no. 17 shaft project—overcoming ‘fourth generation ... · stakeholders. the project is also of...

NO. 17 SHAFT PROJECT—OVERCOMING ‘FOURTH GENERATION’ TECHNICAL CHALLENGES 375

Introduction

Project scopeNo. 17 Shaft is planned to produce 225 000 tonnes permonth of Merensky and UG2 ore from eight operationallevels. At steady state, production will come from 16Merensky half levels and 16 UG2 half levels. Wasteproduction was planned at 13% of reef production.

Three vertical shafts were planned for the project. A linedand equipped 10 m diameter shaft to be sunk to a depth of1922 metres below surface will be used to transport menand materials in and out of the mine and will also be usedas a ventilation air intake. The second shaft is a lined butunequipped 9 metre diameter up-cast ventilation shaft, to besunk to a depth of 1 704m below surface. The third shaft isa lined but unequipped 6.1 m diameter downcast shaft forsuper-cooled air, to be sunk to 1 127m below surface.Ventilation requirements were largely responsible fordetermining the size and number of vertical shafts required.

Mining methodology

The proposed mining methodology is a modification of thecurrent Impala ‘best practice’, adapted to cater for theincreased depth of No. 17 Shaft. The stoping horizon willbe accessed by a vertical shaft system and a network ofhorizontal haulages that will carry people, material andventilation air.

On the reef horizon, a combination of raises and winzeswill be used to establish an on-reef infrastructure. Stopingwill be based on a concept of controlled stopingsequencing, similar to sequential grid mining, which aimsto control levels of stress and seismicity by optimallysequencing mining operations. Stoping operations will beconducted in conventional narrow reef stope panels withscraper cleaning, supported by a combination of pillars orbackfill, mat packs, in-panel elongates and tendons.

Mine design challengesNo. 17 Shaft will be the second shaft to fall into the‘intermediate depth’ mining environment. The project isvery similar to No. 16 Shaft, it represents a change from thewell understood shallow environment (<1 000 m depth) tothe intermediate environment (1 000 m–2 250 m depth).Experience in this environment is limited within theBushveld Complex. Much of the design work conducted forNo. 16 shaft was customised for No. 17 Shaft. However, theincrease in depth presented additional geotechnical andventilation related risks. The geotechnical and ventilationchallenges and solutions are based on the project reports byL.J. Gardner1 and ‘17 Shaft Prefeasibility Summary ofVentilation and Cooling Design Criteria, options, decisionsand constraints’ by Bluhm Burton Engineering Consultants2

respectively.

Geotechnical considerationsA rigorous approach was adopted to predict thegeotechnical environment and anticipated rock massbehaviour in order to have an accurate assessment of thefactors that affect mine design. The geotechnical challengesare typical of the ‘medium depth’ or ‘intermediate depth’mining environment (1 000 metres to 2 250 metres belowsurface, according to Jager and Ryder3). These arecharacterized by the following challenges in the Merenskyand UG2 economic horizons:

• Stress conditions ranging from low to moderately high,with the ratio of horizontal to vertical stress decreasingto below unity

• Rock mass behaviour is dominated by stress-relatedfracturing and displacement, rather than geologicalfeatures

• Most excavations are surrounded by an envelope offractured rock. Although this creates problems withsidewall slabbing, it has the benefit of generatinghorizontal ‘clamping’ stresses, which help to stabilizethe strata. As a result, hangingwall stability increases

ZINDI, L. No. 17 Shaft project—overcoming ‘fourth generation’ technical challenges. Third International Platinum Conference ‘Platinum inTransformation’, The Southern African Institute of Mining and Metallurgy, 2008.

No. 17 Shaft project—overcoming ‘fourth generation’technical challenges

L. ZINDIImpala Platinum Ltd

No. 17 Shaft project was approved by the Impala Platinum Holdings Limited board of directors on12 February 2008. The project was recognized as an important project requiring high levels ofinvestment of approximately R 5.5 billion in 2007 money terms. It presented significant technicalrisks related to mining at depth, but had potentially acceptable financial rewards for Impala and itsstakeholders. The project is also of strategic importance to Impala’s growth strategy. No. 17 Shaftproject is the second shaft at Impala Rustenburg lease area to fall into the ‘intermediate depth’mining environment (1 000 m–2 250 m depth). Experience in this mining environment is limitedwithin the Bushveld Complex. This paper gives an account of how the project team mitigatedtechnical risks related to mining at depth in order to achieve an acceptable business case.

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• Stoping excavations experience significant closure (upto 50%), at widely varying rates

• In areas of high percentage extraction, seismic activitycan be expected.

In addition to the depth-related challenges, the followinggeological structures of note had significant impact on minedesign challenges, listed in some detail by Golenya4 in theproject geological report:

• Hex River fault—this major structural feature strikesNNE-SSW across No. 17 Shaft project. It is a laterallypersistent feature and is defined by a central core zoneapproximately 10 metres wide. Sympathetic faultingcan extend up to 65 metres to the east or west of thecentral core zone. The fault has an average dip of 85 degrees to the east and a constant displacement of 8 metres in the same direction that is normal in nature.The core zone is highly fractured, extensively sheared,altered and has been invaded by serpentinite, chlorite,quartz and calcite filled veinlets. The fault is notnormally associated with water inflows, although onoccasions the core zone can be moist.

The geotechnical risks associated with the Hex River faultinclude:

• Adverse ground conditions that could slowdevelopment and compromise the long-term stability ofexcavations traversing the feature

• The possibility that the structure could at some time inthe future become a source of mining-induced seismicactivity.

For this reason, planned mining rates through the faulthave been reduced, with increased levels of long-termsupport that incorporate yielding capability to accommodateground movement associated with seismic events.

• The ‘keel’ structure lying to the north of the projectarea—this major structure, together with its associatedgeological disturbance, creates two distinct structuraldomains within the No. 17 Shaft project area. These arethe less disturbed southern portion and the structurallycomplex northern area. The other major effect of the

‘keel’ structure is the thinning out of the layering as itapproaches the structure. A geological section of theproject area, taken through a line of explorationboreholes aligned approximately on strike, is shown inFigure 1.From a geotechnical viewpoint, ground conditions inthe vicinity of the keel are likely to be more disturbed,presenting an increased risk of rock-related instability.This was addressed by reducing planned extractionfactors in this area, and will likely also requireadditional support at increased cost.Also, the thinning of the layering implies that tomaintain the required middling between footwallhaulages and the stoping environment, the haulageswill be forced to cross-cut some of the footwall layers.The planned support regime should, however, cater forany complications arising from such crosscutting.

• Potholes—these are common features, randomlydistributed, roughly circular in shape and vary from tensto hundreds of metres in diameter. The occurrence ofpotholes within the No. 17 Shaft project has beenconfirmed by surface exploration drilling and they affectboth the Merensky and UG2 horizons. The percentage of mineable ground possibly lost topotholes is predicted to vary from 10% for normalMerensky ‘A’ reef, through 15% for normal UG2 ‘E’ reefto as high as 35% on the UG2 horizon in the ‘keel’ zone.The geotechnical risks associated with potholes on alocal scale include deterioration in ground conditions inthe vicinity of the pothole, increasing the potential forinstability and requiring additional support. Moreimportantly in the case of No. 17 Shaft is their potentialeffect of disrupting the planned controlled sequencingof stoping operations, resulting in a build-up of stresswith possible seismic activity.

In situ stress regimeIt is important to have a comprehensive understanding ofthe in situ stress regime before addressing the mine designchallenges related to depth.

Figure 1. Showing a strike section through the project area illustrates the impact of the ‘keel’ structure



Several sets of in situ stress measurements wereconducted at Impala, at depths ranging from less than 500metres to more than 1 300 metres below surface. As part ofthe No. 16 Shaft project study, the results of thesemeasurements were analysed and extrapolated for theincreased depths of the fourth-generation shafts.

Based on the results of the No. 16 Shaft project study, thefollowing in situ stress field is expected in the No. 17 Shaftproject area:

• The vertical stress component will approximate theoverburden cover load, based on an overlying rockdensity of approximately 2 750 kg/m3.

• The average k-ratio (i.e. the ratio between vertical andhorizontal stress) will continue to decrease as depthbelow surface increases, as shown in Table I.

• The stress magnitude in the strike direction is likely toremain approximately 20% higher than the stressmagnitude in the dip direction.

Although there will be shear stresses present, their lowrelative magnitudes and inconsistency make themextremely difficult to predict.

Rock property testingIn order to have meaningful prediction of the impact of thein situ stress regime on the mine design the rock propertiesmust be considered. The results of the uniaxialcomprehensive strengths test results for different horizonsare presented in the Table II and further clarification in theAppendix 1 tables. There was no significant differencebetween these results and the presently known rockproperties at Impala, which provided reassurance that therock properties are not changing with increasing depth.

Mine design solutions to ameliorate in situ stress relatedchallenges

Shaft sinking designsAlthough there is minimal design work associated with theindividual shaft barrels at this phase of the project, a

proactive approach was adopted to identifying potentialproblematic areas that may be encountered during thesinking phase. Geotechnical boreholes were drilled downthe length of each shaft barrel.

The core from each hole was then logged and the holeswere further assessed by means of wire-line geophysics and3-dimensional borehole radar. Potentially problematic areaswere identified and the sinking contractor used this at theplanning stage during sinking. This information was alsoused by the sinking contractors to prepare the sinkingtender document.

For the shaft stations, standardized station layouts wereused as far as possible, with a limited number of stationbrow sizes and shapes. Based on work at Nos. 16 and 20shafts, excavation and support methodologies have beenrefined and standardized. The station positions werechecked to ensure that they do not fall into geologicalhorizon boundaries or prominent planes of weakness.

Shaft sinking supportThe intention of support during the sinking phase,particularly in the shaft barrels, is to provide protection tothe personnel working in the shaft bottom from potentialfalls between the bottom and the permanent concrete lining,which is normally lagging 20 to 30 metres above the shaftbottom. In view of the large spans and increased riskassociated with shaft and station intersections and stationbrows, additional penetrating support, as well as arealcoverage is also required in these areas.

The support requirements were split into four distinctsections, namely:

• The shaft barrel• The station brow, catwalk and immediate area• The station and associated development excavations• The Merensky reef shaft pillar pre-extraction.

In addition, all possible additional support requirementswere listed in the report to be priced by the shaft sinkingcontractors.

The shaft barrel supportThe shaft barrel support generally consists of split set-typefriction anchors of various lengths, in cases combined withcladding in the form of welded mesh or expanded metalsheeting.

The station brow, catwalk and immediate area supportThe station brow, catwalk and immediate station area willbe supported by a combination of prestressed cable anchorsand steel tendons, with areal coverage being provided byfibre reinforced shotcrete.

Table IIUniaxial compressive strength values for different geological horizons

Rock type / horizon Strength (MPa) Rock type / horizon Strength (MPa)Bastard pyroxenite 200 FW 8 spotted anorthosite 138Mid 3 mottled anorthosite 185 FW 9 mottled norite 188Mid 2 spotted anorthosite 161 FW 11 spotted anorthosite 155Mid 1 norite 150 FW 12 mottled anorthosite 174Merensky pyroxenite 122 UG2 pyroxenite 104Merensky pegmatoid 131 UG2 chromitite 83FW1 anorthositic norite 152 UG2 pegmatoid 74FW 2 pyroxenite 100* FW 13 spotted anorthosite 166FW 3 / 5 anorthositic norite 147 UG1 pyroxenite 131FW 6 mottled anorthosite 187 UG1 chromitite 57FW 7 anorthositic norite 163 FW 16 mottled anorthosite 162

Table IForecast k-ratios and stress magnitudes for No. 17 Shaft project

Depth below surface—metres 1 200 1 500 1 800 Vertical stress magnitude—MPa 33 41.3 49.5Average k-ratio 1.2 1.0 0.9Strike k-ratio 1.4 1.15 1.0Strike stress magnitude—MPa 46.2 47.5 49.5Dip k-ratio 1.0 0.85 0.8Dip magnitude—MPa 33 35.1 39.6

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The station and associated development excavationsThe shaft stations and associated development will besupported by primary support in the form of steel tendons,with areal coverage being provided by fibre reinforcedshotcrete.

Shaft protection pillarWork done for the No. 16 Shaft project established that theminimum acceptable size for a shaft protection pillar at thisdepth is approximately 500 m x 500 m per shaft on bothreef horizons. This would lock up more than 300 000square metres of ore reserve per reef horizon, as well asdelaying the shaft production build up as haulages aredeveloped to outside the protection pillar boundaries.

As with other recent projects, it was proposed to pre-extract the Merensky horizon protection pillar on the mainand ventilation shafts, making use of satellite pillars toensure shaft stability. The pillar on the UG2 horizon wasleft intact to avoid an extended steel tower.

A flexible ‘skeleton tower’ will be used to suspend themain shaft steelwork in the area where movement is likelyto occur, normally approximately 50 m above and belowreef elevation.

This option has a number of advantages:• The higher-grade Merensky ore is extracted early in the

shaft’s life, providing earlier financial returns, fasteraccess to reserves and an improved stoping build up.

• The risk of intersecting serious geological anomaliesthat require waste stoping is generally lower on theMerensky horizon.

• A smaller portion of each shaft’s barrel is subject todisplacement than if the pillars on both horizons wereextracted, thus less support is required in the barrelsand a shorter steelwork tower is needed in the mainshaft.

• The shafts are effectively destressed, so a smaller pillarwill be required on the UG2 horizon.

A detailed investigation into the protection pillarextraction programme was conducted, based on the workdone for No. 16 Shaft. This investigation has proven thatthe concept of early extraction is definitely viable. Theresulting deformation in the shafts’ barrels can beaccommodated by replacing the shafts’ concrete lining withfibre-reinforced shotcrete, installing additional support andisolating the main shaft steelwork by means of the tower.

Shaft pillar inner cut As part of the shaft pillar pre-extraction, an inner reef cutwill be removed and supported in both shafts during thesinking phase. The inner cuts will measure 20 m on strike x20 m on dip x 1.4 m high.

The process briefly involves extracting the inner cut usingsome suitable mining configuration to be proposed by thecontractor and installing in-stope roof bolts, yieldingelongates and yielding cementitious matpacks (Durapacks)to support the inner cut. Finally, the inner cut will be filledby aerated cement to prevent ground conditions fromdeteriorating.

As the shaft pillar forms part of the on-reef airway, it willbe among the first on-reef areas to be developed,particularly given its proximity to the shaft. The up- anddown-dip stoping of portion of the airway surrounding theshaft will be completed first, followed by both the airwaymining and normal production stoping ‘bomb shelling’away from the shaft area. This concept has been included inthe mine planning modelling exercise, which specifies theproject production build-up.

Regional pillar design Based on the sequential grid mining methodology, theregional support strategy will take the form of large, dip-orientated regional pillars, with pillars separating each raiseline. These pillars are intended to compartmentalize themine, limit the amount of closure in stoping panels andreduce the on-face stress levels and associated seismicactivity.

Where bracket pillars are required on geologicalstructures, these have been incorporated into the overallpillar layout, and offset against the normal pillars asexplained earlier. Every third regional pillar has been leftintact as an aid for ventilation control. As is presentpractice, during stoping operations any large geologicallosses that are encountered (potholes, replacementpegmatoid zones, etc.) will be incorporated into the regionalstability plan, offset against pillars that can be released formining.

The essential design criteria for this pillar layout includethe average stress level within each pillar, the amount ofclosure that occurs in the stopes between the pillars and themagnitude of off-reef stresses generated by the pillars. Thepillars have thus been designed to an industry-acceptedmaximum stress level of 2.5 x the uniaxial compressivestrength of the host rock mass.

Based on work done for the No. 16 Shaft project andsubsequent additional computer modelling, the regionalpillar dimensions have been specified as follows:

• At 1 200 m, 20 m wide pillars• At 1 500 m, 30 m wide pillars• At 1 800 m, 40 m wide pillars

Pillar sizes are increased proportionally at regularintervals between the above benchmarks, as illustrated onFigure 2.

Primary development

LayoutsFrom the shaft stations, the two economic horizons will beaccessed by means of a network of horizontal station cross-cuts, footwall drives and lay-bys or cross-cuts. The stopinghorizon is accessed by travelling ways, and on-reefdevelopment is conducted by raising or winzing. Whenlaying out these excavations, all rock engineeringguidelines applicable to deep-level mining (e.g. inter-excavation spacing, angle of turnouts, angle to intersectdiscontinuities) were applied.

At least one station cross-cut excavation will bedeveloped per level, with most levels having two or more inview of ventilation requirements. As these excavations

Figure 2. Regional pillar configuration and raise spacing



cross-cut through the geological horizons, they willprobably have to have shotcrete applied during thedevelopment phase to ameliorate the effect of stress-induced failure, as illustrated in Figure 3. In addition, inview of their required long-term stability, they will alsoprobably require secondary support in the form of meshingand lacing. The installation of secondary support withyielding capability will be particularly important wherehaulages traverse geological structures that may be, or maybecome, seismically active.

Optimal location below reef to satisfy geotechnicalrequirements has significant impact on excavation volumes,project timing and cost implications. Based on the workdone during the No. 16 Shaft project, a range of gradually-increasing minimum depth below reef was initiallyrecommended (based on the work done during No. 16 Shaftproject), as shown below adapted to local geologicalconditions:

• At 1 200 m, not closer than 21 m below reef elevation• By 1 500 m, this increased to 30 m below reef elevation• By 1 800m, this again increased to 40 m below reef

elevation.However, at the rock engineering workshop held during

November 2006, it became clear that the use of haulageslocated deeper than 30 m in the footwall would severelydelay access to reef, and would enforce the use of cross-cuts, requiring an additional 30% of off-reef developmentper raise line. It was agreed to locate all footwall drivesbelow 1 500 m at 30 m below reef with 100 % of thesedrives to have shotcrete applied and 60 % of the drives tobe meshed and laced. From the footwall drives, Impalapresently makes extensive use of a system of lay-bys toprovide tipping facilities and material storage, withtravelling ways being used to access reef. Similar layoutshave been planned for the upper seven levels of No. 17Shaft. This will result in travelling ways of between 35 and50 m in length.

Given that the travelling ways will be cross-cuttingthrough the footwall layers at an oblique angle, it has beenrecommended that they be meshed and laced to ensurelong-term stability. In addition, due to stress-inducedfailure, it will probably be necessary to apply shotcrete tothese ends during development.

All travelling ways and the cross-cuts will have shotcreteapplied during the development phase, as well as secondarysupport in the form of meshing and lacing. The raises andwinzes will primarily be supported by means of tendons,but a small percentage may require additional primarysupport shotcreting or secondary support in the form ofmeshing and lacing.

Although not normally considered part of the accessdevelopment, rock passes also form a vital part of themine’s infrastructure. Recent experience on platinum minesin the Western Bushveld has shown that as mining depthsexceed 1 000 m, problems are experienced with the stabilityand operability of main/station rock passes. In view of this,space has been planned in the main tip cross-cuts for afourth set of rock passes if required. It is also planned to‘prehabilitate’ the main rock passes by means of additionalsupport prior to use.

Support detailsDevelopment support can be divided into three sections:

• Primary support—installed by the development crew,its main purpose is to provide a safe environment forthe development crew to work in. It typically consistsof tendons of adequate length and load-bearingcapacity (presently hydraulically prestressed units),which should be installed at a density of approximately1.0 square metres per tendon into the hangingwall anddown to the grade line for the sidewalls of excavations. Given the increased depth of the No. 17 Shaft project, itis anticipated that extensive fracturing of thedevelopment end rock walls may occur, particularlywhen developing through anorthositic rock types(almost 70% of development). Present experience hasshown that the optimal method to ameliorate suchfracturing is to spray a 50 mm thick layer of shotcreteonto the exposed rock walls soon after blasting as an inline activity. This action immediately stops furtherfracturing, and allows work to proceed safely.

• Secondary support—normally installed by specialistcontractors, this support is intended to ensure thestability and continued usage of an excavation underinduced stress loading. It generally comprises groutedtendons combined with wire mesh and tensioned lacingrope in different patterns, but can also includeadditional shotcrete.

• Tertiary support—also normally installed by specialistcontractors, this is normally installed in a ‘belt andbraces’ type situation. It varies from additional meshingand lacing patterns to pretensioned grouted cableanchors and yielding steel arches. The use of tertiarysupport should be minimized wherever possible bycareful planning and proper mining strategies. This isthe type of support recommended for excavationsacross the Hex River fault.

• A fourth section can be added, namely the support ofrock passes. This process generally involves accessingthe rock pass and making it safe for work, installingrings of tendons on a set pattern with wire meshcladding, followed by a covering layer of impact- andabrasion-resistant shotcrete. This function is normallycarried out by specialized contractors.

Stoping phase

Closure ratesIn-stope closure typically has two components—the primary(largest) component being based on the rate of face advance,Figure 3. Footwall to footwall shotcrete



while the secondary (smaller) component is time dependant.This combination provides the parabolic closure curvenormally recorded in closure measurement programmes.

A monitoring programme at Northam Platinum Mineduring the 1990s recorded closure rates and magnitudes infour stope panels over a 60-day period. Closure rates variedbetween 1 mm per day and 4.5 mm per day in the stopingpanels. Closure levels in stope panels measured between 50mm and 150 mm, with an average of 100 mm over a 60-dayperiod. Total closure levels of up to 200 mm were measuredin the stope centre gully, despite the stope having beenbackfilled during the project.

A second opinion was obtained from Dr D.F. Malan, anexpert in closure monitoring. He predicted similar rates ofin-stope closure, varying from 1 to 10 mm per day. He alsopredicts that closure recorded in UG2 stopes will generallybe less than in Merensky stopes, due to the lower extractionrates. He has, however, warned that based on previousexperience, the closure rates and magnitudes recorded inmatpack-supported stopes are significantly higher thanthose recorded in pillar- or fill-supported stopes.

SeismicityThe greater mining depth and associated increase in stresslevels mean that rock mass behaviour will be stress related,rather than geology related. This in turn implies an increasein mining-related seismic activity (i.e. the violent release ofstored strain energy), particularly in the deeper areas of theproject.

The majority of current seismic activity is associated within-stope yielding pillars bursting. As mining depthincreases, additional seismic activity could potentially begenerated by movement of geological structures. Thecontrol and management of the rockburst risk associatedwith this seismic activity will be one of the major rockengineering challenges for the No. 17 Shaft project.

Mining strategiesAccording to Jager and Ryder3, strategies to mine in the‘medium’ or ‘intermediate’ depth environment must includethe following measures:

• Regional support systems to control rock massinstability and reduce levels of seismic activity

• The use of appropriate rock engineering design criteriaand computer simulation to evaluate and optimize localmining layouts

• Correct sequencing of mining operations to prevent theformation of potentially seismically active remnants

• The monitoring of seismic activity to determine thesources of seismic activity and the implementation ofeffective measures to reduce the build-up of excessshear stress on problematic geological structures

• Modifying or redesigning support systems toaccommodate the increased closure rates while stillsupplying effective support resistance and energyabsorption capability

• Locating off-reef excavations further below reefelevation to minimize the damage arising from inducedstress levels generated by stoping operations and wherenecessary the greater use of secondary support in off-reef haulages to control such damage.

Controlled stoping sequencingControlling the sequencing of stoping operations is adeparture from Impala’s traditional scattered breast stoping,

but is necessary in view of the increased depth,accompanying stress build-up and associated seismicactivity. The controlled stoping sequencing (CSS) conceptproposed for No. 17 Shaft generally follows the principlesproposed for the sequential grid mining method, withcertain exceptions and allowances based on the Bushveldgeological environment (notably the presence of potholes)and the different rock mass response of the Bushveldmining environment.

Sequential grid mining (SGM) was originally introducedat Elandsrand Gold Mine in the early 1990s, as analternative to replacing scattered stoping methods with longwalling. The primary objective of SGM is to control levelsof stress and seismicity by optimally sequencing breast-mining stoping operations to abut up against regularlyspaced dip-orientated regional barrier pillars. The successof the concept was comprehensively reviewed by Handleyet al.7 in 1998. It has since been implemented on a numberof other gold mines, notably in the Anglogold Ashantioperations.

The idealized scenario for both SGM and CSS is asfollows:

• On a macro scale, the overall direction of mining isoutward from the shaft in a diamond shape, movingfrom raise line to raise line, toward the boundaries ofthe mining lease area. Because the shallowest anddeepest levels are opened up later than the intermediatelevels, they tend to lag slightly, causing the overallmining front to be diamond-shaped. This is consideredto be a favourable overall mining configuration in orderto minimize or eliminate stress induced seismicity.

• On a micro scale (i.e. within the raise line), control ofthe mining sequence involves mining first proceedingtowards the shaft, so that formation of the dip-orientated final pillars takes place at low span (therebylimiting the stress levels). Mining from the same raisethen proceeds away from the shaft, towards thestopping line for the next pillar.

• To assess the practical implications of implementingsequenced stoping operations, visits were conducted tothe Elandsrand and Mponeng mining operations.Discussions were also held with personnel fromNortham Platinum and several external consultants.These visits and discussions have resulted in a set ofguidelines being formulated for CSS. In addition, acomputer planning simulation was commissioned toassess whether CSS can be effectively implementedwithout hindering the production build-up and ongoingproduction requirements.

For Impala, the implementation of controlled stopingsequencing in place of the present scattered mining methodrepresents one of the significant changes to current practice,and will require greater planning, more control, and achange of management mindset.

This change is, however, essential to manage the seismicand rockburst risk associated with deeper-level mining.Another significant change will be the need to over-stopeall near-reef service and access excavations, even if thisentails off-reef stoping.

In-stope support strategyIn-stope needed to support the hangingwall beam up to theBastard Merensky horizon above Merensky stopes, or theuppermost chromitite stringers above UG2 stopes. In viewof the large depth range of the project, the anticipated stress



levels and associated in-stope closure magnitudes, a singlesupport strategy will not be adequate. At least threeseparate strategies for in-stope support (i.e. between theregional pillars) will be required, probably with overlappingareas. These strategies are as follows:

• Yielding in-stope grid pillars (6 m x 3 m) to separatepanels. These pillars are presently used at Impala toprovide localized support resistance, prevent ‘backbreak’ type large falls of ground and reduce closurelevels in panels. They also reduce stress levels on theface, thus reducing the potential for face bursting. Therisk is that as depths, stresses and closure levelsincrease, the pillars themselves could build up toomuch energy and become burst prone. It is unlikely thatthese pillars will function deeper than 1 400 metres.

• Crushing in-stope grid pillars (4 m x 2 m) to separatepanels. The intention with these pillars is that theycrush almost while they are being cut, and so cannotbuild up significant stress until more than 30% closurehas occurred. Although to some degree they fill thesame role as the yield pillars, they provide far lesssupport resistance, resulting on higher face stress levelsand a greater potential for stress fracturing and facebursting. As an optimistic estimate, it is planned to usethese pillars from below the yield pillar zone down to1 600 metres.

• When significant stress fracturing occurs and closurelevels become so great (>40%) that pillars can nolonger function effectively, then conventional artificialsupport is required to allow controlled yielding of thehangingwall. The envisaged stope closure will exceedboth the load and deformation range of availableyielding elongate units. Additional conventionalsupport such as some form of backfill or full-scalematpacks (timber, grout or a combination) will berequired. While both systems have their merits, theyrepresent a significant variation from current Impalasupport practice. Artificial support will be requiredfrom below the crush pillar zone, and will suffice to thebottom of the block (1 850 metres).

The above depth specifications, particularly for thechange-over from yield to crush pillars, should be taken asa guideline only. Depending on the occurrence of pillarbursting, it may be necessary to change from one system toanother at shallower depths than planned.

In-panel support strategyThe in-panel support strategy will also have to be adaptedas mining depth increases, similarly to the in-stope strategy.However, the following key principles are envisaged:

• Bolting of the hangingwall will be required to ensure asafe working area between the mining face and timbersupport. When stress fracturing becomes a significantissue, the bolting may be used in combination withnetting to minimize the effect of small falls. Typicalbolt spacing will be in the order of 1.5 to 1.8 squaremetres per unit.

• Yielding elongates (e.g. Strocams) will be employed asprimary in-panel timber support. These will providecontrolled yield, with the necessary energy absorptioncapability for rockburst-prone areas. Typical spacingfor these units will be 3–4 square metres per unit. Itmust be noted that without pillars, the envisaged stopeclosure will exceed the both the load and deformationrange of the yielding elongate units.

• In areas with in-stope pillars, elongate clusters or someform of pack support (timber, grout or a combination)will be required to provide a barrier line function, andsecure the integrity of gullies and access ways. Typicalspacing for these units will be 8 to 10 square metres perunit.

• In areas without in-stope pillars, additionalconventional support in the form of backfill or full-scale matpacks (timber, grout or a combination) will berequired to supplement the yielding elongates, allowingthe anticipated stope closure to occur in a controlledmanner.

• Additional, longer tendons will also be installed ingullies and access ways, to ensure their integrity. Atgreater depths these may also require some form ofareal support such as netting, shotcrete or thin sprayedliners. Typical tendon spacing for these areas will be inthe order of 1.0 square metre per unit.

Use of backfillThe need for an artificial support system to replace the roleof in-stope pillars, i.e. supporting the middling up to theBastard Merensky horizon or the upper chromitite stringerwas identified.

The analysis showed that although the support resistancerequirements could be met by either a matpack- or a fill-based support system, there were several additionaltechnical benefits to opting for the backfill option. Thesebenefits included safety, logistics, productivity, ventilationeffects and fire risk mitigation.

In addition, the metal retained in in-stope pillars thatcould now be released for mining would offset theadditional cost of the backfill.

In view of these facts, it was recommended that 30% ofthe stoping area in the bottom three levels of the shaft befilled with a 2 MPa cemented backfill mix. The fillingwould take place by hydraulic pressurisation of discretebackfill ‘socks’, contained by wire mesh barricades locatedbetween rows of elongates. Each panel would installbetween 3 and 4 of these ‘socks’ per month.

The design of a suitable mix, as well as the pricing of theappropriate plant, piping and accessories has beenconducted by an external specialist. The presently-envisaged system involves moving tailings from the base ofthe tailings dam to the shaft, then separating out the desiredfraction for fill before pumping the residue back to the dam.The tailings will then be thickened and piped underground,where cement will be added prior to placement.

It will be some time before stoping commences at No. 17Shaft, and much work will have been done on refining theconcept of backfill in the platinum mining industry by then.The project team intends to closely monitor the progress inthis technology, so that the most up-to-date and cost-effective option available at the time can be installed wheneventually required.

The next section of the report addresses the technicalchallenges arising from ventilation and cooling challengesand the mitigation strategies related to mine design of‘fourth generation shafts’.

Mine ventilation and cooling

Ventilation and cooling design criteriaThe ventilation and refrigeration challenges that had to beovercome relate to the virgin rock temperatures that varied



from 41°C in the shallow level (21 level—1 290 metresbelow bank) to 63°C in the lowest working level (28 level –1 820 metres below bank). The typical geothermal gradientin the Bushveld complex is approximately 2.187°C forevery 100 metres of depth. The ventilation and coolingsystems were designed to achieve a reject temperature of28.5°C. The ventilation and cooling design had to cater fora conventional mining method with a total production ofore and waste of 250 000 tonnes per month. This relates to8 Merensky levels stoping at a rate of 2 500 square metresper half level per month and 6 UG2 levels at 2 000 squaremetres per half level per month. This implied that a robustrefrigeration capacity was required to sustain an acceptableworking environment. The detailed ventilation and coolingrequirements were designed and modelled by BluhmBurton Engineering (BBE) consultants using the Vumasimulation software.

Geothermal zonesThe ventilation and cooling requirements for the block wereoptimized by considering two main geothermal zones. Theupper zone comprises working levels from 21 level to 25level, with typical range of virgin rock temperatures of41°C to about 53°C. The lower zone comprised levels 26,27 and 28, with the lowest level with virgin rocktemperatures of about 63°C. Two ventilation and coolingsolutions were considered for the two distinct geothermaldistricts.

Mine ventilation impact on shaft designThe main ventilation and refrigeration parameters for theNo. 17 Shaft mining block were optimized in considerationof the impact they would have on the essential excavationssuch as the main and ventilation shafts. The ventilationshaft system is illustrated in Figure 4.

The results in dramatic increase in cooling requirementsand ventilation rather than hoisting requirements dictatedthe number of shafts and shaft diameters. The challengewas to design an effective system for the range of virginrock temperatures taking into account rock engineeringconstraints and cost implications.

The use of a dedicated fridge shaft for the project wasconsidered after testing several alternatives. Thejustification of an independent fridge shaft was based on acomparative case study against the use of a main shaft andtwo alternatives of cooling infrastructure. In the firstalternative the 10 metre diameter main shaft was used as aintake with a primary ventilation capacity limited to 830kg/s compared to 1 300 kg/s and the surface air coolerwould be limited to 34 MW compared to 56 MW. Howeverthe main shaft option would require underground aircooling in a suite of secondary and tertiary air coolers forrecooling and reuse of air. This could be achieved in twoways: a surface based refrigeration plant with chilled waterpiped to and returned from the underground workings. Thiswould require robust equipment such as turbines and largepump stations. The second alternative would beunderground based refrigeration plant with chilled watercirculated to the underground workings. The first optionrequires underground air cooler duty of 34 MW, giving atotal cooling power capacity of 64 MW compared to the 56MW of the preferred alternative. The second option wouldrequire chilled water flow from surface of 540 litres persecond and associated surface refrigeration plant with acooling capacity of 50 MW. In addition a turbine-generatorstation will be required just above 21 level with a rating of5.5 MWe and water pumps rated at 13 MWe. Additional aircooler duty of 23 MW of secondary and tertiary coolingwill be required. It is evident that either option compared tothe fridge option have higher requirements in terms ofcooling capacity, water handling, and larger underground

Figure 4. Illustration of the No. 17 Shaft systems



excavations, which result in higher costs. The costcomparison was in favour of a surface refrigeration systemwith a dedicated fridge shaft.

The cost comparison was as follows: R 838 million forthe surface system compared to R 951 million for theunderground system.

No. 17 Shaft ventilation and cooling designThe natural downcast air through the main shaft wasdesigned to serve the upper production area between 21 and25 level and the ultra cool air force ventilated through thefridge shaft would serve the lower working levels 26 to 28level.

Ventilation models indicated a global heat-coolingbalance requiring primary ventilation of 1 300 kg/s (5 kg/sper kilo tonne per month production) and a total bulk aircooler duty of 56 MW (220 kW per kilo tonne per monthproduction). Surface exhaust fans on the vent shaft drawexhaust air from the workings in the mine, with downcastair being drawn naturally (9.5 m/s) through the main shaftand ultra cool air force-ventilated at high air speeds (14.5m/s) through the fridge shaft.

Refrigeration infrastructureThe vent shaft will have a large exhaust main fan station onsurface, with four fan-motor sets of 3 MW nominal rating(total 12 MW). Large bulk air coolers will be installed onsurface at the main shaft (34 MW) and fridge shaft (22MW). The refrigeration system will include ice thermalstorage for load damping and energy management. Totalrated electrical power for the refrigeration system will be 17MW. Refrigeration will be required during shaft sinkingand development, and the first refrigeration modules as wellas the fridge shaft bulk air cooler will be installed in2011/12.

Underground mine ventilation designThe underground ventilation system will have eight distinctdistricts. Ventilation will be delivered to the workings viathe network of off-reef haulages, and will be removed viadedicated on- and off-reef return airways. Analyses haveassumed that production will take place on all 8 levelssimultaneously and full production (Merensky plus UG2)will be 255 000 tonnes per month reef and developmentwaste.

The possible synergies with adjacent No. 10 Shaft wereexamined and there was no significant long-term savingsfor the project. However, a cross-connection to bedeveloped from 19 level Merensky south drive to No. 17Shaft 21 level via a raise borehole will allow returnventilation via this connection which will be important forearly development. This cross-connection will also allowthe injection of refrigerated air to the No. 10 Shaftworkings, thus providing a mutual benefit to that part of theoperation. The connection will further serve as a secondoutlet for the No. 17 Shaft operation.

The proposed ventilation and cooling system appears tobe at the limits of its capabilities to satisfy an acceptableworking environment. Any changes in current designworking depths indicated a significant change in ventilationand cooling requirements that would demand a more robustsolution such as ice cooling technology.

Backfill effects on ventilation and cooling needs The use of backfill/grout was proposed for the deeper levelsfor the No. 17 Shaft project. The primary motivation for

using backfil relates to rock support considerations butthere were also very significant benefits related toventilation, cooling and reduction in fire risk.

The application of backfill assists ventilation and coolingby reducing the exposed surface area of hot rock in work-out areas thus reducing heat load and by assisting with in-stope ventilation control. The proposed backfill system willbe applied from below 25 level in the form of ‘paddocks’aligned on dip with ultimate cover of nominally 30%.Specifically the stope-backs 25 to 26 level, 26 to 27 leveland 27 to 28 level will have backfill/grout.

The prediction of the reduced heat load with backfill wasevaluated using the project design VUMA modelling. Themodelling assumed the design parameters used in the fridgeshaft option of primary ventilation capacity of 1 300 kg/sand surface bulk air cooler duty 56 MW.

It is important to note that the proposed refrigerationsystem has zero underground based refrigeration or aircooling equipment. This is a specific feature of the designthat relies on the use of a dedicated ultra-cold ‘fridge’ shaftto deliver cooling to the deeper (25 to 28 level) zone. Theproposed use of backfill in this zone assists in allowingsufficient cooling to be distributed in this manner.

In the proposed cooling system, the primary ventilationrating was based on 5 kg/s per kt/month and the cold airtemperature of 4°Cwb (average No. 17M and No. 17Fshafts). The point to note is that the primary ventilationfactor is relatively high and the cold air temperature is low—there is not much more cooling that can be introduced bythe medium of cold ventilation. Any additional heat load(for whatever reason) will require the introduction of thenext regime of cooling with a step-change in costs andcomplexity. The next stage of introducing cooling will be inthe form of chilled water from surface. While the proposeddesign will satisfy the cooling requirements (withappropriate design safety considerations), if backfill is notapplied the heat load will increase and it will be necessaryto introduce chilled water from surface.

The project design VUMA models were run with thebackfill input removed. The models indicated that withoutbackfill (in the same mine definition) the requiredunderground air cooling would be about 5 MW (air coolerduty) in addition to the proposed surface system. Thiswould require some 100 l/s of chilled air-cooler-water tounderground. The service water flow of 40 l/s (designvalue) would also be chilled and the total chilled water to beproduced on surface would need to be 140 l/s. This wouldfirst be cooled in precooling towers and then chilled by anadditional 10 MW refrigeration plant capacity on surface.

The associated capital penalty will be related toequipment which would typically include: precoolingtowers, additional refrigeration machines, additional surfacedams, insulated shaft column (300 mm), energy recoverystation and dam, insulated chilled water pipe network,underground air coolers, return water dam and pumps,return pump column (300 mm), etc.

The order-of-magnitude capital estimate for thisequipment will be about R50 to 60 million—which is about25% of the currently proposed refrigeration system. Theoperating costs would also increase. For example, theadditional refrigeration plant power (compressors,accessories) will be about 2.5 MW, the additional pumppower underground will be about 3.5 MW (energy recoverywill not be more than 1.5 MW). The use of backfill willallow significant savings in refrigeration related capital andoperating costs. The above is strictly an order-of-magnitude



consideration. There are many secondary issues such as: in-stope vent control, main fan pressures, effects of backfilldrainage, possibility of cooling backfill, etc., which havenot been addressed. But these issues will not change themain order-of-magnitude conclusions.

ConclusionsThe ‘fourth generation’ shafts pose some seriousgeotechnical and ventilation mine design challenges.Adequate mine design information at prefeasibility studyresulted in a mine design of high technical integrity.Identification of the project design risks at the prefeasibilitystudy phase allowed sufficient time to investigate the risksand how to effectively mitigate for these risks. The use ofenhanced support resulted in a working cost increase of upto 10% but reduced the commencement period of theproduction build-up by up to a year. The realized financialvalue significantly outweigh the support cost. There wasalso significant saving by adopting the use of a dedicatedfridge shaft and the use of backfill for the three levels of thedeeper section.

AcknowledgementsI would like to thank the Impala Mine management forallowing me to publish this paper. I would also like toextend my acknowledgements to the Impala Projects

technical team, in particular Les Gardner and Johan Horn,for their contribution.

References

1. GARDNER, L. Project report reference: MPRE/09/07/REM/LJG June 2007.

2. BLUHM BURTON ENGINEERINGCONSULTANTS. No. 17 Shaft PrefeasibilitySummary of Ventilation and Cooling Design Criteria,options, decisions and constraints’.

3. JAGER, A.J. and RYDER, J.A. A handbook on rockengineering practice for tabular hard rock mines, TheSafety in Mines Research Advisory Committee, CredaCommunications, Johannesburg, 1999

4. GOLENYA, F. Extended geology abstract feasibilitycontribution, Internal report reference 17SP/06/07,Impala Platinum Limited, 2007

5. GARDNER, L.J. Support requirements for the sinkinkphase of the No. 17 shaft project, Internal reportreference MPRE/13/06/REM/LJG, Impala PlatinumLimited, 2006

Appendix 1—Geological sucession tables



Appendix 1—Geological sucession tables (continued)

Luke ZindiProject Director: Greenfields, Impala Platinum

Luke is currently employd by Impala Platinum Mines as Project Director: Greenfields. He isresponsible for ensuring the completion of the feasibility study on new mining projects beforepresenting the project to the Board of Directors and the implementation of the green-fields miningprojects at Rustenburg Lease area since January 2006.

Prior to this he was with Anglo Platinum as a senior planning analyst: projects. Luke providedcomprehensive project evaluations to operations and corporate office. He contributed to thetechno-economic inputs as part of a multi disciplinary team working on green-fields project

investigations, feasibility studies and hand over of approved projects. Before this he held several positions in mineral resourcemanagement and mine planning in Anglo Platinum and other metalliferous mines since 1984 after graduation at ComborneSchool of Mines.


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