32 kurnabinna terrace tel : 08 8387 2725 · gcpl room & pillar designs applying the stability...

32 Kurnabinna Terrace Hallett Cove

South Australia 5158

Tel : 08 8387 2725

m: 0407 447 878 e: [email protected]

Geotechnical Consulting Pty Ltd ABN: 49 087 658 213

GBS Gold Australia (Pty) Ltd

TOM’S GULLY

UNDERGROUND PROJECT

GROUND CONTROL MANAGEMENT PLAN

30 AUGUST 2008 Report GCPL-TG-300808

mailto:[email protected]

GCPL-TG-300808 Tom’s Gully Ground Control Management Plan of 89


1.0 SUMMARY

This Ground Control Management Plan (GCMP) document outlines the anticipated geotechnical conditions and presents the stability assessments undertaken to derive geotechnical excavation and support designs appropriate for GBS Gold’s Tom’s Gully Underground Mine. The document details current knowledge of rock mass conditions at Tom’s Gully, the related geotechnical controls on stability, and presents the recommended ground control management strategies which will be employed to maximise ore recovery in a safe and cost efficient manner. As the ore reef dips at less than 200, a room and pillar mining method has been implemented. In-line with the geotechnical conditions anticipated, appropriate room and pillar dimensions are also provided in this report. The GCMP has been compiled at the request of GBS Gold’s Mine Design Team by Geotechnical Consulting Pty Ltd (GCPL). It is intended that the GCMP will be developed as a Major Hazard Standard for improved strata control, for proactive application by GBS employees and contractors and to meet the approval of the NT Mines Inspectorate.

An appreciation of the geotechnical conditions for Tom’s Gully excavation & support designs and for the compilation of this GCMP has entailed the following process:

GCPL Review of previous Geotechnical and Mining Consultant Reports (References 2 to 6);

Inspection and mapping of rock mass exposure underground by GCPL on 21st February, 9th June, 3rd July and 8th August 2008 : undertaken in order to gather existing rock mass and structural information and to gain an appreciation of geotechnical conditions, development and stope stability, first-hand underground;

GCPL review and processing of geotechnical information contained in 60 TGD diamond drillhole logs compiled by Renison Consolidated during 2004. This information had not been fully applied by previous geotechnical consultants – only RQD had been put to use;

GCPL collation, processing and application of laboratory UCS, rock strength estimate, fracture frequency, number of joint set and joint shear strength data (previously not considered);

Existing drillhole geotechnical databases were modified and updated by GCPL to include rock mass classification Q’, GSI, Q, N’, and stope maximum stable unsupported hydraulic radius (HR);

GCPL Rock Mass Classification (Barton & Grimsdad 1994);

Compilation of Q, GSI, rock mass strength and HR contour plots in RockWorks 2006;

Compilation of pillar strength contour plots in RockWorks 2006;

GCPL assessment of Hoek-Brown rock mass strength for application in Phase2 numerical modelling;

GCPL room & pillar designs applying the Stability Graph Method (Hutchinson and Diederichs, 1996), together with Phase2 finite element numerical modelling for confirmation of unsupported stable spans, pillar dimensions and decline stand-off distance from stopes;

Technical discussions on-site with GBS Geologists and Mine Design Engineers;

Subsequent geotechnical analysis, interpretation and reporting by GCPL.

It should be noted that geotechnical recommendations provided in this report are based on the available information and drillhole data outlined previously. Clearly, there are limitations to this data, and for this reason, the recommendations on mine layout and ground support provided, may require modification as actual experience with ground conditions is obtained during mining. Due to the absence of in situ stress and structural information, it is emphasised that whilst every care has been exercised in the process, the geotechnical analyses carried out in this report are based upon conditions currently exposed, the performance of mined stopes and engineering judgement best estimates.

Being a ground control management document, for the sake of completeness, excerpts from earlier geotechnical and mining consultant reports are included in this GCMP document (References 2 to 6).



GCMP CONTENTS

1.0 SUMMARY 3

2.0 OBJECTIVE, SCOPE & SYSTEM STRUCTURE 5

3.0 GEOLOGY, GROUND CONDITIONS & MINING METHOD 9

4.0 MINING PROGRESS TO AUGUST 2008 11

5.0 GEOTECHNICAL MODELS & BASIS OF EXCAVATION & SUPPORT DESIGNS 12

5.1 Geotechnical Data from Resource Diamond Holes 5.2 General Outline of Ground Conditions & Structural Controls on Stability 5.3 Major Structures & Groundwater 5.4 Intact Rock Strength & Laboratory Testwork 5.5 Rock Mass Classification

5.5.1 Barton’s Rock Quality Q 5.5.2 Maximum Unsupported Spans 5.5.3 Design Rock Mass Strength

6.0 STOPE : ROOM & PILLAR DESIGN 31

6.1 Estimated Stress Field & Pillar Loads 6.2 Pillar Strength 6.3 Room Spans 6.4 Phase 2 Numerical Modelling 6.5 Room & Pillar Layout

6.5.1 Mining Block 1 6.5.2 Mining Block 2 6.5.3 Regional Pillars 6.5.4 Integrity of Crown Pillar

7.0 GROUND CONTROL STRATEGIES 47

7.1 Development Primary Support

7.1.1 Roof Beam Theory 7.2 Cable Bolting of Wide Span Intersections 7.3 Grout Mesh Packs 7.4 Ground Support Quality Control 7.5 Excavation Stability Monitoring 7.6 Check-Scaling 7.7 Safety & Duty of Care

8.0 CONCLUSIONS & RECOMMENDATIONS 57

APPENDICES

A CALCULATION OF TGD DRILLHOLE Q-N’-HR B PILLAR STRENGTH CALCULATIONS C TOM’S GULLY STANDARD GROUND SUPPORT SYSTEMS D BOLT TECHNICAL SPECIFICATIONS



2.0 GROUND CONTROL MANAGEMENT PLAN:

OBJECTIVE, SCOPE AND SYSTEM STRUCTURE

The aim of the Ground Control Management Plan (GCMP) is to document current knowledge on ground conditions at Tom’s Gully and to set out the geotechnical systems and excavation design methodologies in place to reduce geotechnical risk during mining, ie -

minimise likelihood of rockfall / excavation failure;

nullify lost time incidents (LTI);

minimise risk to safety and profit; implement the safest, most economic excavation & support geotechnical designs;

optimise blast outcome;

minimise ore dilution. The GCMP ensures systematic planning and implementation of effective and efficient ground control strategies in an attempt to maintain stable / safe excavations during mining. The GCMP also ensures compliance with the relevant Metalliferous Mining Regulations and demonstrates that adequate geotechnical consideration has been taken during the design, planning and development of the mine. The Mines Safety and Inspection Regulations 10.28 (1995) requires that Mine Management are able to demonstrate that they have, in effect, implemented "best practice" in geotechnical engineering during the mining process. The application of sound geotechnical engineering practice in the pursuit of safe, practical, cost-effective solutions to excavation design and stability issues is the basic requirement of Regulation 10.28. The GCMP also reflects GBS Gold's Corporate Occupational Health and Safety objectives. In order to achieve this goal, the approach recommended and currently implemented at Tom’s Gully is outlined in this Ground Control Management Plan (GCMP) document.

More definitively, this GCMP documents Tom’s Gully current geotechnical approach for :

underground mine geotechnical investigations : rock mass and structural data collection and analysis;

geotechnical modelling;

excavation design;

stability analysis;

ground support definition;

support quality control;

excavation deformation monitoring;

operational geotechnical risk reduction activities;

safe operating practice and duty of care;

ongoing monitoring, auditing and improvement to the GCMP.

Moreover, the GCMP is used :

To confirm safe excavation designs on an ongoing basis as mining progresses and as additional exposure becomes available;

To maintain a level of understanding of rock mass conditions during excavation such that any unforeseen conditions are detected at an early stage to enable corrective design measures to be implemented; and

To form contingency plans that can be implemented to minimize the impact of geotechnical issues on production.

The ground control management process is outlined in Figure 1 overleaf :



GEOTECHNICAL, GEOTHERMAL & HYDROGEOLOGICAL SITE INVESTIGATIONS & DATA COLLECTION

DATABASES:• Geological Structures;• Rock Mass;• Ground Water;• Rainfall;• Seismicity;

•Geothermal / Phreatic Pressures;

•Geothermal Temperatures;

GEOTECHNICAL / GEOTHERMAL /

HYDROGEOLOGICAL MODELS:

DESIGN : PIT SLOPE STABILITY / HYDROTHERMAL DEPRESSURISATION:• Excavation Performance Simulations : Lithostatic - Hydrothermal Stress Modelling in Tough2 and FLAC2D;

• Design Steam Relief Wells and Slope Dewatering / Depressurisation Pump-Out and Drainhole Requirements;

• Design Pit Slopes under Static & Seismic Loading : Kinematic / Deterministic Limit Equilibrium /

Probabilistic Slope Stability Analysis;

• Surface Water Management : Drainage Diversion / Seawall / Coffer Embankment Designs.

STABILITY PERFORMANCE MONITORING• Slope Displacement Monitoring : Quickslope Prism ATS,

Condor GPS, Inclinometers, Extensometers & Crack Monitors;

•CSIRO Micro-Seismic Monitoring to provide early

warning of geothermal outburst or seismic event;

• Thermistor and pressure transducer

hydrothermal depressurisation monitoring

to provide early warning of geothermal

outburst.

MINE WATER MONITORING•Rainfall / Precipitation Gauges;

•Surface Water & Drainage;

•Pit Dewatering / Depressurisation

•Ground Water Levels – VWP’s;

GEOHAZARD ID & RISK MANAGEMENT

IS LEVEL OF

GEOHAZARD

RISK (Likelihood

& consequence)

ACCEPTABLE?

STABILISATION / PREVENTION STRATEGY

• Amend Slope / Geothermal Depressurisation Design;

• Improve Surface & Groundwater Drainage;

• Ground Support;

• Implement Cut-Back;• Abandon Area of Pit.

SAFETY & DUTY OF CARE•Safe Operating Practice, SOPs, JSAs;

•Crew Awareness;

•Ongoing Geotechnical Safety Inductions.

NO

YES

MONITORING or CONTINGENCY STRATEGYAccept Risk on basis of ability to predict (monitor), and

accommodate potential failure without endangering

personnel & equipment. Devise a contingency strategy

which protects personnel by specifying responses to an

identified, uncontrolled hazard situation.

Critical Risk Level > 20

High Risk Level < 20

Figure 1 : Ground Control Management Process

2.1 Ground Control Management Responsibilities and Accountabilities

The following paragraphs set out details of the personnel responsible for ensuring that the GCMP is implemented successfully (Table 1).

Manager Underground Mining

The Manager Underground Mining as appointed under the Act must ensure that:

The GCMP is implemented and all regulatory requirements are met;

Resources are made available to implement and update the management plan when required;

Geotechnical aspects are adequately considered in relation to mine design and planning;

Geotechnical aspects are adequately considered;

Ongoing ground control training is available and encouraged for all mining personnel. Areas of need in training of these personnel are identified and auctioned. This is to facilitate an understanding of excavation behaviour, ground control principals and practices;



Standard Work Procedures (SWP’s) relating to excavation stabilisation are implemented, monitored, and modified when needed, collaboratively amongst all relevant personnel;

Annual audit, review and continual improvement of the GCMP is undertaken and satisfactorily completed by the geotechnical engineer or external consultant;

Suitable equipment is supplied and maintained to the specifications required for quality ground control;

Audit, review and quality assurance programs are carried out regularly and documented.

Mining Engineer

The Mining Engineer must ensure that:

Development and room and pillar stope designs are implemented according to plan;


All mine design and planning accounts for the geotechnical recommendations as documented and agreed with between the Geotechnical Engineer, Senior Mine Geologist and Manager Mining.

Mine Geologist

The Mine Geologist must ensure that:

Litho-Structural and hydrogeological information pertinent to the geotechnical aspects of excavation design are collected, analysed, interpreted and communicated. The position of faults, shears and the hangingwall carbonaceous breccia parting seam is particularly important for excavation design at Tom’s Gully, and wireframes should be constructed by geology for application by the geotechnical engineer;

Geotechnical data is logged during the course of any resource diamond drilling;

Information is regularly updated and communicated;

The GCMP is implemented and all regulatory requirements are met.

Geotechnical Engineer (Consultant)

The appointed Geotechnical Engineer or Geotechnical Geologist must ensure that:


Geotechnical data sufficient to provide adequate input for mine design and sequencing is collected, analysed, interpreted and communicated;

Timely advice is given to the planning team for the purpose of mine design and planning;

Excavation performance information is recorded to support mine design and planning;

Geotechnical aspects of proposed mine scheduling are considered, with sufficient notice to consider practical alternatives;

Timely support to the Manager Mining and the mine planning team is given in all geotechnical aspects;

A geotechnical model is maintained, based on the geotechnical setting and the mining environment;

Excavation and Ground support design is based on the geotechnical model;



Excavation stability assessments are to be carried out frequently, monitored and actioned when required;

All rock falls and slips are to be inspected and the standard reporting procedure is to be enacted;

Excavation Stability audits, review and quality assurance programs are carried out regularly and documented;

Ground support designs are sufficient for the current and expected ground conditions;

The GCMP is audited and updated in accordance with current ground conditions.

Mine Surveyor The Mine Surveyor must ensure that:

Excavation Deformation / Convergence monitoring requirements as outlined in the GCMP are complied with.

Mining Contractor

The Mining Contractor must ensure that:

Development and room and pillar stope designs are implemented according to plan;

The work places and their access are maintained in a safe condition;

Ground conditions at each heading are inspected and any loose rocks are barred down;

SWP’s are understood and followed;

Active headings are inspected at least three times during the operating shift and ground conditions recorded whilst monitoring continually throughout the shift;

Ground support is installed to the standard and quality specified in the GCMP, and as directed by the Mining Engineer;

Geotechnical / rock fall hazards are identified and actions are taken to mitigate or eliminate the hazard;

Reports on rock falls, slips and variation in ground conditions are communicated effectively to improve awareness amongst all personnel;

Inactive areas of the mine are cordoned off and access restricted.

OHS Superintendent

The OHS Supervisor must ensure that:

Adequate ongoing training is provided to all contract mining personnel to allow an understanding of ground / excavation behaviour, principals and practices;

SWP’s relating to ground control and excavation stabilisation are implemented, monitored, and modified when needed, collaboratively amongst all relevant personnel; and

Reports on rock falls, slips and variation in ground conditions are addressed and reports distributed as required.

All employees

All employees must ensure that:

No work is undertaken without a plan;

Only work in line with current competence is undertaken;

SWP’s are followed at every work place and active headings are inspected in line with Workplace Shift Inspections;



Ground conditions at each heading are inspected and any loose rocks are barred down;

Ground conditions are monitored during the shift for instability, cracking and general ground movement;

If ground conditions become hazardous, barricade and vacate the area, and immediately notify the Production Supervisor;

Other relevant information in relation to Ground Stability is reported back to the Production Supervisor.

TABLE 1 : APPOINTED GROUND CONTROL MANAGEMENT PERSONNEL AT TOM’S GULLY :

Appointed Positions Appointed Person

Manager Underground Mining Scott Woollard / Rod Bedggood

Mining Contractor Alliance Mining

Mining Engineer Dan Hennessy

Mine Geologist Glen McIlwain & Phil Pojev

Senior Mine Surveyor Paul Middleton

OHS Superintendent Belinda Holt

Geotechnical Engineer / Consultant Geotechnical Consulting Pty Ltd

3.0 TOM’S GULLY GEOLOGY,

GENERAL GROUND CONDITIONS & MINING METHOD

The Tom’s Gully deposit is hosted by the Early Proterozoic Wildman Siltstone, which is a thick unit comprising laminated graphitic shale, carbonaceous and often pyritic siltstone inter-bedded with undifferentiated volcanics and minor dolomitic sediments. The deposit occurs as a 0.5 to 2m thick quartz reef within a mineralised fault zone that dips to the SSW at between 00 to 200. Reef dip generally flattens to the south of 4600N but localised open folding is also encountered in this area. It is thought that the fault was initially developed and was later mineralised. The mineralised fault may be part of an en-echelon vein system which is thought to pinch and swell. (Ref 12). Gold generally occurs as free grains, associated with arsenopyrite. The main mineralised zone is located immediately west of the westerly dipping (600W), NNE-SSW striking Crabb Fault, which is the main structural feature along the eastern margin of the deposit and the Williams Fault to the West. The reef has not been located beyond the eastern contact with the intrusive Mt Goyder Syenite. Western extensions of the reef are possible but exploration to-date has failed to identify the system beyond 800m west of the Crabb Fault. The mineralised part of the reef has a strike length of about 300m and a horizontal dip length of at least 1200m. On the basis of orebody geometry and ground conditions, the area north of 4700N has been designated Block 1 and that south of 4700N, as Block 2. Rosengren (Ref 5) describes Block 1 as having relatively steep dip and poor ground conditions with Rock Quality Designation (RQD) consistently less than 50%, whilst Block 2 to the south has RQD values consistently greater than 50% with good ground conditions. This is confirmed in the GCPL Rock Quality Contour Plots. According to Rosengren, in Block 1 there are adequate natural pillars within the identified economic areas of the reef and hence no additional regional pillars are required. However, a 10 m wide pillar will be required adjacent to the Williams Fault and between any working and the pit. The latter is particularly important with respect to the decline passing beneath the existing pit floor. This is also confirmed in the GCPL Phase 2 numerical modelling of the crown pillar area (Section 6.5.4). The flat dip of the reef has made it virtually essential to utilise room and pillar stoping at Tom’s Gully. As shown in the current mine plan (Figure 2) this involves developing a system of contour drives (5.0 mW x 4.2 mH), along the strike of the reef. The ore between these drives is then extracted partially, leaving a pattern of residual pillars to support the room hangingwall. In previous mining, Block 1 contour drive



spacing has varied from 5 to 8m spacing (edge to edge), with approximately 7 x 6m rooms supported by 4 x 5m pillars. Airleg stoping has been adopted in Block 1. To minimise longhole deviation, and reduce overbreak, Block 2 current design has the contour drives spaced at 14m on dip (edge to edge) and 5 x 5m pillars spaced at 15m on strike (edge to edge). Pillars are staggered between each level, effectively resulting in a room size of 14 x 15m (HR=3.6) and a pillar load tributary area of 19 x 10m.

Figure 2 : Current Mine Plan for Blocks 1 & 2 with Areas Developed in Blue Development is undertaken by twin boom jumbos, utilising resue firing methods to reduce dilution of the ore body. To minimise dilution during development, all drives are mined using a two-pass resue firing sequence. The waste in the lower section of the face is fired and bogged first, followed by a second firing and bogging cycle to remove the largely undiluted ore. Winzes are then developed between the adjacent contour drives, creating free-face slots to fire into. Production long-hole rigs then drill out the stopes in each stoping block. Stope panels are then drilled & fired into the ore drives. Mini remote dozers will be used to push any ore remaining in the stope into the ore drives, which will then be bogged remotely. The mining sequence will progress top down with stopes extracted from the east retreating to the west. The sequence of activities is as follows:

1. Development of the decline and access; 2. Development of contour ore drives; 3. Production drilling; and 4. Extraction of stopes leaving remnant pillars to support the hangingwall.

The main decline (5.0m x 5.0m) has been designed within the western boundary of the reef. An access decline runs parallel to the main decline, 8 to 10m to the east of the main decline, providing a path for return air in the primary ventilation circuit and the second means of egress for the mining block.

BLOCK 1

BLOCK 2



Ground conditions will be continually monitored and reviewed to assess the long term regional pillar requirements. All dilution for the Tom’s Gully orebody has been estimated using the ELOS (Equivalent Linear Overbreak or Sloughing) approach. This is a distance overbreak measure of dilution. An assumed amount of footwall and hanging wall distance of dilution is applied to the cross sectional area of the stope to create a dilution volume that is applied to each mining source. A dilution distance of 0.2 metres of hanging wall and 0.2 metres of footwall dilution has been applied to each stope mining source (Ref 2).

4.0 MINING PROGRESS TO AUGUST 2008

At the time of the GCPL August 2008 visit, in Block 2, decline development had progressed to approx RL815 and Level 825 Access Development was 60% complete with initial development in the A, B, C and D Ore Drives ongoing (Figure 3). Airleg stoping in Mining Block 1 was ongoing with activity in the 905AW & BW rooms. Areas mined in the RL910 and 905 airleg stopes to August 2008 are shown in Figure 4.

Figure 3 : Block 2 Primary Development Schedule for Fourth Quarter 2008

825 ACCESS

DECLINE

A

B

C

D

VENT



Figure 4 : Block 1 : Contour Ore Drive Development and Areas Stoped (August 2008)

5.0 GEOTECHNICAL MODELS & BASIS OF

EXCAVATION & SUPPORT DESIGNS

All of the data gathered as part of the original Renison Consolidated Geotechnical Investigation Programme, as well as the work of earlier geotechnical consultants (Ref 3 to 7) has been applied by GCPL to produce geotechnical models reflecting the range of structural and rock mass conditions that might be anticipated during mining at Tom’s Gully. The following software was utilised in the geotechnical modelling, excavation stability and design process:

RockWare 2006;

Gemcom-Surpac 2006;

Rocscience (2005) DIPS & PHASE2. For incorporation into geotechnical analyses, the as-mined and current design str & dtm files (August 2008) were provided by Mr Dan Hennessy (GBS Gold Mining Engineer), and subsequently imported into RockWare and Gemcom-Surpac by GCPL. 5.1 Geotechnical Data from Resource Diamond Drilling

The localities of the 60 TGD diamond holes drilled by Renison Consolidated during 2004 are shown in Figures 5, 6 & 7. The geotechnical information gathered during this programme, had not been fully applied by previous geotechnical consultants (Ref 2 to 7) – only RQD had been put to use. As part of this study, to derive updated geotechnical models, GCPL have therefore included all laboratory UCS, rock strength estimate, fracture frequency, number of joint set and joint shear strength data.

910 STOPES

905 STOPES

VENT DRIVE

VENT DRIVE

DECLINE



Figure 5 : Mining Block 1 : TGD Resource Holes from which Geotechnical Information was Sourced

Figure 6 : Mining Block 2 : TGD Resource Holes from which Geotechnical Information was Sourced



Figure 7 : TGD Drillhole Traces Superimposed on Current Block 1 & 2 Mining Layout



5.2 General Outline of Ground Conditions and Structural Controls on Stability

Mining Block 1 : Work by previous consultants (Refs 3 to 7), as well as more recent inspections underground by GCPL has confirmed that rock mass quality in Block 1 classifies as poor with a measured Barton’s Q of between 1 and 4. The process utilized in the calculation of Q is outlined in Section 5.5.1, but the contrast in rock quality (Poor in Block 1 & Good in Block 2), is clearly visible in the GCPL contour plot of Rock Quality (Q) derived from the 60 TGD diamond holes, for the Hangingwall in Figure 8. The marked difference in quality may be attributed to the seasonal fluctuation of the ground water table, leaching and weathering within the shallower Block 1 rock mass.

Figure 8 : Contrast in Rock Quality between Mining Blocks 1 & 2 (First Quartile, 10m into Hangingwall)

BLOCK 1

BLOCK 2



In Block 1, the rock mass is generally moderately to highly weathered and is weak to medium strength (ie UCS = 23 to 42MPa). The principal plane of weakness controlling hangingwall stability in-stope, is a flat-dipping, brecciated graphitic / carbonaceous bedding shear seam which occurs variably 0 to 0.5m into the hangingwall and sometimes bisects the orezone itself. These very weak 10 to 300mm thick shear zones are generally comprised of brecciated shale fragments and fine clayey silt gouge and act as conduits for ground water flow. Being lystric, they wander and undulate along dominant bedding, and are at times difficult to pinpoint. Where systematic bolting is not in place (such as is currently the case in the hangingwall of the 905 airleg stopes), overbreak into the hangingwall to the sheared plane of weakness, and rock fall risk becomes difficult to control. (Photo 1). As witnessed in the RL905 stopes, as spans approach 10m, there is general convergence of the backs to a parting carbonaceous bedding shear. This has resulted in sudden failure of the hangingwall. Without systematic bolting and / or some form of temporary propping, there is a very real danger of sudden rock fall and a high risk of exposure to the airleg miners. This aspect therefore requires immediate corrective action by GBS and Alliance Mining. There are also associated external waste dilution issues. For the Block 1 access stopes, check scaling must be ongoing and systematic bolting and / or props (to minimise rockfall risk) will be required to render support to the hangingwall;

However, there is room for improvement in perimeter blasthole accuracy, and a ring of grouted split-sets installed as close to the advancing face as possible, may prevent initial delamination.

Photo 1 : RL905 AW Airleg Stope : Weathered Rock Mass & H/W Overbreak to Parting Bedding Shear

Mining Block 2

As shown in Figure 8, the rock mass in Block 2 is generally of good and better quality with Q values in the range 10 to 80. During GCPL’s recent August 2008 inspections, the decline in Block 2 had progressed to RL815. With the exception of flat-dipping, undulating brecciated carbonaceous shear and calcite zones where conditions are locally poor (and wet), ground conditions along the decline in the competent Wildman



Siltstone remained good. Localised overbreak to continuous sub-vertical jointing striking parallel to the decline (“195-strike-structures”) was observed. Under such conditions, “Ground Support Standard 6” currently being installed to the arched tunnel profile, is deemed to be appropriate. This currently comprises F51 Weldmesh and 2.4m length split-sets installed to within 3 meters of the tunnel invert. (Photo 2).

Photo 2 : Surface Support Standard 6 Installed Along Decline @ RL815

Rock quality along the 825 access drives was also observed as being generally good with the drive shanty backs closely following competent bedding stratification (Photo 3).This is also presently the case at the wider span access / ore drive intersections.

Photo 3 : Mining Block 2 : Ore Contour Drive @ RL825 : Square Profile & Shanty Back in Good Ground



Ground conditions in the competent footwall siltstones of the Block 2 825 Ore Drives A, B and C are therefore generally good, with widely spaced jointing and a measured UCS of 115MPa (ie:strong to very strong rock). During the multiple geneses of deformation, the orebody itself (which will form the room pillars), has behaved in a more brittle fashion and is more closely fractured and jointed as a result. However the rock still classifies as strong with a measured UCS of 84MPa on average. A Digital Schmidt Hammer was used to measure the intact rock strength insitu (Section 5.4). From an excavation design and performance perspective, the poor conditions associated with the flat-dipping, undulating, brecciated carbonaceous shear and calcite zones, require consideration (Photo 4). These “weak seams” of cataclastic rock fragments and clay gouge will in effect reduce overall pillar strength and act as release structures for the convergence of hangingwall and footwall strata into the stope void.

Photo 4 : RL825-B Ore Drive : Back Overbreak to Parting Bedding Shear

5.3 Major Structures & Ground Water

The Block 1 & 2 ore reserves are essentially bounded by two major faults and an aplitic dyke, all striking NNE – SSW, almost perpendicular to the strike of the reef. These are:

Crabb Fault: forming the eastern boundary, orientated approximately 600 WNW. Displacement along this fault is not clear. Poor ground conditions associated with the fault, led to the collapse of the original decline to the underground;

Williams Fault : near the western boundary (there is some mineralisation west of the fault). This fault is steeply dipping and has an apparent displacement of some 10m;

Williams Dyke : orientated approximately 600 NW.

All three structures are probably major aquifers, and any future development through them would require the necessary cover drilling, with possible dewatering and depressurisation prior to tunnelling. (Mine dewatering is currently based on a pump duty of 15L/s, with three pump stations recommended at various positions down the decline with smaller pumps or drainholes feeding these pumps. Pumping to the surface will be either through vertical rising mains or staged pumping through a main line up the decline – Ref 2).



Associated with the major faults is a sympathetic joint set running parallel and striking NNE – SSW (ie 1950). This dominant joint set is synonymous with the “195-structures” referred to in earlier Coffey Reports (Refs 3 to 5). In addition, a series of through going structures which are worth noting due to their potential influence on orebody displacement, ground conditions, and generally on mine planning, were also recently intersected in Block 2 development :

The first is a “195-striking” fault with normal displacement which was intersected in the 830SP and in initial ledging of the 825B Ore Drive. The fault has a dip-slip throw of approximately 1 meter (Photo 5), and is orientated almost parallel to the Crabb and William Faults (Figure 9). The fault strikes through the 825 Access and C Ore Drive Intersection which has a developed span of approximately 10 meters;

The second is a dolerite / lamprophyre dyke which has been intersected at the 825 Access – C Ore Drive Intersection. This strikes along and adjacent to the access drive. Although a wireframe was not available at the time of the GCPL visit, current geological interpretation is that it is sub-vertical, dipping to the east. Ground conditions are fair along the dyke;

The third structure is synclinal – anticlinal folding exposed on the southern sidewall of the 820 Access Drive (Photo 6).

Photo 5 : Normal Fault Intersected in initial ledging of the 825B Ore Drive



Figure 9 : Normal Fault (Navy Blue) Strikes Parallel to Crabb & Williams Faults

Photo 6 : Synclinal – Anticlinal Folding and associated Poor Ground Conditions Exposed on the Southern Sidewall of the 820 West Access Drive

Williams Fault

Crabb Fault

“Normal” Fault



5.4 Intact Rock Strength & Laboratory Testwork As part of the 2004 Dr Rosengren Geotechnical Review for Tom’s Gully (Ref 6), 11 (Eleven) core samples were submitted to James Cook University for Uniaxial Compressive Strength (UCS) tests with strain gauge deformation modulus and density measurements. The results are outlined in Table 2. The outcome of this testwork led Rosengren to deduce that “the results are low in relation to the nature of the rock and indicate the requirement for a conservative approach to pillar design prior to pillar extraction.”

Mining Block

Mining Horizon

Resource Drillhole

Depth (m)

Bulk Density (t/m3)

UCS (MPa)

Youngs’ Modulus Ei (GPa)

1 Hangingwall TGD355A 175.70 - 175.86 2.69 70 5

2 TGD380 236.18 – 236.34 2.76 68 7

Sample Mean 2.72 69 6

1 TGD360 135.69 – 135.85 3.10 38 19

1 TGD362 158.16 – 158.32 2.85 76 17

2 Reef TGD380 236.90 – 237.06 2.72 36 27

2 TGD410 237.68 – 237.84 2.76 68 17

2 TGD386B 241.57 – 241.73 2.91 49 12


1 TGD362 158.30 – 158.46 2.77 58 14

2 Footwall TGD380 239.48 – 239.44 2.70 26 4

2 TGD386B 243.22 – 243.38 2.80 49 8


Population Mean 2.81 54 13

TABLE 2 : Laboratory Strength Test Results (Ref 6)

In order to provide an indication of the reliability of the test results, a programme of Digital Schmidt Hammer R-Rebound - UCS tests were carried out on the in situ rock, underground by GCPL. As reflected in Table 3, 87 such tests were conducted in Block 1 and 57 in Block 2 on footwall, ore and hangingwall siltstone strata. As shown in Table 4, with the exception of the Block 2 footwall, there is reasonably good correlation between the Laboratory and Schmidt Hammer Results. The lab-tests suggest a surprisingly low value of only 26 to 49MPa for the Block 2 footwall units as opposed to 115MPa recorded by the Schmidt hammer. The first quartile values outlined in Table 4 overleaf were generally selected for design purposes. For geotechnical design purposes, application of the first quartile (as opposed to mean) is more realistic, as it is the weaker rock mass that is obviously more prone to failure.



TABLE 3 : Schmidt Hammer R-Rebound Numbers & Equivalent UCS Statistics



TABLE 4 : UCS Statistics from Schmidt Hammer Compared with Lab Results (Brackets)

5.5 Rock Mass Classification

The strength of a jointed rock mass is dictated by the properties of the intact rock blocks and by the freedom of these pieces to slide and rotate under different stress conditions. Determination of the strength of an in situ rock mass by laboratory type testing is limited by sample size – testing of the overall rock mass is impractical. Rock mass strength must therefore be estimated from drillcore, geomechanical observations underground and from test results on individual rock samples and discontinuity surfaces. Rock Mass Classification is used in the quantitative assessment of ground conditions. Over the years, various rock mass classification systems have been developed in an attempt to quantify rock quality and rock mass strength for application in excavation, support and blast design and in dilution control studies. Case histories of excavation performance viewed in the light of mining rock mass conditions have facilitated the development of quantification systems and empirical design charts. These are updated on an ongoing basis taking new cases into account so that the systems are constantly being improved. GCPL have applied the following rock mass classification systems for Tom’s Gully :

Norwegian Geotechnical Institute’s (NGI) or Barton’s Q system;

Geological Strength Index (GSI);

Mathews-Potvin Method for Open Stope Design;

These systems have been applied in the following areas of geotechnical design:

Derivation of Hoek-Brown Rock Mass Strength parameters;

Ground Support Design;

Stope Design;

Ongoing review and validation of current excavation design.

Mining Block

Stats F / W Siltstone

(MPa)

Quartz Ore

(MPa)

Calcite Parting (MPa)

Carbonaceous Brecciated

Shear Gouge (MPa)

H / W Siltstone

(MPa)

1

N 23 35 -

8 21

Min 20 20 1 22

Mean 60 (58) 35 (38-76) 2 41 (70)

Max 150 150 3 76

SD 14 14 1 13

First Quartile 42 23 1 28

2

N 33 16 9 3

Min 28 28 24 1

Mean 115 (26-49) 84 (53) 51 2 (69)

Max 280 270 140 3

SD 18 21 15 1

First Quartile 75 34 36 1



5.5.1 Barton’s Rock Quality Q As discussed briefly in Section 5.2 and as shown earlier in Figure 8, the geotechnical data from 60 TGD drillholes was processed and applied by GCPL in the calculation of Barton’s Q in order to assess the spatial distribution of rock quality and rock mass strength for room & pillar design and layout. The processed Q-data is provided in Appendix A. The reliability of the drillhole data was also verified through window mapping underground by GCPL which has mainly been limited to footwall development with no appreciation of conditions in situ in the hangingwall. For the reef and footwall horizons, positive correlation of drillhole information with the rock mass conditions mapped in situ has been encouraging. In the Q-System, the rock mass is classified according to degree of jointing / 'blockiness', (RQD and number of joint-sets), joint shear strength (roughness, shape and infill) and stress environment in relation to material strength, including the effects of ground water. The Q-System parameters are defined as follows:

SRF

J

J

J

J

RQDQ w

a

r

n

(Equation 1 : Barton 1974);

where,

RQD = Rock Quality Designation;

Jn = Joint set number;

Jr = Joint roughness number;

Ja = Joint alteration number;

Jw = Joint water reduction factor; and

SRF = Stress reduction factor. The process is outlined in Table 5 below, which shows the results from the recent window mapping of the RL825 B & C Ore Drives. Similar values were obtained from nearby drillholes TGD 255 & 258.

TABLE 5 : RL825 Ore Drive Q-Assessment

On the RL825 footwall access level, the rock mass is strong to very strong (115MPa), moderately to widely jointed and is of good quality. Three joint sets were noted, ie:



Bedding : 100 to 170 / 1950 (Dip Angle / Dip Direction);

Reef Strike Jointing 890 / 0300;

Reef Strike Jointing 890 / 2000;

Reef Dip Jointing 760 / 3200. Similar orientations were recorded along the ore zone but bedding and jointing is more closely spaced (0.2 to 1.0m), and rougher. The mapping indicated the following range of conditions:

RQD = 20 to 100%, bedding spacing = 0.2 to 1.0m;

Jn = Number of Joint Sets : (2 to 3 joint sets developed ie : rating = 4 to 9);

Jr = Joint roughness : slickensided undulating (1.5) in bedding shears but generally smooth (2) and rough undulating (3) elsewhere;

Ja = Joint alteration : unaltered joint walls (1) with bands of silty clay gouge (5) along shears;

Jw = Joint water reduction factor : generally dry (1) but with medium inflow along shears;

SRF = Stress reduction factor : Generally favourable stress condition (1) with weak bedding shear zones containing clay gouge (2.5).

In summary:

Work by previous consultants (Refs 3 to 7), as well as more recent inspections underground by GCPL has confirmed that rock mass quality in Block 1 classifies as poor to fair with a measured Barton’s Q of between 1 and 4, and a Bieniawski’s RMR of less than 55;

As shown in Figure 8, the rock mass in Block 2 is generally of good and better quality with Q values in the range 10 to 80 and a RMR of 65 to 85, but with very poor conditions encountered along flat dipping bedding shear zones.

5.5.2 Maximum Unsupported Spans In applying those Q & RMR values, the Bieniawski and Barton-Grimsdad (1993) empirical design charts may be referred to for initial assessment of unsupported spans for development and stopes. Note that in (Figure 10), an ESR = 3.0, has been selected. The Maximum Unsupported span may be assessed from the following equation, where ESR is the Excavation Support Ratio.

Maximum Span (Unsupported) = 2(ESR) Q0.4 (Equation 2 : Barton 1974);

TABLE 6 : Selection of Excavation Support Ratio (ESR) According to Excavation Function

Excavation Function ESR

A Temporary mine openings & non-entry stopes 3-5

B Permanent or re-entry mine openings, water tunnels for hydro power (excluding high pressure penstocks), pilot tunnels, drifts and headings for large excavations

1.6

C Storage rooms, water treatment plants, minor road and railway tunnels, surge

chambers, access tunnels, vent shafts. 1.3

D Power stations, major road and railway tunnels, civil defence chambers, portals &

intersections 1.0

E Underground nuclear power stations, railway stations, sports and public facilities,

factories 0.8



Figure 10 : Initial Assessment of Excavation Unsupported Span & Stand-Up Time from Q & RMR

BLOCK 1 : Q = 1 to 4,

SPAN = 6 to 10m

BLOCK 2 : Q = 10 to 80,

SPAN = 15 to 25m

BLOCK 1 : RMR < 55,

UNSUPPORTED SPAN < 10m

BLOCK 2 : RMR = 65 to 85,

UNSUPPORTED SPAN = 15 to 20m



5.5.3 Assessment of Design Rock Mass Strength Design rock mass strength for the range of conditions anticipated in Mining Blocks 1 & 2 has been assessed directly from Barton’s Q and the Geological Strength Index (GSI) classification systems and application of the Generalised Hoek-Brown Criterion and Mohr Coulomb approximation (Table 7). This has been accomplished using RocLab software by Rocscience. The range of estimated values have been applied in the Phase2 Finite Element Analyses that follow in Section 6.4. Rock mass and shear strengths for the hangingwall, reef, footwall and carbonaceous bedding shear geotechnical domains have been assessed using the Generalised Hoek-Brown Failure Criterion (Hoek et al, 2002), as follows:

1. σ1’ = σ3

’ + σci(mb * (σ3’/σci) + s)a;

where

2. σ1’ = Maximum effective stress at failure;

3. σ3’ = Minimum effective stress at failure, which are the axial and confining effective principal stresses

4. σci = Uniaxial Compressive Strength of intact rock;

5. mb = Hoek-Brown m constant for the rock mass, which is a reduced value of the intact rock mi

In most cases it is practically impossible to carry out triaxial or shear tests on rock masses at a scale that is necessary to obtain direct values of the parameters in the Generalized Hoek-Brown equation. Therefore some practical means of estimating the material constants mb, s and a, is required. According to the latest research, the parameters of the Generalized Hoek-Brown criterion (Hoek, Carranza-Torres & Corkum : 2002), can be determined from the following equations:

6. mb = mi * exp((GSI – 100)/(28-14D));

7. For GSI > 25 (Rock quality better than poor):

8. s = exp((GSI – 100)/(9-3D));

9. a = 0.5, (from original Hoek-Brown Criterion);

10. For GSI < 25 (Rock quality very poor):

11. s = 0;

12. a = 0.65 – (GSI / 200).

The Hoek Brown constant mi was obtained from published data. D is a factor which depends upon the degree of disturbance to which the rock mass may be subjected by blast damage and stress relaxation. It varies from 0 for undisturbed in situ rock masses to 1 for very disturbed rock masses. A value of 0, representative of minimal disturbance through good blasting practice was adopted for Tom’s Gully. For the moderate strength poor to fair Block 1 and high strength good quality Block 2 rock mass conditions observed underground, as well as in drillcore (Figure 8), it is reasonable to assume that the post-peak strength values can be estimated by reducing the in situ GSI value to a lower “post-peak” residual value which characterizes the failed broken rock mass. This reduction in rock mass strength would correspond with the strain-softening behavior to be expected during failure of a poor to average rock mass. Those post-peak characteristics are shown in Table 7. By comparison, a very good quality very high strength rock mass (eg >250MPa quartzites), would generally fail in an elastic-brittle-plastic fashion. When the strength of the rock mass is exceeded, a sudden strength drop occurs, associated with significant dilation of the broken



rock fragments. If confined, the failed rock would exhibit the strength characteristics of rock fill with a residual friction angle of approximately 380 and zero cohesive strength. Analysis of the progressive gradual failure of very poor rock masses suggests that the post failure characteristics are plastic, as the rock continues to deform at a constant stress level, with no volume change. The Geological Strength Index (GSI) values for each domain were obtained from equation 13, and the spatial distribution of GSI anticipated (from the 60 TGD drillholes) is shown in the contour plot below (Figure 11).

13. GSI = 9 * log Q’ + 44.

Figure 11 : Geological Strength Index (GSI) Anticipated in Mining Blocks 1 & 2 (First Quartile, 10m into Hangingwall)



+Mean Values from Digital Schmidt Hammer Tests Underground, **Mean Values from James Cook Lab Testwork.

*Derived from Table 5, Equation 13, and Figure 11. #Ratio MR = Ei / SigC.

TABLE 7 : ESTIMATED RANGE OF ROCK MASS STRENGTH APPLIED IN NUMERICAL MODELLING

ROCK

MASS

DOMAIN

Failure

Criterion /

Mechanism

INTACT ROCK

PROPERTIES

ROCK MASS PROPERTIES

HOEK – BROWN CRITERION

Blast Dist

Factor

MOHR COULOMB APPROXIMATION

COMMENTS

Rock

Density (t/m

3)

c

MPa

Ei

GPa

MR

#

GSI*

MI

UTS MPa

cm

MPa (Global)

Em

GPa

v

mb

s

a

Rock Mass

Cohesive Strength

(MPa)

Rock Mass

Friction Angle

BLOCK 1

Weathered HANGINGWALL

Siltstone

Generalised Hoek-Brown

2.7

28

+

10.5

375

40

25

9

-0.03

-0.01

3.7

2.5

1.7

0.6

0.3

1.056

0.618

0.0013

0.0002

0.511

0.531

0

1.144

0.848

26.7

22.2

PEAK

POST-PEAK

BLOCK 1 Weathered Quartz ORE

2.7

23

+

8.6

375

40

25

9

-0.03

-0.009

3.0

2.0

1.4

0.5

0.3

1.056

0.618

0.0013

0.0002

0.511

0.531

0

0.940

0.697

26.7

22.2

PEAK

POST-PEAK

BLOCK 1 Weathered FOOTWALL

Siltstone

2.7

42

+

15.7

375

40

25

9

-0.05

-0.01

5.6

3.8

2.5

0.9

0.3

1.056

0.618

0.0013

0.0002

0.511

0.531

0

1.716

1.272

26.7

22.2

PEAK

POST PEAK

BLOCK 2 Fresh

HANGINGWALL Siltstone

2.8

69

**

25.9

375

42

25

9

-0.10

-0.03

9.6

6.2

4.7

1.5

0.25

1.134

0.618

0.0016

0.0002

0.510

0.531

0

2.917

2.090

27.3

22.2

PEAK

POST PEAK

BLOCK 2 Fresh

Quartz ORE

2.8

53

**

12.7

375

57

25

9

-0.23

-0.02

10.3

4.8

9.0

1.2

0.25

1.938

0.618

0.0084

0.0002

0.504

0.531

0

2.874

1.605

31.7

22.2

PEAK

POST-PEAK

BLOCK 2 Fresh

FOOTWALL Siltstone

2.8

75

+

28.1

375

54

25

9

-0.26

-0.03

13.6

6.8

10.9

1.7

0.25

1.741

0.618

0.0060

0.0002

0.504

0.531

0

3.866

2.271

30.8

22.2

PEAK

POST - PEAK

Carbonaceous Bedding Shear

Mohr – Coulomb

0.020

20

Ubiquitous Bedding

Barton - Bandis

JRC = 5, JCS = 20MPa, Basic Friction Angle = 30deg



Range of rock mass strength has also been verified through application of the approach devised by Barton (2000) and Singh (1993) which established the following relationship:

14. σcm = 5(Qσc/100)1/3, where is rock density and σc is intact rock strength. The spatial distribution of rock mass strength in the hangingwall (from the 60 TGD drillholes), as calculated using this equation (Appendix A), is shown in the contour plot below (Figure 12).

Figure 12 : Rock Mass Strength (RMS) Anticipated in Mining Blocks 1 & 2 (First Quartile, 10m into Hangingwall)



6.0 STOPE : ROOM & PILLAR DESIGN

Various empirical and numerical modelling methods have been applied by GCPL in the confirmation of appropriate stable unsupported design spans and pillar dimensions for the Block 1 and 2 stoping areas. As presented in the following paragraphs, this has included:

Barton’s Q System for the assessment of unsupported span;

Assessment of critical unsupported spans, anticipated pillar loads and stable pillar dimensions through Phase2 numerical modelling;

Assessment of Rock Mass Strength from Hoek & Brown and Barton (2000) & Singh (1993) Systems;

Assessment of Pillar Strength from Laubscher (1990);

Mathews-Potvin Method for Open Stope Design, and the assessment of hydraulic radius (HR). 6.1 Estimated Stress Field & Pillar Loads In assessing overburden pressure and load on each pillar, reference was made to existing topographic survey information to establish the range in depth of mining below surface. This varies from 10 to 144mbsl in Block 1 and 190 to 274mbsl in Block 2. The current mining layout is shown in context of surface topography in Figure 13 below.

Figure 13 : Overlying Surface Topography and Current Mining Layouts in Mining Blocks 1 & 2 Pillar loads assessed are based on the following principles:

Block 1 (10 – 144mbsl) : Pillar load is a function of rock density, depth below surface and the tributary area between pillars divided by the pillar loading area, with some load transfer through rock arching, onto adjacent pillars;

Block 2 (190 – 274mbsl) : Pillar load is a function of rock density, depth below surface and the tributary area between pillars divided by the pillar loading area, with stress redistribution and load transfer through rock arching, onto adjacent pillars. The load on the pillars is dictated by both the stiffness of the surrounding strata and the stiffness of the pillars themselves;



Phase2 stress modelling was used to assess the extent of stress redistribution, rock arching and resultant pillar loads, for comparison with pillar rock mass strength, and assessment of Factor of Safety.

In the absence of any in situ stress measurements at Tom’s Gully, a gravity stress-field with

has been assumed, applying the bulk densities shown in Table 2. Considering this relationship, the envisaged post-stoping stress state in mining Blocks 1 and 2 has been modelled in Phase2. Phase2 is a 2D elasto-plastic finite element program for simulating stress and strain displacements around underground and surface excavations. 2D models generally give pillar stresses that are a bit higher than those predicted by 3D analysis. Maximum mining induced principal stress levels anticipated in Blocks 1 and 2, are shown in context of the underground excavations – in the Figure 14 to 18 cross – sections below:

Figure 14:BLOCK1 RL905-910 Airleg Stopes : 5x5m Pillar : Loads in the Range 2-14MPa Expected

Figure 15: BLOCK1 RL905 Airleg Stopes : 5 x 5m Pillar : Loads in the Range 9 to 12 MPa, 140mbsl (10m Rooms, 5m Pillars)



Figure 16 : BLOCK2 RL825-770 Stopes : 14x15m Rooms & 5x5m Pillars : Loads in the Range 24-34MPa Expected

Figure 17 : BLOCK2 RL825 Stopes : 5x5m Pillar : Loads in the Range 10-21MPa Expected

Figure 18 : BLOCK2 RL770 Stopes : 5x5m Pillar : Loads in the Range 14-34MPa Expected



6.2 Pillar Strength

Pillars are designed to ensure regional stability, local support in stopes, or to yield under a measure of control. In all cases, the variation in strength of the rock mass must be known, both for the pillar and for the hangingwall and footwall. The shape of the pillar with respect to dominant structure, blasting and stresses is significant. Block 2 pillars have therefore been designed to align with dominant “1950” striking dip structures to minimise overbreak and stress related deterioration. Phase2 stress modelling (Section 6.4), in conjunction with the approach devised by DH Laubscher (1990) Ref 14, has been used to confirm appropriate pillar dimensions and spacings. Most empirical methods for pillar design relate the pillar strength to rock mass strength and the width to height ratio of the pillar (eg : Salamon & Munro, Hedley & Grant, and Laubscher) ie: 15. Pillar Strength Ps = RMS (W0.5 / H0.7); (DH Laubscher 1990) Where : W = 4 x Pillar HR; H = Pillar Height. Using this method, the ratio of pillar strength to pillar load (Section 5.6), is then calculated to give a pillar factor of safety (FoS) – Table 8. The pillar strengths (Ps) calculated from the TGD drillhole reef intersection data are presented in APPENDIX B and in the Figure 19 contour plot. This plot will prove to be a useful tool for mine planning and pillar layout purposes, which will be adjusted in accordance with initial pillar performance in the RL825 stopes.

Mining sigrm W0.5

/ H0.7

Pillar Phase2 Model Pillar

Hole Depth Depth Block (MPa) (5x5m Pillar) Strength Pillar Load FoS

ID From To 5Y (Q

sigci/100)1/3

(2m Stope Ht) (MPa) (MPa)

(5x5m Pillar)

(10m Str x 19m Dip)

Tributary Area

TGD160 130.9 131.9 1 7 1.4 10 6 1.6

TGD168 199 200 2 59 1.4 81 24 3.4

TGD177 189.85 190.45 2 28 1.4 39 24 1.6

TABLE 8 : Example Calculation of Pillar Factor of Safety (FoS)



Figure 19 : Laubscher Pillar Rock Mass Strength (First Quartile) Anticipated in Mining Blocks 1 & 2 (Based on 5x5m Pillars)



6.3 Assessment of Room Spans GCPL have applied the empirical Stability Graph Method (Hutchinson and Diederichs, 1996) together with Phase2 numerical modelling, to assess stable unsupported spans for the stope hangingwall. The Stability Graph Method evolved in Canada but empirical design charts have been adapted specific to conditions and case histories in Australia. The method utilises the first four parameters from Bartons Q system discussed earlier in Section 5.5.1. As shown below, the four parameters are applied in the calculation of Q’, which is adjusted to obtain a modified stability number, N’. Stable stope hydraulic radii (HR) are then derived for the hangingwall from the “Modified Stope Stability Graph” (Figure 20). The stable, transitional and unstable regions in the graph are based upon stope performance data from other mines ie: empirical case history statistics. The relationship between Q, Q’ and N’ is as follows:

SRF

Jwx

J

Jx

J

RQDQ

a

r

n

; (Equation 16);

a

r

n J

Jx

J

RQDQ ' ; (Equation 17);

xAxBxCQN '' (Equation 18).

Table 9 in conjunction with Appendix A and Figure 20 demonstrates the application of the method in the calculation of N’ and stable unsupported hydraulic radius (HR) for the hangingwall in Mining Blocks 1 & 2. For example in the case of the Block 2 drillhole TGD255 hangingwall intersection at 177.7mbsl, the following process is applied in the calculation of HR:

Rock Stress (Factor A), assesses the ratio intact rock UCS to Maximum Mining Induced Tangential Stress (sigma C / sigma max).

Estimated Intact Rock UCS (sigma C) = 75 MPa;

Assessed Sigma max = 13 MPa;

Sigma c / sigma max = 75MPa / 13MPa = 5.8MPa, (Factor A = 0.51);

Discontinuity Orientation (Factor B) assesses the influence of the dominant defect set on stope stability, which in this case is dominant bedding / foliation parallel to planned stoping: (Factor B = 0.3);

Failure Mechanism (Factor C) : potentially gravity fall from the hangingwall, from bedding planes dipping sub-parallel to the stope at approx 00 to 200.

Therefore CHW = 8 - (6 cos 100) = 2.1 for average 100 reef dip; (Factor C = 2.1). N’mean = Q’mean x A x B x C N’ = 8.486 x 0.51 x 0.3 x 2.1 N’h/w = 2.73.

Unsupported span for the hangingwall within the ground conditions (Stability Number N’) anticipated is then assessed from the Modified Stability Graph (Figure 20) and is expressed in terms of stope shape or hydraulic radius (HR).



TABLE 9 : Application of Q’ to Derive the Modified Stability Number N’ for the Empirical Assessment of HR

For the example cited for Block 2, an N’ value of 2.73, results in an HR for the hangingwall of 3.6. This would suggest that for the current layout with a 14m level interval (edge to edge between each ore contour drive), that to maintain stability, 5 x 5m squat pillars should be maintained every 15m along strike. (ie HR = (14x15m) / (2(14+15m)) = 3.6. As shown in Figure 24 and to align with previous recommendations (Ref 3-6), the pillars should be staggered between levels. As extraction progresses, failure of the hangingwall and crushing of pillars is bound to occur, particularly in the lower strength poorer ground depicted by the yellow and orange contour intervals in Figure 19. Depending on hangingwall and pillar performance during initial mining in Block 2, pillar size, spacing and layout may require modification in those areas.

Hole_ID Depth Intact Q' sigvert sigmax

sigCi /

sigmax A B C N' Unsupported

To Rock

(MPa) (MPa) (MPa)

Stable HR

Strength

TGD253 86.8 37 7.917 2.4 6.3 5.8 0.52 0.30 2.10 2.58 3.5

TGD253 87.8 37 10.625 2.5 6.4 5.8 0.51 0.30 2.10 3.42 3.9

TGD253 88.8 75 0.178 2.5 6.5 11.6 1.02 0.30 2.10 0.11 1.0

TGD253 89.8 75 0.482 2.5 6.6 11.4 1.01 0.30 2.10 0.31 1.8

TGD253 90.4 75 0.417 2.5 6.6 11.4 1.01 0.30 2.10 0.26 1.8

TGD255 170.6 37 1.667 4.8 12.4 3.0 0.26 0.30 2.10 0.28 1.8

TGD255 171.6 37 16.667 4.8 12.5 3.0 0.26 0.30 2.10 2.75 3.6

TGD255 172.6 37 6.375 4.8 12.6 2.9 0.26 0.30 2.10 1.04 2.6

TGD255 173.6 37 3.400 4.9 12.7 2.9 0.26 0.30 2.10 0.55 2.3

TGD255 174.6 37 23.500 4.9 12.7 2.9 0.26 0.30 2.10 3.81 4.1

TGD255 175.6 37 23.500 4.9 12.8 2.9 0.26 0.30 2.10 3.78 4.1

TGD255 177.1 25 4.200 5.0 12.9 1.9 0.17 0.30 2.10 0.45 2.1

TGD255 177.7 75 8.486 5.0 13.0 5.8 0.51 0.30 2.10 2.73 3.6

TGD255 178.8 75 0.356 5.0 13.0 5.7 0.51 0.30 2.10 0.11 1.0

TGD256 165.6 75 85.000 4.6 12.1 6.2 0.55 0.30 2.10 29.42 5.0

TGD256 166.6 75 40.005 4.7 12.2 6.2 0.55 0.30 2.10 13.76 5.0

TGD256 167.6 75 20.003 4.7 12.2 6.1 0.54 0.30 2.10 6.84 5.0

TGD256 168.6 75 31.000 4.7 12.3 6.1 0.54 0.30 2.10 10.54 5.0

TGD256 163.6 37 5.167 4.6 11.9 3.1 0.27 0.30 2.10 0.89 2.6

TGD256 169.6 37 31.000 4.8 12.4 3.0 0.26 0.30 2.10 5.17 4.5

TGD256 171 37 6.062 4.8 12.5 3.0 0.26 0.30 2.10 1.00 2.6

TGD256 173.7 37 1.212 4.9 12.7 2.9 0.26 0.30 2.10 0.20 1.6

TGD256 174.7 37 1.212 4.9 12.7 2.9 0.26 0.30 2.10 0.20 1.6

TGD256 175.1 75 28.500 4.9 12.8 5.9 0.52 0.30 2.10 9.33 5.0



Figure 20 : Modified Stability Graph for Assessment of HR from N’

Note that :HR = Area (m2) / Perimeter (m); HR = (stope strike length x stope dip length) 2 (stope strike length + stope dip length)

Reference to the distribution of Bartons Q in Figures 8 & 10, and Equation 2 in particular, also confirms that unsupported spans in the range 15 to 25m will be stable in Mining Block 2. This is discussed further in Section 6.4.3.

MODIFIED STABILITY GRAPH

0.1

1.0

10.0

100.0

1000.0

0 5 10 15 20 25 30HYDRAULIC RADIUS (m)

MO

DIF

IED

ST

AB

ILIT

Y N

UM

BE

R (

N')

WA DATA SET

STABLE

CAVEDN’=2.7

HRHW=3.6



6.4 Phase2 Numerical Modelling This section outlines the approach taken by GCPL in confirming appropriate pillar dimensions and hangingwall spans for Block 1 & 2 stoping. Phase2 has been used for this purpose. Phase2 is a 2D elasto-plastic finite element program for simulating stress and strain displacements around underground and surface excavations. A range of alternatives examining various spans and pillar dimensions were considered in this process. The models constructed assumed the anticipated rock mass strength and mining induced stress conditions outlined earlier in Sections 5.5.3, 6.1 and 6.2. Previous research (Ref 12) suggests that comparisons of roof and pillar safety factor distributions obtained by 3D, 2D and augmented 2D analyses show that the minimum safety factors in the pillar (at the pillar sides) are predicted quite closely by the augmented Phase2D (plane stress) analysis. The same is true of the immediate roof, although the 3D safety factor tends to be higher in the roof (over the room) than that calculated by the augmented 2D technique.

6.4.1 Back-Analysis of Open Pit Wall Failure & Model Calibration

To impart confidence in the rock mass strength values assumed (Section 5.5.3) and to ensure a level of reliability in the modelling results, the Phase2 model was calibrated to reflect the current failure experienced in the SE area of the Tom’s Gully Open Pit. The modelling results are presented in section as strength factor contour plots. Strength factor (SF) is essentially the ratio of rock mass strength to mining induced stress, and is used to simulate excavation stability. Where initiating forces exceed restraining forces, the strength factor diminishes below 1 and fault-induced / rock mass failure will occur. The areas of open pit wall which have either failed in the past or that are at limit equilibrium (FOS=1) are represented by the dark orange contour interval in Figure 21.

Figure 21 : Block 1 Model Calibration with Pit Slopes at Limit Equilibrium

(Evidence of Failure in Existing South Eastern Slopes)



6.4.2 Stoping Block 1 In previous stoping, Block 1 contour drive spacing has varied from 5 to 8m spacing (edge to edge), with approximately 7 x 6m rooms supported by 4 x 5m pillars. To provide an indication of maximum spans and pillar dimensions, this configuration has been replicated and simulated in Phase2. The modelling results confirm that for the rock mass conditions anticipated (Figures 8,10,11,12 & 19):

Continuing with 5x5m square pillars, the RL905 to 890 rooms may generally be stoped to a stable unsupported span of 10m (HR=2.5). (Figures 22a&b). As these stopes will be re-entered for airleg mining, check scaling must be ongoing and some form of surface support and / or props (to minimise rockfall risk) will be required to render support to the hangingwall;

Any stoping in the RL910 rooms should be deferred until the end of mining, and extracted on retreat. This is to ensure crown pillar stability, and prevent storm water flowing into the underground. At that stage, room spans of the order 8 to 10m (HR=2 to 2.5) may be excavated. Check scaling must be ongoing and some form of surface support and / or props (to minimise rockfall risk) will be required to render support to the hangingwall locally.

Figure 22a : Block 1 RL910, 905 & 890 Stopes : 8 x 8m Rooms (Plane Strain Model)

Figure 22b: Block 1 : RL890 Stopes : Pillar Strength Factor = 2 (10m Rooms, 5m Pillars)



6.4.3 Stoping Block 2 Initial Block 2 RL825 trial stoping will be based upon a contour drive spacing of 14m (edge to edge) on dip and 5 x 5m pillars spaced at 15m on strike (edge to edge). Pillars will be staggered between each level, effectively resulting in a room size of 14 x 15m (HR=3.6) and a pillar load tributary area of 19 x 10m (Figure 25). To provide an indication of optimal spans and pillar dimensions, this as well as other configurations were replicated and simulated in Phase2 (Figures 23a-e). The modelling results confirm that for the rock mass conditions anticipated (Figures 8,10,11,12 & 19):

Trial stoping should initially be based on 14 x 15m rooms with 5x5m pillars;

This layout will be optimised as mining progresses, to align with ground conditions encountered in development and in accordance with stope stability performance;

Figures 8, 10, 19 & 24 suggest that there may be the opportunity to increase extraction ratios and implement larger 14 x 25m (HR=4.5) rooms at a later date.

Figure 23a : RL825 : 25m Strike Span Results in Higher Loading on 5x5m Pillars (SF=1.0 to1.3)

Figure 23b : RL825 : 15m Strike Span and 5x5m Pillars (SF=1.3)



Figure 23c : RL770 : 15m Strike Span and 5x5m Pillars (SF=1.2)

Figure 23d : 14m Dip Span and 5x5m Pillars (SF reduces from 1.3 at RL825 to 1.0 at RL770)



Figure 23e : Anticipted Strength Factors (1.3) in 5x5m Pillars During Stoping Initial Block2 825A – 825E

6.5 Room & Pillar Layout 6.5.1 Stoping Block 1 A systematic approach is recommended in future Block 1 stoping, which to date, has largely been carried out in an ad-hoc fashion with no longer-term room & pillar plans readily available. The RL905 to 890 rooms may generally be stoped to a stable unsupported span of 10x10m (HR=2.5) with 5x5m pillars and systematic surface support. Assessed in the light of cut-off grade, it is recommended that the HR-span layout (Figure 25) be superimposed and compared with the as-mined survey drawings, to identify areas where there is opportunity for additional extraction and further stoping. In Block 1, it would appear that most rooms are only 7 x 6m in area, but these may well be stoped to cut-off grade? Generally, pillars should be placed in low grade mining blocks where the reef is narrow and ground conditions are relatively good. 6.5.2 Stoping Block 2

Initial Block 2 RL825 trial stoping will be based upon a contour drive spacing of 14m (edge to edge) on dip and 5 x 5m pillars spaced at 15m on strike (edge to edge). Pillars will be staggered between each level, effectively resulting in a room size of 14 x 15m (HR=3.6) and a pillar load tributary area of 10 x 19m (Figure 24). Using this configuration, the estimated total Extraction Ratio is 87%. This layout will be optimised as mining progresses, in response to ground conditions encountered in development and in accordance with stope stability performance in the interim.



Figures 8, 10, 19 & 25 suggest that south of 4550N, there may be the opportunity to increase extraction ratios and implement larger 14 x 25m (HR=4.5) rooms at a later date. The unsupported span hydraulic radius (HR) contour values shown in Figure 25 should form the basis of the preliminary room & pillar layout for Block2. This, in conjunction with the Figure 8 & 19 contour plots will assist mine planning in the selection of optimal pillar layout in better ground. Generally, pillars should be placed in low grade mining blocks where the reef is narrow and ground conditions are relatively good. 6.5.3 Regional Pillars

To quote Rosengren verbatim (Ref 6) : “In any room & pillar stoping layout, it is necessary to leave a pattern of regional pillars, to provide regional support for the hangingwall. If a reef is extracted over a large area without such pillars, full scale caving of the hangingwall may develop, leading to the sudden crushing of remnant pillars. This situation has occurred in several room & pillar mines, sometimes with disastrous results. In Block 1, there are adequate natural pillars, and no additional regional pillars are required. However, a 10m wide pillar should be maintained in Block 1, on the western side of the Williams Fault.” In Block 2, regional pillars may be required once the quantity and size of low grade pillars expected in the stopes has been established. In-stope pillar performance and ground conditions will be continually monitored and reviewed to assess the long term regional pillar requirements. For preliminary design purposes, it is estimated that 5m wide regional rib pillars, may be required for every 100m of extraction on dip (Ref 5). As shown in (Figures 2,3, 6 and 24), 8 to 10m wide barrier pillars are planned between the decline and ventilation access development with 5m wide rib pillars between the access drive and the stopes. This plan is endorsed geotechnically.

Figure 24 : Preliminary Room & Pillar Layout in Block 2

15m

14m

Pillar Tributary Area = 10 x 19m

8 - 10m

5m



Figure 25 : HR Contour Plot : Should Form the Basis of Preliminary Room & Pillar Layout



6.5.4 Integrity of Crown Pillar

Whilst on site in July this year, GCPL carried out an inspection of the condition of the crown pillar in the vicinity of 910DW & DE stopes (Photo 7). This area is extensively bolted, meshed and polyfibrecreted. No cracking was observed in the shotcrete. Digital schmidt hammer rebound numbers indicated a UCS in the range 21 to 25MPa. (Table 3). The crown pillar area to the east was also inspected subsequently (3rd July 08) by GBS Gold’s Jan Reyneke (GBS Gold). Nothing untoward was reported. An indication of the stability of the crown pillar was also assessed by GCPL in Phase2. As shown in Figure 26, at its narrowest point it is assessed that the 10m crown is marginally stable with a factor of safety (FoS) = 1.04. For this reason, any stoping in the RL910 rooms should be postponed until the end of mining, and extracted on retreat. This is to ensure crown pillar stability, and to prevent storm water from flowing into the underground.

Photo 7 : Crown Pillar Area at 910DW

Figure 26 : At Narrowest Point : 10m Crown Pillar Appears to be Marginally Stable (FoS) = 1.04.



7.0 GROUND CONTROL STRATEGIES Various strategies have been devised to maintain access stability and minimise stope hangingwall failure at Tom’s Gully. This will involve various combinations of:

Decline, Access and Contour Drive Development :

Barrier Pillars;

Primary bolt & mesh surface support systems;

Cable-bolting of wide-span intersections;

Cable-bolting of structurally defined wedges;

Secondary support of very poor ground conditions – fibrecrete and additional cable bolting may be required.

Stopes :

Regional Pillars;

In-Stope Pillars;

Grout Mesh Packs;

Waste backfill;

Possible introduction of drift & fill mining method;

Monitoring of ground response to mining.

The standard ground support systems developed by GBS Gold for Tom’s Gully are presented in Appendix C. These have been reviewed by GCPL, and are deemed to be entirely appropriate for the ground conditions anticipated. 7.1 Development Primary Support In order to confirm ground support requirements for Tom’s Gully decline, access and contour drive development, the range of ground conditions experienced to date and expected in future mining, has been superimposed on the Barton-Grimsdad Support Chart (1993) - (Figure 27). Based upon the range of Rock Quality (Q) logged in TGD drillcore (Figure 8) and that observed from rock mass mapping of existing development underground, the rock has been classified and the required support pressure or equivalent support system required to restrain the rock mass at a known design span, given the engineering function of the excavation, assessed (Figure 27). In Figure 27, the Equivalent Dimension (De) is calculated from the following formula :

ESR

mheightordiameterspanExcavationDe

)(, (Equation 19, Barton 1974);

Where ESR is the Excavation Support Ratio based on excavation function (Table 9).

For example, decline, access and contour drive development will have a span of 5.0m, and therefore an equivalent dimension (De) of : De = 5.0m / 1.6 = 3.1.



This is shown in Figure 27 for the range of conditions expected. As verified in Equation 2, for Q-values in excess of 3.2, and assuming an ESR=1.6 (De=3.1), theoretically, provided there are no unfavourable structural combinations present, an excavation span of 5m or less would remain stable unsupported. However, as a means of minimizing rockfall risk, the DOME / MOSHAB guideline for surface support currently recommends that all primary development headings are to be meshed shoulder to shoulder to within 3.5m of the excavation floor, unless:

Proven to the contrary through geotechnical assessment, or

A frequent check-scaling programme is implemented.

To date, ground condition assessments undertaken at Tom’s Gully have generally confirmed the meshing requirement to provide full coverage and retain smaller rock blocks. Hence, as an absolute minimum, Standard Ground Support System 1 (Appendix C) will be implemented in all access development.

Required Bolt Length

For the planned 5m primary development spans at Tom’s Gully, considering an ESR of 1.6, a guideline for the required bolt or split set length (L), is as follows:

5m Span :

L=2+(0.15 Span) / ESR (Equation 20 : Barton et al 1994).

L = 2 + (0.15 * 5.0m) / 1.6 = 2.5m;

Note that this equation is not intended for application in the case of wide-span intersections, where tall wedges delineated by adverse structures may form. It is emphasised that bolt lengths given are a function of excavation span and ESR only and do not take the effect of dominant structures in the formation of rock wedges and blocks into account. Nor does the estimate take the ratio rock mass strength to insitu stress into account.

As shown in the Tom’s Gully Standard Support Systems (Appendix C), development primary support will typically comprise galvinised weld mesh with galvinised 2.4m length SS47 split sets. The split sets fix the mesh to the crown and shoulders of the excavation. Short 0.6m long split sets are used to pin the mesh along the overlap with the previous sheets by inserting them into the larger diameter SS47 bolts, although this is not always successful with bolts requiring duplication to fix the mesh more effectively. Bolt spacing is dictated by mesh size (3.0 x 2.4m), giving a tributary area of approximately 1.4 x 1.1m. The tendency for split sets to be drilled and installed at an angle not fully optimal, or "dumped" forward should be avoided. The jumbo selected should match the planned development size. "Dumping" the bolts forward by up to 45deg can significantly reduce the bolt effective length by up to 30%, resulting in reduced support into the backs. As split sets are friction bolts, this would also translate into a 30% reduction in load capacity. Bolts should be installed near perpendicular to the walls so as to optimise on the axial tensile (stretch) strength and reduce shearing or guillotining of the bolt. The pull-out resistance (bond strength) of the friction bolt varies according to the rock shear strength and the diameter of the hole drilled. The optimal hole diameter for an SS47 split set is of the order 43 to 45mm.



7.1.1 Roof Beam Theory The Tom’s Gully deposit is hosted by the Early Proterozoic Wildman Siltstone, which is a thick unit comprising laminated graphitic shale, carbonaceous and often pyritic siltstone inter-bedded with undifferentiated volcanics and minor dolomitic sediments. Stratification and the spacing and shear strength of bedding and bedding shears in particular, will have implications on the design and performance of roof spans during stoping. When undermined, the material strength normal to the plane of bedding is significantly reduced. The cohesion between the bedding planes is weakened. The immediate roof strata may become detached from the hangingwall, initiating deflection into the mined void through gravity loading. This is due to the bed itself having to support its own gravitational weight. The effective beam thickness is thus reduced, initiating instability. Three basic modes of roof failure are recognised in room & pillar mining in stratified rock:

Gravity Loading, Initial Roof Sagging and Subsequent Block Release;

Gravity Loading, Roof deflection and Subsequent “Snap” through the intact beam;

Gravity Loading, and Subsequent Shear failure of the Rock Mass.

Roof failure by initial sagging and then by the snap through mechanism are more common modes in larger spans. Shear failure occurs more frequently in poorer more blocky rock masses over shorter spans and under lower confinement. Floor heaving due to redistribution of stress after extraction may also cause instability. Thus the factors responsible for instability in stratified rock due to mining can be summarised as follows:

Post extraction stress redistribution;

Low or zero tensile strength normal to bedding;

Relatively low shear strength along bedding planes.

These conditions need to be taken into consideration when designing roof beams and plates spanning a stope. They are influenced negatively by an increase in span and a decrease in beam or plate thickness. Where access permits, the effective plate thickness may be increased by systematic bolting. Simplified formulae and roof design curves have been derived, to define the limits required for safe roof spans and safe roof beams. These design criteria are based on the “Voissoir Beam Theory.” The assumption is that a beam consisting of a low tension material carries its weight by arching. That is, the vertical load due to gravity approximated by a parabolic arch traced on the beam span, is distributed laterally onto the adjacent pillars or abutments. The over-riding consideration is the unsupported roof span and effective beam thickness. Therefore, to ensure stability, it is advisable to maintain an intact competent strata in the immediate roof to form a beam or plate at a pre-determined span. Systematic rockbolting, installed normal to dominant bedding, serves to reduce the amount of shear displacement between individual roof plates, and effectively increases the beam thickness. This increase is proportional to the amount of pre-tension, angle of installation, length and spacing of the bolts. The systematic bolting of ore contour drives (Appendix C -Ground Support Standard 3) will effectively increase the beam or plate thickness between pillars and limit many of the failure mechanisms discussed earlier.



Figure 27 : Barton – Grimsdad Support Chart : Range of Ground Conditions Expected in Mining Blocks 1 & 2 Suggests that 5m Span Development Could Generally Go Unsupported, but to Reduce

Rockfall Risk the DOME / MOSHAB Guideline to Mesh Backs to within 3.5m of Invert – Has been Implemented

Block 2 Q > 4

Unsupported

Block 1 Q=1 to 4

Unsupported



7.2 Cable Bolting of Wide Span Intersections Figure 28 overleaf illustrates the background requirement for cable bolting of wide-span intersections. Modern Duty of Care provisions require that all personnel-access roof spans exceeding 6m span should be supported by longer tendons. This is particularly the case where full-time geotechnical personnel are not on site to identify adversely oriented structures. As shown, the intent is to support larger potentially unstable wedges delineated by those adversely orientated structures at the wider span intersections. At intersections, standard practice is normally to develop the decline past the intersection and to drill and install cable bolts prior to developing the initial cross-cut / stockpile heading and prior to stripping to full span. This is required to systematically support any structurally delineated tetrahedral rock wedges that may form in the intersection backs. (Figure 28). As specified in Ground Support Standard 7, current practise at Tom’s Gully is to install 5 x 6 meter length twin-strand cables at 2m c-c, central to the intersection. There have been incidents at other mine sites where catastrophic wedge failure from drive backs has occurred without warning and has resulted in considerable damage to jumbo machinery and injury to personnel. Where through-going adversely orientated wedge-forming structures are encountered, installation of cable bolts prior to stripping and ledging-out is therefore a critical requirement at drive intersections. At the time of the July 2008 visit to site, the following recommendations were made by GCPL, outlining the general approach to be taken in assessing the requirement for cable-bolting at wide-span intersections. The requirements for cable bolting at the following development intersections were outlined as follows:

825 Access / Ore Drive B Intersection : Wide 12 meter span previously supported and cable-bolted with 10No, 6m Length, Twin-Strand Cable-Bolts. Irregular intersection profile;

825 Ore Drive A – B Intersection : Drive backs closely following competent bedding stratification with an excellent square profile. No through-going structures identified. Under current conditions, cable bolting is not deemed necessary. Grout existing 2.4 meter length split-sets at intersection, to double bond strength from 4t/m to 8t/m;

825 Access / Ore Drive C Intersection : A normal fault strikes through the 825 Access and C Ore Drive Intersection which has a maximum developed span of approximately 10 meters. Standard Cable Bolt Pattern “Support Standard 7” is therefore required;

825 Access / Ore Drive D Intersection : Drive backs closely following competent bedding stratification with an excellent square profile. Minor truncated structure identified. Under current conditions, cable bolting is not deemed necessary. Grout existing 2.4 meter length split-sets at intersection, to double bond strength from 4t/m to 8t/m;

All other dual – drive wide-span intersections : install and grout 3m length split-sets, in-cycle with standard meshing. Assess whether adverse through-going structures are developed. If developed – install Ground Support Standard 7.

All other multi-drive intersections : install and grout 3m length split-sets. Install Ground Support Standard 7. (Requirement for further cable-bolting at these wider-span intersections currently under review by GCPL).



Figure 28 : Cable Bolting of Structurally Defined Wedges at Wider Span Intersections

21/06/2007 Zapopan Underground

Geotechnical Induction

11

WEDGE AT DRIVE INTERSECTION

JOINT

JOINT

JOINT

GBS Gold Australia Pty LtdGBS Gold Australia Pty Ltd



7.3 Grout Mesh Packs In conjunction with in-stope pillars, grout mesh packs may be used as a yielding support measure to limit stope closure and restrict hangingwall failure. They would be installed mid-span, between pillars, during stoping immediately adjacent to the supported ore drives. It is not expected that the level of risk exposure during installation would be any greater than that currently experienced in Block 1 airleg mining. The packs generally yield at 3-4MPa, which is clearly far less than the expected pillar loads of 2 to 14MPa in Block 1 and 24 to 34MPa in Block 2 (Section 6.1). But gradual crushing of ore pillars is expected as stoping progresses, and the grout packs may be utilised to contain stope closure in a controlled fashion, to maintain access and prevent sudden failure of the hangingwall. 7.4 Ground Support Quality Control

Purpose In order to confirm engineering performance of the various ground support elements installed, a quality assurance-testing programme should be carried out. The prime aim of the testwork will be to quantify:

actual bolt-grout-rockmass bond strength / pull-out capacities being achieved for spilt-sets and cable bolts;

split set & cable bolt grout UCS;

fibrecrete UCS. The UCS specified for shotcrete should be 32MPa at a curing age of 28 days. This has to be demonstrated through ongoing testwork during construction.

And in so doing:

Apply the actual support performance data for optimisation of future support designs.

Ensure that split set and cable bolt grouting material develops a bond strength that matches the yield capacity of the installation steel (ie approximately 12t for split sets and 40t for a double strand cable bolt);

Establish the early bond (12 hour) strength of grouted split sets and cable bolts;

Assess the UCS of fibrecrete after 28 days curing.

Based upon the outcome of the test results, optimisation and improvement to the installation of future support may be achieved as follows :

Support quantity and spacings may be adjusted.

Cable bolts installed in very poor ground may develop unreasonably low bond strength. Under these circumstances, it is important that the toe of the cable bolts be sufficiently embedded into competent ground.

7.4.1 Test Procedures and Specifications

Cable Bolt Pull-Out Tests Ideally cable bolts should be pull-tested at 1 meter grout embedment / bond length to confirm pull-out resistance or bond strength per meter. Pull-testing of fully grouted cable bolts in hard rock is meaningless. If possible, double strand test bolts should be pull-tested to ultimate steel capacity of 50t, over a trial bond length of 1 meter.



Bolt holes are to be drilled straight, perpendicular to the sidewall and as near horizontal as practicable. 2.0m length, plain twin strand cables are to be installed into a 1.6m length ~70mm diameter hole with a 0.5m grouted bond length at the toe. A 1.1m length of cable should be decoupled with pvc sleeving with the remaining 0.4m length left protruding from the hole collar. This facilitates connection of the strands through the jack, and locking-off. A typical configuration is shown in Figure 29 below:

1.6m drill hole

0.5m bonded1.1m debonded

0.4m tail

Steel strand

Grouting cementPVC sleeves

Figure 29 : Cable Bolt Pull-Test Schematic

Split-Set Pull-Out Tests In order to confirm the bond strength that develops per metre of anchorage, 1 metre length test bolts should be installed, perpendicular to the sidewall and as near horizontal as practicable. Pull out testing of split sets and cable bolts should be carried out using a hydraulic jack with a capacity of 10 to 60 tonne. Photo 8 below shows the testing arrangement.

Photo 8 : Jack Configuration for Cable Bolt Pull-Out Testing



Grout Strength Testing

A grout strength testing programme should be introduced to confirm compliance with specification. This will involve sampling of grout directly from the batch-mixer. Cement grout is sampled during grouting

using cylindrical moulds (100mm dia 200mm high) For each curing age stipulated, three samples should be cast for UCS testing in a Geomechanics Laboratory (eg : Ullman & Nolan NT Pty Ltd, Darwin). Alternatively, samples are poured into 50mm diameter (ID) pvc tubes, cured for 7, 14 or 28 days as required, extruded and point-load tested. Cable bolts should be installed with maximum strength grout to maximise the bond strength. Since grout strength increases with decreasing W : C ratio, the lowest W:C ratio grout that is readily mixed and pumped, that has sufficient water content for hydration should be used. Generally, the grout mixture specified for minicage cables should lie between 0.3 and 0.35. This may be increased to 0.4 to improve pumpability. A typical UCS of approximately 30 MPa after 7-days should be achieved. In summary, for split sets and cable bolts both with the equivalent of 1m embedment, the following estimated yield capacities may be used as a guideline for any future test programmes at Tom’s Gully :

Ungrounted SS46 splits sets: 3~4t per metre;

Grounted SS46 splits sets: 8~9t per metre at 12 hours and 12t at 7 days;

Double strand cable bolts: 20~23t per metre at 12 hours and 45~50t at 28 days;

Triple strand cable bolts: 23~30t per metre at 12 hours and 70~74t at 28 days.

These typical results are shown in Figure 30 below:

0

10

20

30

40

50

60

70

80

0 5 10 15 20 25 30

Time after installation (days)

Pu

ll-o

ut lo

ad

(to

nn

es)

double strand cable bolts

triple strand cable bolts

ungrouted split sets

grouted split sets

Figure 30 : Typical Pull-Out Loads for 1m Split Sets and 1m Cable Bolts



7.5 Excavation Stability Monitoring All active development headings and working stopes should be inspected daily for visible instability or potential structurally controlled failures by the Mining Contractor. A geotechnical inspection register, documenting observations should be maintained in the Engineers office. The appointed competent person should complete this immediately after inspections are undertaken. Ore-drive / room convergence monitoring should be ongoing. Convergence survey reference pins may be installed in conjunction with extensometers and SMART (Stretch Measurement to Assess Reinforcement Tension) cable bolts. Based on the stretch of the cable, the load that develops in the bolt can be determined. 7.6 Check-Scaling During the period 1987 to 1996, of the total number of accidents reported in Western Australian metalliferous mines, 25% were rock-fall related, resulting in 26 fatalities. In order to prevent rock fall incidents at Tom’s Gully, an excavation check-scaling programme should be ongoing in all unmeshed sections of primary development (eg :Vent Escape-Way)? and re-entry airleg stopes. The following procedures should be in place

Travelway / Permanent Opening Routine Check Scaling Procedure;

Blast Re-Entry Check-Scaling Procedure;

Dept Of Minerals & Energy (DOME) requirements for Underground Barring Down & Scaling. Scaling procedures should be developed in order to systematically and routinely manage the geotechnical risk associated with loose rocks falling from the excavation perimeter. Scaling may be described as – stabilising and rendering the ground safe by removing loose rock from the backs, walls and face of the surrounding excavation, either manually by sounding and use of a scaling bar, or by machine scaling utilising mechanised scaling equipment. Scaling is to include the removal of accumulated loose scats under pre-existing mesh whenever mesh bulging or unravelling is apparent. At any underground mine, scaling is an ongoing process in each workplace and is only complete upon closure of the operation. In-keeping with the WA Mines Safety Act & Regulations, the purpose of the scaling activity - procedure is as follows :

Routinely identify and prioritise working areas requiring scaling;

Reduce the risk (likelihood & consequence) of rock fall hazards in order to improve safety;

Ensure that appropriate scaling methods, procedures and equipment are used for scaling;

Render safe and suitably prepare the excavation prior to the installation of primary (surface) and secondary support.

7.7 Safety & Duty of Care 7.7.1 Crew Awareness Crew awareness is an essential component in the identification, monitoring and management of geotechnical hazards. Crews are currently alerted of such hazards through direct briefing at safety meetings and at morning shift-change meetings. Dissemination of geotechnical hazard status may in future take the form of a weekly monitoring circular complete with photographs posted on the mine hazard board.



7.7.2 Site Safety Induction Following-on from the previous comments, it is imperative that as part of their induction, new mining personnel become acquanted with :

the danger that geotechnical hazards pose;

recognising hazard controls & potential;

work safe procedures;

existing hazards upon being employed. Appreciation and knowledge of this should be tested as part of the induction process and specific hazards existing at that time should be visited underground to ensure that the new employee understands and is aware of the potential risk. To heighten the awareness of geotechnical hazards and controls on stability underground, a geotechnical induction slide show, along the lines of that delivered at Zapopan should be presented.

8.0 CONCLUSIONS & RECOMMENDATIONS

This Ground Control Management Plan (GCMP) document has outlined the anticipated geotechnical conditions and presented the stability assessments undertaken to derive geotechnical excavation and support designs appropriate for GBS Gold’s Tom’s Gully Underground Mine. The document has detailed current knowledge of rock mass conditions at Tom’s Gully, the related geotechnical controls on stability, and has presented the recommended ground control management strategies which will be employed to maximise ore recovery in a safe and cost efficient manner. As the ore reef dips at less than 200, a room and pillar mining method has been implemented. In-line with the geotechnical conditions anticipated, appropriate room and pillar dimensions are also provided. The main points are as follows:

Mining Block 1 : A systematic approach is recommended in future Block 1 stoping, which to date, has largely been carried out in an ad-hoc fashion with no longer-term room & pillar plans readily available. The RL905 to 890 rooms may generally be stoped to a stable span of 10x10m (HR=2.5) with 5x5m pillars and systematic surface support. To identify areas where there is opportunity for additional extraction and further stoping, it is recommended that the HR-span layout (Figure 25) be superimposed and compared with the as-mined survey drawings. This should be assessed in the light of cut-off grade;

As witnessed in the RL905 stopes, as spans approach 10m, there is general convergence of the backs to a parting carbonaceous bedding shear. This has resulted in sudden failure of the hangingwall. Without systematic bolting or some form of temporary propping, there is a very real danger of sudden rock fall and a high risk of exposure to the Alliance airleg miners. This aspect therefore requires immediate corrective action by GBS and Alliance Mining. There are also associated external waste dilution issues. For the Block 1 access stopes, check scaling must be ongoing and systematic bolting and / or props (to minimise rockfall risk) will be required to render support to the hangingwall;

Mining Block 2 : Initial RL825 trial stoping will be based upon a contour drive spacing of 14m (edge to edge) on dip and 5 x 5m pillars spaced at 15m on strike (edge to edge). Pillars will be staggered between each level, effectively resulting in a room size of 14 x 15m (HR=3.6) and a pillar load tributary area of 10 x 19m (Figure 24). Using this configuration, the estimated total Extraction Ratio is 87%. This layout will be optimised as mining progresses, in response to ground conditions encountered in development and in accordance with interim stope stability performance. South of 4550N, there may be opportunity to increase extraction ratios and implement larger 14 x 25m (HR=4.5) rooms at a later date. The unsupported span hydraulic radius (HR) contour values shown in Figure 25 should form the basis of the preliminary room & pillar layout for Block 2. This, in conjunction with the Figure 8 & 19 contour plots will assist mine planning in the selection of optimal pillar layout in better ground. Generally, pillars should be placed in low grade mining blocks where the reef is narrow and where ground conditions are relatively good. As highlighted in the Figure 19 Pillar Rock Mass Strength Contour Plot, ground conditions in the vicinity of TGD-Diamond holes 417, 419, 427, 429 & 313 (Figure 6) are expected to be



poor with pillar strengths less than 20MPa. With pillar loads approaching 26MPa in those areas, it is anticipated that pillar sizes will need to be increased and / or spans between pillars reduced;

Grout Mesh Packs : In conjunction with in-stope pillars, grout mesh packs may be used as a yielding support measure to limit stope closure and restrict hangingwall failure. They would be installed during stoping, mid-span between pillars and immediately adjacent to the supported ore drives. It is not expected that the level of risk exposure during installation would be any greater than that currently experienced in Block 1 airleg mining. The packs generally yield at 3-4MPa, which is clearly far less than the expected pillar loads of 2 to 14MPa in Block 1 and 24 to 34MPa in Block 2. But gradual crushing of ore pillars is expected as stoping progresses, and the grout packs may be utilised to contain stope closure in a controlled fashion, to maintain ore drive access and prevent sudden failure of the hangingwall;

Regional Pillars : “In any room & pillar stoping layout, it is necessary to leave a pattern of regional pillars, to provide regional support for the hangingwall. If a reef is extracted over a large area without such pillars, full scale caving of the hangingwall may develop, leading to the sudden crushing of remnant pillars. This situation has occurred in several room & pillar mines, sometimes with disastrous results. In Block 1, there are adequate natural pillars, and no additional regional pillars are required. However, a 10m wide pillar should be maintained in Block 1, on the western side of the Williams Fault.” In Block 2, regional pillars may be required once the quantity and size of low grade pillars expected in the stopes has been established. In-stope pillar performance and ground conditions will be continually monitored and reviewed to assess the long term regional pillar requirements. For preliminary design purposes, it is estimated that 5m wide regional rib pillars, may be required for every 100m of extraction down dip. As shown in (Figures 2,3, 6 and 24), 8 to 10m wide barrier pillars are planned between the decline and ventilation access development with 5m wide rib pillars between the access drive and the stopes. This plan is endorsed geotechnically;

Crown Pillar : Whilst on site in July this year, GCPL undertook an inspection of the condition of the crown pillar in the vicinity of the 910DW & DE stopes. This area is extensively bolted, meshed and polyfibrecreted. No cracking was observed in the shotcrete. An indication of the stability of the crown pillar was also assessed by GCPL in Phase2. At its narrowest point, it is assessed that the 10m crown is marginally stable with a factor of safety (FoS) = 1.04. For this reason, any stoping in the RL910 rooms should be postponed until the end of mining, and extracted on retreat. This is to ensure crown pillar stability, and to prevent storm water from flowing into the underground.

It is intended that this GCMP will be developed as a Major Hazard Standard for improved strata control, for proactive application by GBS employees and contractors and to meet the approval of the NT Mines Inspectorate.

Please contact the undersigned should you require clarification on any aspect of this report.

Ian McEnhill (Director - Geotechnical Consulting Pty Ltd)

Disclaimer:

Geotechnical Consulting Pty Ltd (GCPL) warrants that for this study, it has taken reasonable care in accordance with standards ordinarily exercised by members of the profession, generally who practice in the same locality and under similar conditions. GCPL accepts no liability whatsoever in respect of any failure to exercise a degree or level of care beyond such reasonable care. No other warranty, express or implied, is given, save where necessarily incorporated by statute.

GCPL disclaims liability, whether in negligence or otherwise, for any loss or damage claimed or arising from commercial decisions or actions based upon GCPL’s conclusions and recommendations.



REFERENCES

1. Geotechnical Consulting Pty Ltd : Underground Inspection Letter Report : GCPL-TG-040708; 2. Heath Gerritsen Mine Design Pty Ltd : Technical Report on the Mineral Reserves of the Tom’s Gully Gold Mine (December 2007); 3. Coffey Mining Consultants(Dr PM Dight & G Auld) Report MINEHERD00035AA (28 August 2006); 4. Coffey Mining Consultants(Dr PM Dight) Site Visit Report PZ00035/01 (14 November 2005); 5. Coffey Mining Pty Ltd (2007), Email, Geotechnical Questions about Tom’s Gully; 6. Kevin Rosengren & Associates Pty Ltd : Tom’s Gully Underground Mine,

Geotechnical Review (February 2006); 7. Tennent Isokangas Pty Ltd Consulting Mining Engineers: Feasibility Study Mining Report for

Renison Consolidated Mines NL (December 2004); 8. Rocscience Inc,Toronto, Canada, Phase2 Version 6.0, 2006.

Finite Element Numerical Modeling Software; 9. Dr Evert Hoek's 1999 Course Notes; 10. N Barton & V Choubey (1976) : The Shear Strength of Rock Joints in Theory and Practice; 11. N Barton (NGI 1973) : Review of a New Shear-Strength Criterion for Rock Joints; 12. 3D Mine Pillar Design Information from 2D FEM Analysis WG Pariseau & WK Sorensen (International

Journal for Numerical & Analytical Methods in Geomechanics, Volume 3, 145-157), 1979; 13. Tim Richards (Sirocco Resources), MIM’s Flitch Plan & Structural-Ore Genesis Interpretation

(21 September 1997); 14. DH Laubscher : A Geomechanics Classification System for the Rating of Rock Mass in Mine Design

(1990).



APPENDIX A

TGD Diamond Hole Q, N’ & HR Calculations for Hangingwall



APPENDIX B

TGD Diamond Hole : Pillar Rock Mass Strength Calculations



APPENDIX C

Ground Support Standards & Specifications



APPENDIX D

Rockbolt Technical Specifications



GEOTECHNICAL CONSULTING PTY LTD

Standard Terms of Engagement

1. Geotechnical Consulting Pty Ltd (GCPL) : Responsibilities GCPL will :

(a) carry out the project with due care and skill and in accordance with the project plan; (b) use reasonable efforts to provide all deliverables to the Client by the dates specified

in the Services Agreement Document.

2. Confidentiality and Restraint of Trade

Each party (GCPL and the Client) agrees to keep confidential all information that the other party discloses to it and specifies as confidential unless :

(a) the discloser agrees in writing that it can be released;

(b) the information is in the public domain;

(c) the information is required to be disclosed by law or persuant to legal proceedings.

3. Limitation of Liability

In this section we set out, and the Client accepts the limitations which apply to Geotechnical Consulting Pty Ltd’s (GCPL) liability to you should you (the Client) have reason to make a claim against GCPL. The limitations and exclusions are accepted by both parties to be fair and reasonable, given the services being provided, the sums to which GCPL is entitled and the availability (and cost) of insurance. 3.1 The Client agrees that GCPL’s liability for any loss or damage suffered by you (whether

direct, indirect or consequential) in connection with our engagement or in any way arising out of our performance of the Geotechnical Services, including (without limitation) liability for any negligent act or omission or mirepresentation of GCPL, shall be limited to the amount of the professional fees paid to GCPL in respect of the Services, and you agree to release GCPL from all claims arising in connection with the Services, to the extent that GCPL’s liability in respect of such claims would exceed the amount of those professional fees.

3.2 If GCPL is liable for a breach of any warranty implied by section 74 of the Trade Practices Act, 1974, then GCPL’s liability under that section will be limited to supplying the Geotechnical Services again or the payment of the cost of having the Services supplied again, whichever GCPL, in it’s absolute discretion elects.



4 Indemnities

4.1 You agree to indemnify and hold harmless GCPL against any and all losses, claims, costs,

expenses, actions, demands, liabilities or any other proceedings whatsoever incurred by GCPL in respect of any claim by a third party arising from or connected to any breach by you of your obligations under this Agreement.

4.2 GCPL shall not be liable for any losses, claims, expenses, actions, demands, damages,

liabilities or any other proceedings arising out of any reliance on any information provided by the Client which is false, misleading or incomplete. You agree to indemnify and hold harmless GCPL from any such liabilities we may have to you or any third party as a result of reliance by GCPL on any information provided by the Client or any of the Client’s representatives.

5 Variation

No variation of this agreement will be valid unless confirmed in writing by authorised signatories of both parties on or after the date of signature of the Geotechnical Consulting Pty Ltd Services Agreement Letter.

6 Disclaimer

Geotechnical Consulting Pty Ltd (GCPL) warrants that for this study, it has taken reasonable care in accordance with standards ordinarily exercised by members of the profession, generally who practice in the same locality and under similar conditions. GCPL accepts no liability whatsoever in respect of any failure to exercise a degree or level of care beyond such reasonable care. No other warranty, express or implied, is given, save where necessarily incorporated by statute.

GCPL disclaims liability, whether in negligence or otherwise, for any loss or damage claimed or arising from commercial decisions or actions based upon GCPL’s conclusions and recommendations

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