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MURR/MCDA Ch 5 Mineral Processing-final 25 October 2010

Table of Contents

5 Mineral Processing Review ........................................................................................ 5-1 5.1 Introduction ....................................................................................................................... 5-1

5.2 Ore Sources and Mineralogy ............................................................................................ 5-1

5.3 Previous Operations ......................................................................................................... 5-2

5.4 Metallurgical Investigations ............................................................................................... 5-2 5.4.1 Bearden Potter Phase 1 ..................................................................................................... 5-3

5.4.2 Bearden Potter Phase 2 ..................................................................................................... 5-3

5.4.3 Foskor ................................................................................................................................ 5-7

5.4.4 Resource Development Inc. ............................................................................................... 5-9

5.4.5 IMMT ............................................................................................................................... 5-10

5.5 Mineral Processing Facilities .......................................................................................... 5-13 5.5.1 Phase 1 ........................................................................................................................... 5-13

5.5.2 Phase 2 ........................................................................................................................... 5-15

5.5.3 Review of the IMMT Flow Sheet ...................................................................................... 5-15

5.5.4 Further Issues .................................................................................................................. 5-21

5.6 Projected Production Performance ................................................................................. 5-22

5.7 Operating Costs .............................................................................................................. 5-24 5.7.1 Phase 1 ........................................................................................................................... 5-24

5.7.2 Phase 2 ........................................................................................................................... 5-27

5.8 Capital Costs .................................................................................................................. 5-28 5.8.1 Phase 1 ........................................................................................................................... 5-28

5.8.2 Phase 2 ........................................................................................................................... 5-31

5.9 Conclusions and Recommendations for Further Feasibility Study Investigations .......... 5-31 5.9.1 Metallurgical Characterization of the Deposits ................................................................. 5-31

5.9.2 Confirmation of IMMT Flowsheet at Pilot Scale ................................................................ 5-32

5.9.3 Design Allowance for Process Variability ......................................................................... 5-34

5.9.4 Review Concentrate Production Strategy ........................................................................ 5-35

5.9.5 Review Fertilizer Production Strategy .............................................................................. 5-35

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List of Tables Table 5.1: TICAF Plant Results ................................................................................................ 5-2

Table 5.2: Bearden and Potter Phase II Material Balance ........................................................ 5-6

Table 5.3: Foskor Pilot Plant Results – South Valley ............................................................... 5-7

Table 5.4: Foskor Pilot Plant Results – West Valley ................................................................. 5-8

Table 5.5: Size Grade Analyses ............................................................................................. 5-11

Table 5.6: Feed, Product and Waste Size Distribution ........................................................... 5-19

Table 5.7: Projected Annual Production ................................................................................. 5-23

Table 5.8: Phase 1 implied volumes of SSP from different components ................................ 5-23

Table 5.9: Waste Products and Size Distribution ................................................................... 5-24

Table 5.10: Opex Estimate of 50tph Apatite Beneficiation Plant .............................................. 5-24

Table 5.11: Opex Estimate of 250tpd Sulphuric Acid Plant (DRAMP) ...................................... 5-25

Table 5.12: Opex Estimate of 100tpd Phosphoric Acid Plant ................................................... 5-26

Table 5.13: Opex Estimate of 200tpd SSP and 150tpd TSP Fertilizer Plant ............................ 5-26

Table 5.14: Opex Estimate of one 250tph Apatite Beneficiation Plant ..................................... 5-27

Table 5.15: Capex Estimate of 50tph Apatite Beneficiation Plant ............................................ 5-28

Table 5.16: Capex Estimate of 250tpd Sulphuric Acid Plant .................................................... 5-29

Table 5.17: Capex Estimate of 100tpd Phosphoric Acid Plant ................................................. 5-30

Table 5.18: Capex Estimate of 200tpd SSP or 150tpd TSP Fertilizer Plant ............................. 5-30

Table 5.19: Capex Estimate of one 250tph Apatite Beneficiation Plant ................................... 5-31

List of Figures Figure 5.1: Schematic Flow Sheet - Bearden Potter ................................................................ 5-5

Figure 5.2: Schematic Flow Sheet - IMMT ............................................................................. 5-12

Figure 5.3: Phase 1 Scheme of Arrangement ........................................................................ 5-14

Figure 5.4: IMMT Flow sheet .................................................................................................. 5-17

Figure 5.5: Conceptual plant layout ........................................................................................ 5-18

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5 Mineral Processing Review 5.1 Introduction

This section of the report reviews the metallurgical and mineral processing aspects of the project undertaken to date. It also presents independent capital and operating cost estimates prepared by Dowding Reynard and Associates (“DRA”) of the process route developed by the Indian Institute of Minerals and Materials Technology (“IMMT”) on behalf of Nilefos during earlier investigations. Further work and investigations recommended for completion during the Feasibility Study are also outlined.

5.2 Ore Sources and Mineralogy The Sukulu Carbonatite Complex consists of five major zones: • Central carbonatite series; • Incomplete ring of magnetite-apatite-mica rock; • Ring of feldspathoidal mixed rocks; • Ferritized granite; and • Granitic-gneissic country rocks. In the central carbonatite series, sovite is the principal carbonatite. Lenses of clear crystalline calcite are common. Carbonate minerals form the majority of the carbonatite with accessory amounts of apatite, mica, magnetite, pyroxene, tremolite, garnet, quartz, hematite, barite and pyrochlore. Rare occurrences of the minerals perovoskite, zircon, anatase, baddeleyite, pyrrhotite, pyrite, chalcopyrite and galena have been noted. Local patches of pyrolusite are known. Through the process of physical and chemical weathering, carbonatites comprised primarily of calcium carbonate slowly dissolve. Certain accessory minerals such as apatite, crandallite, magnetite, rutile and pyrochlore are resistant to weathering and remain to form a residual soil along with other decomposition products such as clays and iron hydroxides. Through the process of weathering, the hard rock source containing around 3.5% P2O5 is naturally upgraded to 10-15% P2O5 in the residual soil, contributing to the economic viability of Sukulu. At Sukulu such residual soils have accummulated in three valleys, the North, South and West Valleys. These will be mined by open cast mining techniques and fed to the concentrator. Based on field observations, the residual soils contain about 30% apatite, 25-30% granular magnetite, 30% fine grained goethite and 7% quartz as well as minor amounts of ilmenite, zircon, pyrochlore and baddeleyite which are variable across the deposit. Secondary apatite and mica have been reported at depth. There are anomalous CaO / P2O5 variations with depth which have been ascribed to the presence of an aluminium phosphate, crandallite.



The ore is described as a soft soil which is easily compacted and becomes quite sticky when wet. This is an important observation considering the two rainy seasons per year. Bedrock is the original unweathered carbonatite, low in P2O5. Some pebbles and boulders of this hard material can be expected near the bedrock.

5.3 Previous Operations Tororo Industrial Chemicals and Fertilizer Ltd (“TICAF”) previously operated a mine, concentrator and single super phosphate (“SSP”) fertilizer facility from 1962 to 1978. Ore was mined with front end loaders, loaded onto trucks and stockpiled at the remote mine plant. This plant scrubbed, partially dewatered and sized the ore and partially removed the magnetics from the ore. The product flowed downhill via pipeline to the flotation plant. Flotation feed passed through another magnetic separation step, was ground to -105 micron and concentrated by flotation. The apatite product was cycloned directly as a wet feed to the fertilizer plant. Over its life approximately 160 000 tonnes of apatite concentrate (40-41% P2O5) was produced from 2.16Mt of ore with a feed grade of around 12% P2O5. The capacity of the previous beneficiation plant was 2.5tph of apatite. The production capacity of the fertilizer facility was about 6 000tpa P2O5 as bagged, granular SSP for domestic consumption and export to Kenya and Tanzania. Average plant results achieved are summarised in Table 5.1.

Table 5.1: TICAF Plant Results

Weight % P2O5 Fe2O3

Stream Analysis %P2O5

Distribution %

Analysis % Fe2O3

Distribution %

Feed 100.0 13.0 100.0 32.0 100.0 Magnetic Reject 27.0 2.4 5.0 79.6 67.2 Slime Reject 35.0 10.4 28.0 29.0 31.7 Flotation Feed 38.0 22.6 66.1 0.9 1.1 Flotation Concentrate 19.3 41.3 61.3 0.3 0.2 Flotation Tailings 18.7 4.0 5.8 1.6 0.9

Operational difficulties were experienced with typical P2O5 recoveries of 25-35% being reported by TICAF. The TICAF definition of ‘recovery’ was not reported. Much of the problem was reportedly caused by losses to mud balls discarded as oversize to waste, entrainment of apatite with magnetic waste and losses to slimes.

5.4 Metallurgical Investigations In pursuing the development of the phosphate reserves in Eastern Uganda, the Government of Uganda proposed a phosphate fertilizer project based on the exploitation of the Sukulu Hills (‘Sukulu’) phosphate deposit. The project was run through the parastatal corporation TICAF. In 1982 Bearden Potter were appointed to undertake a detailed engineering study of the project, which was completed in two phases. SRK reviewed the reports of both phases of study, the results of which are summarised here.

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Nilefos subsequently invited Foskor Ltd (“Foskor”) of South Africa to participate in the project. Foskor tested some Sukulu material through its pilot plant in Phalaborwa. SRK reviewed the Foskor report as well as commentary on the Foskor report by Madhvani International S.A. (“MISA”). In 2007, SRK (USA) undertook a preliminary assessment to validate original data ahead of a planned Feasibility Study. As part of this assessment, Resource Development Inc. (‘RDI’) undertook a programme of laboratory beneficiation studies. SRK reviewed the draft SRK (USA) report. More recently the IMMT of Bhubaneswar, India was commissioned by Nilefos to review the RDI flowsheet with a view to improving recoveries and to develop a commercial flow sheet for the beneficiation of Sukulu phosphate ore and for the production of phosphoric acid. SRK reviewed the IMMT reports, the results of which are summarised here.

5.4.1 Bearden Potter Phase 1 During August 1982, 15 holes were drilled to confirm the reliability of previous exploration. The samples were collected in 1.52m intervals although in some cases the lithology dictated separating the normal intervals into subintervals. The samples were subsequently composited and split into three subsamples for moisture determination, chemical analysis and preliminary beneficiation testing. Testwork was conducted on a single bulk sample to develop a conceptual process for the Phase 1 study. It was determined that basic beneficiation techniques for apatite recovery including liberation via scrubbing followed by upgrading with magnetic separation, desliming and flotation were dictated for Sukulu ore. Ten drill hole samples were also tested to develop metallurgical relationships for prospect data.

5.4.2 Bearden Potter Phase 2 During September 1983, 12 one meter diameter by up to 18.3m deep pits were excavated in the South Valley of the Sukulu Hills deposit. The key objective was to provide data to further substantiate the previous reserve estimate data and to generate samples for beneficiation and fertilizer manufacturing tests. Samples were collected in 1.52m intervals. Samples from the 1.52m intervals were also composited into 6.1m increments of about 10kg for processing. Thirty such samples were generated. A summary of key results on the composite samples is shown in the Appendix. It is seen that, whilst there is considerable variability, on average 59% of the sample mass was less than 325# or 44 micron. Pit SL3 was sampled to produce a 6.5 tonne sample. This sample was processed to create 560kg of apatite for fertilizer manufacturing tests plus 50kg for beneficiation optimization tests.



Detailed results of the beneficiation optimization tests were reported in the Bearden Potter Phase II Technical Report. The following is a summary of the results for each optimization test: • Raw Soil Scrubbing: Apatite was liberated from the raw soil (ore) with high attrition

scrubbing without grinding. Because of the large quantity of -44 micron fines, it was considered critical to include multiple stages of desliming with scrubbing. Based on bench scale testwork, scrubbing time could be as low as 5 minutes.

• Flotation Feed Scrubbing: Scrubbing of flotation feed to polish apatite did not appear necessary.

• Conditioning: Tests showed that conditioning at 75% solids for two to three minutes was optimum.

• Starch Depressant: Starch addition to either the conditioner or to the cleaner flotation cells at about 400-500g/t produced apatite concentrate at above 40%P2O5, less than 1%Fe2O3 and P2O5 flotation recoveries of over 95%. These results however, were contingent upon having a reasonably clean, magnetically separated flotation feed.

• Size Distribution/ Recovery: Sizing prior to flotation may improve the P2O5 recovery by one or two percent. Size versus recovery data showed excellent performance for all minus 48# (300 micron) feed.

• Collector Use Levels: Collector (fatty acid: fuel oil at 1:1) use levels of 400-500g/t of concentrate for each reagent seemed to be optimum. Sodium silicate was not required as a dispersant.

• Wet High Intensity Magnetic Separation: WHIMS tests showed that an apatite concentrate of 32% P2O5 at less than 1% Fe2O3 could be produced with 90% recovery. When amine flotation was used alone, it yielded a 38.3% P2O5 product with a 70% P2O5 recovery. The combined magnetic/ flotation P2O5 recovery (90%/70%) gave an overall recovery of 63%.

• Waste Fines Recovery: Using magnetic separation and flotation, a recovery from the coarser fraction (+15 micron) of the waste fines produced a 37.9% P2O5 apatite at 0.62% Fe2O3 and at 87% P2O5 recovery. Collector usage however, was nearly three times that at on the +44 micron size fraction.

• Fines Settling Tests: Settling rates of -44 micron waste were 5-9cm/h. Commercial flocculent increased the settling rate to 230-600cm/h.

Key observations were that apatite could be adequately liberated by scrubbing (without grinding) and that iron minerals could be depressed with starch. Beneficiation tests in both phases of testwork confirmed that more than half of the P2O5 content of the ore was in the size range below 44 micron. Work carried out on this fraction showed an almost uniform distribution of P2O5 and Nb2O5 throughout the entire size range. A not so favorable result was that 75% of the slimes reported as -10.5 micron, complicating conventional beneficiation techniques. Mineralogical studies found most of the phosphate minerals in these ultrafines to be crandallite.



Hydrosizing, wet high intensity magnetic separation and flotation were shown to be successful in treating a coarser (+15 micron) fraction of the fine slimes. This was associated with high reagent usage and considering the wide variation in the proportion of slimes (45-75%) and the fact that this process route was not regarded as proven technology, it was elected not to process the -44 micron fraction but reject it to waste. The opportunity for later treatment would obviously remain. The proposed process route arising out of the Bearden Potter investigations is shown schematically in Figure 5.1.

Figure 5.1: Schematic Flow Sheet - Bearden Potter



The material balance used for this study was developed from the Phase II prospect pitting programme and is shown in Table 5.2.

Table 5.2: Bearden and Potter Phase II Material Balance

Unit Weight % %P2O5 Distribution

%P2O5 Oversize +840 micron 5 5 2.43 Magnetics 10 5 4.85 Slime -44 micron 60 8 46.60 Flotation Tailings 15 5 7.28 Flotation Concentrate 10 40 38.84

It is noted that under this arrangement, a P2O5 recovery of 39% is indicated, with a concentrate product yield of 10%. In order to convert ore tonnage to concentrate tonnage for reserve estimation, Bearden and Potter used the Phase II metallurgical data to develop multiple regression models to estimate the ratio of concentration. Four formulae were developed, one for the total South Valley and one for each of the three upper levels: Total South Valley • R/C Total = 155.7571 – (114.4202 * EXP1) Level 1 (0.0-6.1m) • R/C Level 1 = 163.1083 – (120.4972 * EXP1) Level 2 (6.1-12.2m) • R/C Level 2 = 170.6781 – (125.7468 * EXP1) Level 3 (12.2-18.3m) • R/C Level 3 = 59.88196 – (39.49309 * EXP1) Where EXP1 = (Ore %P2O5)0.1 From a processing perspective this also highlights the variability in concentrate generation that can be anticipated from the various areas. This is illustrated in the following example assuming an ore feed grade of 14% P2O5. Total South Valley • R/C Total = 6.8 Level 1 (0.0-6.1m) • R/C Level 1 = 6.2 Level 2 (6.1-12.2m) • R/C Level 2 = 7.0 Level 3 (12.2-18.3m) • R/C Level 3 = 8.5



5.4.3 Foskor Following laboratory studies undertaken on grab samples in November 2000, Foskor’s geologists selected approximately thirty tons of samples for further investigation and processing through their pilot plant in Phalaborwa. Samples were taken from pits dug by mechanical digger to a depth around 2m on approximately 200m square grids in each of the North, South and West valleys. After creating a single composite sample from the bags of North valley material, the grade was found to average 8.8% P2O5. This was considered to be too low and this material was therefore excluded from further investigation. Each bag comprising the South and West valley samples were then analysed separately, with individual bags being selected to achieve composites samples with grades of 10.3% P2O5 and 11.2% P2O5 respectively. Unfortunately this resulted in the total mass of sample being reduced from thirty tons to twelve tons for testing in the pilot plant. The Foskor pilot plant comprised the following unit processes: • Repulping to the required density; • Attritioning; • Scrubbing in a rod mill; • Magnetic separation; • Desliming of the non-magnetic fraction in hydrocyclones; • Thickening of the hydrocyclone underflow; • Conditioning, pH adjustment and flotation. Results for the pilot plant runs on material from the South and West valley are summarized in Table 5.3 and Table 5.4 respectively.

Table 5.3: Foskor Pilot Plant Results – South Valley

Weight % P2O5 Fe2O3

Stream

Analysis %P2O5

Distribution %

Analysis % Fe2O3

Distribution %

Feed 100.0 9.8 100.0 37.0 100.0 Magnetic Reject 29.7 3.5 10.6 75.2 60.4 Slime Reject 18.2 9.0 7.5 29.2 14.3 Flotation Feed 52.1 15.4 81.9 17.9 25.3 Flotation Concentrate 17.2 38.4 67.4 4.9 2.3 Flotation Tailings 34.9 4.3 14.5 24.4 23.0



Table 5.4: Foskor Pilot Plant Results – West Valley

Weight % P2O5 Fe2O3

Stream Analysis %P2O5

Distribution %

Analysis % Fe2O3

Distribution %

Feed 100.0 9.0 100.0 38.9 100.0 Magnetic Reject 21.1 4.7 11.1 75.2 40.8 Slime Reject 22.6 7.5 18.8 29.2 16.9 Flotation Feed 56.3 11.2 70.1 29.2 42.3 Flotation Concentrate 12.4 37.2 51.4 7.2 2.3 Flotation Tailings 43.9 3.8 18.7 35.4 40.0 It is seen that on average Foskor achieved 57.2% P2O5 recovery with a yield of 15.6% to a concentrate with a grade of 37.8% P2O5. These results were lower than those reportedly achieved in the earlier TICAF operation. Foskor acknowledged that there were certain shortcomings in the pilot plant, notably the non-optimal operation of the cyclones used for desliming. Notwithstanding this they concluded that the ore responds well to flotation and that with optimized desliming, a high P2O5 grade concentrate with low Fe2O3 should be obtainable. In addition the concentrate produced was evaluated in the Foskor phosphoric acid plant. On average a total P2O5 efficiency of 98.6% was achieved. The gypsum filtration rate averaged 6.1 ton P2O5/m2/day with a maximum of 10.8 ton P2O5/m2/day. It was noted however that due to high iron content in the concentrate, X-compound formation was experienced when concentrating phosphoric acid to 54%. They expressed the view that it would not be possible to produce saleable phosphoric acid from concentrate containing only 30% P2O5. MISA reviewed the Foskor report and was generally critical of the procedures followed and the conclusions reached. They identified a number of limitations: • The Foskor pilot plant was set up to evaluate feed to their own process and was not ideal

for evaluating alternative sources such as Sukulu which have different characteristics. In SRK’s view the principal issue was that the Foskor pilot plant was not fully optimized to meet the requirements of Sukulu ore.

• MISA elaborated at length on the impact of -44µm material on the assessment of the deposit grade. They stated that the 44µm cut-off had been carried forward in the history of Sukulu including the TICAF operation and the Bearden Potter studies. In MISA’s view, much of the Foskor work was interesting but not meaningful because Foskor’s quoted feed grades were for total samples. By contrast MISA maintained that feed grades quoted in earlier work excluded the -44µm fraction. SRK has reviewed the Bearden Potter reports but not the underlying raw data. From its review of these reports SRK is not of the same view as MISA that the Bearden Potter study excludes the -44µm fraction in the metallurgical feed grades. SRK acknowledges however that a large proportion of naturally occurring -44µm material may not have reported to the sample because of the sampling technique employed. This is a very important issue and one that must be resolved one way or the other as the Sukulu deposit is characterized by a very high proportion of -44µm material.



• In further support of their hypothesis regarding -44µm material, MISA quote the concentration ratios as determined by Foskor of 3.1 for the South valley and 4.5 for the West valley compared to values of 9 to 10 predicted by Bearden Potter. In SRK’s opinion the Foskor concentration ratios quoted by MISA are not for the overall plant but for flotation section only and therefore represent a possible misinterpretation.

• The grades of the North valley sample may not have been as low as indicated. Rather the sample may have had a higher proportion of -44µm material than the other valleys. MISA accordingly considered the rejection of low grade North valley sample to be unjustified. SRK agrees that the proportion of -44µm material may well have contributed to the observed low grade of the North valley sample. It is unfortunate that this sample was not processed as this would have avoided speculation as to why the grade was found to be low.

• The flotation pH of 10 employed by Foskor was too high. This resulted in the dissolution of caustic soluble components which subsequently re-precipitated as fine hematite and alumina in both the concentrate and tailings. This contributed to the high iron content of concentrates and the subsequent formation of X-compound. Previous studies had recommended flotation pH in the range 9.0 to 9.2 to avoid this problem. SRK concurs that the flotation pH may well have been a problem and that this highlights the shortcomings of operating a pilot plant that has not been optimized to meet the feed characteristics.

• The non-optimal operation of the cyclones contributed to an excessive use of reagents. MISA expressed the view that whilst the Foskor results were of interest, they did not enhance the understanding of processing Sukulu ore in a substantive manner. The project has in the meantime moved on and the proposed flow sheet differs significantly from the previous TICAF flow sheet, the Bearden Potter flow sheet and the Foskor flow sheet. The above concerns are therefore of little relevance to the current project.

5.4.4 Resource Development Inc. In 2007, SRK (USA) undertook a preliminary assessment to prepare an updated resource statement ahead of a planned Feasibility Study. As part of this assessment, RDi undertook a programme of laboratory beneficiation studies. SRK (USA) supervised the collection of 24 samples from hand dug pits in the South, North and West valleys. Initial moisture determinations found the samples to vary between 9% and 30% moisture with an average of 16%. With the exception of one very low value sample the remaining samples ranged between 10% P2O5 and 19% P2O5 with an average of 13%P2O5. A composite of the 24 samples was produced for mineralogical and beneficiation investigations. The composite was found to contain magnetite (40%), apatite (18%), crandalite (10%), quartz (12%), hematite (5%), vermiculite (5%) and others (5%). Crandallite is an aluminum phosphate that is considered detrimental to the marketing of the apatite for the manufacturing of phosphoric acid by sulfuric acid methods. It had been detected in previous



investigations but was thought to report to the ultra-fines particle size fractions that would not affect the marketing of the coarser sized apatite concentrate. Beneficiation studies followed the general practices of previous investigators and included washing, screening, removal of -45µm slimes, magnetic separation and froth flotation of the apatite from a -850µm +45µm size fraction. The final concentrates produced in the laboratory studies did not meet the desired grade of greater than 40% P2O5. Additionally, the yield of apatite concentrate was lower than expected, and was only in the range of 7% of the feed by weight. Beneficiation studies to recover apatite from the -45µm fraction were not successful. An examination of the settling characteristics of the -45µm fractions determined that the material was rapid settling, but only reached a final density of 24% solids by weight. Following completion of the beneficiation studies, a flowsheet was developed that was similar to previous studies. A pilot plant was manufactured and shipped to site with the objective of confirming process performance at larger scale and providing product for testing by potential customers. To date the pilot plant has not been commissioned.

5.4.5 IMMT In October 2007, IMMT issued its report entitled “Development of the Commercial Process Flow sheet for Beneficiation of Low Grade Phosphate Ore and Production of Phosphoric Acid”. Testwork was reportedly conducted on 15 tonnes of test pit material received from Nilefos. The source of these samples was reportedly the same as the SRK(USA) geological samples. A review of the sample intervals however shows them to be shallow, with the majority being less than 10m, but with some going to about 20m. Characterisation Studies Representative subsamples were extracted from all samples received and submitted for chemical analysis by XRF. The analytical results largely reflected grades and compositions determined in previous investigations. XRD investigations confirmed the major minerals in all three valleys to be magnetite and fluorapatite. Bench Scale Beneficiation Studies Individual samples were combined on a valley basis and tested separately. Size grade analyses were undertaken as summarized in Table 5.5.



Table 5.5: Size Grade Analyses

Size North Valley South Valley West Valley (micron) Wt% Fe% P2O5% Wt% Fe% P2O5% Wt% Fe% P2O5% +3 000 12.33 55.01 7.42 8.42 48.58 7.56 15.03 47.19 7.76 +2 000 5.49 53.05 6.52 4.01 54.17 7.90 18.34 43.56 9.86 +1 000 5.83 56.97 6.55 7.74 55.29 6.81 6.34 45.52 10.70 +500 9.59 54.45 8.37 5.43 48.58 9.23 3.84 44.12 11.05 +300 6.04 40.77 16.28 5.50 33.51 17.34 4.51 34.67 12.84 +150 4.69 27.93 22.34 5.37 22.34 22.71 4.94 27.65 16.65 +75 8.78 22.90 25.22 7.07 18.43 25.85 3.89 25.13 19.19 +45 5.48 16.80 26.12 8.22 16.19 26.75 5.71 25.13 18.25 -45 41.77 19.54 13.99 48.24 19.54 12.96 37.40 25.97 13.92 Head 100.00 33.10 13.98 100.00 28.28 14.43 100.00 34.71 12.47

Bulk 27.92 13.82 36.86 11.42

The size grade analyses indicate that the coarse fractions tend to have higher iron values than the finer fractions, whilst the reverse holds for phosphate values. IMMT undertook a series of bench scale scrubbing, sizing, magnetic separation and flotation tests and ultimately defined a process route similar to that schematically represented in Figure 5.2. The IMMT flow sheet is seen to be considerably more complex than that proposed by Bearden Potter. The following key differences are obvious: • Bearden Potter rejected coarse oversize to waste after scrubbing whilst this was milled in

the IMMT flow sheet. • Bearden Potter rejected -44µm slimes to waste whilst this was processed in the IMMT

flow sheet. • Bearden Potter used high energy scrubbing and attritioning rather than low energy

scrubbing and ball milling employed by IMMT. • Bearden Potter employed Wet High Intensity Magnetic Separation as opposed to Low

Intensity Magnetic Separation used by IMMT. • IMMT used two stages of flotation producing two grades of apatite concentrate as

opposed to one stage used by Bearden Potter.



Figure 5.2: Schematic Flow Sheet - IMMT The -44µm slimes fraction can contain up to half of the P2O5 in the feed and its rejection to waste in the Bearden Potter flow sheet is probably the most significant difference from the IMMT flow sheet. Under this arrangement IMMT anticipate producing two apatite concentrates. The first is expected to have a grade of 40-42% P2O5 and a yield of approximately 15%. The second is expected to have a lower grade of 33-34% P2O5 and a yield of 6%. The combined yield of 21% is significantly higher that the 10% anticipated by Bearden Potter. The process route developed was seen as appropriate to set up a commercial pilot scale plant to generate data for a commercial plant and to assess process efficiency.



Settling Study on Sukulu Tailings A series of bench scale settling tests were carried in a 1 litre measuring cylinder using two Indian flocculants (DK Set 3110 and SN-1) and a UK flocculant (Magnafloc). The two Indian flocculants were found to outperform Magnafloc and they were used in the balance of the investigation. Of the two Indian flocculants, SN-1 was considered to be the better. Later studies used the technique of lowering the pH to assist settling. Leaching of Phosphate Rock With Sulphuric Acid The concentration of P2O5 in solution increased with time and pulp density and a density of 40% was considered as optimum. At this density the concentration of P2O5 in solution was 106g/l and the leach efficiency was measured at 61%. The leaching of iron also increased with time and pulp density but the calcium concentration was seen to drop above 20% pulp density due to precipitation as CaSO4. Pilot Plant Beneficiation The pilot plant study was conducted on about 3 tonnes of South Valley ore. It was undertaken with equipment available in the laboratory that had different capacities. Continuous operation of the pilot plant was therefore not possible. The results of the pilot plant investigation were reported as a series of mass and water balances across the various pieces of equipment but with no mention of P2O5 grades in any stream. Whilst the results may be of benefit in specifying equipment and pump sizes, they are of no value in terms of assessing metallurgical performance. Metallurgical results were provided to SRK on request but it is not certain whether these were derived from a process simulation or from the pilot plant study.

5.5 Mineral Processing Facilities The mineral processing facilities will be implemented in two phases. The first phase will include a concentrator and associated sulphuric acid plant, phosphoric acid plant and fertilizer manufacturing facility to produce SSP and triple super phosphate (“TSP”) fertilizer. The second phase will significantly expand the concentrator capacity to produce apatite concentrate for export.

5.5.1 Phase 1 The Phase 1 (IMMT) concentrator will have a feed capacity of 50tph or 320ktpa, producing around 67ktpa of concentrate. Conventional technology will be applied including the following unit processes: • Scrubbing; • Magnetic separation and storage of magnetic product; • Screening; • Milling;



• Magnetic separation and storage of magnetic product; • Two stage flotation; • Dewatering of flotation concentrate for fertilizer manufacture; • Tailings storage. Phase 1 will include the following facilities for the production of SSP and TSP: • Sulphur burning sulphuric acid plant with capacity of 250tpd or 82.5ktpa; • Hemihydrate phosphoric acid plant with capacity 100tpd phosphoric acid or 54ktpd P2O5; • Fertilizer manufacturing facility with capacity 200tpd SSP/ 150tpd TSP or 66ktpa SSP /

49.5ktpa TSP. The scheme of arrangement for Phase 1 is schematically shown in Figure 5.3.

Figure 5.3: Phase 1 Scheme of Arrangement

Plant feed is likely to show high variability in size distribution as well as the proportion of magnetite. Stockpiling and blending may be necessary to minimize such variability but it is recommended that the plant equipment and ancillaries be sized to handle higher variability than may be typical. The exact range of variability should be defined in the next phase of study. The feed to the flotation circuit is expected to be relatively consistent in grade and physical character.



The study assumes that elemental sulphur will be imported for the production of sulphuric acid. There are two basic technology alternatives for the production of phosphoric acid namely dihydrate and hemihydrate, although there are also other hybrid arrangements. The dihydrate acid digestion system produces dihydrate gypsum (CaSO4.2H2O) and phosphoric acid at 28-30% P2O5. The hemihydrate system produces hemihydrate gypsum (CaSO4. ½H2O) and phosphoric acid at 43-45% P2O5. This system is reportedly more difficult to operate but the higher acid strength results in lower evaporation requirements to produce 54% P2O5 used for fertilizer production. It also produces purer gypsum that may have some market value. It has been assumed for the purpose of the study that the hemihydrate process will be installed.

5.5.2 Phase 2 The Phase 2 (IMMT) concentrator will have additional feed capacity of 3 x 250tph or 4.8Mtpa, producing 1Mtpa of concentrate. Conventional technology will be applied including the following unit processes: • Scrubbing; • Magnetic separation and storage of magnetic product; • Screening; • Milling; • Magnetic separation and storage of magnetic product; • Two stage flotation; • Dewatering of flotation concentrate for bagging and export; • Tailings storage.

5.5.3 Review of the IMMT Flow Sheet In order to assess any flow sheet developed for Sukulu, one must first understand the nature of the ore both physically and chemically. The Sukulu ore is of igneous origin, with the associated volcanic protrusion subsequently being eroded over the aeons. The local topography then contributed greatly in that the product of weathering could collect in three main valleys. These weathered soils consist principally of fine apatite and iron oxide, mostly in the form of magnetite, hematite and goethite. The fine size distribution and the high proportion of fine iron oxides are key aspects when considering the beneficiation of this ore. On average more than 50% of the ore in the size range 0 to 44µm and about 40% of the available apatite reports to this size fraction while about 25% of the iron is contained in this particle size range. Furthermore the apatite and magnetite are evenly distributed throughout the particle size ranges. Bearden Potter results also show that milling the ore does not help as both the magnetite and apatite just shift into the smaller size fractions with no differential to use as a means of separating the two.



Two ways of beneficiating this ore are therefore apparent. Firstly one could remove the fine fraction (-44µm) ahead of magnetite removal and apatite recovery from the coarse fraction. Secondly one could reduce the size of the total feed to below 44µm ahead of magnetite removal and apatite recovery. The latter option was the one favoured by IMMT. The IMMT circuit can basically be divided into two main parts. In the first part the RoM ore is treated to firstly separate the fines associated with the ore. The coarse fraction is then milled to the desired fineness. Both these streams are then passed through Low Intensity Magnetic Separators (“LIMS”) to remove the magnetic fraction. In the second part of the circuit the two streams are combined and subjected to two stages of froth flotation to recover apatite. The first flotation stage renders a high quality product with very clean apatite (“P1” concentrate). The tailings from this circuit are then further subjected to flotation to yield a lower quality apatite concentrate (“P2” concentrate). Both concentrates are then again subjected to a LIMS step for further upgrading and then dewatered on belt filters. The IMMT flow sheet is schematically shown in Figure 5.4. A typical plant layout is depicted in Figure 5.5.



Figure 5.4: IMMT Flow sheet

SUKULU PHOSPHATE PROJECTSCHEMATIC PROCESS PLANT LAYOUT

Project No.

400092

Date:

11 12 2009

Compiled by:

BENE

Scale:

1:10 000

Fig No.

Date: 11 12 2009 Revision: A

Date:

Compiled by:

Scale:

Fig No.

5-5

J:\Proj\400092_Sukulu_phosphate\8CAD\FIGURES\Figure 5-5.dwg, mcdo



The use by IMMT of a screw classifier to achieve the attritioning is considered to be a good idea in that while it is cleaning the larger crystal surfaces it is also classifying the particles and separating the very fine material from the coarse material in a single step. Having reviewed the IMMT flow sheet however, SRK has concerns about the following observations: • Apatite, or phosphate rock, is a relatively cheap commodity and as such any recovery system

must be as simple and cost effective as possible. The IMMT flow sheet is fairly complex compared to conventional flowsheets and each additional unit process is a potential source of phosphate loss.

• This flow sheet was designed from a series of batch tests carried out at laboratory scale and did not account for circulating loads. In the commercial plant it will be necessary to control flotation pH at around 8 to 9 in order to optimise apatite recovery over magnetite recovery and to halt iron salt precipitation. Subsequently it will be required to adjust the pH of the pulp in the thickener to 6 to achieve optimum settling rates. The effect that recirculation of the lower pH thickener overflow will have on flotation is unknown and was not determined in the batch test procedures. Furthermore, it is well known in semi soluble salt flotation, that constituent ions of the semi soluble salt in the aqueous phase will depress flotation recovery. In the case of apatite flotation phosphate ions in the aqueous phase depress fatty acid adsorption.

• A further observation relating to the flow sheet development is the absence of P2O5 values in the reported pilot plant results. Without such information it is not possible to calculate recoveries or losses across individual unit processes. Metallurgical results were provided to SRK on request but it is not certain whether these were derived from a process simulation or from a pilot plant study.

• A key observation relates to the fineness of the ore and the extent to which this has been compounded by the further size reduction in the IMMT process. The size distribution of principal process streams are summarized in Table 5.6.

Table 5.6: Feed, Product and Waste Size Distribution

80% Passing

µm 50% Passing

µm 20% Passing

µm % Passing 16

µm RoM 1 000.0 50.0 2.6 35.0 Coarse Mags 382.0 113.0 49.0 4.5 Fine Mags 23.9 8.5 1.7 67.1 Prim Float Feed 24.1 7.4 1.1 69.0 P1 Conc 37.0 25.0 18.0 17.0 Sec Float Feed 7.6 2.3 0.7 99.0 P2 Conc 7.6 2.3 0.7 98.0 Tails 7.6 2.3 0.7 99.0

It is seen that 80% of the primary flotation feed is less than 24µm in size. Moreover 50% is below 7µm. Furthermore it is seen that according to the Modsim simulation, 80% of the primary flotation concentrate is less than 37µm in size, with 50% being below 25µm. Stated differently, the primary flotation concentrate is coarser than the primary flotation feed, implying that the ultra fine particles were not floated in the column flotation cells. These ultra fine particles are included in the secondary flotation feed which is indicated to be 80% less than 8µm. Secondary flotation comprises a



conventional flotation cell in series with a column cell. It is not apparent to SRK why particles that did not float in primary flotation should float in secondary flotation without any intermediate surface preparation. SRK is also concerned with the efficiency of flotation of such ultra fine particles. Problems of entrainment of phosphate particles by the very fine iron particles are also likely to be experienced in the LIMS step. • IMMT predict the product size of the P1 concentrate to be 80% passing 37µm whilst the P2

concentrate is predicted to be considerably finer at 80% passing 8µm. There is a real risk of losses in filtration, conveying and transport to distant customers of such fine product.

• Similarly tailings are seen to be 80% passing 8µm, posing severe tailings storage challenges.

In SRK’s view the IMMT flow sheet incorporates two features not found in conventional flotation circuits. Firstly, there is no desliming ahead of flotation. Secondly there is no attempt to change the surface characteristics of the ore between primary and secondary flotation. Notwithstanding these reservations, bench scale testwork undertaken by IMMT showed this process route to be very successful, indicating a combined concentrate yield of virtually double that previously shown. Performance of the pilot plant was reportedly also successful although no metallurgical results were reported to confirm this at the larger scale. On the basis of the superior bench scale test results and the reported performance of the pilot scale SRK has accepted the IMMT flow sheet for the purpose of the scoping study. We recommend strongly however, that the metallurgical performance of this process should be demonstrated in a continuous pilot plant before finalisation of the feasibility study. A key motivation for a continuous pilot plant is to assess the impact of recycle streams but it will also offer the opportunity to assess scale up issues and generate concentrate from different parts of the orebody. Issues such as concentrate filtering and drying should also be included in the pilot plant studies in order to quantify product moisture and the potential for losses of fine material.

The IMMT process route will generate less tailings because of the higher yield. The tailings will however be considerably finer which will impact on water retention and tailings stability. These aspects will need to be investigated as will water usage including the recovery and recycle of water. The pilot plant will assist greatly in understanding these issues. It was shown that fine material consumes a disproportionate amount of reagents and furthermore it was reported that the undesirable crandallite reports to the very fine fractions. The IMMT process route does not deslime ahead of flotation whereas the Bearden Potter route rejected all -44µm material to separate storage. There may well be another more optimal cut off size that is not worth processing. SRK recommends that the pilot plant have the option to reject material at say 15 to 25µm to assess any potential benefit.



5.5.4 Further Issues A fundamental question concerns the relationship between the P2O5 assay grade in the ore/ plant feed and the ultimate production of concentrate at the desired specification. Key aspects that impact on this will include the following: • The size distribution of the ore and the associated deportment of P2O5 across the various size classes. • The proportion of the magnetic component in the ore. • The proportion of P2O5 contained in apatite as opposed to crandallite. • The efficiency of the process in recovering P2O5. Bearden Potter addressed this aspect by using the Phase II metallurgical data to develop multiple regression models to estimate the ratio of concentration in order to convert ore tonnage to concentrate production tonnage. The P2O5 grade of the ore was the independent variable used in these regression models. These models obviously assumed the selected process of the time and are not applicable to the currently proposed IMMT flowsheet. Equivalent regression models, possibly using other dependent variables such as size and/or magnetite distribution, will have to be developed for the IMMT flowsheet during the next phase of study. This will require an extensive programme of test work if these models are to be statistically valid and the orebody will have to be characterized in terms of the selected dependent variables. The deposit is known to exhibit considerable variability, particularly in terms of grade, distribution of magnetics and the proportion of fines. Such variability can impact greatly on process performance. Most of the investigations to date have been undertaken on composite samples and in SRK’s view there is an inadequate understanding of the variability in metallurgical response spatially throughout the deposit. It is recommended that a programme of variability testing be included in the feasibility study. The current project proposes to produce two concentrates, a primary concentrate at 40-42% P2O5 and a secondary concentrate at 33-34% P2O5. During Phase 1 of the project the primary concentrate will be used in the production of phosphoric acid and TSP fertilizer whilst the secondary concentrate will be used in the production of SSP fertilizer. During Phase 2 of the project both concentrates will be exported. It is recommended that a cost benefit analysis be undertaken to assess the incremental return on the investment required to produce the secondary concentrate. The marketability of the secondary concentrate in particular should also be established during the feasibility study. An alternative to the dual concentrate approach may be to produce a single, lower grade concentrate suitable for fertilizer production. The economics of the Sukulu phase 2 operation will be dependent upon transport costs. The gross tonnage of concentrate transported can within limits be controlled in the flotation circuit. The trade-off of increasing the concentrate grade however, will be lower recovery. In order to optimize the concentrate grade it will be necessary to understand the concentrate grade/ recovery relationship and this should be addressed during the feasibility study.



SSP fertilizer is manufactured from apatite concentrate and sulphuric acid, whilst TSP fertilizer is manufactured from phosphoric acid and apatite concentrate. Manufacture of SSP only would thus avoid the production of phosphoric acid thereby reducing capital and operating costs and also avoiding the need to dispose of gypsum. Unless the manufacture of TSP is an absolute requirement of the licence, consideration should be given to only producing SSP or alternatively delaying the production of TSP. The key process risks are thought to be in beneficiation rather than in the manufacture of sulphuric acid, phosphoric acid or fertilizer. The reason for this is that beneficiation is principally dependent on the inherent characteristics of variable ore types whereas given a concentrate of known specifications, the production of phosphoric acid and fertilizer is relatively straight forward. In this light it is suggested that consideration be given to delaying the acid and fertilizer production in Phase 1 until concentration has been successfully demonstrated. Variations in the CaO/ P2O5 ratio have been observed at depth and these have been ascribed to the presence of crandallite, an aluminium phosphate. This needs to be further investigated during the feasibility study. The source of sulphur for the manufacture of sulphuric acid should be confirmed during the feasibility. The plant and associated facilities should be designed to cater for an annual rainfall of approximately 1 500mm per year received in the two rainy periods from March to June and September to November.

5.6 Projected Production Performance Projected annual production for Phase 1 and Phase 2 is summarized in Table 5.7.



Table 5.7: Projected Annual Production

Unit Phase 1 Phase 2 Concentrator Plant Feed t 320 000 4 800 000 P2O5 Grade %P2O5 13.9 13.9 P1 Concentrate t 48 317 724 752P2O5 Grade %P2O5 41.0 41.0 P2 Concentrate t 18 883 283 248 P2O5 Grade %P2O5 33.0 33.0Combined Concentrate t 67 200 1 008 000 P2O5 Grade %P2O5 38.8 38.8 Plant Discard t 252 800 3 792 000P2O5 Grade %P2O5 7.3 7.3 Total P1 Concentrate Yield % 15.1 15.1 Total P2 Concentrate Yield % 5.9 5.9Total Combined Concentrate Yield % 21.0 21.0 Total P2O5 Recovery % 58.7 58.7 Phosphoric Acid P1 Concentrate Utilisation t P1 35 754 P1 ConcentrateP2O5 Content tP2O5 14 659 Sulphuric Acid Utilisation tH2SO4 37 760 Gypsum Generation t CaSO4 66 288 Phosphoric Acid P2O5 Content tP2O5 13 120 Phosphoric Acid Production t H3PO5 24 296 Triple Super Phosphate P1 Concentrate Utilisation t P1 12 562 P1 ConcentrateP2O5 Content tP2O5 5 151 Phosphoric Acid Utilisation tP2O5 13 120 Phosphoric Acid Utilisation t H3PO5 24 296 TSP Production t TSP 39 360 TSP P2O5 Content tP2O6 15 744 Single Super Phosphate 0 P2 Concentrate Utilisation t P2 18 883 P2 ConcentrateP2O5 Content tP2O5 6 231 Sulphuric Acid Utilisation tH2SO4 11 200 SSP Production t TSP 31 360 SSP P2O5 Content tP2O5 5 331 Sulphuric Acid 0 Sulphur Utilisation t S 16 500 Sulphuric Acid Production (100%) tH2SO4 48 960

It is important to note that at the rated capacities of the process plant components, the implied volumes of SSP that can be produced relative to the individual component capacities are not in balance (see Table 5.8). Table 5.8: Phase 1 implied volumes of SSP from different components

Plant Component Production Capacity Component Production

Volume

Equivalent SSP Production

(ktpa) (ktpa) Concentrator 50tph x6 400hr/yr 67 200 concentrate 111 600 Sulphuric Acid Plant 250tpd x 345 day/yr 86 250 acid 176 020 SSP/TSP plant (producing SSP only) 200tpd x 345 day/yr 69 000 SSP 69 000 If the production capacity of SSP plant were to be increased to 300tpd, it would then be possible to produce around 105 000tpa of SSP, which would then effectively match the capacity of the concentrator. Sulphuric acid that is surplus to requirements can be sold.



Waste product generation and size distribution associated with the two phases of production are summarized in Table 5.9. Table 5.9: Waste Products and Size Distribution

Arisings

(tpa)

80% Passing

(µm)

50% Passing

(µm)

20% Passing

(µm)

Passing 16µm

(%) Phase 1

Coarse Magnetics 82 392 382.0 113.0 49.0 4.5 Fine Magnetics 44 401 23.9 8.5 1.7 67.1Flotation Tails 126 007 7.6 2.3 0.7 99.0 Gypsum 66 288

Phase 2Coarse Magnetics 1 235 884 382.0 113.0 49.0 4.5 Fine Magnetics 666 012 23.9 8.5 1.7 67.1 Flotation Tails 1 890 104 7.6 2.3 0.7 99.0 Gypsum 0

1 235 884 382.0 113.0 49.0 4.5 It is seen that the “fine magnetics” and “flotation tails” in particular are extremely fine. No information was found in the test reports reviewed on the gypsum crystal size. Results of a literature search however, suggest that the mean average diameter of gypsum crystals is likely to be less than 10µm.

5.7 Operating Costs DRA Mineral Projects (“DRAMP”) was requested to provide estimates for capital and operating costs of the Phase 1 and Phase 2 process plants. These estimates were based upon the beneficiation plant flowsheet developed by the IMMT. DRAMP regard the design, costing and implementation of mineral processing plants as their principal area of expertise. They accordingly commissioned Desmet Ballestra in partnership with BCS Holdings Company to estimate the capital and operating costs of the sulphuric acid, phosphoric acid and fertilizer plants.

5.7.1 Phase 1 The operating cost estimate for the 50tph beneficiation plant is summarized in Table 5.10. An independent consultant commissioned by Nilefos has reviewed these costs and found them to be acceptable. Table 5.10: Opex Estimate of 50tph Apatite Beneficiation Plant

Component Annual Cost (US$’000)

Labour 597.7 Beneficiation Plant 597.7 Spares 445.2Mechanical & EI 445.2 Consumables 1 326.9Reagents 217.9 Steel Balls 97.4 Steel Lining 972.1 Bulk Bags 0 Water 39.4 Concentrate Transport 0 Power 1 193.0Beneficiation Plant 1 193.0 Base Annual operating Cost (US$’000) 3 562.7Unit cost (US$/t concentrate) 53.02



The labour estimate was based on a minimal structure allowing for a plant manager but excluding other senior management positions. The following budget rates of remuneration for Uganda as advised by Nilefos were assumed: • Manager – expatriate rate • Staff – US$1 250 per month • Worker – US$250 per month Annual Mechanical and E&I spares were estimated at 4% of the total mechanical equipment cost. Consumables costs were based upon budget prices received from reputable South African suppliers. DRAMP used in-house knowledge to estimate steel ball consumption and mill liner consumptions. Power consumption was estimated based upon 75% of the mechanical equipment installed power at a cost of US$0.12/kWh consumed. A raw water cost of US$0.40/t was assumed for the plant. The operating cost estimate for the 250tpd sulphuric acid plant as compiled by DRAMP is summarised in Table 5.11. On the basis of 345 plant operating days per annum, the total acid that can be produced is 86 250tpa. The labour cost was estimated based on a complement advised by Desmet Ballestra and utilising budget rates of remuneration for Uganda advised by Nilefos. Spares costs were estimated at 4% of the plant capital cost. Consumables costs were based upon sulphur and water consumption advice from Desmet Ballestra, a sulphur price of US$180/t delivered Tororo and a water price of US$0.40/t. Desmet Ballestra estimated the absorbed power at 740kW which was costed at US$0.12/kWh consumed.

Table 5.11: Opex Estimate of 250tpd Sulphuric Acid Plant (DRAMP)


Labour 907.5 Sulphuric Acid Plant 907.5

Spares 540.5Mechanical & EI 540.5Consumables 4 966.6Sulphur 4 966.6 Power 703.4Sulphuric Acid Plant 703.4 Base Cost (US$’000) 7 118.0Unit cost of acid produced (US$/t acid) 82.53

Based on first principles and typical industry usage factors, Nilefos’ independent consultant calculated a unit cost for sulphuric acid of US$84.88/t acid. Although this was done for a larger plant capacity, SRK considers this to be in good agreement with the cost shown in Table 5.11 and the DRAMP cost has thus been used for evaluation purposes. The operating cost estimate for the 100tpd phosphoric acid plant is summarised in Table 5.12. The phosphoric acid operating costs were factored from an original estimate for a 50tpd production rate. The labour cost was estimated based on a complement advised by Demit Ballestra and utilising the budget remuneration rates for Uganda as advised by Nilefos. It was further assumed that engineering and management personnel would be common for the acid and fertilizer plants. Spares costs were



estimated at 4% of the plant capital cost. Desmet Ballestra estimated the installed power at 2 500kW of which it was assumed 60% would be absorbed at a cost of US$0.12/kWh consumed.

Table 5.12: Opex Estimate of 100tpd Phosphoric Acid Plant


Labour 678.0 Phosphoric Acid Plant 678.0 Spares 1 092.5Mechanical & EI 1 092.5 Consumables 79.2

79.2 Power 1 425.8Phosphoric Acid Plant 1 425.8 Base Cost (US$’000) 3 275.5Unit cost (US$/t acid) 94.94

Nilefos’ independent consultant did not comment specifically on the operating cost for the phosphoric acid plant. This is included here for completeness as TSP was part of the original exploitation concept for Sukulu, although the results of this study suggest that TSP should not be produced, and hence phosphoric acid would not be required. The operating cost estimate for the 200tpd SSP and 150tpd TSP fertilizer plants is summarised in Table 5.13. The labour cost (assumed to be the same for both SSP and TSP processes) was estimated based on a complement advised by Demit Ballestra and utilising the budget remuneration rates for Uganda as advised by Nilefos. It was further assumed that engineering and management personnel would be common for the acid and fertilizer plants. Spares costs were estimated at 4% of the plant capital cost. Desmet Ballestra estimated the installed power at 50kW/t product of which it was assumed 75% would be absorbed at a cost of US$0.12/kWh consumed.

Table 5.13: Opex Estimate of 200tpd SSP and 150tpd TSP Fertilizer Plant


SSP Only(69.0ktpa)

SSP + TSP (31.4ktpa SSP) (39.4ktpa TSP)

Labour (Fixed) 678.0 678.0 Labour Fertilizer Plant 678.0 678.0 Spares 610.9 610.9 Mechanical & EI 610.9 610.9 Consumables 4 384.1 3 486.9 SSP

Diesel oil 4 356.5 1 980.0 Water 27.6 12.5

TSP Diesel oil 1 485.0 Water 9.4

Power 326.7 259.9 SSP 326.7 148.5 TSP 111.4 Base Cost (US$’000) 5 999.7 5 035.7 Unit conversion cost of fertilizer (US$/t product) 86.95 71.21



It should be noted that a major component of the consumable cost in the SSP/TSP conversion cost in Table 5.13 is for diesel oil. SRK has been unable to get any clarity on the reason why diesel oil is included in the conversion cost, but suspects that this may be related to diesel-fired drying of the finished product. If the use of the diesel oil can be obviated, by for example the use of the discard heat from the sulphuric acid plant for product drying, then the unit cost in Table 5.13 reduces to US$23.82/t of SSP produced. Based on the 67ktpa of concentrate produced in Phase 1, Nilefos’ independent consultant argued that it would be possible to produce 107ktpa of SSP (compare this to Table 5.8). Working from first principles and applying typical industry usage factors, he estimated that the unit conversion cost for SSP would then be US$26.81/t. Although this was done for a larger plant capacity, SRK considers this to be in reasonable agreement with the adjusted cost above.

5.7.2 Phase 2

The operating cost estimate for the 250tph beneficiation plant is summarised in Table 5.14.

Table 5.14: Opex Estimate of one 250tph Apatite Beneficiation Plant


Labour 920.0 Beneficiation Plant 920.0 Spares 1 247.5Mechanical & EI 1 247.5 Consumables 15 947.5Reagents 3 228.8Steel Balls 1 454.3 Steel Lining 7 232.2 Bulk Bags 3 421.6Water 610.6 Concentrate Transport 0 Power 9 995.2Beneficiation Plant 9 995.2

Base Cost (US$’000) 28 110.2 Cost assumptions were similar to those for the 50tph beneficiation plant. It should be noted that the cost of rail transport of the concentrate to the port of Mombasa is not included in the estimate.

From information provided by Nilefos and Nilefos’ independent consultant, SRK can find no reason why diesel oil should be included in the SSP conversion cost. For evaluation purposes, SRK has decided to remove this component from the SSP conversion cost. The reader is cautioned that should this be proved incorrect, it has a significant effect on the project economics.



5.8 Capital Costs DRAMP compiled mechanical equipment lists from the IMMT process flowsheet and design documentation. Budget quotations were obtained from reputable South African vendors for both the equipment included in the 50tph and 250tph beneficiation plants. Based on recent contracts, the total mechanical supply cost was aligned to the percentage for the mechanicals (27.5%) and the other cost components such as civil construction, steel, plate, piping, electrical etc. were factored from the mechanical supply cost. The total project cost included EPCM costs but made no contingency allowance. Separate allowances were made for earthworks and buildings. It should be noted that no layout work was done in this concept study. All costs were estimated in SA Rands and converted to US Dollars using an exchange rate of ZAR8.00 = US$1.00. Desmet Ballestra in association with BCS Holdings Company provided order of magnitude costs for the supply of the sulphuric acid plant, the phosphoric acid plant and the two fertilizer plants. These estimates excluded delivery to site, civil designs and earthworks, mechanical and E&I erection, site supervision, utilities and interconnections between plant components (conveyors). Desmet Ballestra’s cost estimates for the acid plants were quoted in Euros and converted to SA Rands at ZAR11.50 = €1.00. In SRK’s view the basis of the capital cost estimation is appropriate for the level of a scoping study.

5.8.1 Phase 1 The capital cost estimates for the various components of the Phase 1 processing plants are summarised in Table 5.14 to Table 5.17, as follows: • Phase 1 50tph beneficiation plant (Table 5.15); • 250tpd sulphuric acid plant (Table 5.16); • 100tpd phosphoric acid plant (Table 5.17); and • the 200tpdSSP / 150tpd TSP fertilizer plant (Table 5.18).

Table 5.15: Capex Estimate of 50tph Apatite Beneficiation Plant Component Factor Cost

(US$million)Civil Works 11.5% 4.8 Steel Supply 10.5% 4.4 Steel Erection 2.5% 1.0 Platework Supply 7.0% 2.9 Platework Erection 1.0% 0.4 Mechanical Supply 27.5% 11.4 Mechanical Erection 4.5% 1.9 Piping 8.0% 3.3 Electrical & Instrumentation 11.5% 4.8 Transport 1.0% 0.4 EPCM 15.0% 6.2 Pre-Production Expenses 0.0% Consumables/ Spares 0.0% Total Plant (US$million) 100.0% 41.5

Capital costs for earthworks (US$7.0million) and buildings (US$4.3million) which had been included by DRAMP, have been removed from Table 5.15, as they are non plant-related costs and



are seen to duplicate similar costs carried elsewhere (see Section 8.14). No contingency amount has been included in the estimates set out in Table 5.15. Nilefos’ independent consultant estimated a constructed cost for a Phase 1 beneficiation plant at ±US$32million, including supply, erection and EPCM. The plant configuration and location differed from that presented in this report and also did not provide for the Phase 2 expansion, so this cost estimate is not directly comparable to the figure in Table 5.15. It does suggest though that it may be possible to realise certain savings in the capital cost for the Phase 1 beneficiation plant, depending on plant location and configuration. This will need to be examined during the Phase 2 feasibility study.

Table 5.16: Capex Estimate of 250tpd Sulphuric Acid Plant

Component Cost (US$ million)

Desmet Ballestra Solid sulphur feeding Sulphur melting and feeding Sulphur combustion, SO2/SO3 conversion Air drying, SO3 absorption Compressed/ Instrument air production Demi water production unit Auxilliary steam production Cooling water system Sulphuric acid tank farm Budget price FOB country of origin (€ 9.4 Million) Budget price FOB country of origin 13.5BCS Holdings Company Materials supply process plant & utilities 1.9 Civil earthworks 3.3 Mechanical, E&I erection 7.0 Auxilliaries 0.2Shipping to RSA 0.5Shipping to Uganda 2.2 Cranage 0.9 Vehicles 0.4 Accommodation 2.7 Budget price for plant erection 19.1Total Sulphuric Acid Plant (US$million) 32.6

Nilefos’ independent consultant estimated a constructed cost for a 300tpd sulphuric acid plant at ±US$19million, including supply, erection and EPCM. He pointed out that for small acid plants, there is little difference in the cost for a 200, 250 or 300tpd plant, so going for a larger plant involves no financial penalty and offers improved flexibility. His cost excluded civil earthworks (carried elsewhere), had a much lower shipping/transport and erection cost and did not provide for utilities. This cost estimate is thus not directly comparable to the figure in Table 5.16. It does suggest though that it may be possible to realise certain savings in the capital cost for the sulphuric acid plant. This will need to be examined during the Phase 2 feasibility study.



Table 5.17: Capex Estimate of 100tpd Phosphoric Acid Plant

Component Cost(US$ million)

Desmet Ballestra Prayon licence and know how Basic and detailed engineering Budget price FOB country of origin (€ 19 million) Budget price FOB country of origin 27.3BCS Holdings Company Materials supply process plant & utilities 3.8 Civil earthworks 6.6 Mechanical, E&I erection 14.2 Auxilliaries 0.5 Shipping to RSA 0.5 Shipping to Uganda 2.2 Cranage 0.4 Vehicles 0.2 Accommodation 2.7 Budget price for plant erection 31.2TOTAL (US$million) 58.5

Nilefos’ independent consultant did not comment specifically on the capital cost for the phosphoric acid plant. This is included here as TSP was part of the original exploitation concept for Sukulu, although the results of this study suggest that TSP should not be produced, and hence phosphoric acid would not be required. Table 5.18: Capex Estimate of 200tpd SSP or 150tpd TSP Fertilizer Plant

Component Cost (US$ million) 200tpd SSP 250tpd TSP

Desmet Ballestra Prayon licence and know how Basic and detailed engineering Budget price FOB country of origin (€10.6 million) Budget price FOB country of origin 12.2 15.3BCS Holdings Company Materials supply process plant & utilities 1.7 2.1 Civil earthworks 2.9 3.7 Mechanical, E&I erection 6.4 7.9 Auxilliaries Shipping to RSA 0.2 0.2 Shipping to Uganda 1.1 1.1 Cranage 0.2 0.2 Vehicles 0.2 0.2 Accommodation 1.4 1.4 Budget price for plant erection 14.1 16.9TOTAL (US$million) 26.4 32.2

DRAMP advised it is technically feasible to combine the SSP and TSP fertiliser plants into a single process area which would make use of common equipment. Thus both TSP and SSP fertiliser could be manufactured on a batch basis. The capital cost of a combined SSP/TSP plant would be in the order of 25% more than a single, stand-alone SSP plant, as can be seen in Table 5.18.



Nilefos’ independent consultant expected that the constructed cost for a 200tpd SSP plant, including a granulation and bagging plant, would be approximately US$14million, including supply, erection and EPCM. This was based on a Banfield Den system, which may use different technology to that offered by Desmet Ballestra. His cost excluded civil earthworks (carried elsewhere), had a much lower shipping/transport and erection cost and did not provide for utilities. This cost estimate is thus not directly comparable to the figure in Table 5.18. It does suggest though that it may be possible to realise certain savings in the capital cost for the SSP/TSP plant. This will need to be examined during the Phase 2 feasibility study. The capital estimates as provided by DRAMP have been accepted for evaluation purposes, although it is possible that certain savings in capital costs may be possible.

5.8.2 Phase 2 The capital cost estimates for one Phase 2 250tph beneficiation plant are summarised in Table 5.18. Certain adjustments incorporated into the capital cost estimate for the 50tph beneficiation plant as compiled by DRAMP have been applied on a pro-rata basis to the capital estimates in Table 5.19. Table 5.19: Capex Estimate of one 250tph Apatite Beneficiation Plant

Component Factor Cost(US$ million)

Civil Works 11.5% 8.1 Steel Supply 10.5% 7.4 Steel Erection 2.5% 1.8 Platework Supply 7.0% 4.9 Platework Erection 1.0% 0.7 Mechanical Supply 27.5% 19.3 Mechanical Erection 4.5% 3.2 Piping 8.0% 5.6 Electrical & Instrumentation 11.5% 8.1 Transport 1.0% 0.7 EPCM 15.0% 10.5 Pre-Production Expenses 0.0% Consumables/ Spares 0.0% Total Phase 2 Beneficiation Plant (US$million) 100.0% 70.3

The total capital cost of the Phase 2 plant expansion for 3 additional modules is estimated at US$211 million.

5.9 Conclusions and Recommendations for Further Feasibility Study Investigations

5.9.1 Metallurgical Characterization of the Deposits A fundamental question concerns the relationship between the P2O5 assay grade in the ore/ plant feed and the ultimate production of concentrate at the desired specification. This question is fundamental to the definition of the reserve but also to the variability in metallurgical response across the deposit. Key parameters that may impact on these aspects will include the following:



• The size distribution of the ore and the associated deportment of P2O5 across the various size classes;

• The proportion of the magnetic material in the ore; • The purity of the apatite component; • The proportion of P2O5 as apatite as opposed to crandallite; • The efficiency of the process in recovering P2O5. The Sukulu deposits are known to exhibit considerable variability, particularly in terms of grade and the proportion of fines. Most of the investigations to date have been undertaken on composite samples and in SRK’s view there is an inadequate understanding of the variability in metallurgical response spatially throughout the deposit. In order to complete such metallurgical characterization and variability analysis it is recommended that the following be undertaken: • Design a bench scale test procedure that represents the finally selected flowsheet. This should

include the main unit processes namely, size classification, pre-flotation LIMS, milling, two stage flotation, post-flotation LIMS.

• Extract approximately 30 drill core samples selected spatially throughout each of the three deposits.

• Subject each sample to the defined bench scale test. • Analyze data with the objective of correlating concentrate production with key input variables. • For the purpose of the scoping study it has been assumed that thirty samples will be tested from

each of the three deposits. The criteria for selection of these samples will have to be defined early in the next phase of study. It has further been assumed that these samples will be extracted from drill core required for geological purposes.

• It is estimated that this programme could be completed within the latter half of the ten week pilot plant programme outlined below, once the process flowsheet has been finalized and approved.

5.9.2 Confirmation of IMMT Flowsheet at Pilot Scale On the basis of the superior bench scale test results and the reported performance of the pilot scale tests undertaken by IMMT, SRK has accepted the IMMT flow sheet for the purpose of the scoping study. SRK recommends strongly however that the metallurgical performance of this process should be demonstrated in a continuous pilot plant before finalisation of the feasibility study. The objective of the pilot plant programme will be to confirm the process route, develop process design criteria and generate concentrate for market assessment. A key requirement of a continuous pilot plant is to assess the impact of recycle streams but it will also offer the opportunity to assess scale up issues, determine reagent consumptions and generate concentrate from different parts of the orebody. Issues such as concentrate filtering and drying should also be included in the pilot plant in order to quantify product moisture and the potential for losses of fine material. The pilot plant must allow full mass and component balancing in terms of tonnage, P2O5, Fe and size grading. The IMMT process route does not include a desliming step but there may well be an optimal cut off size that is not worth processing. In order to assess the impact of fines on the process efficiency, SRK recommends that the pilot plant include the flexibility to reject material at say 15 to 45µm to



assess any potential benefit of not processing the total size spectrum. This could be achieved with a sizing cyclone that could be bypassed if the total feed is to be processed. Exactly what constitutes pilot scale operation is subject to interpretation. IMMT have indicated that they would be prepared to guarantee their process subject to pilot plant operation at 10-20tph. At the lower end of the scale other international laboratories have proposed pilot plant investigations at 0.25-1.0tph. Also available are mini-pilot plants that operate continuously at less than 10kg/h. In assessing the appropriate scale of operation there are a number of key criteria to consider: • The operation must be continuous with recycle of intermediary streams as per the final plant. • The stream flows must be large enough such that samples extracted for analysis do not impact

greatly on circuit stability. • The throughput must be high enough to produce concentrate for market assessment if required,

in a reasonable period of time. In SRK’s view, IMMT’s proposal of a 10-20tph pilot plant would be extravagant particularly when Phase 1 of the project envisages construction of 50tph beneficiation plant. That said however, we consider that it is of paramount importance that IMMT give their unqualified support to the pilot investigation and its findings as it is their process that is being evaluated. On balance, having considered these and the above listed criteria, SRK recommends the use of an intermediate scale pilot plant (0.25-1.0tph) for Feasibility Study investigations. The following programme of work is envisaged: • 20t sub-samples with plant configured per finally approved flowsheet. Collect 3 x 30t bulk

samples that are representative of the average life of mine ore characteristics of each deposit. • Split out the following representative sub-samples for each of the three deposits:

- 1 x 5t sub-sample plus 1 x 20t sub-sample. - Combine and blend the three 5t sub-samples and split into 2 x 7.5t sub-samples.

• Extract representative sub-samples from the balance of the bulk samples to conduct the following bench scale tests for each of the three deposits: - Optical mineralogy; - Scanning electron microscopy; - Head assay, XRF scan, XRD; - Size/ grade analysis (P2O5 and Fe); - Determination of optimal desliming size (from no desliming to desliming at say 15 and

45µm); - Scrubbing assessment at optimal desliming size; - LIMS assessment at optimal desliming size; - Grinding indices at optimal desliming size; - Batch rougher/ cleaner flotation at optimal desliming size; - Settlement tests at optimal desliming size;

• Pilot plant operation on 1 x 7.5t sub-sample with no desliming (as per IMMT flowsheet). • Pilot plant operation on 1 x 7.5t sub-sample with desliming at optimal desliming size

determined in bench scale tests.



• Finalise flowsheet on basis of above two pilot plant runs. • Pilot plant operation on 3 x Data collected from the pilot runs will be used to develop full mass and component balances in terms of tonnage, P2O5, Fe and size grading. The concentrates collected will be subjected to assay and XRF scans. Waste products will be available for further testing as required by tailings and waste storage design engineers. In this regard it is noted that the IMMT process route will generate less tailings than the conventional route but these will be finer which will impact on water retention and tailings stability. These aspects will need to be investigated as will water usage including the recovery and recycle of water. A deliverable of the Preliminary Assessment conducted by SRK (USA) in 2007 was the design and supply of a pilot plant for on site testing of Sukulu ore. This was completed and a 113kg/h (250 lb/h) pilot plant was delivered but never installed and commissioned. The obvious question arises as to its suitability for the currently proposed pilot plant campaigns. SRK has reviewed the pilot plant configuration plus equipment supplied and concluded that it is not ideal for the currently proposed programme. Our principal concern is that the SRK (USA) pilot plant was designed in terms of a different processing philosophy. Under this philosophy, coarse and fine fractions were rejected, with only an intermediate size fraction being processed. By contrast, the IMMT philosophy is to process the total feed size package and accordingly very little of the supplied equipment will be appropriately sized. Furthermore the IMMT circuit includes milling and column flotation which were not included in the SRK (USA) flowsheet. It may be possible to reconfigure the SRK (USA) pilot plant to match the finally approved flowsheet but in SRK’s view this would best be assessed in the future should there be a requirement for a plant optimisation tool. Of greater importance presently is IMMT’s unqualified support of the pilot investigation and its findings. They have already expressed reservations with reconfiguring the SRK (USA) pilot plant, further supporting the use of another facility. Finally, an alternative approach presents itself should it not be required to generate large quantities of concentrate for market assessment. A company called Canadian Process Technologies Inc markets a mini pilot plant. Typically this would operate at about 10kg/h. SRK has reviewed the available modules and considers that they could be arranged to match the IMMT flowsheet with some additional equipment and a little ingenuity. This would however, need to be ratified by IMMT. It is estimated that the required modules could be purchased for US$1.5million but this would largely be offset by the lower sample transport and processing costs. A further advantage of the mini pilot plant is that it would be available during the rest of the mine’s life for optimisation studies. For the purpose of the scoping study we have assumed the use of an off site pilot plant but it is recommended that applicability of the mini pilot plant be assessed early in the Feasibility Study.

5.9.3 Design Allowance for Process Variability Plant feed is likely to show high variability in size distribution as well as the proportion of magnetite. Stockpiling and blending may be necessary to minimize such variability but it is recommended that the plant equipment and ancillaries be sized to handle higher variability than may be typical. The exact range of variability should be defined following the variability analysis in the next phase of study.



5.9.4 Review Concentrate Production Strategy The current project proposes to produce two concentrates, a primary concentrate at 40-42% P2O5 and a secondary concentrate at 33-34% P2O5. It is recommended that a cost benefit analysis be undertaken to assess the incremental return on the investment required to produce the secondary concentrate. The marketability of the secondary concentrate in particular should also be established during the feasibility study. An alternative to the dual concentrate approach may be to produce a single, lower grade concentrate suitable for fertilizer production. The gross tonnage of fertilizer and concentrate transported can within limits be controlled by adjusting the concentrate grade. The trade-off of increasing the concentrate grade however, will be lower recovery. In order to optimize the concentrate grade it will be necessary to understand the concentrate grade/ recovery relationship and this should ideally be addressed during the feasibility study.

5.9.5 Review Fertilizer Production Strategy SSP fertilizer is manufactured from apatite concentrate and sulphuric acid whilst TSP fertilizer is manufactured from phosphoric acid and apatite concentrate. Manufacture of SSP only would thus avoid the production of phosphoric acid thereby reducing capital and operating costs and also avoiding the need to dispose of gypsum. Unless the manufacture of TSP is an absolute requirement of the licence, consideration should be given to only producing SSP or alternatively delaying the production of TSP. The key process risks are thought to be in beneficiation rather than in the manufacture of sulphuric acid, phosphoric acid or fertilizer. In this light, consideration should be given to delaying acid and fertilizer production in Phase 1 until concentration has been successfully demonstrated. The source of sulphur for the manufacture of sulphuric acid should be confirmed during the feasibility.



APPENDIX PHASE II COMPOSITE SAMPLE RESULTS

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