was replaced by sodium sulphite (Na2SO3) which was used in copper flotation (and only tested shortly in nickel circuit). When pyrrhotite was no more depressed with special chemicals some other measures were taken to prevent excessive flotation of pyrrhotite: a weaker collector, sodium ethyl xanthate (SEX) was used; dosage levels of collector were substantially dropped; alcalic pH was applied in whole nickel circuit and high pH in last cleanings; a 5-stage cleaning circuit was introduced; and regrinding was used in the cleaning circuit. The grinding circuit was the same as in pilot phases 1 and 2, but a smaller feed rate, 0.85t/h, was used in order to have a finer feed to flotation and to save the ore sample for achieving a longer operation time. The feed point of milk of lime was moved from ball mill to rod mill to avoid the oxidation of ore by using a more reducing pulp environment. The flotation flow sheet in phase 3 is presented in Figure 8. The copper flotation circuit and the intermediate thickening between Cu- and Ni flotation circuits were as previously, but the nickel flotation did undergo modifications compared to pilot phases 1 and 2. Cy O/F

CuCC4 to thickener

Cu Rougher FlotationNi Rougher Flotation

2 x 150 l 2 x 150 l

2 x 500 l 2 x 500 l 2 x 500 l

2 x 50 l

2 x 50 l

2 x 150 lNi Cleaner Flotation 1





CuRC

CuCC1CuCT1

CuCT2

NiRC1

NiCC1

NiCT2

NiCC2

CuRT

CuCT3

CuCC2

1 x 50 l



CuCC3

CuCT4

1 x 50 l

Lamellathickener

LCy O/F to Cu flotation through grinding circuit

DewateringCyclone

LCy U/FLamella U/F

2 x 500 l

2 x 150 l

NiT

NiRghF1 NiRghF2 NiRghF3 NiRghF4

NiRC2 NiRC3NiRC4

Re-GrindingMill

2 x 150 l

2 x 50 l

2 x 50 l



NiCT3/1

NiCT4

NiCT5NiCC5

NiCT1

NiCC3/1

NiCC4

SulfurF

SulfurC

SulfurT

2 x 150 l

Figure 8. Flow sheet in pilot phase 3 (Jan-Feb 2007) Ni rougher flotation line consisted of three 2 x 500 liter flotation machines, followed by a scavenger cell, 2 x 500 liter. The three rougher concentrates were combined and fed to the first cleaner stage, 2 x 150 liter, and the scavenger concentrate and the first cleaner tailings were returned back to the rougher flotation. The first cleaner concentrate was fed to the second cleaner, 2 x 150 liter, and the second cleaner tailing was returned back to the first cleaning. The rougher flotation, scavenger flotation and the first and the second cleaners operated at natural pH (which was 10-9.5 in rougher flotation and 9-8.7 in the first and second cleaning).


The second cleaner concentrate was conditioned with a milk of lime before it was fed to the re-grinding mill (Sala SAM 7.5 tower mill). The mill discharge was addressed to the third cleaner flotation which had a cell volume of 2 x 150 liter. The third cleaner tailing was returned back to nickel roughing. The third cleaner concentrate was pumped to the fourth cleaner, 2 x 50 liter. The fourth cleaner tailing was returned one step backwards and the fourth cleaner concentrate was fed to the fifth cleaner, 2 x 50 liter. The fifth cleaner tailing was returned to the fourth cleaner, and the fifth cleaner concentrate was pumped to the concentrate thickener. Milk of lime was used in cleanings 3, 4 and 5 to depress pyrrhotite. The 3rd cleaner operated at pH 10, cleaner 4 at pH 10.5 and cleaner 5 at pH 11. The reagents used in nickel flotation were sodium ethyl xanthate (SEX) as a collector, MIBC frother, CMC depressant, and milk of lime for pH control. As only one tailing was discharged from flotation an extra flotation stage (2x150 liter) was added to produce two tailings, one with a high sulphur content and the other with low sulphur. This was done by floating the remaining sulphides after addition of collector and frother whereby the high sulphur tailings were obtained as froth product and low sulphur material re-mained as tailings of this cell. The day by day operation in pilot phase 3 is described Appendix 8. 5.3 Equipment Flotation equipment The flotation circuit consisted of the following cells and conditioners: Copper flotation

• Two (OK/JA-RO NO-C20, d 0.75 m x 1.2 m, impeller d 0.14 m, 2.2 kW) 200 liters conditioning tanks before flotation

• Two OK-0.15 (1.75 m x 0.82 m, 2.2 kW) flotation machines, total volume 2 x 2 x 150 liters, copper rougher flotation

• OK-0.05 (1.33 m x 0.67 m, 1.1 kW) flotation machine, volume 2 x 50 liters, copper cleaner flotation 1




Nickel flotation

• Two (OK/JA-RO NO-C20, d 0.75 m x 1.2 m, impeller d 0.14 m, 2.2 kW) 200 liters conditioning tanks before flotation

• Four OK-0.5 (2.6 m x 1.12 m, 5.5 kW) flotation machines, total volume 4 x 2 x 500 li-ters, nickel rougher and scavenger flotation


• OK-0.15 (1.75 m x 0.82 m, 2.2 kW) flotation machine, volume 2 x 150 liters, tailings flotation for scavenger tails

• OK-0.15 (1.75 m x 0.82 m, 2.2 kW) flotation machine, volume 2 x 150 liters, nickel cleaner flotation 1





The larger cells were used in roughing, scavenging and the first cleaning stages, the smaller ones in later cleanings. Auxiliary equipment A lamella thickener was used for copper tailings to control the feed pulp density of nickel flota-tion. A vertical mill (Sala SAM 7.5) was applied for regrinding in nickel flotation circuit. For pumping of slurries LPK-50, LPN-50 and LPSN-32 slurry pumps were used. When needed, the speeds of the pumps were controlled by frequency controllers. The concentrates were pumped to Supaflo thickeners (1 m by diameter) during the pilot opera-tion. The thickened concentrates were filtered by Ceramec capillary disc filter or by Larox pressure filter. Separate thickening and filtration tests were also done and are reported in this report. The flotation reagents were fed with process computer controlled diaphragm and Bredel-type peristaltic pumps. Each pump was calibrated before putting into operation, and re-calibration was done once in every working shift. The pH electrodes in conditioners and flotation cells were calibrated daily. Low pressure blower was used for feeding the flotation air, the flow rates were measured and adjusted by rotameters. The pulp levels in flotation cells were adjusted manually. The process samples were handled as follows:

• Adequate amounts of samples were collected from the slurry flows either manually or with automatic slurry samplers.

• Wet samples were weighed.

• Wet samples were filtered, dried and homogenized on hot plate.

• Dry samples were weighed for solid content determination.

• Dry samples were divided for assaying.


• Assay samples were packed into plastic bags and coded. The sample codes were put in an Excel-file with the weight and identification information.

• Analyses were made with XRF, FA and AAS (see chapter Chemical Analyses).

• An on-line analyzer (Courier-30) was used in pilot phase 3 to provide Ni-, Cu- and Fe-assays of nickel concentrate.

5.4 Flotation Chemicals The flotation chemicals used in pilot tests are presented in Table 6 and in Appendix 4/1. The number of chemicals was reduced in pilot phase 3 compared with phases 1 and 2. During phase 3 sulphuric acid, copper sulphate and TETA were not used at all. Sodium metabisulphite (MBS) was used during phases 1 and 2 both in the last cleaning of cop-per circuit and together with TETA in nickel cleaning circuit. In pilot phase 3 sulphite was not used in nickel circuit and in copper circuit MBS was replaced with sodium sulphite. Table 6. Flotation chemicals in pilot tests Chemical Task Solution % Pilot Phase 3 (Jan-Feb 2007) Ca(OH)2 Calcium hydroxide pH control 2,5 AP 3418A Sodium dithiophosphinate Collector in Cu-flotation 1 SEX Sodium ethyl xanthate Collector in Ni-flotation 5 or 0,625 PAX Potassium amyl xanthate Collector in tailings flotation 5 MIBC Methyl isobutyl carbinol Frother in Cu & Ni flotations 1 Flotanol C-7 Polyglycol Frother in tailings flotation 1 CMC (FF-BDA) Carboxy methyl cellulose Depressant in Cu & Ni flotations 1 Na2SO3 Sodium sulphite Depressant in Cu-flotation

(replaced MBS) 5

Na2S2O3 (MBS) Sodium metabisulphite Depressant in Cu-flotation 5 (was later replaced by Na2SO3) Pilot Phases 1&2 – Additional reagents H2SO4 Sulphuric acid pH control in Ni-roughing CuSO4 Copper sulphate Activator in Ni-roughing SIPX Sodium isopropyl xanthate Collector in Cu&Ni-flot (tested) DF 245 Sodium isobutyl dithio-

phosphate Collector in Cu-flotation (tested)

Na2S2O3 (MBS) Sodium metabisulphite Depressant in Cu-& Ni-flotation TETA Triethylene tetramine Depressant in Ni-flotation


The consumption of flotation chemicals is presented in Table 7 as average figures from 6-7 Fe-bruary, 2007. The chemical feed rates to different points of flotation circuits are shown in Fig-ure 9 and Appendix 4/2. Compared to pilot phases 1 and 2 the chemical consumptions were much lower in phase 3 when nickel flotation was run without special depressants (TETA and sulphite) at alkaline pH and with a changed flow sheet. Still lower collector dosages were used at the end of phase 3, but the figures in Table 7 may be more representative for needs of collectors. The consumption of CMC is very low in Table 7 and reservation could be done for somewhat higher addition levels. Sodium sulphite was used only in copper circuit, but might be useful also in nickel cleanings. The feed rates of chemicals to each flotation stage are presented in Appendix 5 for each day of pilot run. Table 7. Consumption of flotation chemicals on 6-7 Feb, 2007 Chemical Circuit Average consumption 6-7 Feb 2007 g / t Copper Flotation Lime Cu-flotation ab. 1300 AP 3418A Cu-flotation 11 MIBC Cu-flotation 27 CMC Cu-flotation 12 Na2SO3 Cu-flotation 65 Nickel Flotation Lime Ni Cleanings ab. 200 SEX Ni Rgh Flotation 118 SEX Ni Cleanings 29 MIBC Ni Rgh Flotation 47 MIBC Ni Cleanings 5 CMC Ni Cleanings 11 Total Cu- & Ni-Flotation Lime Ab. 1500 AP 11 SEX 147 MIBC 79 CMC 23 Na2SO3 65 Tailings Flotation PAX Ni tails flotation 50 Flotanol C-7 Ni tails flotation 8


KEVITSA Chemicals 6-7 Feb 2007PILOT 2007

LSYNi-Flotation NiRC total SEX 119 g/t

Cu-Flotation LY MIBC 47 g/tCycl OF Ca(OH)2, pH 11

AP 3418A: 5 g/t AP 3418A: 4 g/t LSA pH 9,8 pH 9,7 pH 9,4MIBC 22g/t AP 3418A: 1.8 g/t SEX 35 g/t SEX 26 g/t SEX 32 g/t SEX 26 g/t

LA MIBC 4,7 g/t MIBC 9,4 g/t MIBC 7 g/t MIBC 26,5 g/t SEX 50 g/t2x 150 L 2x 150 L Flotanol 8,1 g/t

Feed NiFeed 2x 500 L 2x 500L 2x 500L 2x 500L Low STRM BM NiT 2 x150L

0,85 t/h 2x 200 L CuT 2x 200 L

NiRC1 NiRC2 NiRC3 NiRC4High ST

CuRC NiRC1-3Water Ca(OH)2 CuCT1

850 g/t pH 9,1SEX 17 g/t 2 x 150L NiCT1MIBC 2 g/t

Ca(OH)2, pH 12CMC 12 g/t 2 x 50 L NiCC1MIBC 4,4 g/t

40 L CuCC12 x 150L NiCT2

CuCT22 x 50 L

NiCC250 L CuCC2

Tower MillCuCT3 Ca(OH)2, pH 10

no reagents 1x50L 2 x 150 L NiCT3CMC 7,4 g/t

CuCC3 SEX 6,4 g/tCopper Flotation Total: MIBC 2,9 g/t NiCC3 Nickel Flotation Total:Lime ab. 1300 g/t Lime ab. 200 g/tAP 3418A 11 g/t Na2SO3 65 g/t 1x50L CuCT4 Ca(OH)2, pH 10.5 SEX 147 g/tMIBC 27 g/t CMC 4,1 g/t NiCT4 2 x 50L MIBC 52 g/tCMC 12 g/t SEX 6,2 g/t CMC 12 g/tNa2SO3 65 g/t CuCC4 NiCC4

Tailings Flotation:Ca(OH)2, pH 11 2 x 50L NiCT5 SEX 50 g/t

Flotanol 8 g/t

NiCC5 Figure 9. Chemical feeds in pilot plant on 6-7 Feb, 2007 6 PILOT PLANT OPERATION 6.1 Operating Parameters The pilot plant was run totally 50 days, most of time by selective flotation for producing sepa-rate copper and nickel concentrates. Bulk flotation was run during four days (Nov 28 – Dec 1, 2006). Usual daily operation was 10-12 hours in two shifts, but also longer periods of 20-36 hours were run in three shifts. The main process operation parameters of each day are presented in the tables of Appendix 6. The dosage levels (g/t) of flotation chemicals and pH values have been compiled in Appendix 5. A more comprehensive summary with the daily operating hours, process flow sheets and the process conditions is presented in Appendices 7 and 8. In short the days and periods of the pi-lot run with the main operating parameters were as follows:


Flotation Method Date Main New Details of Tests Phase 1 (September-October 2006) No Flotation 4.9. General parameters in grinding No Flotation 5.9. Grinding fineness tuning Selective Flotation 7.9. Flotation startup, open cleanings Selective Flotation 8.9. Flotation startup, filling Ni circuit Selective Flotation 11.9. Flotation tuning, closure of cleaning circuits Selective Flotation 12.9. Finer grinding from 75 % to 81 % -75 µm Selective Flotation 13.9. Volume increase in Cu and Ni cleanings Selective Flotation 14.9. Intermediate thickening for Cu tails Selective Flotation 15.9. Water circulation in Cu circuit, MBS to Cu cleaning Selective Flotation 18.9. No regrinding in Ni circuit Thickening of Ni-tailings was started Selective Flotation 19.9. Volume increase in Ni roughing (from 3 to 4 cells) Selective Flotation 20.9. Open Ni cleaning, volume increase in 1st Ni cleaning Selective Flotation 21.9. Water circulation in Ni circuit from tailings thickener

SEX was tested in Cu flotation Selective Flotation 22.9. Collector mixture Aerophine+SEX in Cu flotation Selective Flotation 25.9. Back to Aerophine in Cu flotation, Ni rougher pH 6 MBS+TETA from Ni 1st cleaner to 3rd cleaning Selective Flotation 26.9. Finer grind by feed rate from 1.0 to 0.85 t/h. Longer retention times and higher reagent dosages in Ni-circuit Selective Flotation 27.9. Ni cleaning 1 was closed again Selective Flotation 28.9. Ni cleaning without TETA was tested in short Selective Flotation 29.9. Ni float without TETA, MBS, CuSO4 and sulphuric acid.

SIPX was replaced by SEX. Alcalic pH in Ni-circuit. Selective Flotation 4.10. As previous, separate cleaning for 1st NiRghCon. Bigger cells in Ni-cleanings 2 and 3 (2x150 L in both) Selective Flotation 5.10. As previous (i.e. no special chemicals), separate cleaning discarded. Cell volumes halved in Ni cleaners 2 and 3.


Flotation Method Date Main New Details of Tests Selective Flotation 6.10. SIBX as collector in Ni float, MBS, TETA in use again Selective Flotation 10-11.10. Three-shift run, SIBX or SIPX as collector in Ni float Selective Flotation 12.10. Kitinen water in use (2.1 m3/h), closed Ni cleaning, Process conditions like 19-20.9. Selective Flotation 13.10. Kitinen water, open Ni cleaning. Phase 2 (November-December 2006) Selective Flotation 13-14.11. Volume increase in Ni roughing, 4th Ni cleaning in use,

regrinding in use, open Ni-rougher flotation, MBS+TETA in use, no water circulation from Ni-tails

Selective Flotation 16-17.11. Regrinding bypassed, 4th Ni cleaning discarded, volume

increase in 1st Ni cleaning (open circuit) Selective Flotation 21-22.11. Ni roughing divided, regrinding for scavenger concen-

trate and 1st Ni cleaner tails, TETA not used and other reagents were changed accordingly (see 29.9) Selective Flotation 23-24.11. Ni roughing divided, regrinding for scavenger concen-

trate and 1st Ni cleaner tails, TETA not used Selective Flotation 24.11. 3-shift run from 23.11 continued. TETA was added

(about 75 g/t) Bulk Flotation Bulk Flotation 28.11. Bulk flotation, closed circuit with 3 cleanings, pH 6 in

roughing, scav conc and 1st cleaner tails to thickener, re-grinding and back to heads

Bulk Flotation 29.11. Bulk flotation, closed circuit, pH 4 in roughing Bulk Flotation 30.11. Bulk flotation, closed circuit with 2 cleaners (3rd cleaner

discarded), pH 4 in roughing Bulk Flotation 1.12. Bulk flotation, closed circuit, lime added starting from

rod mill, pH 10,5 in roughing


Date Main New Details of Tests Phase 3 (January-February 2007) Selective Flotation 22-23 Jan Start-up, filling the circuits, tuning of grinding fineness, calibration of reagent pumps, pH controls, particle size analyzer 24 Jan Running with the new circuit configuration and reagents adopted for phase 3, Na2SO3 to Cu-cleaning 4 (instead of MBS), no sulphite to Ni-cleanings 25 Jan MBS to Cu-cleaning 4, Na2SO3 was tested 88 g/t to Ni-cleaning 3, MIBC was replaced by Flotanol C-7 in Ni-circuit, inceased pyrrhotite circulation, and was changed back to MIBC 29 Jan Adjustment of reagent dosages and pH levels to improve the results, slightly more MBS to Cu-cleaning 4 30 Jan Adjustment of reagent dosages, in Cu-rougher flotation pH from 11 to 11.5 and smaller dosage of collector. Later, collector from Cu-cleanings was taken off. Dosage of SEX was increased in Ni-roughing. 1 Feb Na2SO3 back to Cu-cleaning 4 (instead of MBS), collector dosages further reduced in Cu-circuit, rougher pH back to 11. In Cu-cleanings, lime addition only to cleaner 1 (to pH 12), other cleaning without lime. An additional flotation stage for final tailings to make high and low sulphur tailings 6 Feb An intermediate regrinding was tested for Ni-rougher tailings 2 and the product continued to rougher cell 3. This was to test cleaning of mineral surfaces from suspected surface oxidation. No benefits was seen from intermediate grinding. 7 Feb The extra grinding was discarded. Otherwise similar run as on 6 Feb. 8 Feb The 6th cleaning stage was tested in Ni-circuit. pH-levels in Ni-cleanings 3 and 4 were slightly inceased, 10.5 to 10.7 and 11.0 to 11.2, respectively, and also CMC-dosages were increased in the same stages. 13 Feb Ni-circuit without regrinding, the 6th cleaning was taken off. 14 Feb Regrinding in use again, the dosage level of SEX was reduced further. 15-16 Feb Period of 21.5 hrs in three shifts.


6.2 Process Water Process water in the pilot tests was pumped from the tailings pond of GTK pilot plant. This water contains ions dissolved from different tailings materials accumulated during the years in the tailings area. As such it does not represent the water of Kevitsa tailings, but anyway con-tains elevated concentrations of ions typical for sulphide tailings. The ion composition of this water as well as that of Kitinen river water and GTK tap water is shown in Table 8. The recirculated water was needed as surplus water to grinding circuit and flotation cell laun-ders. Otherwise water to grinding circuit was obtained from intermediate thickener of copper tailings. The thickening of nickel tailings and water circulation to beginning of nickel circuit was also tested during phase 1 in September 2006. It was, however, abandoned because dilution water was not needed after lamella thickener and also it was suspected that the slimes in tailings had an adverse effect in flotation. About 60 m3 of Kitinen river water was transported to GTK in the autumn 2006 and this was used as process water on 12-13 October, 2006 (replacing the water from tailings area). No evi-dent changes were observed in flotation with Kitinen water. Table 8. Water analyses

1 2 3Pilot Kitinen Tap

recirculated river water6 Nov 2006 June 2006 June 2006

pH 7,66 7,44 7,65Li mg/L 0,01 < 0,01 0,02Na mg/L 37,92 0,84 2,38Mg mg/L 12,87 0,93 4,80Al mg/L < 0,01 0,08 < 0,01P mg/L < 0,01 0,01 0,02S mg/L 44,89 0,58 3,69K mg/L 4,38 1,43 4,92

Ca mg/L 40,03 2,32 12,30Mn mg/L 0,02 < 0,01 < 0,01B mg/L 0,04 < 0,01 0,02Fe mg/L 0,01 0,49 < 0,01Co mg/L 0,01 < 0,01 0,01Ni mg/L < 0,01 0,03 0,01Cu mg/L 0,01 < 0,01 0,01Zn mg/L < 0,01 < 0,01 0,01As mg/L 0,12 < 0,01 < 0,01Cd mg/L < 0,01 < 0,01 < 0,01Sr mg/L 0,16 < 0,01 0,06Ba mg/L 0,06 < 0,01 0,01Pb mg/L 0,01 < 0,01 < 0,01Cr mg/L < 0,01 < 0,01 < 0,01Cl- mg/L 27,6 0,89 13,70

NO3- mg/L 1,2 23,9 0,04SO42- mg/L 112 - -

Notes:1 Recirculated from tailings water pond of GTK pilot plant2 Kitinen river water used in pilot tests 12-13 Oct, 20063 GTK tap water


6.3 Chemical Analyses Process samples Pulp samples from pilot plant were collected for two purposes:

1) Samples from most important process streams to follow and control the flotation. These samples included 5 samples: flotation feed (cyclone overflow), copper concentrate, copper tailings, nickel concentrate and nickel tailings. In some cases only the feed, con-centrate and tailings of nickel flotation were collected as the copper circuit was running rather steadily. The samples were collected either from each hour or every second hour. Usually Cu, Ni and Fe were analysed (AAS).

2) Samples for calculating the metallurgical results and balances. These covered all of the process streams, i.e. concentrates and tailings of all flotation stages, in some cases also individual cells. These samples were collected once or twice per day after the process had stabilized (usually near the end of day). For these samples Cu, Ni and Fe were ana-lysed by AAS. Full assays, also including precious metals, sulphur and XRF-analyses, were done for selected sets of samples.

Pulp samples of flotation feed, concentrates and final tailings were collected by automatic sam-plers. The other pulp samples were collected manually, usually three sample rounds per hour. The samples were filtered and dried after which subsamples were split for analysis. The main samples were stored dry. On-line analyses Courier 30 online analyzer was used during pilot phase 3 to monitor the assays of Ni, Cu and Fe of final nickel concentrate. The analyzer was calibrated beforehand with concentrate sam-ples from earlier pilot phases. Concentrates The copper and nickel concentrates were thickened and then filtered by ceramic capillary disc filter or Larox pressure filter during which samples of filter cake were collected and then ana-lysed. Filter cakes from different periods of pilot run (usually one or two days per period) were analysed and stored separately. Water analysis Samples were collected from water phase of tailings and concentrates which were analysed by ICP giving assays of 21 elements. The same analysis were also done for process waters. Assay methods The following methods were used for analysis: Cu, Ni, Fe:

1) AAS after nitric acid dissolution, giving total assays of elements (used for regular analysis)


2) AAS after brome methanol dissolution, giving these metals in sulphide minerals Sulphur:

Combustion infra-red analysis by Eltra-analyzer Pt, Pd, Au: Fire Assay followed by dissolution and assaying by AAS 40 elements (including e.g. Mg): XRF In addition ICP was used for water analysis to yield assay of 21 elements 7 RESULTS Most of the pilot tests were run by producing two concentrates, copper and nickel, in a selec-tive flotation flow sheet. Therefore the results of selective flotation will be presented first. Af-ter this the results of bulk flotation will be handled. The main results of both flotation concepts are presented in the tables of Appendix 9/1. The same numbers are also shown in Tables 9, 10, 11 and 12 in the text. 7.1 Selective Flotation The results of selective flotation have been compiled in the tables of Appendix 9 and graphs of Appendix 10. The main results are also presented as numbers and graphs in the text below. 7.1.1 Copper Flotation The results of copper flotation are shown in Table 9 and Figure 10. The grade-recovery plot of Figure 10 shows that generally copper recovery was 75-84 % and the grade 27-32 %. The av-erage results in phases 1, 2 and 3 were: grade 28,5-30,5 % Cu and copper recovery 77-80 % (Table 9). Copper flotation pointed out quite successful taking into account the rather low grade of the ore (phase 1: 0,51 % - phase 2: 0,55 % - phase 3: 0,45 % Cu). Nickel losses to copper concentrate were small, usually 1,5-2,5 % of total nickel in the feed. It was also important to keep the nickel content of copper concentrate low enough, preferably below 0,4 %, and anyway below 0,8 %. During pilot phase 1 the content was slightly too high, 0,81 % Ni as shown in Table 9. During phases 2 and 3 the mass pull in copper cleanings was restricted in order to keep copper grade high and nickel content low. It is seen that Ni was 0,50 % during phase 2 and 0,61 % during phase 3. Concerning the precious metals the target was that gold would report to Cu-concentrate and PGM’s to Ni-concentrate. It is seen in Table 9 that 25-39 % of gold, but only 8-11 % of plati-num and 12-16 % of palladium floated to copper concentrate. The recovery of gold dropped to 25 % during phase 3 (from the earlier 37-39 %) which may be connected with lower copper grade in the ore and thus lower mass recovery of copper concentrate. The grade-recovery plot of gold in copper concentrate is shown in Figure 11.


Table 9. Copper concentrate in selective flotation

Mass Cu RCu Ni RNi Fe S Pt RPt Pd RPd Au RAu MgO Phase wt% % % % % % % g/t % g/t % g/t % %

1 Sept-Oct

2006 1,42 28,48 78,6 0,810 3,0 30,6 30,1 3,91 8,1 4,42 16,0 4,37 37,8 2,8

2 Nov 2006 1,44 30,50 80,3 0,504 1,8 2,50 10,0 3,17 13,3 4,41 39,3

3 Jan-Feb

2007 1,18 29,28 77,3 0,612 2,0 30,9 34,1 2,96 10,7 3,19 11,6 3,92 25,0 2,3

Kevitsa Pilot 2006-2007Cu-Flotation

74

76

78

80

82

84

86

24 26 28 30 32Cu, %

Cu-

Rec

over

y, %

Phase 1 (Sept-Oct2006)

Phase 2 (Nov 2006)

Phase 3 (Jan-Feb2007)

Figure 10. Cu-recovery vs. grade in copper concentrate


Kevitsa Pilot 2006-2007Au in Copper Concentrate

10

20

30

40

50

60

2 3 4 5 6 7Au, g/t

Au-

Rec

over

y, %

Phase 1 (Sept-Oct2006)

Phase 2 (Nov 2006)

Phase 3 (Jan-Feb2007)

Figure 11. Recovery and grade of gold in copper concentrate 7.1.2 Nickel Flotation The feed of pilot plant contained 0,40 % Ni during phases 1 and 2 (2006) and 0,36 % Ni during phase 3 (2007). The results are presented for periods without and with the special reagent, TETA. During phase 3 in January-February 2007 this chemical was not used. The main results of nickel flotation are shown in Tables 10 and 11, Ni-grade/recovery graph in Figure 12. The details can be found in Appendices 9 and 10. Table 10. Nickel concentrate in selective flotation

Mass Cu RCu Ni RNi Fe S Pt RPt Pd RPd Au RAu MgO Fe / Phase wt% % % % % % % g/t % g/t % g/t % % MgO

1 Sept-Oct 2006 9 best days (TETA) 2,55 1,85 9,2 10,95 66,8 32,6 22,6 8,51 50,5 6,85 48,8 1,25 18,4 10,1 3,5

Sept-Oct, 3 days (withoutTETA) 3,21 1,45 9,5 8,11 63,2 40,3 26,9 5,06 45,1 4,44 41,0 0,98 14,2 7,9 5,5

2 23-24 Nov 2006 (without TETA) 4,21 1,25 12,3 6,69 70,5 44,0 4,16 47,9 4,20 50,9 0,85 21,9 6,5 6,8

24 Nov 2006

(TETA) 3,50 1,22 8,8 7,83 68,1 41,9 6,15 57,1 4,97 48,6 0,77 15,8 6,6 6,4

3 Jan-Feb 2007 1,97 1,85 11,2 12,16 65,3 44,3 35,8 10,69 58,1 8,04 42,8 1,56 15,0 2,7 16,3


Kevitsa Pilot 2006-2007Ni-Flotation

16,66; 63,8

15,35; 64,6

12,45; 65,510,85; 65,7

9,32; 68,0

7,21; 69,0

60

70

6 8 10 12 14 16 18 20Ni, %

Ni-R

ecov

ery,

%

Phase 1 (Sept-Oct 2006)

Phase 2 (Nov 2006)

Phase 3 (Jan-Feb 2007)

Averages with different grade levels

Daily Results

Figure 12. Ni-recovery vs. grade in nickel concentrate The highest grades of nickel concentrate were achieved in phase 3 (Jan-Feb 2007), average 12,2 % Ni, although special chemicals (MBS and TETA) were not used. The recovery of nickel was 65,3 %. The quality of concentrate was also otherwise better than earlier as Fe was high (44,3 %) and MgO much lower (2,7 %) than earlier thus yielding a good Fe/MgO-ratio (16,3). It is also seen that, as the mass of concentrate was lower, the grades of precious metals were higher in phase 3. The recoveries of precious metals kept at the same level as earlier being 58 % fo Pt, 43 % for Pd and 15 % for Au. The results with different Ni-grade levels of concentrate during phases 1-3 are shown in Table 11. It is seen that as the grade increased from 7,2 % to 16,7 % Ni the recovery decreased from 69,0 to 63,8 %. In phase 3 the recovery of nickel did not actually change much although the grade of concen-trate varied between 9 and 18 %. With higher grades (14-18 %) the recovery was 64-65 % and with lower grades (9-12 %) 64-67 %. It was seen in phase 3 that rather high grade Ni-concentrates could be produced without the use of special chemicals and the content of silicates could be dropped at the same time. With concentrate grades of 12-18 % Ni the recovery of nickel remained at 64 to 66 % which is somewhat lower than target (about 70 %). The main reason was concluded to be mineralogical: it pointed out that the pilot feed contained 460 to 600 g/t Ni in silicates and also some nickel in goethite, i.e. the content of nonfloatable nickel was significantly higher than normally in the ore (275 g/t Ni in silicates in the average main ore).


Table 11. Different levels of nickel concentrate during phases 1-3

Mass Cu RCu Ni RNi Fe MgO Fe/MgO Grade level wt% % % % % % %

Ni 6-8 % 3,87 1,30 10,9 7,21 69,0 39,0 8,8 5,2

Ni 8-10 % 2,82 1,40 9,4 9,32 68,0 36,6 10,4 3,5

Ni 10-12 % 2,19 1,79 9,8 10,85 65,7 40,6 7,6 7,9

Ni 12-14 % 2,01 1,96 11,8 12,45 65,5 38,3 7,4 9,5

Ni 14-16 % 1,48 2,20 10,4 15,35 64,6 41,7 2,7 15,4

Ni 16-18 % 1,44 2,72 9,1 16,66 63,8 37,4 4,0 11,5

The results in different phases of pilot campaign (Table 10) will be still discussed in the fol-lowing. Tests with TETA (about 300 g/t) During the 9 best days of phase 1 Ni-grade was 10,95 % with a recovery 66,8 % as shown in Table 10. During these days TETA was used to depress pyrrhotite in nickel flotation. The mass pull of concentrate was 2,55 wt-%. Iron content was low, 32,6 % Fe, and MgO was high, 10,1 %, giving a low Fe/MgO-ratio 3,5 (as the target was higher than 5). So it was seen that Fe was depressed effectively, but the content of silicates was high. Final tailings contained 0,127 % Ni. Tests without TETA The most successful period without TETA was run in phase 3 and the results were already dis-cussed above. Some periods without this chemical were also run earlier and they are com-mented here. During phase 1 there were three days when TETA was not used. The recovery of nickel was 63,2 % and grade 8,11 %. The nickel recovery remained low when the floatability of excess pyrrhotite was restricted by using a weaker collector (ethyl xanthate), a smaller dosage of it, and high pH in Ni-cleanings. Fe was now higher, 40,3 % and MgO lower, 7,9 %, thus giving a better Fe/MgO ratio 5,5. There was also a period without TETA during phase 2; 23-24 November (3-shift run). The tar-get was to increase the recovery of nickel and also to get more iron to Ni-concentrate. As seen in Table 10 the mass pull was 4,21 wt-%, Ni-recovery 70,5 %, grade 6,69 % Ni and content of Fe 44,0 %. MgO was 6,5 % and Fe/MgO 6,98. So the recovery of nickel and Fe/MgO-ratio improved, but nickel grade remained poor. Tests with small addition of TETA As nickel grade was low during the last period of phase 2 (23-24 Nov 2006) some measures were taken to avoid excessively high flotation of pyrrhotite. This was done by adding some TETA (75 g/t, as the normal dosage had been 300 g/t), increasing CMC dosages and control-


ling the mass pull (air flow rates to cells). The results of this period are also shown in Table 10. It is seen that mass pull dropped to 3,50 wt-%, grade increased to 7,83 % Ni while nickel re-covery decreased slightly to 68,1 %. Content of Fe dropped to 41,9 % as did also Fe/MgO ratio (6,4). So the achieved Ni-grade was still low. Copper and precious metals in nickel concentrate According to Table 10 the copper recovery to nickel concentrate was 9-12 % meaning that total recovery of copper was 90 % or more. Copper grade in nickel concentrate was usually 1,5-2 %. The recoveries of precious metals to nickel concentrate were: Pt about 50 % (45-58 %), Pd about 45 % (41-51 %) and gold about 15 % (14-22 %). So it was seen that PGM’s were float-ing mainly to nickel concentrate and gold predominantly to copper concentrate. The grade/recovery graphs of platinum and palladium in nickel flotation are presented in Fig-ures 13 and 14.

Kevitsa Pilot 2006-2007Pt in Nickel Concentrate

20

30

40

50

60

70

80

2 4 6 8 10 12 14 16Pt, g/t

Pt-R

ecov

ery,

%

Phase 1 (Sept-Oct 2006)

Phase 2 (Nov 2006)


Figure 13. Recovery and grade of platinum in nickel concentrate Fe and MgO in nickel concentrate The content of MgO and the ratio Fe/MgO in nickel concentrate are plotted vs. Ni-recovery in Appendix 10/4. The graphs visualize how MgO dropped and ratio Fe/MgO improved in pilot phase 3 when compared with phases 1 and 2 of the pilot run. In phase 3 the average analysis were 44,3 % Fe and 2,7 % MgO giving Fe/MgO = 16,3 (Table 10).


Kevitsa Pilot 2006-2007Pd in Nickel Concentrate

10

20

30

40

50

60

70

80

2 4 6 8 10 12Pd, g/t

Pd-

Rec

over

y, % Phase 1 (Sept-Oct 2006)

Phase 2 (Nov 2006)


Figure 14. Recovery and grade of palladium in nickel concentrate 7.2 Bulk Flotation Bulk flotation was run during four days from 28th November to 1st December, 2006. The target was to improve the recoveries of all valuables, on the cost of grades - of course, compared with the selective flotation. The downstream processing method for bulk concentrate would be leaching. The results of bulk flotation are shown in Appendix 11 and are also presented in Table 12 and Figures 15, 16 and 17. During the first three days of bulk flotation it was seen that it was difficult to get nickel recov-ery substantially higher than what it was in nickel flotation during the selective scheme. Nickel rougher flotation was run at low pH (5,5 – 4,5) during those days of bulk flotation. The lowest content of nickel in tailings was 0,102 %. For the last day of bulk flotation another concept was taken into use by running the circuit in alkalic conditions. Lime was added starting from the rod mill. Clearly lower tailings grades and higher recoveries were obtained. During the first hours Ni-content in tailings was 0,061-0,062 % when high amount of concentrate was floated with a low grade of Ni (starting from 2,54 % Ni). After this the mass pull was restricted and the results shown in Table 12 were achieved during the last 4 hours. The mass pull of bulk concentrate was 9,27 wt-%, Cu-recovery 94,9 % and Ni-recovery 84,4 % with the grades 6,06 % Cu and 3,65 % Ni. Tailings contained 0,069 % Ni.


The recoveries of precious metals, as also shown in Table 12, were high: platinum 78 %, palla-dium 90 % and gold 77 %. When a high amount of concentrate was floated rather much of silicates were also coming to concentrate as shown by the content of MgO, 12,9 %. Fe-content was 30,3 % and ratio Fe/MgO 2,35. There were, however, no specific limiting values for MgO as discussed with the representatives of client during the trial. Table 12. Results of bulk flotation

Mass Cu RCu Ni RNi Fe S Pt RPt Pd RPd Au RAu MgO Fe / Phase wt% % % % % % % g/t % g/t % g/t % % MgO

2 Dec 1, 2006 9,27 6,06 94,9 3,65 84,4 30,3 21,7 2,99 77,6 3,26 90,0 1,36 77,0 12,9 2,35

The development of the results during the four days of bulk flotation is shown in Figures 15, 16 and 17.

Kevitsa Bulk Flotation 28.11-1.12.2006Grades of Concentrate

0

2

4

6

8

10

12

14 16 18 20 10 12 14 16 18 20 10 12 14 16 11 13 15

%

Cu

Ni

Cu-%

Ni-%

28 29 30 1 Dec

Figure 15. Cu- and Ni-grades in bulk concentrate during Nov 28 – Dec 1, 2006


Kevitsa Bulk Flotation 28.11-1.12.2006Recoveries

60

70

80

90

100

14 16 18 20 10 12 14 16 18 20 10 12 14 16 11 13 15

%

Cu-Rec

Ni-Rec

Cu-Rec

Ni-Rec

28 29 Nov 30 1 Dec

Figure 16. Cu- and Ni-recoveries in bulk flotation during Nov 28 – Dec 1, 2006

Kevitsa Bulk Flotation 28.11-1.12.2006Grades in Tailings

0,000

0,050

0,100

0,150

0,200

14 16 18 20 10 12 14 16 18 20 10 12 14 16 11 13 15

%

Cu

Ni

Cu-%

Ni-%

28 29 30 1 Dec

Figure 17. Cu- and Ni-content in tailings of bulk flotation during Nov 28 – Dec 1, 2006 7.3 Floatable and Nonfloatable Nickel Series of ore and tailings samples were selected for which nickel in sulphides and silicates was analysed. Total Ni was analysed by AAS after nitric acid dissolution, sulphide Ni after brome


methanol dissolution and Ni in silicates was calculated by subtracting sulphide Ni from total Ni. The analyses are presented as averages in Table 13 and in more detail in Appendix 12/1-2. Table 13. Nickel analyses of ore samples, pilot feed and flotation tailings

Samples / Date Note Total Sulphide Silicate Silicate Ni Ni % Ni % Ni % % of total Drill core samples: 20 samples (received Feb ’07) Ore samples 0,290 0,266 0,0237 8,2 Selective Pilot: Pilot sample 1 (Sept-Oct ‘06) Feed ore 0,376 0,331 0,046 12,1 Pilot sample 2 (Nov ‘06) Feed ore 0,382 0,335 0,047 12,4 Pilot sample 3 (Jan-Feb ‘07) Feed ore 0,348 0,287 0,061 17,4 Laboratory tests: Drill core samples (Kevitsa 8) Tailings 0,041 0,009 0,032 77,4 Pilot sample 1 (Sept-Oct ‘06) Tailings 0,067 0,014 0,054 79,8 Pilot sample 3 (Jan-Feb ‘07) Tailings 0,083 0,022 0,061 73,3 Selective Pilot: Pilot sample 1 (Sept-Oct ‘06) Tailings 0,138 0,074 0,064 46,4 Pilot sample 2 (Nov ‘06) Tailings 0,137 0,076 0,061 44,3 Pilot sample 3 (Jan-Feb ‘07) Tailings 0,139 0,066 0,073 52,5 Bulk Pilot: Pilot sample 2 (Dec ‘06) Tailings 0,069 0,018 0,051 74,0

Note: The drill cores and pilot feed samples were analysed by using a sensitive calibration of AAS The first analyses in Table 13 are averages of 20 drill core samples which were selected by Kevitsa Mining Oy to represent the Kevitsa main ore. It is seen that with a total nickel content of 0,290 % the assay of silicate Ni is 237 ppm or 8,2 % of total Ni. So the analysed silicate nickel corresponds to the expected level of about 250 ppm (275 ppm is used in the estimates for average main ore). In the feed of pilot plant the content of silicate Ni was higher, 460-610 ppm or 12-17 % of total Ni. The share of silicate Ni was largest in pilot sample 3 which was run in January-February 2007. The silicate Ni in flotation tailings of drill core samples (not the same as the analysed 20 cores) was 320 ppm as shown in the table (variation being 230-390 ppm among 6 drill core samples). These samples do not represent the pilot feed, but originate from drill holes no. 40, 42 and 49 selected for Ni feed grade/recovery study. The flotation tailings of pilot samples contained silicate Ni 610-730 ppm in pilot tests and 540-610 ppm in laboratory tests (analysed by normal calibration of AAS). The sulphide Ni in tailings of laboratory tests was 90, 140 and 220 ppm being highest in tests with pilot sample 3. In pilot tests the sulphide Ni was lowest in bulk flotation tailings, 180 ppm, but was significantly higher in selective flotation tailings, 660-760 ppm. So, it was seen that much more sulphide nickel remained in the tailings of pilot plant compared to laboratory tests. It must be, however, noted that in bench scale tests the tailings refer to the rougher tail-ings only as cleaner tailings were not circulated in batch tests.


In pilot plant the sulphide Ni was lowest (660 ppm) in the tailings of phase 3 according to as-says in Table 13. It is seen in Appendix 12/2 that in pilot phase 3 the content of both sulphide and silicate Ni in tailings increased towards the end of pilot run. The feed ore may have been more oxidised at the end of campaign as suspected during the pilot run. Recovery of Total and Sulphide Nickel The contents of different nickel phases (total, sulphide and silicate) in feed, Ni-concentrate and tailings are shown in Table 14 for three periods of pilot tests. The same results are presented for selected days of each period in Appendix 12/3. The recovery of sulphide nickel to Ni-concentrate was calculated and is also shown in Table 14 in addition to recovery of total nickel. It was supposed that Ni in concentrate is only sulphide nickel like shown by the phase analyses of concentrates. The average recoveries of total nickel were 65-69 % as seen in the table. The recovery of sul-phide nickel was 76,1 % in phase 1 of pilot run and improved to 78,8 % in phases 2 and 3. The flotation result was the best in phase 3 noting that the grade of concentrate was highest and, particularly, the feed ore had the lowest total Ni and the highest silicate Ni. Table 14. Total and sulphide nickel recoveries in pilot plant phases 1-3

Feed ore Ni Concentrate Ni Tailings Ni tot Ni sulf Ni sil Ni tot Ni tot Ni sulf Ni tot Ni sulf Ni sil % % % % Rec % Rec % % % %

Selective Pilot: Phase 1 (Sept-Oct ’06) 0,398 0,351 0,047 11,0 66,8 76,1 0,127 0,079 0,049 Phase 2 (Nov ’06) 0,408 0,361 0,047 7,3 69,2 78,7 0,119 0,069 0,050 Phase 3 (Jan-Feb ’07) 0,359 0,299 0,060 12,2 65,3 78,8 0,120 0,058 0,062 7.4 Material Balances The material balances in the flotation circuits were calculated by using the BILCO calculation program specifically designed for this purpose. Calculations were done for 18 days of pilot campaign and they were based on samplings and analyses of all pulp streams inside the cir-cuits. The calculated material flows and analyses are presented in Appendix 13. Balance calculations are based on the measured feed rate (t/h) of pilot plant and analyses of Cu, Ni, Fe, Au, Pd and Pt in each flow. As a result of calculation all analyses have to match a bal-ance situation in flotation circuit. The calculation thus changes many of the original analyses. The analyses and recoveries may thus differ to some extent from the numbers presented as the especial results of pilot tests in this report. Balance calculations and the related results show what is happening in each stage of flotation circuit and give clues how to develop the process operation. Balances also serve as a basis of process design of full scale operation. Some comments are given in the following concerning the different phases of pilot run.


September-October 2006 (Phase 1) The balances were calculated for five days (between 5-12 October) when nickel circuit was run with 3 cleanings, the tailings of 1st cleaning was returned to nickel roughing, and regrinding was not used. Rather high dosages of chemicals, like collector SIPX, were used and MBS+TETA was used to depress pyrrhotite. Nickel rougher flotation was either done at pH 5 or pH 9,5-10 (natural after copper flotation). The exception was 5th of October when MBS and TETA were not used and ethyl xanthate was as a collector. The following comments can be done on the basis of data in Appendices 13/1-5:

- Rather good results could be obtained for nickel, as an example 11th of October: 11,52 % Ni and recovery 67,5 %. This was obtained by depressing pyrrhotite with TETA and as a result the content of iron was low, 31,8 %. The drawback was that nickel concen-trate contained 20 to 30 % silicates and ratio Fe/MgO was low. This was a general trend in pilot tests during phases 1 and 2 (September-November 2006).

- Ni and Fe were floating slowly in rougher circuit. Although most of nickel floated in rougher cell 1 much of it still floated in the last rougher cell (nr. 4). The same holds true for precious metals.

- Ni, Pd and Pt were floating in the same ratio in nickel cleanings as about 60 to 64 % of these metals floated to final concentrate from the feed of cleaning circuit (11th October, for example). It is generally noted that flotation of Pd and Pt was closely correlated with the recovery of nickel. At the same time, Au floated much less than these elements in nickel circuit. Fe was depressed in cleanings so that usually only 10 to 20 % was re-covered (and part of Fe is contained by pentlandite and chalcopyrite in addition to pyr-rhotite).

- On 5th of October, when TETA was not used, the iron content in final concentrate was higher, 41,8 % Fe, but nickel recovery was rather modest, 62,7 %, with a grade of 10,0 % Ni.

November-December 2006 (Phase 2) Selective flotation The selective flotation tests during phase 2 are represented by a 3-shift run on 23-24 November in Appendices 13/6-7. MBS and TETA were not used, ethyl xanthate was as a collector and high pH (about 12) was applied in nickel cleanings to depress pyrrhotite. Nickel rougher flota-tion was done at pH 9,5-10 (natural). The circuit configuration had been changed so that rougher concentrates 1-3 were addressed to cleanings whereas rougher concentrate 4+5 and cleaner tailings 1 were pumped to regrinding and circulated back to rougher flotation. The fol-lowing observations can be done:

- The grade of nickel concentrate kept low, 6-9 % Ni, in these tests without TETA (as an example 8,65 % Ni with recovery of 66,3 % on 24th November). With these nickel grades the concentrate contained more Fe than earlier, 39-44 %. However, the content of silicates kept still high in concentrate.

- A careful mass pull was done in nickel rougher flotation, especially when TETA was not used in cleanings and, as a result, Ni was floating slowly in rougher cells. Part of nickel was not recovered until in rougher cell no.5. The long flotation time, however, brought a lot of silicates to circulation and part of them ended up to final concentrate.


- Like in phase 1, Pd and Pt were floating in nickel cleanings in the same ratio as nickel as the recoveries from cleaner feed to final concentrate were 72 % for Ni, and 65 % for Pd and Pt. The corresponding recovery was 55 % for Au and 36 % for Fe (24th Novem-ber).

Bulk flotation Bulk flotation tests were performed during four days from 28th November to 1st December 2006. The balance calculations are presented for two last days in Appendices 13/8-9. On 30th November the bulk rougher flotation was run at pH 4,5 and on 1st December at pH 10,5. Much higher recoveries were achieved at pH 10,5. The rougher concentrates were cleaned in two stages to yield the final bulk concentrate. For bulk flotation on 1st December the following ob-servations can be done:

- The mass pull of concentrate was 8,8 wt-% which is much higher than totally for two concentrates in selective flotation. A high recovery of metals was targeted whereby only a small part of iron was depressed in cleanings and the content of silicates in con-centrate was not an issue of concern.

- The recoveries of Ni and precious metals were substantially higher than in selective flo-tation whereas the grades of bulk concentrate were much lower.

- A rather low tailings grade, 0,102 % Ni, was already achieved after two rougher flota-tion stages, but still a long flotation time (totally 6 rougher stages) was needed before the final grade 0,072 % Ni was obtained.

January-February 2007 (Phase 3) The selective flotation tests were continued from 24th January to 16th February 2007. The whole period was run without MBS+TETA and the configuration of nickel circuit and use of chemicals were adjusted to make it possible to produce decent nickel grades without special chemicals. Nickel rougher flotation was shortened to original four stages from which rougher concentrates 1-3 were addressed to cleanings. Cleaning circuit was extended to include five stages with the idea to depress silicates in the first two cleanings after which the second cleaner concentrate was reground before the three last cleanings in which the excess pyrrhotite was depressed. It was, however, important to keep the iron content of concentrate high in order to achieve a good Fe/MgO ratio. The balance calculations are presented for 9 days in Appendices 13/10-18. The following con-clusions can be drawn:

- Rather good grades of nickel concentrate were achieved although special chemicals were not used. The grade level during the whole phase 3 varied from about 9 to 18 % Ni and the iron content was high at the same time, from 47 to 40 % Fe (as presented in Appendix 9).

- The content of silicates in nickel concentrate was significantly lower than earlier, the average MgO being 2,7 % compared to earlier averages 6,5 – 10 % MgO. The ratio Fe/MgO was now 16,3 compared to earlier best averages 3,5 – 7,9 (Appendix 9).

- In nickel rougher flotation, almost all of valuable elements were floated already in the first rougher cell (much more than during pilot phases 1 and 2). Rougher stages 2, 3 and


4 were needed to improve the recovery although this meant a substantial extension of flotation time.

- Nickel grade did not usually improve much in cleaning stages 1 and 2 which mainly served as silicate depression phases. The second cleaner concentrate was directed to re-grinding (which was used during the whole phase 3 except on 13th February).

- In cleaner stages 3, 4 and 5 the grade of concentrate was improved by depressing the excessive pyrrhotite. It is seen that Fe was usually higher in tailings than in concen-trates of these cells.

- Like in the earlier selective flotation tests, Pd and Pt floated in nickel cleaning circuit in the same ratio as nickel, i.e. about 60-70 % of Ni, Pd and Pt was recovered to final con-centrate from the feed of cleaning circuit. Au was depressed to higher extent in nickel circuit. For Fe the recovery was only 20 to 30 % because of pyrrhotite depression.

- In pilot run without regrinding on 13th February the grade of concentrate remained at 8,2 % Ni and Fe-content was high, 49,7 %. MgO was low also now, 3,55 %. It is obvi-ous that pyrrhotite was not depressed effectively when regrinding was not used.

- It was concluded that the meaning of regrinding was more in creating suitable pulp chemical conditions than in liberation or size reduction. The iron from regrinding cre-ated chemically reducing conditions (negative electrode potentials) which helped in de-pressing pyrrhotite.

- Circulation of Ni and Fe to beginning of nickel circuit could be rather high although mass circulation kept rather low. For example, in tests on 13-16 February, the calcu-lated nickel circulation was 25 to 52 % and iron 27 to 41 % when mass circulation was only 5,0 to 6,7 %.

7.5 Recovery correlations Metal recoveries from balance calculations were used to plot correlations presented in Figures 18-22. The recoveries, which are total recoveries to copper and nickel concentrates or to bulk concentrate, are also shown in Table 15. It is seen that the recoveries of precious metals improved when the recoveries of Cu, Ni and Fe increased (Figures 18, 19 and 20). Copper recovery seems to have a strong effect on the recov-eries of Pd and Au (Figure 18), but this effect is probably indirect; the copper recovery in-creases together with recoveries of Ni and Fe, or when more nickel concentrate has been floated. The recovery of palladium seems to increase strongly with the recoveries of Ni and Fe (Figures 19 and 20) although data is very scattered. Also platinum seems to correlate with Ni whereas gold correlates more with Fe. In circuit analysis (chapter 7.4) it was seen that Pd and Pt be-haved in nickel circuit in the same way as nickel; the recoveries of Ni, Pt and Pd decreased in the same ratio in the nickel cleanings. The recovery of nickel increases with the recovery of Fe although the dependency is not very strong as Ni-recovery improves about 10 % (65 75 %) when Fe-recovery increases from 20 to 47 % (Figure 21). Copper recovery depends more on Fe-recovery than on Ni-recovery (Figure 22).


Table 15. Total recoveries to concentrates from material balance calculations

Date Mass RCu RNi RFe RPt RPd RAu Wt% % % % % % %

5 Oct 4,28 90,7 65,5 30,5 58,0 70,5 50,1 6 Oct 4,56 90,0 71,6 23,6 69,1 56,2 40,4

10 Oct 4,53 89,9 70,2 28,0 56,2 88,4 44,8 11 Oct 3,81 89,7 70,0 22,9 58,2 80,4 45,8 12 Oct 3,67 89,9 64,4 23,2 55,9 43,3 57,0 23 Nov 5,53 92,9 64,2 39,3 59,4 88,0 81,6 24 Nov 4,93 92,5 70,9 38,4 56,7 79,1 99,4 30 Nov 7,69 93,0 74,5 46,8 75,4 96,0 67,9 1 Dec 8,77 95,2 82,8 46,7 73,1 96,0 66,4 24 Jan 3,01 89,9 65,5 21,0 54,8 51,7 37,5 1 Feb 3,28 89,2 68,9 24,3 58,2 50,6 34,7 7 Feb 2,74 88,1 65,9 20,2 55,1 75,9 40,6 16 Feb 2,64 86,2 57,2 19,5 42,9 32,3 50,0

Note: 30 Nov and 1 Dec are bulk concentrates

0

10

20

30

40

50

60

70

80

90

100

84 86 88 90 92 94 96 98

R Cu(%)

R A

u,Pd

,Pt (

%)

RAu%

RPd%

RPt %

R Au

R Pd

R Pt

Figure 18. Recoveries of Au, Pt and Pd vs. total recovery of Cu


0

10

20

30

40

50

60

70

80

90

100

50 55 60 65 70 75 80 85R Ni (%)

R A

u,Pd

,Pt (

%)

RAu%

RPd%

RPt %

R Au

R Pd

R Pt

Figure 19. Recoveries of Au, Pt and Pd vs. total recovery of Ni

0

10

20

30

40

50

60

70

80

90

100

10 15 20 25 30 35 40 45 50

R Fe (%)

R A

u,Pd

,Pt (

%)

RAu%

RPd%

RPt %

R Au

R Pd

R Pt

Figure 20. Recoveries of Au, Pt and Pd vs. total recovery of Fe


50

55

60

65

70

75

80

85

90

15 20 25 30 35 40 45 50

R Fe (%)

R N

i (%

)

R Ni

Figure 21. Recovery of Ni vs. recovery of Fe

0

10

20

30

40

50

60

70

80

90

85 86 87 88 89 90 91 92 93 94 95 96

R Cu (%)

R N

i,Fe

(%)

RNi%

RFe%

R Ni

R Fe

Figure 22. Recoveries of Ni and Fe vs. recovery of Cu


7.6 Flotation Times The flow rates of process streams and flotation times were calculated for three selective flota-tion days and one bulk flotation day. These are presented in Appendix 14. The residence times of flotation stages are also presented in Tables 16 and 17. The corresponding times in labora-tory tests are given and the ratios between pilot and laboratory times were calculated. In bulk flotation the residence times in pilot plant were quite moderate as shown in Table 16. It is obvious that the normal scale-up factor 3 would be usable in this case (ratio between pilot and laboratory times). Table 16. Flotation times in bulk flotation

Lab tests Factor1 Dec 2006 Pilot / Lab

min minBulk Roughg 71,0 25,0 2,8Bulk Clng 1 56,1 13,0 4,3Bulk Clng 2 8,1 11,0 0,7

Flotation times in pilot plant

In selective flotation the times in pilot circuit tended to be significantly longer than on bench scale as seen in Table 17. Only in copper rougher flotation the time in pilot plant was short be-ing only slightly longer than in laboratory. In nickel circuit rougher/scavenger flotation was 5 times longer than in laboratory. The cleaning stages were significantly longer in pilot plant compared to bench scale. Probably the main reason is that there were not small enough cells to be used as cleaners with moderate residence times. The available cell sizes 2x150 litre and 2x50 litre units had excessively vol-ume for small flows in cleaning circuits. Each cleaning stage had a unit of two cells because the use of one cell only would easily cause bypass flows (short circuiting). Table 17. Flotation times in selective flotation

Lab tests Factor24 Jan 2007 1 Feb 2007 13 Feb 2007 Pilot / Lab

min min min minCu-Roughg 8,7 9,8 5,5 1,6 - 1,8Cu-Clng 1 20,9 48,2 4,0 5 - 12Cu-Clng 2 27,6 70,8 3,0 9 - 24Cu-Clng 3 11,8 27,2Cu-Clng 4 31,7 37,2

Ni-Rough+Scav 133,5 125,7 132,0 25,0 5Ni-Clng 1 43,8 69,3 47,7 5,0 9 - 14Ni-Clng 2 154,1 154,0 129,7 4,0 32 - 38Ni-Clng 3 98,8 76,1 74,0 3,0 25 - 33Ni-Clng 4 49,3 30,9 34,7Ni-Clng 5 67,1 49,4 67,3

Tailings Flot. 10,4 9,7

Flotation times in pilot plant


7.7 Concentrates Produced During the pilot run the concentrates were collected either directly to 200 litre plastic drums or they were thickened in two Supaflo thickeners (diameter 1 m) from which the underflows were collected to drums. Collection of products was done by keeping the products of each running period (usually each day) separate. The concentrates of each pilot phase were filtered afterwards with Ceramec capillary disc filter or Larox pressure filter. The filter cakes of each period (one or two days) were sampled for analysis during collection to plastic drums. The weights, moistures and grade levels of col-lected concentrates are presented as a summary in Table 18. The day by day figures for copper, nickel and bulk concentrate can be found in Appendix 15/1-3. More exact moisture numbers are presented in chapter 9 (dewatering tests). Chemical analyses including the main elements, precious metals and XRF assays are presented for bulk concentrates in Appendix 15/3 and for some copper and nickel concentrates in Appen-dix 15/4. Table 18. Produced concentrates in pilot tests Product Gross weight

of filter cake Moisture Grade range

kg % Copper concentrate 6272 6,8 23,6 – 32,3 % Cu Nickel concentrate 10391 7,7 5,8 – 14,5 % Ni Bulk concentrate 1732 6,9 3,4-4,0 % Ni

5,0-7,1 % Cu

The size distributions of copper and nickel concentrates from 1st of February, 2007, are shown in Table 19. It is noteworthy that although the concentrates are rather fine by particle size their moisture in filter cake is low. Table 19. Size distributions of concentrates (Feb 1, 2007)

Screen Cu-concentrate Ni-concentrate µm % passing % passing

90 99,8 99,8 75 99,5 99,6 45 93,6 94,5 32 85,1 86,1 20 68,1 65,3


7.8 Samples Outside Samples were also collected during the pilot run for tests outside GTK. These samples are de-scribed in Table 20 and in more detail in Appendix 16. The ore type samples for grinding tests were rock samples collected after the primary jaw crushing in the pilot plant. The rocks were divided to two ore types, olivine pyroxenite and metaperidotite, by John Pedersen / Kevitsa Mining. The drill core samples were received from GTK Rovaniemi and were sent as such to test work. The bulk concentrates and small samples of nickel concentrate were shipped as filter cakes containing 6-7 % moisture. Flotation tailings were sent as slurry samples as agreed with the receiver. Samples for geo-chemical tests comprised of two samples, high sulphur and low sulphur tailings. Tailings sam-ple for geomechanical tests represented combined tailings. The assays of the shipped bulk concentrates and nickel concentrates are given in Appendices 15/3-4 and the flotation tailings in Appendices 16/2-3. Table 20. Samples sent for outside tests Sample Description Sample Purpose size Ore type samples Rocks 100-200 mm 464 kg Grinding tests Drill cores 440 kg Bulk concentrate Filter cake 1612 kg Metallurgical tests Nickel concentrate Filter cake 1.2-1.5 kg

(3 samples) Metallurgical tests

Flotation tailings Slurry samples 4 x 15 L Geochemical tests (High and low sulphur) Flotation tailings Slurry samples 20 & 50 L Geomechanical tests

7.9 Ion Concentrations in Tailings and Concentrates Concentrations of different ions in the water phase of process samples were analysed by ICP. The analyses of filtrate waters of concentrates and flotation tailings are presented in Appendix 17. The range of Ni- and Cu-analyses are summarised in Table 21. The main process conditions having an effect on the solubility of nickel and copper from min-eral surfaces were observed to be pH in nickel rougher flotation (acidic pH 5-6 vs. alkalic pH about 10) and the use of special chemicals, metabisulphite (MBS) and TETA in nickel clean-ings.


Table 21. Concentrations of Ni- and Cu-ions in water phase of concentrates and tailings Ni-Concentrate Ni-Tailings Cu-Concentrate Conditions Ni Cu Ni Cu Ni Cu mg/l mg/l mg/l mg/l mg/l mg/l

1 Alcalic Ni-flotation Without MBS+TETA 0-0,20 0-0,13 0-0,03 0-0,03

2 Alcalic Ni-flotation MBS+TETA in use 3,5-8,7 13,0-26,1 0,02-0,07 0,02-0,06 0,03-2,1 0-0,2

3 pH 5-6 in Ni-roughing MBS+TETA in use 0,5-2,2 0,5-26,9 1,5-9,5 0,04-11,8

The concentrations of Ni and Cu kept rather low in the filtrate water of copper concentrate al-though MBS was used in the final copper cleaning. Higher concentrations of nickel than cop-per could be observed (Table 21). pH in copper cleanings was close to 12. The analyses of Ni and Cu were very low in filtrate waters of nickel concentrate (<0,2 mg/l) and final tailings (<0,03 mg/l) when the circuit was run alkalic, and MBS and TETA were not used. These were the process conditions in phase 3 of the pilot run (Jan-Feb 2007) and during some periods in pilot phases 1 and 2 (2006). The metal concentrations in flotation tailings kept also low when MBS and TETA were used, but the whole nickel circuit was alkalic. The effect of MBS+TETA was, however, clearly seen in the filtrate water of nickel concentrate in which high levels of Ni and Cu were analysed. It is thus obvious that MBS and TETA caused the dissolution of metals from mineral surfaces. When nickel rougher flotation was run at acidic pH, and MBS and TETA were used there were significantly higher concentrations of Ni and Cu in final tailings also, probably because ions dissolved by chemicals were not precipitated at acidic conditions.


8 ESTIMATE FOR AVERAGE ORE The results of pilot tests were used to estimate the nickel recoveries of selective flotation for an average Kevitsa main ore containing 0,30 % Ni. The content of nickel in silicates was set to 0,0275 % or 275 ppm as agreed with the client (note that the analysis of 20 drill core samples selected from the Kevitsa ore was 237 ppm). Silicate Ni (275 ppm) represents 9,2 % of total nickel in the average ore which is less than in the feed ores of pilot tests (12,1-17,4 % of total Ni, Table 13). The recoveries of sulphide nickel in pilot plant were also used in the estimations for average ore. This can be justified in two ways. First, the content of sulphide nickel in the average ore (0,273 %) does not differ that much from the sulphide nickel in the feed of pilot plant (0,351; 0,361 and 0,299 % in phases 1, 2 and 3, respectively). Secondly, it was seen in laboratory tests that the feed grade variation of sulphide Ni did not have any distinguishable effect on the flota-tion recovery of sulphide nickel. The recovery of sulphide nickel (about 76 to 79 %) in pilot phases 1, 2 and 3 was shown in Table 14. The recovery estimates for average ore are presented in Table 22. It is seen that with the nickel concentrate grades of pilot tests, 11,0; 7,3 and 12,2 % Ni, the estimated recoveries are 68,9; 71,2 and 71,4 %. The difference between the estimate and the pilot plant recovery is highest for pilot phase 3; 6,1 %. The content of silicate nickel was highest in the feed of pilot phase 3 and for that reason there is the highest potential for improvement when compared to the average ore. Table 22. Recovery estimates for average ore (0,30 % Ni, 275 ppm silicate Ni)

Ni-Conc Estimate Pilot Differ- Pilot & Flotation Tailings Reference grade Aver.ore plant ence Aver.ore Estimate for average ore Pilot phase Ni tot Ni tot Ni tot Ni tot Ni sulf Ni tot Ni sulf Ni sil % Rec % Rec % Rec % Rec % % % % Phase 1 (Sept-Oct ’06) 11,0 68,9 66,8 2,1 76,1 0,089 0,060 0,029 Phase 2 (Nov ’06) 7,3 71,2 69,2 2,0 78,7 0,080 0,051 0,029 Phase 3 (Jan-Feb ’07) 12,2 71,4 65,3 6,1 78,8 0,081 0,053 0,028

The results of Table 22 are presented in more detail in Appendix 18/1-2. It is noted that in ad-dition to recoveries shown in Table 22 part of nickel reports to copper concentrate to which the nickel recoveries were 2,4; 3,3 and 2,3 % in pilot phases 1, 2 and 3. The flotation of nickel was most successful in pilot phase 3 despite of the highest content of nonfloatable nickel in the feed ore. The daily results of pilot phase 3 and estimates for average ore are shown in Figure 23 and are also presented in Appendix 18/1. The regression line of total Ni in Figure 23 gives the grade / recovery numbers shown in Table 23. For the average Kevitsa main ore, when the grade increases from 10 to16 % Ni in nickel concentrate, the estimated recovery of total Ni changes from 72,0 to 70,3 % and the recovery of sulphide Ni from 79,5 to 77,7 %. (For sulphide Ni the recovery numbers are those achieved in pilot tests, picked from the regression line of sulphide Ni here).


Table 23. Estimated results of nickel flotation for average Kevitsa main ore

Total Ni Sulphide Ni Ni-Concentrate Recovery Recovery

% Ni % %

10 72,0 79,5 12 71,5 78,9 14 70,9 78,3 16 70,3 77,7

Kevitsa Pilot and Estimates for Average Ore / Ni-Flotation

10,77; 72,612,10; 73,2

11,40; 70,9

12,52; 69,4

15,35; 70,817,30; 69,910,50; 71,1

9,39; 71,8

10,12; 73,0

y = -0,2838x + 74,859

60

70

80

8 10 12 14 16 18 20Ni, %

Ni-R

ecov

ery,

%

Pilot results Phase 3

Estimate based on Pilot phase 3

Linear (Pilot results Phase 3)

Linear (Estimate based on Pilot phase 3)

Estimates for average ore (0,3 % Ni and 0,0275 % Ni in silicates)

Figure 23. Estimate of nickel flotation for average Kevitsa main ore on the basis of results from pilot phase 3. The results and estimate of Figure 23 are repeated in Figure 24 in which also theoretical maxi-mum recoveries are plotted. This maximum curve is shown for pilot phase 3 feed (0,35 % Ni, 600 ppm in silicates) and for average Kevitsa ore (0,295 % Ni, 275 ppm in silicates). The theo-retical recoveries were calculated for two concentrates: 1) Ni-concentrate consists only of pent-landite and all pentlandite is recovered to it, and 2) All pentlandite and pyrrhotite is recovered to nickel concentrate. It is assumed that no other minerals come to concentrate. Nickel losses are caused by nickel in pyrrhotite and silicates in case 1) and only silicates in case 2). Figure 24 shows that theoretically there is a lot of room for improvement in nickel recovery, about 20 % from the estimated figures to theoretical maximum (which is about 90 %).


Kevitsa Pilot, Estimate for Average Ore and Theoretical Curves / Ni-Flotation

0,295; 100,0

31,6; 89,27,9; 90,7

8,6; 82,9

0,35; 100,0

31,6; 81,7

0

10

20

30

40

50

60

70

80

90

100

0 5 10 15 20 25 30 35Ni, %

Ni-R

ecov

ery,

%

Pilot results Phase 3

Estimate based on Pilotphase 3

Theoretical: Kevitsa ore0,295 % Ni, 275 ppm insilicates

Theoretical: Pilot feed0,35 % Ni, 600 ppm insilicates

Linear (Pilot resultsPhase 3)

Linear (Estimate basedon Pilot phase 3)

Feed

Pentlandite in NiCon

Pentl+Pyrrh in NiCon

Figure 24. Results of pilot phase 3, estimate for average ore and theoretical maximum recover-ies. 9 DEWATERING TESTS ON CONCENTRATES Dewatering tests comprised of thickening and filtration tests for copper and nickel concentrates produced in pilot plant. These tests were done with bench scale equipment after the pilot runs. A separate report is attached as Appendix 19 from which the essential results are repeated here. Concentrate samples The samples for the settling and thickening tests were from pilot run on Feb 1, 2007 with the particle size distributions given earlier in Table 19 and the following analyses: Copper concentrate (CuCC4): 29,87 % Cu, 0,53 % Ni, 31,2 % Fe

Nickel concentrate (NiCC5): 1,89 % Cu, 11,8 % Ni, 46,7 % Fe

The samples for filtration tests were from Feb 14 (copper) and Feb 8, 2007 (nickel): Copper concentrate (CuCC4): 29,8 % Cu, 0,53 % Ni, 30,8 % Fe

Nickel concentrate (NiCC5): 1,64 % Cu, 9,6 % Ni, 46,2 % Fe


According to size distributions of Table 19 the concentrates were very fine: Cu-concentrate 68 % -20 µm and Ni-concentrate 65 % -20 µm. Methods The methods presented in the following table were used. Table 24. Methods used in dewatering tests of concentrates Test type Equipment Test variables Settling test Graduated glass cylinder

of 2 litres Feed pulp densities 20, 30, 40 % Flocculant dosages 0 and 6 g/t

High rate thickening test Supaflo bench scale high

rate thickener (d 94 mm) Feed pulp densities 15, 20, 30 % (solids feed flow constant) Flocculant dosages 5-28 g/t

Capillary vacuum filtra-tion

Ceramec capillary filter plate (2 dm2)

Feed pulp densities 49 and 69 % Filtration and drying times

Pressure filtration Larox PF 0.1 (10 dm2) Feed pulp densities 75, 65, 55 %

Pumping times 10, 30, 60 sec

Results The tables of test results from Appendix 19 are repeated here. The flocculant in the tests was Fennopol N200. Table 25. Results of settling tests

Cu-Concentrate Ni-Concentrate

Flocculant addition g/t - - - 6g/t 6g/t 6g/t - - - 6g/t 6g/t 6g/t

Feed pulp density % 20,0 30,0 40,0 20,0 30,0 40,0 20,0 30,0 40,0 20,0 30,0 40,0

U/F pulp density % 65,0 68,0 69,0 63,3 67,5 69,4 63,3 72,1 73,0 71,5 72,1 72,8

Settling Time ts min 10,7 15,7 22,0 8,0 13,0 18,3 8,0 12,3 22,0 7,6 11,7 18,3

Specific Settling Area m2/t/h 2,83 2,46 2,42 1,79 1,88 2,02 2,14 1,94 2,50 2,02 1,87 2,05


Table 26. Results of high rate thickening

Cu-Concentrate Flocculant addition g/t 5,0 15,6 25,4 4,4 14,0 22,2 3,9 12,2 20,7

Feed pulp density % 30,1 29,1 27,1 22,1 21,1 21,1 19,1 18,1 17,1

U/F pulp density % 58,4 66,3 68,1 60,7 64,7 70,1 62,5 64,9 65,0

Rise Rate m/h 3,67 3,74 3,93 6,18 6,16 6,23 8,51 8,50 8,60

Solids Loading t/m2/h 1,46 1,39 1,35 1,66 1,54 1,56 1,86 1,77 1,67

Specific Settling Area m2/t/h 0,69 0,72 0,74 0,60 0,65 0,64 0,54 0,56 0,60

Ni-Concentrate Flocculant addition g/t 5,4 16,0 25,7 5,1 14,7 21,7 4,9 13,9 28,1

Feed pulp density % 28,1 27,1 27,1 21,1 20,1 21,1 15,1 16,1 13,1

U/F pulp density % 67,9 68,2 68,1 67,3 68,2 68,1 65,5 67,2 68,3

Rise Rate m/h 3,80 3,91 3,98 5,74 6,33 6,31 8,51 8,71 8,65

Solids Loading t/m2/h 1,34 1,35 1,35 1,44 1,47 1,59 1,47 1,55 1,23

Specific Settling Area m2/t/h 0,75 0,74 0,74 0,70 0,68 0,63 0,68 0,65 0,81

Table 27. Ceramic capillary filtration test results

Cu-Concentrate Ni-Concentrate Feed Capacity Moist. Feed Capacity Moist. (g/l) kg/m2/h (%) (g/l) kg/m2/h (%)

2100 731 11,4 2100 882 14,5

1600 299 12,4 1600 280 15,8

Table 28. Pressure filtration test results

Cu-Concentrate Ni-Concentrate Feed Capacity Moist. Feed Capacity Moist. (g/l) kg/m2/h (%) (g/l) kg/m2/h (%)

2340 635 9,7 2340 623 10,2

2000 656 7,6 2000 727 6,9

1700 574 7,0 1700 682 7,2


10 CONCLUSIONS Ore Sample The pilot plant test work was carried out for Kevitsa ore in which the grades were 0,40 % Ni and about 0,52 % Cu during phases 1 and 2 (2006) and a bit lower, 0,36 % Ni and 0,45 % Cu in phase 3 (2007). The feed grades kept rather steady from day to day thus indicating the ho-mogenous nature of the ore sample. The homogenisation stockpile of pilot plant also helped in keeping the feed grades invariable. The ore represents the upper parts of Kevitsa deposit as it was taken from a test pit after re-moval of overburden and oxidised upper layer of ore. The top layers probably always have been subject to some oxidation although the ore samples were selected so that there would not occur any clear weathering. Oxidised surfaces could be, however, seen in some rocks. It is ob-vious that at least ore sample no. 3 which had remained at the bottom of heap in the pit, and was the last one shipped to GTK, contained some oxidised material. In mineralogical study a small part of copper and nickel was found to occur in goethite in this sample. The nickel grade of pilot sample, 0,40 % and 0,36 %, was higher than the target, 0,30 %, the expected average grade of Kevitsa main ore. Part of nickel in the ore occur in silicates and it pointed out that in the feed of pilot tests Ni in silicates was 460-605 ppm (12-17 % of total Ni) which is higher than the average for Kevitsa main ore, about 275 ppm (237 ppm or 8,2 % of total Ni was analysed for 20 drill core samples). So, in the pilot feed there was nonfloatable Ni in silicates and some of it also in goethite. As the feed grade of pilot tests was higher than the average for Kevitsa ore it was suspected that nickel recovery in pilot tests would be higher than what it would be for average ore. After phase analysis of nickel it was, however, seen that the content of floatable, or sulphide, nickel was not much higher in pilot feed compared to average ore. Especially in pilot phase 3 the grade of sulphide nickel was only slightly better than in the average ore. A comparison is pre-sented in Table 29. Table 29. Nickel phases in the feed of pilot plant and Kevitsa average ore

Ni total Ni sulphide Ni silicate % % % Pilot phase 1 0,398 0,351 0,047 Pilot phase 2 0,408 0,361 0,047 Pilot phase 3 0,359 0,299 0,060 Average ore 0,300 0,273 0,0275

The high content of silicate nickel in the feed of pilot plant caused the inevitable drop of recov-ery in nickel flotation. Another factor decreasing the recovery was that there was at least some degree of oxidation in the ore sample. Nickel in oxidised form may origin from two sources; from oxidised minerals containing nickel, like goethite, and secondly from oxidised surfaces of sulphide mineral, pentlandite. It should also be noted that in phase 3 of the pilot campaign the ore sample was more difficult to process than in the earlier phases of pilot run; the ore was more oxidised and the content of silicate nickel was highest in phase 3. In addition, it seemed that during phase 3 the ore


changed more difficult towards the end of the test work (possibly containing more oxidised ore). Copper Flotation In the selective flow sheet copper flotation succeeded in pilot plant much better than what was expected on the basis of laboratory and minipilot test work. High grade copper concentrate, generally about 30 % Cu, was produced with copper recovery of 77-80 %. Nickel content in copper concentrate could be dropped to level of 0,5 – 0,6 % Ni. Nickel losses to this product were small, 1,5-2,5 %. The main targets of copper flotation, to produce a smelter grade concen-trate and to remove most of copper from the feed of nickel flotation were thus fulfilled. The recovery of gold to copper concentrate was 38-39 % during pilot phases 1 and 2, but dropped to 25 % in phase 3. Possibly strong depressing actions for pentlandite, by minimising the use of collector for instance, caused the depression of gold in copper circuit in pilot phase 3. Another factor was that smaller amount of copper concentrate was floated when the copper grade of feed was lower in pilot phase 3. Small amounts of platinum and palladium also re-ported to copper concentrate with recoveries of about 10 % and 13 %, respectively. Copper flotation was operated in a closed circuit which made it possible to have a high recov-ery of copper while, at the same time, pentlandite and other minerals were depressed effec-tively by applying a 4-stage cleaning, very low dosages of reagents (especially collector), high pH levels and sulphite as a depressant in the last cleaning. The depressant was originally so-dium metabisulphite (MBS, about 100 g/t), but it was later replaced by sodium sulphite (finally about 70 g/t). The total consumption of dithiophosphinate collector, AP 3418A, could be de-creased to 7-8 g/t towards the end of pilot campaign. The dosage level of CMC depressant was increased to about 115 g/t during the last days. Nickel Flotation Nickel flotation was found to be very challenging, especially in terms of nickel recovery. The highest recoveries, 70-71 %, were achieved in phase 2, but the grade was low, 6-7 % Ni. With grade levels of 8-12 % Ni the best recoveries were 67-69 %, just occasionally about 70 %. The higher grades, 12-18 % Ni, were achieved with nickel recoveries 64-67 %. In phase 3 of pilot tests the grades of concentrate were 9-18 % Ni and the recoveries 64-67 % being almost independent of grade. The target, or expectation, for nickel recovery had been 70 % or more. There were a couple of reasons for nickel losses in tailings. The first was the high amount of nonfloatable nickel, mostly in silicates and partly in oxidised form in goethite. The analysed content of nickel in silicates in the feed of pilot plant was 460-605 ppm or 12-17 % of total nickel. This is much higher than the expected average in Kevitsa main ore, 275 ppm (or 9,2 % of total grade 0,3 % Ni). It was also found that the amount of silicate Ni increased during the pilot campaign as it was 460 ppm in phases 1 and 2, and 605 ppm in phase 3 (the assays of pi-lot feed samples). Another reason for nickel losses was that part of free liberated pentlandite was lost to tailings. This may have been caused by passivation of pentlandite surfaces due to oxidation during flo-tation. The recovery of sulphide nickel was 76 % in pilot phase 1 and 79 % in phases 2 and 3


which can be regarded as satisfactory numbers although there remains a lot of room for im-provement. The mineralogical analysis of pilot tailings on February 1, 2007, revealed that total nickel in tailings, 0,123 %, was distributed between the following phases: silicates 0,071 (58,3 % of total Ni in tailings), pentlandite 0,035 (28,2 %) and goethite 0,017 (13,5 %). These numbers give a recovery of 66 % for total nickel in Ni-concentrate. The recovery of sulphide Ni (pentlandite) is 86 %. (These are numbers from the mineralogical analysis. The recoveries of sulphide nickel presented earlier in Table 14, 76-79 %, were based on nickel phase analysis in which the total nickel is divided only to two phases, sulphide and silicate nickel. In this analysis nickel in goe-thite is probably included in the numbers of both sulphide and silicate nickel). The quality of nickel concentrate improved in phase 3 as the content of iron increased and sili-cates dropped at the same time. The content of nickel was rather high, 9-18 %, the average Fe 44,3 %, MgO 2,7 % and ratio Fe/MgO 16,3. The recoveries of precious metals to nickel concentrate were in phase 3: Pt 58 %, Pd 43 % and Au 15 %. It was found that Pt and Pd were floating in the same ratio as Ni from the feed of cleanings to final concentrate. The overall recoveries of Pt and Pd to nickel concentrate were lower than the recovery of nickel because precious metals were floating more to copper con-centrate and more of them were lost to tailings than nickel. Gold was floating in nickel circuit with significantly lower recovery than Pt or Pd. Pyrrhotite was depressed rather strongly in nickel circuit so that the recovery of Fe in cleanings was much less than the recoveries of Pt and Pt. It seems that flotation of precious metals has much stronger correlation with Ni than with Fe. The quality of nickel concentrate could be improved in phase 3 without the use of special chemicals, TETA and MBS. In addition, sulphuric acid and copper sulphate was not used and generally the consumption of chemicals could be decreased. In order to improve the quality of nickel concentrate the following measures were taken in phase 3: the use of more selective collector (ethyl xanthate) and reduced dosage levels of col-lector and other chemicals (frother and CMC), alkalic circuit, increased number of cleanings (5), and regrinding in cleaning circuit. The purpose was to prevent excessive flotation of pyr-rhotite and silicates while keeping the nickel recovery as high as possible at the same time. The idea was to get silicates depressed in two first cleanings and to depress pyrrhotite at high pH in three last cleanings after regrinding. The meaning of regrinding seemed to be in creating the reducing pulp environment (by iron from grinding) which helped to depress pyrrhotite. It was concluded that there was not a question of mineral liberation in this case. The product of regrinding was very fine which seemed not to deteriorate the selectivity of flotation in subse-quent cleaning stages. In nickel rougher circuit almost all of valuables were already floating in the first rougher cell (of total of four cells). This was the situation especially in pilot phase 3. The circulation of nickel and iron was rather high in nickel circuit which might have enhanced the fast flotation in roughing. Although most of valuables already floated in the first rougher cell, three more cells were needed to float more of the recoverable material to improve the recoveries.


The operation of nickel circuit was successful only with circulations, although the circulation of mass was not high. Operation in open circuit, in which the 1st cleaner tailings were taken out, was not successful in pilot plant. The retention times in all stages of nickel circuit were long compared to bench scale tests (longer than bench scale times multiplied by the normal scale-up factor of 3). The long reten-tion times in cleanings can be partly explained by available cell sizes; the cells were rather large for small pulp flows in cleaning circuit. TETA and MBS were used to depress pyrrhotite in pilot tests during phases 1 and 2. It was seen that a good selectivity could be achieved between pentlandite and pyrrhotite when these chemicals were used. Generally pyrrhotite was depressed so effectively that iron content of concentrate remained low. The amount of silicates was high at the same time and therefore, the ratio Fe/MgO was too low. The depression of pyrrhotite was achieved without regrinding when TETA was used. In phases 1 and 2 (autumn 2006) some days were run without TETA, but those tests yielded too low nickel grades in concentrate. The reagent dosages (especially collector) were obviously too high and the circuit configuration was not optimal for running without special chemicals. The drawback of using TETA was that it dissolved metals like nickel and copper from mineral surfaces which was seen as high concentrations of these elements in water phase of tailings and nickel concentrate. This might constitute an environmental problem to be taken care of. TETA is also rather expensive reagent. Its use in flotation may be still protected by a patent of INCO. Bulk Flotation In bulk flotation the purpose was to produce a concentrate with high metal recoveries for downstream treatment by leaching. The bulk tests were run at the end of phase 2 from Novem-ber 28 to December 1, 2006. In tests at low pH (5,5-4,5) the recoveries of nickel were not much higher than in selective flo-tation. The tailings grade did not decrease below 0,1 % Ni. The recoveries improved substan-tially when alkalic conditions were applied during the last day of bulk flotation. It is possible that surface oxidation of pentlandite had taken place in acidic conditions and could be pre-vented by adding lime starting from the rod mill and using high pH in flotation. The tailings grade was 0,069 % Ni and in mineralogical study no pentlandite was detected in tailings. The content of nickel in tailings thus indicated the level of nonfloatable Ni. The bulk flotation in alkalic conditions yielded high recoveries for all metals including pre-cious metals. Correspondingly, the grades were lower than in selective flotation, because the mass pull to concentrate was high, 9,3 wt-%. The metal recoveries were: Cu 94,9 %; Ni 84,4 %; Pt 77,6 %; Pd 90,0 %; Au 77,0 %. The grades were 6,06 % Cu and 3,65 % Ni. With a high mass pull, there were also rather much silicates in the product as MgO was 12,9 %.

metallurgical testwork on kevitsa ore in pilot scale

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