BIRD IN HAND GOLD PROJECT · 2019-06-19 · FINAL REPORT BIRD-IN-HAND GOLD PROJECT GEOTECHNICAL...

BIRD IN HAND GOLD PROJECT MINING LEASE PROPOSAL MC 4473

ABN | 66 122 765 708 Unit 7 / 202-208 Glen Osmond Road | Fullarton SA 5063

APPENDIX M1

GEOTECHNICAL ASSESSMENT

FINAL REPORT

BIRD-IN-HAND GOLD PROJECT


For

TERRAMIN AUSTRALIA LIMITED

Job No. 2241_G Mining One Pty LtdLevel 9, 50 Market Street

Melbourne VIC 3000 Ph: 03 9600 3588

Fax: 03 9600 3944

Doc No. 4819_V3.docxDate: June 2017Prepared by: M. Bijelac

B Roache

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TABLE OF CONTENTS

EXECUTIVE SUMMARY ............................................................................................................................... i

1 INTRODUCTION AND SCOPE ......................................................................................................... 1

1.1 Scope of Work .......................................................................................................................... 1

2 BACKGROUND AND PROPOSED MINING METHOD .................................................................... 4

3 GEOLOGY ......................................................................................................................................... 5

3.1 Key Geotechnical Considerations ............................................................................................ 5

4 GROUNDWATER MODEL ................................................................................................................ 6

5 GEOTECHNICAL MODEL .............................................................................................................. 10

5.1 Test Pits and Geotechnical Drilling......................................................................................... 10 5.1.1 Test Pits .............................................................................................................................. 10 5.1.2 Drilling and Logging ............................................................................................................ 11 5.2 Sampling and Laboratory Testing........................................................................................... 13 5.2.1 Clay/Highly Weathered Domain ......................................................................................... 13 5.2.2 Rock Domain ...................................................................................................................... 18 5.3 Rock Surface .......................................................................................................................... 20 5.4 3D Rock Quality Model ........................................................................................................... 21 5.5 Structural Model...................................................................................................................... 23 5.5.1 Decline ................................................................................................................................ 23 5.5.2 FW Access Infrastructure ................................................................................................... 25 5.5.3 Major Structures ................................................................................................................. 27 5.6 Rock Mass Domains ............................................................................................................... 29 5.7 Stress Model ........................................................................................................................... 32 5.7.1 Introduction ......................................................................................................................... 32 5.7.2 Material Properties ............................................................................................................. 32 5.7.3 In-Situ Stress Conditions .................................................................................................... 33 5.7.4 Model Sequencing and Grid Locations .............................................................................. 33 5.7.5 Model Results ..................................................................................................................... 35

6 GEOTECHNICAL DESIGN .............................................................................................................. 37

6.1 Surface Stability ...................................................................................................................... 37 6.1.1 Raw Water Dam and IML Stability ...................................................................................... 37 6.1.2 Boxcut Design ..................................................................................................................... 40 6.1.3 Pre-Cast Concrete Arches (BEBO Structure) .................................................................... 42 6.1.4 Laydown Area – Upper Decline Spiral Interaction .............................................................. 43 6.1.5 Site Stabilisation and Erosion Control ................................................................................ 44 6.2 Infrastructure Positioning ........................................................................................................ 47 6.2.1 Vent Shaft Positioning ........................................................................................................ 47 6.2.2 Lower Decline and Vent Return .......................................................................................... 48 6.2.3 Influence of Major Structures .............................................................................................. 48 6.3 Ground Support Design .......................................................................................................... 49 6.3.1 Empirical Support Design Based on Q Classification ......................................................... 49 6.3.2 Wedge Analysis .................................................................................................................. 52 6.3.3 Proposed Ground Support Design ..................................................................................... 59 6.3.4 Cablebolting for Wide Spans .............................................................................................. 60 6.4 Roadheader Suitability ........................................................................................................... 61 6.4.1 Introduction ......................................................................................................................... 61 6.4.2 Weathering and Strength .................................................................................................... 62

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6.4.3 Roadheader Suitability Evaluation...................................................................................... 62 6.4.4 Excavator Suitability Evaluation ......................................................................................... 63 6.4.5 Tunnel Support ................................................................................................................... 63 6.4.6 Conclusions ........................................................................................................................ 63 6.5 Cemented Rock Sill Pillars ..................................................................................................... 64 6.5.1 Suggested Way Forward for the Proposed CRF Sill Pillars ............................................... 64

7 MINING INDUCED GEOHAZARDS ................................................................................................ 66

7.1 Surface Subsidence ............................................................................................................... 66 7.1.1 Types of Mining Related Subsidence ................................................................................. 66 7.1.2 Mining Method and Size of Underground Voids ................................................................. 67 7.1.3 Surface Stability Assessment – Underground Mining ........................................................ 69 7.2 Earthquake and Mining Induced Seismicity ........................................................................... 71 7.2.1 Naturally occurring earthquakes ......................................................................................... 71 7.2.2 Mining Induced Seismicity .................................................................................................. 72

8 RISK ASSESSMENT CONSIDERATIONS ..................................................................................... 74

9 SUMMARY, CONCLUSIONS AND RECOMMENDATIONS .......................................................... 75

REFERENCES ........................................................................................................................................... 77

TABLE INDEX

Table 5-1: Test pit details ........................................................................................................................... 10

Table 5-2: Drill hole details ......................................................................................................................... 12

Table 5-3: Summary of Atterberg Limits testing ......................................................................................... 14

Table 5-4: Summary of Emerson Class testing .......................................................................................... 15

Table 5-5: CU testing results - clays .......................................................................................................... 17

Table 5-6: Correlation between derived strength and logged strength estimate ....................................... 17

Table 5-7: Summary of UCS test results .................................................................................................... 19

Table 5-8: Comparison of logged UCS estimate versus tested strength ................................................... 19

Table 5-9: Summary of Young’s Modulus testing results ........................................................................... 19

Table 5-10: Summary of Hoek Triaxial testing results ............................................................................... 19

Table 5-11: Descriptions associated with the Q System ............................................................................ 22

Table 5-12: Structural sets for Upper Level Decline Domain ..................................................................... 25

Table 5-13: Structural sets for the FW ....................................................................................................... 27

Table 5-14: Material Properties Used for MAP3D Modelling ..................................................................... 33

Table 5-15: Assumed In-Situ Stress Field .................................................................................................. 33

Table 5-16: Summary of Maximum Sigma 1 Stress Magnitudes from MAP3D Modelling ......................... 36

Table 6-1: Provisional cut slope design (<3m height) ................................................................................ 38

Table 6-2: Summary of Results for Modified Infrastructure Design ........................................................... 39

Table 6-3: Summary of shaft option assessment ....................................................................................... 47

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Table 6-4: Summary of Q values for each domain and infrastructure type ................................................ 50

Table 6-5: Ground support estimates based on empirical design chart ..................................................... 52

Table 6-6: Bolt Properties of Types Considered for Analysis ..................................................................... 53

Table 6-7: FW access infrastructure 0° plunge – 2.4m Split Sets @ 1.5m spacing .................................. 56

Table 6-8: FW Access Infrastructure 0o Plunge – 2.4m Resin Bolts @ 1.5m Spacing .............................. 56

Table 6-9: Results summary showing wedges with FOS below 1.5 – no support ..................................... 58

Table 6-10: Results summary showing wedges with FOS below 1.5 – split set support at 1.5m spacing 59

Table 6-11: Proposed ground support estimates – 4 to 6m wide spans .................................................... 60

Table 7-1: Design guidelines for crown pillar design life (after Carter et al, Ref. 16) ................................. 71

FIGURE INDEX

Figure 1-1: Bird-in-Hand project layout – cross section looking north

Figure 2-1: Orebody cross section showing cut and fill drives

Figure 4-1: South west view of mine design and HW and FW faults

Figure 4-2: Plan at 200m RL +/- 10m showing faults

Figure 4-3: Water inflow predictions at approximately 150m depth


Figure 5-1: Plan view of pit locations relative to underground mine design (portal and upper decline)

Figure 5-2: East west view showing drill holes and mine design

Figure 5-3: Spatial location of tested samples

Figure 5-4: Summary of Atterberg Testing on Plasticity Chart

Figure 5-5: p’ – q’ Triaxial Testing summary graph

Figure 5-6: Normal – shear stress analysis of Hoek Triaxial testing

Figure 5-7: Depth of “top of rock” from ground surface level (scale distance in m)

Figure 5-8: Cross section along upper portal/decline showing soft materials (scale distance in m)

Figure 5-9: Visual examples of RQD values

Figure 5-10: View of the 3D geotechnical model for Q’

Figure 5-11: Structural patterns along the upper decline path, with stereonets

Figure 5-12: Stereonet showing structural patterns and sets for Upper Level Decline Domain

Figure 5-13: Location of FW access drill holes considered for assessment

Figure 5-14: Stereonet showing structural patterns and sets for FW

Figure 5-15: Logged faults/shears along the drill hole (cross section looking north)

Figure 5-16: Section showing major structures

Figure 5-17: Median/35th percentile Q’ conditions along upper decline

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Figure 5-18: Typical ground condition core photos of upper decline

Figure 5-19: Footwall Q’ conditions using ore zone defining drill holes

Figure 5-20: Rock mass domains

Figure 5-21: MAP3D Model Geometry Construction

Figure 5-22: MAP3D View Showing Generalised Sequencing

Figure 6-1: Stability cross section locations

Figure 6-2: Modified Surface Infrastructure Design (May 2017)

Figure 6-3: Location of Raw Water Dam and IML for Modified Design

Figure 6-4: Slope directions used in the stability assessment

Figure 6-5: Boxcut slope design based on maintaining stability

Figure 6-6: Plan Showing Boxcut and Portal

Figure 6-7: Phase 2 Analysis – Sigma 1 Stress Contours

Figure 6-8: Phase 2 Analysis – Differential Sigma 1 Stress Contours Due to 60 tonne Loading

Figure 6-9: Shafts assessed by Mining One (looking north)

Figure 6-10: Possible faulting influence on shaft option positions

Figure 6-11: Ground support guidelines for mine drives of 4 to 6m span (Ref. 3)

Figure 6-12: 5m x 5m Arched Profile with 2.4m Long Bolts Spaced at 1.5m

Figure 6-13: Shantyback Profile for Ore Drive Flatbacking – 1.5m Bolt Spacing

Figure 6-14: Maximum left wall wedge for stope access drives (110° trend)

Figure 6-15: Maximum right wall wedge for cross-cuts at 40o trend

Figure 6-16: Sump mixing schematic example (from Sainsbury and Sainsbury, 2014)

Figure 7-1: Discontinuous Subsidence

Figure 7-2: Generalised schematic cross section of overhand cut and fill mining

Figure 7-3: Typical void when filling with a standard loader

Figure 7-4: Voids left when mining cut and fill orebody with CRF sill pillars

Figure 7-5: Scaled crown pillar span chart and serviceable life classes (after Carter et al, Ref. 16)

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APPENDICES

A. Geotechnical Cross Sections

B. Geotechnical Rock Mass Domains

C. Ore Body Rock Mass Quality

D. Raw Water Dam and IML Stability

E. Boxcut Stability Analysis and Design

F. Site Stabilisation and Erosion Letter Report

G. Shaft Location Assessment

H. Wedge Analysis of Mine Design

I. Roadheader Assessment Option

J. Risk Assessment Appraisal

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EXECUTIVE SUMMARY

Mining One conducted a geotechnical assessment for the Bird-In-Hand Gold (BIH) project during 2016. The BIH project is located approximately 3km from Woodside in the Adelaide Hills, South Australia and has been mined historically, dating back to the 1800’s.

The BIH ore zone is relatively narrow and moderately dipping (about 50°), with the ore zone striking approximately north-south (020o trend). The first mining production level is 155m below surface level and is planned to be mined to 445m below ground surface. Access to the mine is planned to be by decline. The proposed mining method is overhand cut and fill, where the underground voids are filled progressively, so no large voids are created during mining or left following mine closure.

The geotechnical assessment was based primarily on drill hole data collected by Terramin and mining plans, also provided by Terramin. The geotechnical assessment found that the clay material thickness at the surface varies significantly across the site. The rock conditions vary between Poor to Good and bedding/foliation is dominant in a generally blocky rockmass. Major structures will have a very important influence on stability of underground openings. Systematic grouting is expected to be required to control water inflows and this will improve the rockmass in terms of rock quality with the grout filling structures and making the rockmass more competent, when compared to the ungrouted state.

The mine was separated into Geotechnical Domains for the geotechnical assessment and are known as the Clay/Highly Weathered, Upper Level Decline, Middle Level and Lower Level Domains, with part of the Middle Level Domain and the Lower Level Domain including the ore body. Ground support derived for each Domain is dependent on the expected water inflows/pressures and ground conditions, but fibrecrete and either a resin bar bolt or a grouted splitset are the recommended ground support choices for development and production mine areas. Cablebolts will be required for wide spans in excess of 6m span.

The boxcut position allows access to rock conditions at shallow depth below surface, but some blasting may be required to break the rock at the base of the boxcut. When underground mining commences, the rock conditions are expected to be too hard for a roadheader, meaning drill and blast methods will be required. The alternative of moving the decline to position it within the Clay/Highly Weathered Domain will not allow roadheader excavation methods as the material is expected to be too soft to cut effectively.

Underground infrastructure is positioned to consider the mining method, induced stress conditions, ground conditions, and proximity to surface infrastructure. Shafts to surface will be influenced by the thickness of the surficial clay materials, with all current shaft position options likely to require lengthy pre-sinks with precast concrete liner within the clay materials.

Induced stress conditions are expected to be manageable for the current mining method and depth of mining. Recommended ground support systems utilise methods that are commonly used at many underground mines in Australia and are widely accepted for their success in controlling rock fall risk.

The surface stability of excavations was assessed for two designs provided by Terramin (2016 design and a May 2017 design). For the 2016 design, all the excavations assessed were found to be stable for the purpose designed for. This assessment is also seen for the May 2017 design for the peak production and closure design. However, for the construction phase of the

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IML infrastructure, a 5m vertical wall is designed as part of the stripping of the surface prior to the mullock being introduced as fill. This 5m high vertical wall in clay/highly weathered material is close to failure (Table 6 2). It is recommended that this batter be designed at approximately 40o to 45o, as recommended as a guideline in Table 6 1.

The most significant excavation assessed by Mining One was the boxcut and this is intended to be a temporary excavation with a structural tunnel liner inserted into the boxcut, with the boxcut then backfilled to ground surface level.

Surface stability due to underground void collapse and mining induced vibrations were assessed and both found to be unlikely. The mining method utilising backfill in the excavated voids has a large influence on maintaining surface stability as there are no large underground voids either during or following mining. The shallow depth of the mine, low induced stress change, small mining area, small blast sizes and blocky rockmass will mean the sudden release of energy due to mining is unlikely.

Mining One recommends the following be completed to verify the geotechnical model and design prior to commencement of mining:

Further work on the major structure model is required to gain as much confidence in the position and description of the major structures as reasonably possible. The major structures are very important to both the hydrogeology study (underground water inflows) and the stability of underground openings. Without depressurisation of the hangingwall faults and other prominent water bearing structures, Mining One consider that there would be a significant risk to the operation and to underground employees from ground failure (due to high water pressures) and flooding.

Confirmation of the grouting plan and expected water inflows/pressure into the underground openings and applying this information to check the influence of final assessment water inflows and grouting extents on the ground support design.

Dedicated investigation drill holes and geotechnical assessment is required for each shaft to ground surface.

The structural tunnel liner to be placed in the boxcut will require final engineering confirmation of the liner load bearing capacity and construction specifications for the footings and backfill types/compaction requirements.

CRF sill pillars will need proper design for each sill pillar location and width. The design work will be required to commence when there is access to the mullock on the IML stockpile, to allow for mixing of samples using mullock from the BIH mine.

Michael Bijelac Ben Roache Senior Geotechnical Engineer Geotechnical Manager

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1 INTRODUCTION AND SCOPE

Mining One Consultants Pty Ltd (Mining One) were engaged by Terramin Australia Limited (Terramin) to undertake a geotechnical assessment of the mining plan at the Bird-in-Hand Gold Project (BIH). The BIH project is located approximately 3km from Woodside in the Adelaide Hills, South Australia and has been mined historically, dating back to the 1800’s.

The BIH ore zone is relatively narrow and moderately dipping (about 50°), with the ore zone striking approximately north-south (020o trend). The first mining production level is 155m below surface level and is planned to be mined to 445m below ground surface. Access to the mine is planned to be by decline. The proposed mining method is overhand cut and fill mining, where the underground voids are filled progressively, so no large voids are created during mining or left following mine closure. Due to regulatory requirements, it is currently anticipated that mining will occur below the water table to cause minimal impact to groundwater levels at nearby sensitive receptors. A cross section layout of the project is shown in Figure 1-1.

Figure 1-1: Bird-in-Hand project layout – cross section looking north

1.1 Scope of Work

Mining One has completed the study with the help of Landloch (surface stabilisation and erosion experts) to deliver the project Scope of Work. Mining One has addressed the mine geotechnical aspects of the project, while Landloch has addressed items related to acid mine drainage (AMD) from the temporary surface integrated mullock landforms (IML) and erosion prevention design.

The scope of work interpreted by Mining One from information provided by Terramin is as follows. Scope items are numbered sequentially for ease of reference.

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Mine Geotechnical

Mine geotechnical study to address the following aspects:

1) Process the remaining drilling data and testing results for inputs into the design.

2) Geotechnical domaining of the project area.

3) Ground support design for mine areas, including:

a) The portal face and upper (near surface) decline.

b) The decline, accesses and ore drives (cut and fill stopes) in each domain.

4) Stress modelling of the underground mine voids, to use as general discussion and back up justification for placement of declines, shafts, stress development in Cemented Rock Fill (CRF) sill pillars etc.

5) Comment on shaft positioning and construction methodology.

6) Assessment of potential for surface subsidence due to mining, while operations are underway and following closure.

7) Describe the potential for natural geohazards impacting on the underground mine.

8) Include a section on the use of a Bebo style of entrance and detail a generalised diagram and plan with this to describe the construction in general terms.

9) Address the stability of all surface waste storages and boxcut.

10) Make recommendations for further mine geotechnical work programs prior to commencement of mining.

11) Assessment of the use of a roadheader in the upper decline. This assessment will focus on the appropriateness of using a roadheader, expected cutting rates and identification of problems that may be experienced.

12) Input into a risk assessment and write up of a separate report section for this. This will likely require a visit to Adelaide to attend a risk assessment workshop.

Landform Design – AMD and Erosion (Landloch)

13) Understand the mine plan, and review the available data.

14) Address methods of stabilisation and erosion control. Detail in a brief plan.

15) Address water movement through stockpiles and AMD potential.

16) Assessment of post-completion chemical and physical stability.

17) Address source, pathway and ultimate fate of potential mobile contaminants.

18) Write up a small report to insert into the submission for the mine Work Plan.

19) Make separate recommendations relating to the forward work program (items that will need to be resolved prior to commencing mining).

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In general, the work requested is effectively a desktop review of current plans and of the data supporting those plans. Rob Loch’s (Principal Consultant at Landloch) review would:

a) Identify any critical data gaps;

b) Assess current plans within the constraints of available data; and

c) Outline future actions or work needed to address any areas where data or investigation is needed to achieve any essential enhancements in planning or methods.

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2 BACKGROUND AND PROPOSED MINING METHOD

The BIH project is located in an area of historic gold mining, dating back to the late 1800’s. Water inflow was intersected in the original BIH mine, resulting in flows that were uneconomic at the time to control, and the mine was abandoned. There has been no underground mining at BIH for about the last 100 years. The historic workings are shown in red within Figure 1-1 and are indicated to extend to about 120m below surface level.

A scoping study was completed by Terramin for the BIH project in October 2013 (Ref. 1). A basic geotechnical assessment was made and concluded mining could proceed utilising a cut and fill mining methodology. Since that time, further drilling, laboratory testing and test pit data has been collected for the current study, as well as more advanced mine plans made by Terramin to define the potential BIH operations.

The proposed cut and fill mining method has been assessed by Terramin, and a preliminary mine design has been developed by Terramin. The design has the following details:

Targeting multiple reefs, up to 15m wide in wider reef or multiple reef locations;

Single central level accesses;

4 lifts per central stope access;

CRF sill pillars required per central access block; and

Drift and fill with CRF for wider areas.

The generalised proposed sequence for the four cut and fill lifts per access are as follows:

1. Decline down at 1 in 7 to access the bottom cut and fill lift (green drives in Figure 2-1);

2. Strip the floor out from bottom lift and backfill with CRF to produce a sill CRF pillar (blue drives in Figure 2-1);

3. Access into the 2nd cut and fill lift above the bottom lift (orange drives in Figure 2-1), this leaves a final crown sill pillar below the CRF; and

4. Mine and backfill the final sill pillar in the block below the CRF of the lift above.

Figure 2-1: Orebody cross section showing cut and fill drives

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3 GEOLOGY

The following regional geological description was provided by Terramin (shown in italics below).

“Primary gold mineralisation at Woodside is located within a synclinal structure in Adelaidean strata at the transition between the Warrina and Heysen Supergroup with the fold axis plunging at an angle of about 45° to the east. The syncline is cut on the eastern side by a major north-south-striking, east-dipping Nairne Fault, along which younger Cambrian metasediments of the Kanmantoo Trough were thrust over older Adelaidean Strata. The majority of gold was mined from vein structures hosted by the lower Umberatana Group. This group unconformably overlies clastic metasediments of the Burra Group, which forms the upper part of the Warrina Supergroup. The Burra Group contains only minor mineralisation in the Woodside area”.

The local geological description was provided by Terramin and was added to by Mining One (shown in italics below).

“The hangingwall of the ore zone is the Tarcowie Siltstone of the Umberatana Group which is characterised by fine grained interbedded metasandstones and metasiltstones. These rock types display varying quantities of biotite resulting in a banded appearance. In places where they have decomposed and become unconsolidated they can cause difficulties during drilling, often resulting in the need to case off with HQ rods and continue with NQ sized drilling.

The Brighton Limestone which hosts the Bird in Hand mineralisation is stratigraphically beneath the Tarcowie Siltstone and is typically characterised by upper and lower coarse grained and recrystallised marble units with fine grained calcareous siltstones between.

The blue-grey Tapley Hill Formation which is located in the footwall of the Brighton Limestone and also where the mine infrastructure is located (such as the decline and shafts) is a partly dolomitic siltstone often is finely laminated, cross-bedded or shows sedimentary features. In areas it is rich in pyrite and organic matter. Many of the major goldfields in the Mt. Lofty Ranges are partly hosted by the Tapley Hill Formation or in stratigraphic units close to it”.

The complex local geology has meant that a three dimensional model of each lithology has not been constructed. Defining the geotechnical (stability) performance of the rock has not relied on the rock type but instead focused on the available structural and rock quality information collected during core drilling, as discussed in Section 5.

3.1 Key Geotechnical Considerations

Geological issues that are expected to contribute to the stability performance of the mine are:

The dip of the bedding/foliation. The bedding/foliation is interpreted to be dipping at 46°/101° (dip/dip direction) throughout the site.

There is a lack of geotechnical drilling for the deeper part of the mine. This is planned to be addressed by Terramin in future drilling, and will assist in confirming the ground conditions at depth.

The major structure model is critical to the success of the project, due to the adverse stability influence major structures have on underground mining voids. Terramin are planning to continue refining the structural model over time.

The depth of clay material, which is defined by the “top of rock” surface, is critical for positioning underground infrastructure.

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4 GROUNDWATER MODEL

The BIH project is located in an area surrounded by farms and local wineries. Groundwater bores used by local farmers and wineries is prevalent and groundwater is an important environmental issue for the project. It is understood that the mine will depressurise the hangingwall to prevent water from entering the mining operations initially and then re-injecting water back into the aquifer to save pumping and water treatment costs. Grouting would then be used to manage any inflows encountered on top of this strategy. It is understood from historical mining anecdotes that mining of the upper areas was wet and that water seepage into some mine openings was significant.

To investigate the hydrological system at BIH, Terramin have required assistance from groundwater specialist consultants and a grouting expert. It is understood that this work is currently ongoing and likely to finish outside the timeframe of this report.

The major rock units defining the groundwater water model can be summarised as follows:

Footwall (FW) Metasediments (Tapley Hill) – This is where the decline and FW infrastructure are located;

Marble/Limestone unit hosting the reefs and orebody;

Hangingwall (HW) interbedded Metasiltstones and Metasandstones; and

HW faults.

Three major faults have been provided as 3D wireframes to Mining One and these are shown in Figure 4-1 and Figure 4-2 as a long section and plan respectively.

It is understood from Terramin that from pump testing completed to date, the Lower HW fault is a significant water bearing structure. This fault appears to define the southern extents of the orebody and will be exposed in the drives at the southern end.

The groundwater modelling specialists engaged by Terramin have developed a regional model to assist in the assessment of potential impacts to environmental receptors. This model was also used to predict the potential inflows into the mine drives in a range of scenarios.

Figure 4-3 and Figure 4-4 (provided by Terramin) show the predicted inflows from a 10m local area that is un-grouted (assumed all other drives are grouted) at levels 150m and 300m depth respectively (provided by Terramin’s groundwater consultant and assuming no depressurisation). The inflows have been predicted for the following scenarios:

Base Case – decline and drives open and grouting is 70% effective;

10m segment of FW drive (Tapley Hill) open and un-grouted, while the rest of the drives are grouted to 70% effectiveness;

10m segment of Marble/Limestone/Ore drive open and un-grouted, while the rest of the drives are grouted to 70% effectiveness; and

10m segment of HW fracture zone (Lower HW Fault) drive open and un-grouted, while the rest of the drives are grouted to 70% effectiveness.

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The results appear to indicate the following:

The Lower HW fault will produce extremely large amounts of water when exposed and will need to be carefully managed both from inflow and stability purposes. Based on the predicted results the pressure head could be a few hundred metres;

The marble/limestone/orebody unit will produce approximately half of the water that the Lower HW structure will produce in the upper levels, but not much water inflow is predicted in the lower levels;

For the FW Metasediments where the decline and FW infrastructure is located, no increase in flow is predicted when compared to the base case of effectively grouting all the drives to 70% effectiveness; and

The base case model indicates that when all drives are grouted to 70% effectiveness, the total mine inflow will still be up to a maximum of 30L/s.

The implications of these groundwater inflows and pressures on the geotechnical design particularly for ground support are as follows:

Ideally de-pressurisation of the Hangingwall and ore zone marble/limestone would be essential if there is a hydraulic connection of hundreds of metres. If this is not possible due to environmental issues, then grouting of all water producing drives should be conducted and investigated in terms of effectiveness;

It appears that the decline and most of the FW infrastructure should have relatively low inflow and pressure, however will likely be in a wet environment which may affect the corrosion rate of ground support; and

Groundwater pressures will need to be incorporated into the ground support design.

During this study period, no allowance for the effectiveness of grouting the drives to limit/reduce/eliminate water inflow and pressures has been included in the analysis. The regional groundwater model results were only available to Mining One near the completion of reporting. It is understood that Terramin are engaging a grouting expert to assess the level of grouting required to operate the mine successfully. These results are not factored into this report.

Figure 4-1: South west view of mine design and HW and FW faults

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Figure 4-2: Plan at 200m RL +/- 10m showing faults


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Without depressurisation of hangingwall and other prominent water bearing structures, Mining One consider that there would be a significant risk to the operation and to underground employees from ground failure (due to high water pressures) and flooding.

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5 GEOTECHNICAL MODEL

5.1 Test Pits and Geotechnical Drilling

5.1.1 Test Pits

Test pits were used initially to assist with defining the “top of rock” surface in the area of the proposed boxcut. When drilling identified a deep weathering profile immediately to the south east of the proposed boxcut position, the test pits were extended to better define the top of rock surface and locate rock material for the potential decline location. Mining One logged test pits TP1, TP2, TP3 and the Sumps, while Terramin provided the logging results for the remainder of the pit locations (refer to Figure 5-1).

Other excavations in the area of the boxcut were used to assist with defining the top of rock, such as shallow heritage investigation pits and drill rig sumps. A list of the available test pits and sumps is shown in Table 5-1.

Table 5-1: Test pit details

Pit ID Northing

(mN) Easting

(mE) Elevation

(mRL) Rock Depth

(m) Max Depth

(m) Pit Type

TP1 308460.5 6129911 412.8 0.9 3.2 Mining One Test Pit

TP2 308493 6129898 417.7 1.7 2.8 Mining One Test Pit

TP3 308535.7 6129926 417.7 1.7 5 Mining One Pit

SUMP 308546.7 6129896 415.723 1.5 2.5 Drill rig sump

SUMP_2 308553 6129857 418.8666 1.5 2.5 Drill rig sump

PIT1 308582 6129895 418.18 3.4 3.5 Terramin Test Pit

PIT2 308551.3 6129827 422.492 No rock 5.4 Terramin Test Pit

PIT3 308513.9 6129823 419.321 2 2.6 Terramin Test Pit


PIT6a 308518.7 6129754 424.308 2.5 2.7 Terramin Test Pit

PIT8 308594.2 6129864 420.532 No rock 6.7 Terramin Test Pit


PIT21 308473 6129840 418.182 1 1.5 Terramin Test Pit

PP02 308515 6129845 417.16 0.6 0.7 Heritage Pit

PP03 308513 6129855 418.65 No rock 0.85 Heritage Pit

PP04 308511 6129865 415.76 0.75 0.85 Heritage Pit



PP07 308503 6129894 413.7 0.85 0.95 Heritage Pit

PP08 308502 6129906 413.5 0.87 0.97 Heritage Pit

PP09 308507 6129911 412.99 0.9 1 Heritage Pit


PP11 308517 6129893 415.269 0.93 1.03 Heritage Pit

PP12 308521 6129881 417.027 0.8 0.9 Heritage Pit

PP13 308520 6129872 418.5 0.7 0.8 Heritage Pit

PP14 308524 6129865 419.5 0.76 0.86 Heritage Pit

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Pit ID Northing

(mN) Easting

(mE) Elevation

(mRL) Rock Depth

(m) Max Depth

(m) Pit Type

PP15 308523 6129852 421.6 0.74 0.84 Heritage Pit

pp16 308532 6129835 421 0.8 0.9 Heritage Pit

PP17 308543 6129835 423.5 0.98 1.08 Heritage Pit

PP18 308533 6129851 423.2 0.85 0.95 Heritage Pit


PP20 308441 6129899 409.2 0.6 0.7 Heritage Pit


Figure 5-1: Plan view of pit locations relative to underground mine design (portal and upper decline)

5.1.2 Drilling and Logging

The geotechnical drilling program was conducted and completed by Terramin which included sampling and logging of the drill core. Mining One provided the following assistance to Terramin for the data collection process:

Initial input into the quantity and design of the geotechnical drill holes;

Initial training for Terramin geologists to complete geotechnical logging; and

Sampling and testing requirements.

Table 5-2 shows the drill hole details for the geotechnical drilling program, while Figure 5-2 shows where the holes are located spatially in relation to the proposed mine design. Holes BH024, 028, 029 and 044 (lower two thirds of ore body definition) were not part of the geotechnical program. These holes were requested by Mining One to be included in the geotechnical assessment because there was limited geotechnical data below the upper part of

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the stoping area. Mining One logged these holes from core photos with reference to the geology logs and RQD estimates already completed by Terramin.

Table 5-2: Drill hole details

Hole ID Northing (mN) Easting (mE) Elevation (mRL) Max Depth (m) Azimuth (o) Dip (o)

BH047 6129800.8 308642.0 426.6 99.7 347.0 -58.0

BH048 6129829.0 308598.2 422.7 91.0 121.0 -59.0

BH049 6129892.7 308541.1 415.7 40.0 360.0 -90.0

BH050 6129730.9 308727.1 433.7 150.0 158.0 -72.0

BH051 6129709.5 308991.7 454.5 190.6 336.0 -86.0

BH052 6129847.8 308554.8 419.5 132.5 263.0 -53.0

BH053 6129788.4 308697.3 427.0 120.0 265.0 -54.0

BH054 6129698.0 309005.9 455.4 213.9 36.0 -86.0

BH055 6129584.0 308915.8 447.2 222.9 307.0 -71.0

BH056 6129693.6 309004.5 455.2 218.0 360.0 -90.0

BH057 6129653.5 309041.1 452.4 225.0 275.0 -83.0

BH058 6129693.0 309003.2 455.1 246.3 119.0 -84.0

BH059 6129648.9 309035.8 452.2 251.8 287.0 -87.0

BH024 6129562.3 309152.9 434.7 381.9 295.0 -70.0

BH028 6129541.1 309251.2 444.4 395.0 284.0 -69.0

BH029 6129593.5 309106.2 433.9 267.2 295.0 -72.0

BH044 6129495.5 309206.0 437.0 470.7 296.0 -87.0

Figure 5-2: East west view showing drill holes and mine design

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5.2 Sampling and Laboratory Testing

The following sample types were collected by Terramin for testing:

Atterberg Limits (Clay/Highly Weathered Domain);

Emerson Class Number (Clay/Highly Weathered Domain);

Particle Size Distribution (Clay/Highly Weathered Domain);

Consolidated Undrained Triaxial, CU (Clay/Highly Weathered Domain);

Unconfined Compressive Strength, UCS (Clay/Highly Weathered Domain and Rock Domains);

Modulus (Clay/Highly Weathered Domain and Rock Domains); and

Hoek Cell Triaxial (Rock Domains).

Sample locations for CU, UCS, Modulus and Hoek Cell are shown in Figure 5-3, and indicate that most of the samples have been collected from the upper part of the decline. All of the testing was completed by TRILAB in Brisbane (NATA accredited testing).

Figure 5-3: Spatial location of tested samples

5.2.1 Clay/Highly Weathered Domain

5.2.1.1 Nature of Clay Material

To ascertain the characteristics of the variable depth of clay material, three tests were used:

Atterberg Limits;

Emerson Class Number; and

Particle Size Distribution (PSD).

These three tests were completed on samples collected from two holes (BH048 and BH050) at varying downhole depths. The test results for the Atterberg Limits are summarised in Table 5-3. These results are also graphed on a Plasticity chart shown in Figure 5-4.

The test results for the Emerson Class Number are shown in Table 5-4.

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In addition, particle size distribution (PSD) testing was completed for two holes with numerous downhole depths tested as follows:

BH048 – Tested at 7.4, 14.7, 24.2, 30.3, 39.4 and 45.5m downhole depths; and

BH050 – Tested at 5.2, 14.6, 24.6, 33.7, 42.0, 52.8, 62.0, 71.7, 79.2, 91.2m downhole depths.

Table 5-3: Summary of Atterberg Limits testing

Hole ID Hole

Depth (m) Liquid

Limit (%) Plastic

Limit (%) Plasticity Index (%)

Linear Shrinkage

(%)

Moisture Content

(%)

BH048 7.40 35 21 14 4.0 3.8

BH048 14.70 39 26 13 2.5* 4.6

BH048 24.20 37 24 13 3.5* 1.9

BH048 30.30 34 23 11 3.0* 2.0

BH048 39.4 35 22 13 5.5* 3.2

BH048 45.50 34 21 13 5.0* 3.3

BH050 5.20 27 17 10 2.5* 1.4

BH050 14.60 35 20 15 5.5 7.3

BH050 24.60 24 17 7 3.0 1.9

BH050 33.70 62 34 28 8.0 10.3

BH050 42.00 23 15 8 2.5 0.7

BH050 52.80 39 22 17 3.5* 1.6

BH050 62.00 53 29 24 3.0 7.0

BH050 71.70 42 24 18 2.5 0.8

BH050 79.20 33 24 9 3.0 5.1

BH050 91.20 34 23 11 2.5* 3.3

* Cracking occurred

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Figure 5-4: Summary of Atterberg Testing on Plasticity Chart

Table 5-4: Summary of Emerson Class testing

Hole ID Hole Depth (m) Description Emerson Class Number

BH048 7.40 SILT- pale grey/white 2



BH048 30.30 SILT- pale brown 2








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Hole ID Hole Depth (m) Description Emerson Class Number






5.2.1.2 Triaxial Testing

Consolidated undrained triaxial testing of five samples were completed from drill holes BH048 and 053. The triaxial testing comprised of 3 stages per test ranging from 200kPa to 1500kPa. Results of the tests for the five samples are shown in Table 5-5.

An interpretation of the results for each test sample is shown in Table 5-6 with Cohesion, Friction Angle and UCS derived. In addition, the correlation between logged strength estimate and UCS derived from testing is shown. It indicates that three of the samples (highlighted red) have been overestimated in terms of logged strength, while the other two samples (highlighted yellow) tested at the lower end of the logging estimate strength range.

Figure 5-5 shows the results of the triaxial testing on a p’ – q’ graph and shows an upper and lower strength range for four of the five tests. The upper and lower range is as follows:

Lower Range – Cohesion = 25.6kPa,

Friction Angle = 24.7°,

UCS = 80kPa.

Upper Range – Cohesion = 107kPa,

Friction Angle = 26°,

UCS = 343kPa.

The deeper sample from BH053 (63.4m hole depth) indicates a higher strength of 839kPa, which is likely to indicate that the strength increases with depth. This strength is still within the R0 strength logging category, but is approaching the lower end of the R1 strength category cut-off (1MPa).

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Table 5-5: CU testing results - clays

Table 5-6: Correlation between derived strength and logged strength estimate

Sig 1 (kPa) Sig 3 (kPa)

200 632 154 1.01

399 1037 301 1.69

599 1342 409 3.90

351 1001 234 1.24

701 1742 484 2.04

1051 2279 699 3.43

97 373 80 1.80

195 625 172 2.54

295 831 260 3.26

151 315 100 1.51

298 598 212 2.42

448 870 336 2.97

498 1988 332 0.69

999 3271 791 1.17

1500 4183 1074 3.07

SILTY SAND‐white 2.17 17.0 9716040137 BH053 63.4

1.98 25.3 9716040136 BH053 21.6CLAYEY SILT‐

yellow/brown/white

27.7 9716040135 BH048 13.3CLAYEY SILT‐

yellow/brown1.95

SANDY SILT‐mottled

pale grey/yellow‐

brown

1.95 22.8 9716040134 BH048 40.9

1.99 25.6 9816040133 BH048 25.4Clayey Silt

yellow/brown

Principle Effective Stresses

Strain (%)DescriptionWet Density

(t/m3)

Initial

Moisture

Content (%)

B Response

(%)

Effective

Pressure (kPa)

Lab

Sample No.Hole ID

Hole

Depth

(m)

BH053 63.4 SILTY SAND‐white Logged as lower end strength category

Hole IDHole

Depth (m)Description

BH048 25.4Clayey Silt

yellow/brown

BH048 40.9

SANDY SILT‐mottled

pale grey/yellow‐

brown

289 29.5 0.99 R0/R1 R1 1 ‐ 5

0.085 S2 R1 1 ‐ 5 Logged as 4 categories above tested strength28 23.8BH053 21.6CLAYEY SILT‐

yellow/brown/white

R0 0.4 ‐ 1 Logged as 2 categories above tested strength55 25.8 0.18 S3BH048 13.3CLAYEY SILT‐

yellow/brown

Logged as lower end strength category112 27.9 0.4 S4/R0 R0 0.4 ‐ 1

0.2 S3/S4 R0 0.4 ‐ 1 Logged as 1 to 2 categories Above tested strength61 28.1

Strength

Logged

As

ISRM

Strength

(MPa)

CommentDerived

Cohesion

(kPa)

Derived

Friction

Angle (o)

Derived

UCS

(MPa)

Equivalent

ISRM

Strength

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Figure 5-5: p’ – q’ Triaxial Testing summary graph

5.2.2 Rock Domain

A total of eight UCS tests were completed on fresh core samples from four different holes. A summary of the testing results is shown in Table 5-7. The results indicate that most of the samples are classified as “Very Strong Rock”, apart from two samples which failed on shear and of which reached “Medium Strong” category.

All of the samples when logged were estimated to be of R4 strength (Strong Rock), however typically the testing has indicated that most of the samples are likely to be of R5 strength (Very Strong), apart from the two samples which failed on shear (Table 5-8).

The Young’s Modulus testing has indicated that the intact stiffness is in the range of 50 to 80GPa as shown in Table 5-9.

Five samples across three holes were also tested using the Hoek Cell triaxial method, with each sample undergoing one stage of confinement testing. The confinement pressure ranged from 5 to 40MPa, with the results shown in Table 5-10. Figure 5-6 summarises the Hoek triaxial testing on a shear-normal plot. The results indicate a poor correlation between the samples with a Mi value of approximately 50, which shows a poor fit of the individual data. This is likely due to samples being tested across different holes with likely slightly different geological types being compared.

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Table 5-7: Summary of UCS test results

Table 5-8: Comparison of logged UCS estimate versus tested strength

Table 5-9: Summary of Young’s Modulus testing results

Test Type

Hole_ID Hole Depth (m)

Vertical Depth (m)

Wet Density (t/m3)

Failure Mode UCS (MPa)

Young’s Modulus (GPa)

Tangent Secant

MOD BH049 38.7 38.7 2.68 Conical 124 78 74

MOD BH049 27.4 27.4 2.64 Shear 40 67.8 61.8

MOD BH052 61.3 50.1 2.8 Disintegration 152 77.3 77.3

MOD BH052 82.7 67.7 2.77 Shear 103 78.9 73.8

MOD BH047 96.2 83.1 2.67 Shear 41 51.9 52.2

Table 5-10: Summary of Hoek Triaxial testing results

Test Type Hole_ID Hole

Depth (m)

Peak (MPa) Residual (MPa) Young’s Modulus

Sig 3 Sig 1 Sig 3 Sig 1 Peak (GPa)

HOEK BH049 21.8 5 181 5 25.1 18.8

HOEK BH049 37.2 10 158 10 83.6 15

HOEK BH052 62.9 20 308 20 98.3 18.3

HOEK BH052 83.3 30 276 30 151 17.2

HOEK BH047 95.3 40 190 40 186 11.5

UCS BH055 161 151 2.7 Disintegration 184

UCS BH055 168 158 2.74 Disintegration 186

UCS BH055 188.5 177 2.73 Disintegration 181

UCS & MOD BH049 38.7 38.7 2.68 Conical 124

UCS & MOD BH049 27.4 27.4 2.64 Shear 40

UCS & MOD BH052 61.3 50.1 2.8 Disintegration 152

UCS & MOD BH052 82.7 67.7 2.77 Shear 103

UCS & MOD BH047 96.2 83.1 2.67 Shear 41

UCS (Mpa)Wet Density (t/m3) Failure ModeTest Type Hole_IDHole

Depth (m)

Vertical

Depth (m)

UCS BH055 161 184 R5 R4 50 ‐ 100 Logged as 1 category below Lab Strength

UCS BH055 168 186 R5 R4 50 ‐ 100 Logged as 1 category below Lab Strength

UCS BH055 188.5 181 R5 R4 50 ‐ 100 Logged as 1 category below Lab Strength

UCS & MOD BH049 38.7 124 R5 R4 50 ‐ 100 Logged as 1 category below Lab Strength

UCS & MOD BH049 27.4 40 R3 R4 50 ‐100 Logged as 1 category above Lab Strength

UCS & MOD BH052 61.3 152 R5 R4 50 ‐ 100 Logged as 1 category below Lab Strength

UCS & MOD BH052 82.7 103 R5 R4 50 ‐ 100 Close to higher end of R4 which was logged

UCS & MOD BH047 96.2 41 R3 R4 50 ‐ 100 Logged as 1 category above Lab Strength

Equivalent

ISRM

Strength

Strength

Logged

As

ISRM

Strength

(MPa)

CommentTest Type Hole_ID

Hole

Depth

(m)

Lab

UCS

(MPa)

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Figure 5-6: Normal – shear stress analysis of Hoek Triaxial testing

5.3 Rock Surface

A top of rock surface was developed by Mining One, based on the drilling and test pit data. This surface is shown in Figure 5-7. In the northwest of the site at the boxcut location, the rock depth is shallow and within 4m of the ground surface level. The rock depth increases to the east and south east.

Figure 5-8 shows a section of the “Soft Material” model, together with the boxcut and decline design. It shows that the soft material domain boundary dips at a similar angle to the decline path with a cover distance of greater than 22m, but as little as 8m in the upper decline. The decline has a rock cover of at least 20m when the decline reaches half of the first spiral, as shown in Figure 5-8. To reach half way around the first decline spiral, the distance is approximately 90m.

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Figure 5-7: Depth of “top of rock” from ground surface level (scale distance in m)

Figure 5-8: Cross section along upper portal/decline showing soft materials (scale distance in m)

5.4 3D Rock Quality Model

A 3D rock quality model was constructed for the site, using the drillhole database provided by Terramin. A rock quality model using the Rock Quality Designator (RQD), which measures the frequency of natural rock structures in each logged interval, and using the Q System (Q´).

The RQD model contains a larger dataset compared to the Q’ model, as many of the older drill holes have recorded RQD values. The RQD model can be interpreted with the assistance of

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Figure 5-9 which shows examples of RQD in relation to the appearance of typical drill core in the core tray.

Figure 5-9: Visual examples of RQD values

Q´ values can be interpreted with the assistance of Table 5-11, which describes the ground condition in relation to the Q value.

Table 5-11: Descriptions associated with the Q System

Ground Conditions Q

Exceptionally Poor 0.001 - 0.01

Extremely Poor 0.01 - 0.1

Very Poor 0.1 - 1

Poor 1 - 4

Fair 4 - 10

Good 10 - 40

Very Good 40 - 100

Extremely Good 100 - 400

Exceptionally Good 400 - 1000

A structural trend of 46°/101° was applied to the model to allow for an anisotropic influence (bedding) at BIH, and can been seen in a general overview of the model in Figure 5-10. Geotechnical cross sections for BIH are presented in Appendix A.

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Figure 5-10: View of the 3D geotechnical model for Q’

5.5 Structural Model

An assessment of the structural conditions for the footwall (FW) infrastructure positions was completed by using the following data:

Major structures model provided by Terramin; and

Structural logging from the geotechnical drill holes provided by Terramin.

A summary of the likely structural conditions are outlined in Sections 5.5.1 to 0.

5.5.1 Decline

The decline route and geotechnical drill holes provided to assess the ground conditions are shown in Figure 5-11. The holes comprise of the following:

BH049, 052, 048, 047, 053, 050 and 055.

No drill holes with structural logging were available for the middle and lower part of the decline spiral below BH055.

Stereonets showing the structural patterns from each hole are also shown in Figure 5-11, and these indicate the following:

The bedding is consistently dipping at a moderate angle towards the east for all of the drill holes along the decline route;

Faults and shears are irregular in terms of orientation and dip; and

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The joint patterns are also too irregular from hole to hole, and within the holes to form strong joint sets.

Given the consistent orientation and dip nature of the bedding along the decline route and the irregular nature of the jointing and faults/shears, one structural domain has been defined along the decline route for the design assessment. Three structural sets have been defined as part of the domain and these are defined in Table 5-12, while Figure 5-12 shows the structural patterns and defined sets within a stereonet.

Figure 5-11: Structural patterns along the upper decline path, with stereonets

BH055

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Figure 5-12: Stereonet showing structural patterns and sets for Upper Level Decline Domain

Table 5-12: Structural sets for Upper Level Decline Domain

Set No. Defect Type Dip/Dip Dir’n (°) Infill Type Roughness

1 Mostly Bedding 46/101 Mostly Clean Mostly PS

2 Joint 18/258 Mostly Clean PR and PS

3 Joint 25/334 Clean and CA Mostly PS

PR – planar rough, PS – Planar smooth, CA - Calcite

5.5.2 FW Access Infrastructure

The FW access infrastructure structural conditions were assessed by using drill holes shown in Figure 5-13. These holes have been drilled from the hanginwall (HW) and cross the orebody into the FW with varying hole depths. The lower holes comprising of BH029, 24, 28 and 44 however did not contain any structural logging. As a result only the upper holes were useful and comprised the following:

BH051, 054, 057 and 059. BH056 and 058 also had no structural logging.

Of the four holes listed above, only two holes (054 and 057) had structural measurements through the orebody reef zones and into the FW.

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The structural patterns from these two holes combined are shown in stereonet format in Figure 5-14, and the following conclusions are made:

Although limited in quantity, bedding measurements were consistent with the decline measurements, being moderate dipping towards the east (measured from HW core);

No bedding measurements were logged in the footwall rockmass from the two holes; and

The jointing patterns are relatively shallow to moderate dipping, with five possible sets defined.

Table 15-3 describes the indicative joint sets derived from the patterns in the stereonet. Given the limited data, some of the joint sets are described as “weak” in terms of confidence.

Figure 5-13: Location of FW access drill holes considered for assessment

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Figure 5-14: Stereonet showing structural patterns and sets for FW

Table 5-13: Structural sets for the FW

Set No. Defect Type Dip/Dip Dir’n (°) Infill Type Roughness

1 Mostly Joints 14/055 Mostly CH Mostly PS

2 Mostly Joints 47/035 Mostly Clean Mostly PR

3 Mostly Joints 30/261 Mostly Clean an CA

Mostly PR

4 Mostly Joints 37/196 No dominant infill

PR

5 Mostly Joints 50/311 Mostly Clean PR

PR – planar rough, PS – Planar smooth, CA – Calcite, CH - Chlorite

5.5.3 Major Structures

Mining One were provided with a 3D wireframe showing HW and FW major structures described by Terramin as faults. These wireframes together with the geotechnical logging of the available drill holes were assessed.

Figure 5-15 shows a longitudinal section of the decline route with the traces of the geotechnical drill holes. Along the drillhole traces where faults/shears have been logged both in the structural and interval logging (rockmass), red indicators are marked to show the logged position.

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The faults/shears provided by Terramin as 3D wireframes are shown in section in Figure 5-16 together with the mine design and drill hole traces and also with the location of faults/shears logged. It shows the following:

The mineralised reef zones are typically striking 020°:200° and are moderately dipping towards the east;

The two HW faults and the FW fault are parallel to each other, striking approximately 070o:250o, and dipping to the south at a moderate angle;

The lower HW fault and the FW faults appear as bounding structures to the north and south for the orebody; and

The lower HW fault is expected to have significant amounts of water pressure and flow once intercepted by drives in the southern area (Refer to Section 4).

Figure 5-15: Logged faults/shears along the drill hole (cross section looking north)

Orebody

Location

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Figure 5-16: Section showing major structures

5.6 Rock Mass Domains

To assess the interval data from the core logging, the Modified Rock Tunnelling Quality Index Q’ was determined from the rock mass interval logging undertaken by Terramin. The Modified Rock Tunnelling Quality Index Q’ does not incorporate the active stress quotient of Jw/SRF (water pressure and mining induced stress). Therefore the Modified Rock Tunnelling Quality Index Q’ is calculated using the following expression:

Ja

Jr

Jn

RQDQ '

Where;

RQD is the Rock Quality Designation for each interval;

Jn is the number of defect sets;

Jr is the Joint (defect) roughness; and

Ja is the Joint (defect) alteration number which is determined by the type and thickness of infill.

Mining One’s ratings in this section do not account for any stress/water effects.

The quotient RQD/Jn assesses the blockiness of the rockmass, while Jr/Ja defines the inter-block shear strength. The active stress quotient incorporating water pressure and mining induced stress is addressed in Section 0.

Ground conditions along the upper decline route were assessed by determining Q’ for each logging interval for each hole, and statistically determining the median and 35th percentile values. Figure 5-17 visually shows the typical ground conditions (median and 35th percentile values) that can be expected along the upper decline route. The Q/Q’ classification system is log based with Table 5-11 defining ground condition categories and equivalent Q values. The significance of using the 35th percentile is discussed in Section 6.3.1.2.

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From the portal to approximately 308,700mE, the Q’ values indicate Fair to Good ground conditions, while the decline further down towards the spiral is indicated to be of poorer quality rock based on holes BH050 and BH055. Ground conditions in this part of the decline are estimated to be Poor to Fair based on Q’. This is one category lower than the upper part of the decline. Figure 5-18 shows typical rock core trays of each hole down the decline as a comparison.

No geotechnical holes were available to assess the ground conditions below 250RL for the decline spiral which is located to the south-west of the orebody. As a result, the ore reef defining holes drilled from the HW were used to assess the FW conditions when they intercepted and passed the orebody. The locations of the pierce points of the holes with the main orebody can be seen in Figure 5-19.

The FW ground conditions in the upper part of the ore defining holes (above 200RL) indicate quite a few holes with Poor to Fair ground conditions to be expected, similar to the lower part of the main decline discussed above. As a result it was considered appropriate to combine these two areas as one geotechnical domain. Below 200 RL, the deeper FW ground conditions appear to improve, however this is only based on four drill holes.

Figure 5-20 shows the derived FW geotechnical domains based primarily on Q’ for the rock domains. The geotechnical domains are summarised as follows:

1. Clay/Highly Weathered Domain (strength R0 and lower);

2. Upper Level Decline Domain (portal to approx. 308,700m E);

3. Middle Level Domain (east of 308,700m E to 200m RL and above); and

4. Lower Level Domain (below 200m RL).

Appendix B shows the variation in parameters such as Q’, RQD, logged strength etc, for each of the three rock domains as histograms, and Appendix C shows summary pictures of the expected orebody conditions.

Figure 5-17: Median/35th percentile Q’ conditions along upper decline

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Figure 5-18: Typical ground condition core photos of upper decline

Figure 5-19: Footwall Q’ conditions using ore zone defining drill holes

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Figure 5-20: Rock mass domains

5.7 Stress Model

5.7.1 Introduction

To assess for the likely mining induced stress conditions based on the mine design, a 3-dimensional stress analysis model was constructed of the BIH mine using the software program MAP3D (Ref. 4). MAP3D is a Rock Mechanics and Stress Analysis program for the Mining, Civil and Geomechanics Industries.

Map3D is a fully integrated three-dimensional layout (CAD), visualisation (GIS) and stability analysis package using the Boundary Element Method (BEM) of stress analysis.

Map3D is suitable for building and modelling rock and soil engineering design problems involving both irregular 3D massive excavations, tunnels and tabular shapes. It can be applied to the analysis of underground layout and mining sequence problems, as well as the assessment of pillar designs, stope span stability and fault stability.

Like most other stress analysis programs, it can be used in two ways:

1. Given detailed and accurate input data, to perform precise computations and make firm predictions regarding the response of a system such as an underground mine or mine block; or

2. Given qualitative or uncertain data with respect to the rock properties, distributions of different rock units, faults and in situ conditions a comparative study can be completed for various mining methods and sequences to determine the most appropriate for the conditions.

5.7.2 Material Properties

To Mining One’s knowledge, no detailed 3D geological wireframe model has been developed for BIH to date. The MAP3D model constructed for the BIH project was simplified to consist of one major host unit (geological unit). The material properties estimated for the host unit are summarised in Table 5-14 and are based on:

Lower Percentile intact UCS strength (Table 5-7);

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Median Rock Mass Quality Strength – Q’ (Section 5.6); and

Mi constant of 10 (sediments) given the Hoek triaxial test results provided unreliable mi constant results (i.e. close to 50).

Table 5-14: Material Properties Used for MAP3D Modelling

Tension

Cut-Off (MPa)

mi UCS

Intact (MPa)

Q’ Equivalent

GSI

Hoek-Brown Parameters

Young’s Modulus

(Rockmass)

(GPa)

Poisson’s Ratio

m s

-0.6 10 100 10 64 2.77 0.0183 22.4 0.25

5.7.3 In-Situ Stress Conditions

To Mining One’s knowledge, no stress measurements have been completed in the Adelaide Hills region. Previous work by Mining One for the Angas Zinc Mine (Ref. 5) assumed a regional stress condition in the Adelaide Hills with the major principal stress (sigma 1) direction being east-west. As this is our best estimate of stress conditions, Table 5-15 summarises the in-situ stress conditions used in the BIH modelling. A horizontal to vertical stress ratio of two has been assumed to exist at the mine at depth.

Table 5-15: Assumed In-Situ Stress Field

Stress Component Magnitude (MPa/m) Trend (o) Plunge (o)

Sigma 3 vertical 0.027 x Vertical Depth - 90

Sigma 2 Horizontal 1.5 x Sigma 3 000 0

Sigma 1 Horizontal 2 x Sigma 3 090 0

5.7.4 Model Sequencing and Grid Locations

The constructed MAP3D model geometry is shown in Figure 5-21 and is based on 3D wireframes provided by Terramin. Eleven mining steps in total were used to determine the likely stress magnitudes along the decline, FW infrastructure and the ore drives (flatbacking or shanty mining). Figure 5-22 shows the simplified mining sequence adopted to obtain the areas of potentially elevated stress levels.

The provided mine design has each stope access drive declining/inclining to four flatback lifts. The generalised sequence for the four flatback lifts are as follows:

1. Decline down at 1 in 8 to access the bottom lift (green lift in Figure 5-22);

2. Strip the floor out from bottom lift and backfill with CRF to produce a sill CRF pillar (blue lift in Figure 5-22);

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3. Access into the 2nd Flatback lift above the bottom lift (orange lift in Figure 5-22), this leaves a final crown sill pillar below the CRF; and

4. Mine the final sill pillar in the lift below the CRF.

For model simplification, the model sequence mined 3 flatbacking levels (x 4 lifts in each level), using a top down sequence of level excavation, progressively mining the decline slightly ahead of the levels being mined.

2D grid locations for the stress analysis results were placed along the decline (vertical), across the FW infrastructure, and also on the flatbacking llifts to determine indicative stress magnitudes during the mining sequence.

Figure 5-21: MAP3D Model Geometry Construction

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Figure 5-22: MAP3D View Showing Generalised Sequencing

5.7.5 Model Results

The results of the MAP3D modelling are summarised in Table 5-16. The predicted Sigma 1 stress magnitudes are summarised as follows:

Low stress magnitudes are predicted in the Upper Level Decline Domain;

For the Middle Level Domain, the stresses in the backs for the decline and FW infrastructure are in the range of 7 to 15 MPa. This increases for the Lower Level Domain from 14 to 37MPa. Wall stresses are significantly less in magnitude;

Once the first flatback drive is mined in the Middle Level Domain, typical Sigma 1 stresses in the backs are in the range of 10 to 15 MPa. This increases to 15 to 30 MPa for the Lower Level Domain; and

The final sill pillar below the CRF has an average Sigma 1 stress magnitude range of 12 to 25 MPa for the Middle Level Domain. This increases to 20 to 46 MPa for the Lower Level Domain.

Mining One interprets this to mean that there is the potential for stress driven deterioration of the rockmass around the development openings at the base of the mine. The rockmass response is likely to be progressive, rather than instantaneous and due to the fast turnover of ore drives, may not be a significant enough require frequent rehabilitation prior to excavating the next cut and fill lift.

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Table 5-16: Summary of Maximum Sigma 1 Stress Magnitudes from MAP3D Modelling

Geotech Domain

In‐Situ Sigma 1

Stress

Magnitude

(MPa)

Infrastructure Opening FaceInduced Sigma 1 Stress

Magnitude (MPa)

Back 0 ‐ 7

Wall 0 ‐ 4.5

Back 7 ‐ 14

Wall 5 ‐ 9

Back 10 ‐ 13

Wall 7 ‐ 9

Back 10 ‐ 15

Wall 6 ‐ 8

Back 10 ‐ 15

Wall 6 ‐ 8

Back 12 ‐ 25

Wall 8 ‐ 10

Back 14 ‐ 30

Wall 9 ‐ 18

Back 15 ‐ 27

Wall 11 ‐ 18

Back 19 ‐ 37

Wall 10 ‐ 20

Back 15 ‐ 30

Wall 10 ‐ 21

Back 20 ‐ 46

Wall 10 ‐ 28

12 ‐ 23

Cross‐Cut

Stope Access

Decline/Incline

Ore Drive

First Level

Ore Drive

Final Level

Cross‐Cut

Stope Access

Decline/Incline

Ore Drive

First Level

Ore Drive

Final Level

Lower Level

Decline

Upper Level Portal/Decline0 ‐ 5

Middle Level

Decline

5 ‐ 12

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6 GEOTECHNICAL DESIGN

6.1 Surface Stability

6.1.1 Raw Water Dam and IML Stability

A preliminary surface infrastructure design was provided by Terramin for Mining One to assess the raw water dam and IML stability. In May of 2017, the surface infrastructure design was reassessed by Terramin and modified. Mining One has assessed both designs and the results of the assessments are in the following sections.

6.1.1.1 Preliminary Surface Infrastructure Design

Surface stability across the preliminary surface infrastructure design site was assessed at two locations, the first at the integrated mullock landform (IML) and the second at the cut slope for the raw water dam. The stability cross section locations are shown in Figure 6-1.

Figure 6-1: Stability cross section locations

The Raw Water Dam, located to the southwest of the site will be cut into the side of a shallow dipping slope, creating an embankment 11m high. The stability of the section was assessed using two dimensional limit equilibrium techniques. The results of the analysis and the material parameters used are defined in Appendix D, and show that a high and acceptable Factor of Safety of greater than 3 is achieved for overall slope stability. Refer to Appendix D, Figure D1.

For excavated slopes across the site that are up to a maximum 3m height, the cut profiles shown in Table 6-1 are recommended.

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The IML will essentially be keyed into the natural ground surface and form a temporary storage of waste rock. The IML is indicated to be stable when dry, but will need to be managed for operational crest stability due to the potential for shallow skin failures because of a lack of cohesion between the rock particles (i.e. minimal compaction).

Table 6-1: Provisional cut slope design (<3m height)

Soil Unit Temporary (<3months)/ Permanent

Estimated Slope Angle (Max. Ht. 3m)

Slope (deg.)

In-situ Clay Temporary 1H : 1V 45°

Permanent 1H : 0.5V 26.6°

Extremely Weathered Rock

Temporary/Permanent 1H : 2.14V 65°

Notes: 1. Following excavation of cutting, recommended slope angles may have to be flattened (or retaining structures installed) due to localised groundwater inflows or poor ground

conditions.

2. Stable slope angle may need to be revised subject to discontinuities and rockmass type / condition exposed by initial excavation.

6.1.1.2 Modified Surface Infrastructure Design (May 2017)

The modified surface infrastructure design completed in May 2017 is shown in plan view in Figure 6-2. In the modified design raw water dam and the IML are in similar positions as shown when comparing Figure 6-1 and Figure 6-3. There are differences in the batter angles for the modified designs however, and these changes have been assessed. In addition, three stages of the design and construction phase have been provided and these have been reviewed. These include:

1. Construction Phase;

2. Peak Production Phase; and

3. Closure Phase.

Table 6-2 summarises the results of the assessment for the raw water dam and the IML stability of the modified design. It shows that the peak production design for both the raw water dam and the IML stability indicate stable conditions with the Factor of Safety (FOS) in excess of approximately 2.

However, for the construction phase of the IML infrastructure, a 5m vertical wall is designed as part of the stripping of the surface prior to the mullock being introduced as fill. This 5m high vertical wall in clay/highly weathered material is close to failure (Table 6-2). It is recommended that this batter be designed at approximately 40o to 45o, as recommended as a guideline in Table 6-1.

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Table 6-2: Summary of Results for Modified Infrastructure Design

Infrastructure Phase Wall Water

ConditionsGeometry FOS

Figure in Appendix D

Raw Water Dam

Peak Production

East

Dry 18o batters,

11m high (2 batters)

3.95 D3

Saturated 2.75 D4

North Dry 18o batter,

6m high

6.03 D5

Saturated 4.89 D6

IML

Construction West Dry 5m

vertical cut/wall

1.04 D9

Peak Production

West Dry 21o batter, 6m high

1.97 D7

North Dry

21o batters,

13m high (2 batters)

1.96 D8

Figure 6-2: Modified Surface Infrastructure Design (May 2017)

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Figure 6-3: Location of Raw Water Dam and IML for Modified Design

6.1.2 Boxcut Design

The BIH boxcut is designed to provide access to the decline and is planned to be backfilled around a self-supporting structure. The boxcut will be a temporary excavation, to allow the portal to be positioned within rock material and then permanently backfilled. This will re-establish the natural ground surface level and eliminate the risk of slope instability within the boxcut.

The size and depth of the boxcut design are related to the following:

The target depth was to achieve 5m of rock material in the crown of the portal face; and

The boxcut batters were designed to be stable for the duration of excavation and for the installation of the self-supporting structure within the boxcut.

Mining One used the following design methodology for the boxcut design:

1. The top of rock surface defined the expected materials within the boxcut. Shallow softer materials are planned to be offset from the boxcut by using a 5m wide berm. Limit equilibrium stability modelling was completed to assess the overall stability of the excavated boxcut (Appendix E); and

2. Structural data collected during the drilling program was used to undertake a kinematic assessment of the boxcut walls. During this assessment it was found that bedding will control the stability of the western wall, and this wall was laid back to match the dip of bedding and reduce the likelihood of undercutting it and destabilising the slope.

A number of test pits and a drill hole (BH049) were used to define the ground conditions at the boxcut site.

The stability analysis used the slope orientations, shown in Figure 6-4.

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Figure 6-4: Slope directions used in the stability assessment

The stability assessment results are shown in Appendix E, which defines the structural inputs to the kinematic assessment, displays the kinematic assessment results and the limit equilibrium stability assessment results.

The recommended boxcut slopes from that assessment are shown in Figure 6-5.

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Figure 6-5: Boxcut slope design based on maintaining stability

The preparation for the first underground mining cut should include the use of spiling bars, for the first 6m of underground development. The spiling product should be approved by a geotechnical engineer, but a 3m long grouted product, installed in layers with a 1m overlap and a full ring shotcrete arch installed inside each spiling bar collar position is recommended. Spiling bar to be spaced at 500mm collar spacing, shoulder to shoulder.

6.1.3 Pre-Cast Concrete Arches (BEBO Structure)

The surface mining lease area for the BIH project is fairly limited and as a result Terramin propose to utilise the available space to maximise the required infrastructure to operate the mine. It is proposed by Terramin that once the boxcut is established, pre-cast concrete arches (BEBO) will be placed in the boxcut to form an artificial portal. Once in place it is planned to backfill over the arches back to the original surface level. This will then provide more surface area for infrastructure such as a laydown area.

The pre-cast arches will be required to be manufactured by an outside source, with appropriate footings to be constructed by Terramin (or appointed contractor) to the manufacturer’s requirements.

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Figure 6-6 shows a plan of the boxcut and portal. Preliminary discussions by Mining One with one of the potential suppliers of the BEBO pre-cast structures (Humes/LafargeHolcim) has indicated that the following loads could be viable above the backfilled structure:

A minimum of 0.5m of compacted backfill is required above the crown;

For areas of the structure with 0.5m fill above the crown, a maximum weight of 30 tonnes could be applied; and

It is likely that areas of the structure with greater than 0.5m backfill above the crown could support greater loads, but the supplier has not supplied any loads until profiles and detailed assessment could be made by them.

Figure 6-6: Plan Showing Boxcut and Portal

6.1.4 Laydown Area – Upper Decline Spiral Interaction

As shown in Figure 6-6, the proposed laydown area may be above the portal and upper spiral decline. The depth from the surface to the decline location is approximately 10m as shown in cross section in Figure 6-7.

A 2D stress analysis was completed using the properties and stress field used for the MAP3D modelling. The section location is shown in Figure 6-6, while Figure 6-7 shows the maximum Sigma 1 stress contours once the spiral decline is developed. It indicates that the decline position closest to the surface may have a stress magnitude of about 0.5MPa.

Applying a 60 tonne load which could represent a fully loaded truck on surface was modelled and this is shown as Figure 6-8. It indicates that only a 0.03MPa increase in stress around the portal opening is calculated. This is considered a very small increase in stress and unlikely to cause any stability issues from the loading on surface.

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Figure 6-7: Phase 2 Analysis – Sigma 1 Stress Contours

Figure 6-8: Phase 2 Analysis – Differential Sigma 1 Stress Contours Due to 60 tonne Loading

6.1.5 Site Stabilisation and Erosion Control

This part of the project was completed by Dr Rob Loch from Landloch (soil scientists), and as part of his assessment he visited the BIH site in October 2016 to familiarise himself with the site and collect the required data to complete his assessment.

A letter report has been completed by Landloch which summarises the work completed and findings of their work. This letter report is shown in full in Appendix F. Some soil testing work is still to be completed for Landloch to complete the work which is being organised by Terramin.

A summary of the main points from the draft report are detailed below. For further detail, photos and figures, refer to Appendix F.

6.1.5.1 Site Setting and Overview

The site is located within an area of significant agricultural/horticultural development. As such, the environment - although far from degraded or unstable - is not pristine, with some areas of bare soil bordering roads, farm tracks, stock water points, etc. Equally, some major surface water flow lines in the general area show points of instability and riparian vegetation is generally present but not extensive.

0.6 0.6

Sigma 1min (stage): -0.04 MPa

max (stage): 0.28 MPa

-0.05-0.03-0.02 0.00 0.02 0.03 0.05 0.07 0.09 0.10 0.12 0.14 0.15 0.17 0.19 0.21 0.22 0.24 0.26 0.27 0.29

44

42

04

00

38

03

60

-100 -80 -60 -40 -20 0 20 40 60 80 100 120 140

kfechner

Sticky Note

Thought we were cutting landloch off? Need to remove this sentence because no work pending for Rob Loch

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It is reasonable to expect that the BIH mining operation would ensure that the quality of water discharged from its site would be of equivalent or better standard than runoff from surrounding areas. This can be achieved by:

Maintaining high levels of vegetative cover on all non-work (undisturbed) areas;

Managing potential areas of sediment or pollutant source to minimise impacts on runoff water quality;

Containing or treating low-quality water to prevent its uncontrolled discharge; and

Managing flow paths to both ensure their stability and to optimise runoff water quality.

6.1.5.2 Storm Water Plan

The stormwater plan developed by Tonkin provides a useful and practical basis for management of water quality on and leaving the site, and will be used to support some following suggestions and comments.

It quite usefully separates flows of differing quality from various sources and directs them to appropriate areas for treatment. There are aspects of the plan that may need to be adapted or considered in greater detail when being installed, but the overall concept is assessed by Mining One to be sound.

6.1.5.3 Key Source Areas

The main key sources areas investigated are shown below.

Integrated Mullock Landform (IML)

The proposed IML was assessed on the basis of the following areas:

Construction – The current design requires the excavation of an area under the footprint of the IML. This is planned to reduce the overall height of the final IML and utilise a smaller overall footprint than previously designed. The IML is a temporary structure and Terramin will need a closure plan to rehabilitate the excavated foundation;

Avoidance of Mass Failure – The IML should be constructed so that drainage from south to north is not impeded or ponded. Where stockpiling creates a barrier to drainage movement (generally at the outer edge of the landform), there is potential for water to pond against that barrier during major rainfall/runoff events, and result in mass failures and relatively large-scale mud flows that can overtop drainage networks;

Minimising erosion and sediment loads – Generally, it would be advisable to avoid having runoff flows concentrate and then move along some steep pathway to exit the IML, as the potential for scour along that flow line would be quite high. A situation where a large sediment load entered the drain at a defined point would have considerable risk of a large deposited fan causing the drain to overtop and discharge low-quality runoff to the adjoining landscape, but likely to be contained however in the defined runoff zones.

Seepage – It is not clear whether seepage from the IML will be of concern. Provided reports and data from Terramin were unable to locate clear information on potential for Acid Mine Drainage (AMD). There are references to AMD associated with historical mining, which would suggest that there may be a potential issue with the ore (which is planned to be on surface for short periods of time and be stored in a covered ROM bin prior to truck loading), but the quality of the mullock may be of less concern. Apart from AMD, it seems unlikely that any seepage generated would contain serious contaminants,

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but - given the depth from which the mullock will be generated – there is potential for the seepage to be saline; and

Dust - Dust from the IML may well become a concern if works are being carried out during prolonged dry periods. The same compounds used for erosion control could also be used to control dust.

Landscape Mounds

General comments regarding the currently landscaped mound on site are as follows:

Although establishment of trees on the bunds is undoubtedly a priority, establishment of grass to stabilise slopes and to cover the bare soil should be the initial target for revegetation works;

Soil amelioration and fertilisation will be necessary for optimal results, so some sampling and analysis of materials placed in and on the bunds is recommended;

Stripping and management of topsoil should be planned as part of the construction of the landscape mounds;

Seeding methods such as hydroseeding or hydromulching may be necessary to achieve the desired outcomes of vegetation establishment and erosion control; and

Lower batter gradients may be needed for the higher bunds.

Exiting Points of Erosion

Although erosion rates on the site are visibly not high, there are some points where localised flows have caused on-going scour. These include:

The line of flow concentration draining the western catchment, and

A large rill or very small gully along the tree line in the southern creek catchment, possibly caused by a cattle track along the fence line.

Management of these areas of minor scour will be important as some of them are close to discharge points and, therefore, any erosion that develops will be more likely to impact on the quality of water discharged from the site.

Flow Pathways

The flow pathways outlined in the stormwater design provided by Tonkin appear to be appropriately sited, and include suitable controls to manage water quality.

6.1.5.4 Forward Work

As at December 2016, Items to be considered as forward work include:

Development of a topsoil management plan covering:

o Soil stripping;

o Construction of the IML and landscape mounds; and

o Planning for eventual site closure;

Development and refinement of guidelines for establishment of vegetation on the landscape mounds;

Consideration of landform designs for the IML;

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Development of erosion and dust management guidelines for the IML; and

Development of a management plan for grassed areas on the site.

Landloch was advised that a range of bushfire risk management practices will be an integral part of the forward work, and control of fire risk on mine sites is strongly endorsed.

6.2 Infrastructure Positioning

6.2.1 Vent Shaft Positioning

Two vent shaft position options have been developed by Terramin, and each was assessed by Mining One to determine the most suitable in terms of the expected ground and stress conditions. There are also a number of possible options for escape way rises. Mining One have assessed each option for depth of soft surficial materials (soil like), rock quality and stress. The position of each shaft assessed by Mining One is shown in Figure 6-7.

The pictorial results for each shaft option are shown in Appendix G. A summary of each option is presented in Table 6-3.

Figure 6-9: Shafts assessed by Mining One (looking north)

Table 6-3: Summary of shaft option assessment

Shaft Option Soft Surficial Materials

Rock Quality Stress Change (Sig.1 & 3)

Early Decline Return

21m Extremely Poor to Poor

Insignificant stress change due to mining

Option 1 112m (vent shaft)

127m (egress)

Very Poor, Poor Insignificant stress change due to mining

Option 2 69m (vent shaft)

113m (egress)

Poor to Good Insignificant stress change due to mining

Option 3 97m (egress ) Poor to Good Insignificant stress change due to mining

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The shaft assessment indicates that the depth of soft surficial materials and the interpreted rock quality will be more significant influences on stability than stress change. The soft surficial materials will be the most significant factor in selection of shaft positions, due to the likelihood of lengthy pre-sinks being required for the shafts.

6.2.2 Lower Decline and Vent Return

The lower decline and vent return were assessed for stress change and for rock quality. The results are shown in Appendix G, Figure G3 and show that:

The footwall decline is appropriately located with regard to rock quality, with conditions indicated to be Poor to Good. It should be noted though that currently there is limited Q´ logging for the lower decline; and

There is expected to be only minor stress change in the lower decline.

6.2.3 Influence of Major Structures

Major structures have a significant influence on the stability of raisebored shafts. Peck and Lee (Ref. 6) have empirical evidence that for poor quality rock to significantly impact raise wall stability, zones of poor quality rock need to be greater than 3m in vertical thickness. For zones shorter than this, over-break may be experienced but continual unravelling does not tend to occur. Faults often contribute to these poor rock quality zones, and have the potential to cause large scale shaft instability and failure.

Figure 6-10 shows the positions of faults interpreted by Terramin. There is potential for these faults to intersect all shaft options.

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Figure 6-10: Possible faulting influence on shaft option positions

6.3 Ground Support Design

The ground support design has been derived by using the following techniques:

1. Empirical support charts based on the Q classification system;

2. Structural wedge analysis based on the defect sets defined in Section 5.4; and

3. Experience from other mine sites.

The following sections detail the ground support design process completed for the project.

6.3.1 Empirical Support Design Based on Q Classification

6.3.1.1 Rock Mass Quality – Q System

The three rock geotechnical domains have been based on the blockiness and inter-block shear strength of the rockmass, termed Q’ (Section 5.6). The equation for Q’ is as follows:

Ja

Jr

Jn

RQDQ '

Q’ alone only defines the rock mass quality without any active stress component applied to the rockmass such as water pressures, mining induced stress, and weaknesses in the rockmass such as faults and shears. Two parameters which are part of the Q classification system define the active stress that the rockmass will encounter. These are:

Jw – Joint Water Pressure; and

SRF – Stress Reduction Factor (mining induced stress and/or major fault/shear influences on the rockmass).

The quotient Jw/SRF is multiplied by Q’ to determine the final Q value of the rockmass under active stress.

Section 4 has discussed the groundwater model and the regional modelling that has been applied to the local underground mining environment. It has highlighted that a key risk to the design process will be the water pressure and the success of the curtain grouting method for areas such as the orebody and were the lower hangingwall fault is intercepted. The technical aspects of achieving successful grouting of the rockmass around the drives has not been addressed in this report and is to be assessed by a grouting expert appointed by Terramin. However none of the technical specifications of this method were available to Mining One during this study.

As a result of the uncertainty regarding groundwater pressures and the grouting method and impact on the pressures and flows, Mining One has assumed the following two cases for design purposes:

Dry conditions (i.e. low inflows and/or pressures); and

Medium or Large inflows (based on Q system categories).

Mining One has assumed that dry (low pressure) conditions could be encountered in headings were no significant faults or shears or broken zones are encountered. Most of the water

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pressure and flows will likely be found in broken zones and faults and shears. For the FW infrastructure a “medium” water inflow factor has been assumed when significant structures are intercepted and a “large” water inflow factor for the ore zone which has large continuous faults as part of the FW and HW. Corresponding Jw factors of 0.66 and 0.5 respectively have been applied (Table 6-4). If there are any breaches of the grout curtains and the aquifers are extensive with high hydraulic connectivity (which is possible in this geological environment) the pressures could be extreme, and the ground support may be ineffective.

The Stress Reduction Factors (SRF) for each heading type, and each domain, has been determined from the MAP3D stress model and also from the influence of major structures such as faults and shears in the rockmass. A SRF of 2.5 has been applied to the portal area due to low cover depths and low stress magnitudes.

For the remaining decline and FW infrastructure, the stress environment is not high due to the depth of the orebody being less than 500m, and as a result the SRF applied was 1.

However for the ore drives (flatbacks), higher induced stress will be encountered and the rockmass will likely have fault/shears within it. As a result a higher SRF of 5 has been applied to the orebody and immediate FW and HW.

Table 6-4 summaries the Jw and SRF factors applied, and the corresponding effect on the Q rock mass quality value.

Table 6-4: Summary of Q values for each domain and infrastructure type

6.3.1.2 Ground support estimates based on Q system

Empirical ground support estimates based on the Q classification system have historically been determined using the Grimstaad and Barton support chart (Ref. 1). This chart has

Dry 1 2.5 4.4

Medium 0.66 1.7 2.9

Dry 1 6.3 10.9

Medium 0.66 4.2 7.2

Decline/Stockpile Dry 1 2.5 5

Cross Cuts

Stope Access Dec/Inc Medium 0.66 1.7 3.3

Dry 1 0.58 1.42

Large 0.5 0.29 0.71

Dry 1 1.66 3.34

Large 0.5 0.83 1.67

Dry 1 3.92 5

Large 0.5 1.96 2.5

Decline/Stockpile Dry 1 7.5 12.8

Cross Cuts

Stope Access Dec/Inc Medium 0.66 5.0 8.4

Dry 1 0.82 0.9

Large 0.5 0.41 0.45

Dry 1 1.48 1.7

Large 0.5 0.74 0.85

Dry 1 0.188 1.4

Large 0.5 0.094 0.7

Domain

Q'

Infrastructure SRF

Q

35th Percentile Median35th

PercentileMedian

Jw

58.3 16.7

Upper Level 6.3 10.9

Portal ‐ Approx 90m 2.5

Decline/Stockpile 1

Lower Level

7.5 12.8

Ore Drive FW

Middle Level

2.5 5.0

2.9 7.1 Ore Drive HW

Water

Inflow

1

4.1 4.5 Ore Drive HW

57.4 8.5 Ore Drive Ore

0.94 7.0

Ore Drive Ore

19.6 25.0 Ore Drive FW

1

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predominantly been developed from case studies within the civil engineering field rather than mining, which is not applicable to design of ground support in a mine.

A recent paper presented at the Ground Support 2016 conference (Ref. 3) has introduced an alternative empirical support chart which has been developed solely from mining cases studies from Australia and Canada. The support chart estimate is shown in Figure 6-11, and is applicable for headings which are 4 to 6m in span.

Rocks and soil strengths are known to vary spatially, with statistical variability typically characterised by coefficients of variation (standard deviation divided by mean) of 30-35% (Wiles, 2006, Ref. 7).

This means that failure paths will tend to encounter a range of shear strengths, with the overall shear strength being an average of what is encountered along the failure path. However, failure paths will also tend to find their way through weaker parts of the rock mass, but not the weakest (minimum strengths) because those are generally too widely dispersed to form a continuous failure path. Exceptions to the latter are likely for continuous weak zones such as faults and weak bedding structures or units.

Modelling of spatial variability and experience (Jefferies et. al. 2008, Ref. 8) suggests that lower percentiles, such as the 25th, 30th or 35th percentile values, tend to represent the overall shear strengths encountered by failure paths which find the weaker, but not weakest (minimum) strengths because the minimum strengths are too widely dispersed in the profile.

Both median and 35th percentile rock mass conditions are presented here, to show the variability, while for design purposes the 35th percentile value was used.

Based on the empirical design chart the bolt spacing for the decline and FW infrastructure is recommended at about 1.4m spacing, with 50mm fibrecrete to gradeline or shoulder. For ore drives a 1.2m bolt spacing is recommended when faults with significant water is encountered, with thicker fibrecrete to the floor. For dry conditions the fibrecrete is recommended to gradeline at 50mm thickness with a 1.4m bolt spacing (Table 6-5). Bolt spacings for substituting mesh instead of using fibrecrete are listed for all situations except for the first 90m of decline development, where fibrecrete use is preferred due to the reduced rock cover overhead.

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Table 6-5: Ground support estimates based on empirical design chart

Figure 6-11: Ground support guidelines for mine drives of 4 to 6m span (Ref. 3)

6.3.2 Wedge Analysis

A wedge analysis was completed for the following types of infrastructure:

1. Decline (both plunging at 8° and flat (stockpiles);

Surface Support Bolt Density (bolt/m2) Bolt Spacing (m)

Dry 50mm Fibrecrete Gradeline 0.5 1.4

Medium 50mm Fibrecrete Gradeline 0.5 1.4

Dry 50mm Fibrecrete Shoulder 0.45 1.5

Medium 50mm Fibrecrete Shoulder 0.45 1.5

50mm Fibrecrete Gradeline 0.5 1.4

Mesh to Gradeline 0.7 1.2





75mm Fibrecrete Floor 0.65 1.25

Mesh to Floor 0.85 1.1

50mm Fibrecrete Shoulder 0.45 1.5

Mesh to Shoulder 0.65 1.25

50mm Fibrecrete Shoulder 0.45 1.5

Mesh to Shoulder 0.65 1.25



75mm Fibrecrete Floor 0.65 1.25

Mesh to Floor 0.85 1.1

Dry

Medium

FW Infrastructure

Dry

Large

Ore Drive Ore

Dry

Medium

FW Infrastructure

Dry

Large

Ore Drive Ore

Middle Level

Lower Level

Upper Level

Decline

Portal ‐ Approx 90m

Decline/Stockpile

Domain InfrastructureWater

Inflow

Designed on 35th Percentile Q Conditions

Potvin & Hadjigeorgiou 2016, GGSO Paper

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2. FW access drives (cross-cuts, stockpiles and stope access drives); and

3. Flatback Ore drives.

The structural analysis completed in Section 5.5.1 for the decline path identified three sets (1 strong bedding set, and two joints sets). The mean orientation of the three sets, as shown in Table 5-12 was used for the wedge analysis of the decline.

For the FW access infrastructure and ore drives, the structural analysis identified five weaker joint sets (Table 15-3) and these were used for the wedge analysis.

For the decline and FW access drives, an arched profile of dimensions 5.0mW x 5.0mH was used for the structural wedge analysis as shown in Figure 6-12. A shanty profile for the flatbacking ore drives is proposed as shown in Figure 6-13. The wedge analysis was based on this profile for the ore drives.

Two types of ground support were also considered being:

1. 2.4m long, 46mm Split Set at 1.5m bolt spacing; and

2. 2.4m long, 24mm Solid Resin Bolt at 1.5m bolt spacing.

The bolt strength properties used in the wedge analyses are shown in Table 6-6. The ring spacing considered was 1.5m for both bolt types as a first pass.

No shear strength testing of the defect surfaces (direct shear testing) were completed as part of the testing program. Infill type and planarity characteristics from the defect sets defined in Table 5-12 for the decline indicate that most of the defects are planar smooth and contain no infill. Defect strength properties used in the wedge analysis for all joint sets were as follows:

Cohesion 0kPa; and

Friction Angle 25o.

Table 6-6: Bolt Properties of Types Considered for Analysis

Bolt Type Tensile Capacity

(tonnes) Plate Strength

(tonnes) Bond Strength

(t/m)

46mm Split Set 18 10 4

24mm Solid Resin Bolt

21 10 30

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Figure 6-12: 5m x 5m Arched Profile with 2.4m Long Bolts Spaced at 1.5m

Figure 6-13: Shantyback Profile for Ore Drive Flatbacking – 1.5m Bolt Spacing

6.3.2.1 Decline

The wedge analysis was completed for the decline route by varying the trend of the decline by 20o, between 0o and 360o. In addition, two plunge cases were assessed being horizontal and the other plunging at 8o.

The detailed results of the analysis are shown in Figures H1 to H6, in Appendix H. Only wedge results are shown in the Tables if the FOS is 1.5 or below. All other wedges are not discussed.

The results show that for a plunging decline at 8° there are two decline trends where the footwall has wedges with the FOS below 1.5 when Split Sets are used. These are:

Decline trend 20o, FOS = 1.1 and wedge weight = 205t; and

Decline trend 260o, FOS = 1.1 and wedge weight = 628t

No hangingwall wedges are shown to be occurring with an FOS below 1.5. Back wedges are small in weight and can be contained by surface support.

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When Resin Bolts are used at the same pattern, the footwall wedges increase to a FOS of at least 1.42.

A similar result is also shown for a non-plunging decline (flat), with Resin Bolts showing the FOS of all wedges (apart from small back wedges) to be above 1.5.

6.3.2.2 FW Access Infrastructure

Based on the mine design provided by Terramin, the typical cross-cut, stockpile and stope access directions were used to complete a wedge analysis. These are shown in Table 6-7 and Table 6-8.

For the stope access drives which have typically been designed at an approximate trend of 110o, the wedge analysis has shown that in the arched backs only small volume wedges are predicted which can be stabilised/captured by surface support. For the walls, up to 11 tonne wedges are predicted and using Split Sets at 1.5m spacing, the FOS is about 1. By using Resin Bolts the FOS for this sized wedge increases above 1.5. The maximum size wedge for the wall is 1136 tonnes, which is shown in Figure 6-14 but a wedge of this shape is unlikely to be unstable due to stress clamping across the wedge increasing the FOS substantially.

For Cross-Cuts, smaller wedges in the walls and backs are predicted and these are supportable using shotcrete surface support options. In the arched backs, a 21 tonne wedge is predicted for north trending drives, however pattern support using Split Sets increases the FOS to 1.3, while if Resin Bolts where to be used the FOS increases above 1.5. A wedge weight of 84 tonnes is predicted in the walls and using Split Sets at 1.5m spacing the FOS is unstable and below 1. By using Resin Bolts at the same spacing the FOS increases to above 1.5. A number of large wedges above 300 tonnes are also predicted for the walls which are again likely to be stable due to the stress clamping effect across the wedge surfaces. Figure 6-15 shows the 338 tonne wedge in relation to the drive and 1.5m Resin Bolts. The wedge apex is 8m into the wall.

For stockpiles driven at a trend of approximately 126°, small arched back wedges are formed between the bolts and are supportable using shotcrete surface support. Very large wedges are also predicted to be possible for the walls (>800 tonnes), and as discussed earlier stress clamping across the wedges will provide sufficient resistance to prevent instability. In addition, the three structure sets would need to be continuous to form this sized wedge.

For the right wall, two other potential smaller wedge sizes exist, each weighing 22 and 80 tonne. Using Split Set bolts at 1.5m spacing, the FOS is 0.77 and 1.07 respectively. Changing the bolt type to Resin Bolts, the FOS increase to 1.37 and greater than 1.5 respectively.

Appendix H shows the detailed wedge analysis results.

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Table 6-7: FW access infrastructure 0° plunge – 2.4m Split Sets @ 1.5m spacing

Note: Cells intentionally left blank

Table 6-8: FW Access Infrastructure 0o Plunge – 2.4m Resin Bolts @ 1.5m Spacing


FOS Weight (t) FOS Weight (t) FOS Weight (t)

stope access 110 1.12 1136 0.39 0.22 1.01 11.2

0.62 0.12 0.6 0.19 1.07 890

0.52 338

0.8 84

0.43 0.26 1.3 21 1.36 305

1.07 3606

1.11 1575 0.39 0.12 0.62 1118

1.06 80

0.77 22

InfrastructureTrend Drive

(o)

Infrastructure with 0 Degree Plunge ‐ Split Sets 1.5m Spacing

Left Wall Back Right Wall

cross cut

0

40

stockpile 126


stope access 110 1.22 1136 0.39 0.22

0.62 0.12 0.6 0.19 1.07 890

1.04 338

0.43 0.26 0.4 0.09 1.1 3606

0.39 1.8

1.2 1575 0.39 0.12 1.19 1118

1.37 22


(o)

Infrastructure with 0 Degree Plunge ‐ Resin Bolts 1.5m Spacing


cross cut

40

0

stockpile 126

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Figure 6-14: Maximum left wall wedge for stope access drives (110° trend)

Figure 6-15: Maximum right wall wedge for cross-cuts at 40o trend

6.3.2.3 Flatback Ore Drives

The typical ore drive is striking towards 020o, and this trend was used for the wedge analysis. From the structural analysis (Section 5.5.2) five weakly developed joint sets were defined. These five sets were used to complete the wedge analysis, with 10 three-side wedge combinations possible. Table 6-9 shows the results of the wedge analysis when no support is included and only shows wedges with the FOS below 1.5.

Perspective

4

Perspective

2

7

8

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The analyses indicate that all potential wedges for the left wall (FW) have an FOS above 1.5. For the backs and right wall (HW) numerous wedges are formed with an FOS below 1.5. Once ground support is included in the form of Split Sets at 1.5m bolt and ring spacing, the wedges with an FOS below 1.5 reduce significantly (Table 6-9).

Two very large wedges are geometrically possible but very unlikely to fail as the structures are unlikely to be that continuous and planar, and the likely benefits of confining stresses will maintain stability.

One of the wedges predicted which has an 80 tonne weight has an FOS below 1.0 even with the Split Set support. By replacing the Split Set support with Resin Bolt support, all of the wedges identified in Table 6-9 and Table 6-9 have a FOS greater than 1.5, except for the two extremely large and very unlikely wedge geometries.

Table 6-9: Results summary showing wedges with FOS below 1.5 – no support


FOS Weight FOS Weight FOS Weight

123 0.43 1

124

125 0.43 2.2 0.64 0.74

134 0 1.6 0.8 202

135 0.39 7.4

234 0 1.6 0.8 112

235

345

145 0 3.5 0.39 0.75

245 0 2 0.39 80

0.39, 548t

1.07, 7161t

Joint Comb

Left Back Right

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Table 6-10: Results summary showing wedges with FOS below 1.5 – split set support at 1.5m spacing


6.3.3 Proposed Ground Support Design

The structural wedge analysis has indicated that 2.4m long Resin Bolts at a bolt and ring spacing of 1.5m supports all of the reasonably size wedges formed in the backs, and walls based on the structural sets defined from the logging data. Split set bolts at the same length and density does not adequately support every wall wedge combination (Refer to Section 6.3.2).

The life of the mine is approximately 5 to 6 years, and it is likely that the mine environment could be wet due to the proposed groundwater control strategy of not dewatering. This may have a significant impact on the corrosion aspects of the chosen bolt for the decline access and infrastructure which are classed as life of mine structures. It is recommended that fully encapsulated Resin Bolts be used to protect against corrosion and offers higher bond strength which reinforces the rockmass and better controls wedge instability.

Table 6-11 summarises the recommended ground support estimates required for each domain and each heading type. For the ore drives, given that they are relatively short term openings, it may be viable to use a Stiff Split Set bolt which offers high tensile strength and relatively high bond strength.

The support estimates for both the middle and lower levels are identical and can be combined for simplification. For spans wider than 6m it is recommended that longer support in the form of Cablebolts be used.

FOS Weight FOS Weight FOS Weight

123

124

125

134 1.22 202

135

234 1.25 112

235

345

145

245 0.87 80

0.52, 548t

1.07, 7161t

Joint Comb

Left Back Right

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Table 6-11: Proposed ground support estimates – 4 to 6m wide spans

6.3.4 Cablebolting for Wide Spans

In areas where wide spans are created, fully grouted Cablebolts are required. This includes intersections, but also ore drive areas where the width will exceed 6m.

Intersections

Intersections with spans greater than 6m require secondary Cablebolt support. Cable bolt support should comprise 6.0m length twin strand bulbed Cablebolts (15.2mm) on a 2.0m x 2.0m spacing. The number of Cablebolts required should be determined by using the parabolic arch loosening concept, due to the expected variable ground conditions, and loosening over time. The required calculations to determine the number of intersection Cablebolts are:

v = (π x d3) / 24

m = v x ρ

B = (m / c) x FOS

Where:

v = volume

d = diameter of the paraboloid (or the biggest circle that can be drawn in the intersection in plan view)

Surface

Support

Type

Bolt Spacing

(m)

Surface Support

Recommendation

Dry Fibrecrete 3m long, 1.3 50mm Floor

Medium Fibrecrete 3m long, 1.0 75mm Floor

Dry Fibrecrete 1.4 50mm Gradeline

Medium Fibrecrete 1.2 50mm Floor

Fibrecrete 1.4 50mm Gradeline

Mesh 1.2 Mesh to Gradeline

Fibrecrete 1.2 50mm Floor

Mesh 1.0 Mesh to Floor













Ore Drive Ore

Ore Drive Ore

FW Infrastructure

Ground Support Estimates based on 35th

Percentile Q Rock Mass Quality Values

Dry

Medium

Dry

Large

Dry

Medium

Dry

Large

Domain InfrastructureWater

Inflow

Lower Level

Middle Level

Upper Level

Decline

Portal ‐ Approx 90m

Decline/Stockpile

FW Infrastructure

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m = mass (tonnes)

ρ = density (t/m2)

c = capacity of the Cablebolt, and taken as 50t for the ultimate strength (breakage) of a twin strand Cablebolt

FOS = commonly set at 1.5

B = number of Cablebolts required

The number of Cablebolts required for common intersection diameters (spans) are:

12m – 18 Cablebolts.



Ore Drive Wide Spans

Ore drives in some locations will require drift and fill techniques to enable extraction of the majority of the ore. As the placed fill will not reach the back position, the span formed when taking the adjacent drive will require Cablebolting. The calculations to determine the number of wide span Cablebolts are:

m = 2/9 ρw2

Normally Cablebolts are spaced 2m apart, so the required ring spacing (rs) is:

rs = (c x n) / (m x FOS)

Where:

w = drive width (m)

n = number of Cablebolt rows across the drive

The number of Cablebolts required for a 15m wide span would be:

A 15m wide span, say 6 Cablebolts per ring (spaced across the back), with the ring spacing (rs) of 1.5m. Cablebolt length to be calculated by w / 3 + 3 or 8m long in this case.

6.4 Roadheader Suitability

6.4.1 Introduction

Terramin are considering two design options for sighting the upper decline for the BIH project. These are:

1. OPT_v5, if the majority of the decline is positioned in rock; and

2. “clay” option where the upper portions of the decline pass through a significant thick overlying clay stratum before encountering the underlying rock unit.

In addition to traditional drill and blast techniques for developing the decline, Terramin are investigating the roadheader or excavator option. One of the benefits of using a roadheader for developing the decline would be that vibrations from drill and blast techniques would be eliminated, which would help Terramin with easing the local communities concerns regarding impacts from the BIH mine.

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To evaluate the roadheader and excavator suitability method for the decline, Mining One engaged McMillen Jacobs Associates (MJA) to complete this aspect of our study as they are experts in the field. MJA have completed this part of the study and have provided Mining One with a Technical Memorandum letter report as draft for inclusion into our report. The full letter report is included in Appendix I.

A summary of the findings from the MJA letter report is detailed in the sections below.

6.4.2 Weathering and Strength

Mining One provided MJA with the following data to complete their assessment:

Borehole logs and core photos for BH047, BH048, BH052 and BH53;

Soil classification file;

Summary of Lab Testing – BIH; and

UCS and Triaxial Test Results.

The general site geology is medium-strong to very-strong rock (40 to 190 MPa) overlain by a bowl of stiff, mostly low plastic (CL) clayey soils.

The overlying clays, which have typically greater than 80% clay fraction, generally increase in strength with depth and the consolidated-undrained triaxial tests report strengths starting at 193 kPa with 151 kPa of confinement and increasing to over 1555 kPa at 1500 kPa of confinement.

The rock structure and blockiness is important in the cuttability assessment for roadheaders. If the rock is too blocky and very strong then damage to the picks is an increased likelihood. A review of the borehole cores indicates that there will be sections of the rock decline where pick damage is likely.

6.4.3 Roadheader Suitability Evaluation

The rock cutting ability of roadheaders is heavily controlled by the machine weight and the installed power as well as the intact rock strength and the rock structure. Larger machines generally have more power delivered to the cutting drum and are more effective at cutting stronger materials. The upper limit rock strength that can be efficiently cut by road headers is typically around 120 MPa (for relatively intact rock) for the largest machines.

Even the largest roadheaders available from Mitsui Miike cannot efficiently cut the high strength rock (up to 190 MPa) that may be encountered along OPT_v5 decline. In addition, such a machine may not be available in Australia within a reasonable timeframe.

If the decline is relocated into the clay stratum, and the clay is suitably hard to stiff, a roadheader could be used to excavate the clay. A machine with a cutting area greater than the required decline section would be required to prevent benching. A significant risk of using roadheaders in clay is the clogging of the cutting drum teeth. Most of the clay sampled is low plasticity but high plastic clays are also present. Additionally, roadheaders are typically difficult to relocate quickly and the 2.9m width of the MRHS200 (the small machine capable of excavating a full face) would prevent most ground support equipment from accessing the face to install ground support between advances. These factors would significantly reduce excavation rate. The stability of the invert would also be a concern in the clay materials and an aggregate road base would have to be continuously advanced (and maintained) up to the cutting face.

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6.4.4 Excavator Suitability Evaluation

An alternative to using a roadheader in the “clay” decline option is to use a specialised short boom excavator fitted with different excavation tools which are common for soft ground tunnels and for short tunnels in weak or highly fractured rock. For a tunnel of this size an excavator around the 20 tonne size would seem feasible.

The short boom, although reducing the reach, helps prevent the boom articulation from hitting the crown above the excavator. These machines generally have quick hitches that allow efficient change between excavating tools such as drum cutters and rock hammers. The rate of excavation is significantly slower than a roadheader in rock formations but these machines are significantly cheaper to operate. The excavator has the flexibility to cut a full heading where the roadheader cannot and are very good in profile control.

The availability of these machines is good and a smaller excavator is generally a standard piece of kit when a roadheader is used to help with clean up and excavating selective structures like sumps and hitches.

6.4.5 Tunnel Support

For the OPT_v5 decline, a combination of bolts and shotcrete or bolts and mesh would be the likely ground support measures. Such support systems are standard for mine designs and can be prepared by most geotechnical engineers experienced in underground design.

For the “clay” decline option a change in support philosophy is required and would likely require a reinforced shotcrete shell and possibly pre-support measures such as spiling and/or face dowels. Design of such a support system will require a more detailed structural lining design.

The most significant time/cost factor will be the sequencing of the works. For both excavation methods in clay a shotcrete liner will be required. Smaller roadheaders will not be able to cut a full face in the clay and will require benching. This makes the installation of the concrete sprayed liner difficult as the load will come onto the liner and will need to be directed into the ground that is capable to taking these temporary loads.

A sequencing issue that will affect both the roadheader and excavator excavation methods is the strength gains in the concrete sprayed liner required prior to taking the next advance. It is likely the full face advance will be limited to 1 to 2 m prior to installing the shotcrete. Depending on the design thickness the shotcrete will need to be installed in a number of layers, this can be done in stages/advances. The concrete will require a certain strength prior to allowing the next cut is taken. This can be addressed in using accelerators in the shotcrete but will need further assessment.

6.4.6 Conclusions

Roadheaders will likely be unable to efficiently mine the very strong rock at the site and would experience slow cutting rates and high pick use. Additionally, roadheaders will likely encounter difficulties associated with clogging of the cutter head when excavating in clays.

If it is decided to locate the decline entirely in the rock, then drill and blast is the only practical option for excavating the decline as the rock is too strong for even the largest roadheaders.

Alternatively, if the decision is made to move the decline into the clay area then a possible approach for mining the decline would involve:

Excavate the section of decline in clay with an excavator;

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Construct an underground magazine at an appropriate depth and location along the decline; and

Complete the decline utilising drill and blast methods.

6.5 Cemented Rock Sill Pillars

Cemented rock sill pillars will be used when mining the orebody in a number of positions. Detailed design of the cemented rock fill (CRF) mix will occur much closer to the commencement of mining, due to CRF requiring to be tested with the actual waste rock, cement slurry, mixing technique and environmental conditions as planned. CRF with the same cement content may have different strengths as the rock type, particle size grading, mixing bay set up and mixing methodology, as well as curing conditions will influence the CRF strength.

Based on Mining One’s previous work on designing CRF for underhand exposure, a minimum of 5% cement by weight is recommended for drive widths of up 6m. CRF sill pillars should be planned to avoid the wider areas of the orebody where possible, as the thickness of placed CRF will need to increase with increasing drive widths. Compaction of the CRF will occur during placement when the floor is benched out due to progressively being run over by the loader (bogger).

6.5.1 Suggested Way Forward for the Proposed CRF Sill Pillars

Mining One makes the following suggestions related to CRF sill pillar planning:

Proceed to use 5% by weight CRF sill pillars up to 6m width, with the thickness at a minimum of 5m;

Where CRF sills are planned to be wider than 7m, cement contents should be greater and CRF sill thickness should increase. Allow for 8% cement by weight and a sill thickness of 10m for planning purposes;

Confirm with laboratory testing the UCS gain over time of the CRF mix. This will confirm if stability is achievable for the CRF sills. A trial mix will be required to achieve this;

Further optimisation of the CRF mix is possible, to allow increased efficiencies and reduced cement costs;

Prepare a backfill plan that details the fill plan clearly. Consider creating diagrams describing the sump mixing process. An example is shown in Figure 6-16.

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Figure 6-16: Sump mixing schematic example (from Sainsbury and Sainsbury, 2014)

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7 MINING INDUCED GEOHAZARDS

7.1 Surface Subsidence

Surface subsidence is defined as the downward movement of the natural surface level in response to changes beneath the ground surface. It is normal for the ground surface level to move in response to groundwater level changes over time or following settlement of backfilled materials following excavation without compaction. The surface subsidence assessed by Mining One relates to underground mining and the creation of voids causing the ground surface to subside.

7.1.1 Types of Mining Related Subsidence

Underground mining and the types of failure that can have an effect on the surface are summarised below:

Continuous Trough Subsidence: involves a smooth subsidence profile at the ground surface and is usually associated with the mining of thin, flat deposits (like coal) overlain by weak sedimentary rock which need to fail to fill the mined void. This type of subsidence is not applicable to BIH due to the moderate dipping nature of the orebody and the strong strength rock around the orebody. It is normal for the ground level to fluctuate naturally due to the seasonal changes in groundwater storage level.

Discontinuous Subsidence: involves large surface displacements over limited surface areas, and may happen suddenly or progressively. Types of discontinuous settlement are shown in the sketches in Figure 7-1 and include:

Crown holes or Pillar collapse arising from the collapse of the roofs of shallow and open underground workings;

Chimney failure, piping or funnelling caused by progressive failure of an unsupported open mining cavity through the overlying material to the ground surface. The surface subsidence can be of a similar plan area to the underground void; and

Plug Failure where a block of material moves downwards into the unfilled open void.

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Figure 7-1: Discontinuous Subsidence

7.1.2 Mining Method and Size of Underground Voids

The cut and fill mining method removes the ore by following an alternating process of mining, and then filling. First a level is mined, and then filled prior to mining the level above, see Figure 7-2. Since it is a requirement to always work on the filled drive, very close to 100% of the ore body is filled during mining. With this method, only small voids are left underground.

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Figure 7-2: Generalised schematic cross section of overhand cut and fill mining

The voids formed in the orebody when using this mining method will only be present at the top, or the last level of the mined orebody. Backfilling all the ore drives will mean that a void of about 1m height will be present between the top of the filled drive and the roof, as shown in Figure 7-3. When mining up to preplaced engineered sill pillars, the same small height voids will be left following filling, as shown in Figure 7-4.

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Figure 7-3: Typical void when filling with a standard loader

Figure 7-4: Voids left when mining cut and fill orebody with CRF sill pillars

7.1.3 Surface Stability Assessment – Underground Mining

The ground surface stability and subsidence related to underground mining will be managed by limiting the size of the underground openings and promptly backfilling stopes to ensure there are no long term ground surface stability effects. Due to the small voids created during mining using a cut and fill mining method the stability of the Crown Pillar (and resulting surface stability) has been assessed by an empirical method to compare the Crown Pillar at BIH to case studies

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of Crown Pillar stability from around the world. The crown pillar requirement for stability is defined as long term which means a 1000 year design life.

The Scaled Crown Pillar Span method is an empirical method of assessing stability based on a database of over 500 cases from worldwide sources collated by Carter (Ref. 16). The graphical database is shown in Figure 7-5 and data for the minimum Crown Pillar dimensions and mean rock mass quality parameters for the BIH Hangingwall are plotted (shaded in yellow). The BIH orebody falls into the Class E to F category which is described in Table 7-1; Crown Pillars with less than 5% probability of failure, suitable for medium term use. However this assumes the void will be unfilled, yet in the case of BIH, the void will essentially be filled and negate the possibility of surface effects.

The Carter database was developed to determine the acceptable geometry of Crown Pillars above open (i.e. unfilled) spans of tunnels, caverns drifts and stopes. In cases where Crown Pillars may fail and cause some surface subsidence, backfilling of the underlying opening is the standard method of remediation. In the present case, where mined stopes are almost entirely filled, the Carter method can be used to check whether the Crown Pillar might be subject to failure, however the usual hazards associated with such failure (caving, excessive settlement) cannot occur because cave propagation is restricted by the presence of the fill.

Figure 7-5: Scaled crown pillar span chart and serviceable life classes (after Carter et al, Ref. 16)

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Table 7-1: Design guidelines for crown pillar design life (after Carter et al, Ref. 16)

Other areas including shafts and decline development in close proximity to the surface will likely be required to be backfilled at mine closure. During operations, these openings will be monitored and in most cases, supported with rock bolts and support, to prevent deterioration. Within the deeper extents of the mine, standard development heights are not capable of reaching the surface level due to bulking of failed material effectively choking the void and arresting the failure progression.

Historical mine voids exist in the area, and have been known to collapse, especially around shafts and shallow underground workings. The collapses are generally due to the workings not being filled in and left to deteriorate over time. The historical mine voids are generally small, and as a result are a low risk of collapse and small in scale. The water contained within these historical mine voids must be managed to prevent inrush to the underground workings.

7.2 Earthquake and Mining Induced Seismicity

7.2.1 Naturally occurring earthquakes

BIH has a Peak Ground Acceleration of about 0.057g for a 500 year return period (Geoscience Australia website). By Australian standards, this represents a moderate seismic hazard.

It shows that a ground acceleration of 0.057g (where g is the acceleration due to gravity) has a 10% chance of being exceeded in a 50 year period. This corresponds to a 500 year average recurrence interval for this intensity of ground shaking.

For comparison, the highest earthquake hazard zones in Australia have predicted ground accelerations of 0.10g to 0.15g in southern Gippsland, Victoria, and parts of central Northern Territory and Western Australia. It is common practice to account for earthquake motions in slope designs in areas of high earthquake hazard, and/or for long-term projects where there is more likely to be a significant event during the project period.

Earthquake hazard maps present the ground acceleration for a standard 10% probability of exceedance in a 50 year period, which is used for assessment of long-term stability. The ground

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acceleration that has a 10% probability of exceedance for the Project duration of about 5 years will be one tenth of 0.057g, or 0.0057g.

Mining One considers this to be a negligible acceleration that would not normally be accounted for in stability assessments in an underground mine. The reasons for this are the limited surface exposure and the lack of evidence of underground mines having been damaged by natural earthquakes in the past. There have been a number of studies conducted by others to assess the impact of earthquakes on the stability of underground openings (Refs. 17 to 23). From a review of these references, Mining One has concluded that:

In a very large magnitude earthquake located in the region, some underground void damage could be expected;

Generally less damage is seen in deeper voids;

Voids in soft materials have greater damage due to earthquakes than in rock; and

There were no reports of catastrophic mine failures as a result of natural earthquakes.

The CRF and rock fill are expected to remain stable and in place following a natural earthquake.

7.2.2 Mining Induced Seismicity

Mining induced seismicity is usually associated with the ground around the mining voids responding to a redistribution of stress due to mining. This type of seismicity is usually caused by rock breaking due to over stressing, known as rock burst or due to the release of energy due to movement on a natural structure (i.e. faults). These events are localised to the underground opening and known as a shakedown event. These types of seismic events are a hazard to underground employees and infrastructure.

Other potential impacts of mine induced seismicity are vibrations that in some cases can be felt on the ground surface. Mines such as Mt Charlotte beneath Kalgoorlie and Stawell Gold Mine have surface monitoring and strategic mine plans to help avoid causing mining induced seismic events being felt by members of the public, but are mainly in place to monitor blast vibrations.

In the case of the BIH Mine plan, there are a number of aspects that suggest that mine induced seismicity is highly unlikely. These are:

The mine is relatively shallow, with the maximum mine depth of 450m. Mine induced seismicity usually requires greater mine depths before becoming noticeable, causing vibration generating events. Rock mass strength to mining induced stress ratio is large enough to prevent failure from occurring;

The dimensions of the mine are small, and the underground openings will be filled during mining. Stress builds up to greater levels when large orebodies are excavated. Also, vibrations can be felt in some cases when unfilled voids collapse, which cannot occur at BIH due to the relatively confined mining method;

The blast sizes will be small, with only development sized firings. Stress related redistributions at the time of blasting will therefore be small compared to those from large stope firings;

The stress model indicates that there will be stress change following mining, but at levels unlikely to be noticeable during mining. It is thought that the ground control issue facing BIH will be low stress, where the ground lacks sufficient clamping and higher capacity rock bolts may be required;

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Mining sill pillars has been checked in the stress model, and due to the relatively shallow mine depth there are levels of stress build up that are not considered likely to cause observable mine induced seismicity; and

The rockmass at BIH has been described as Poor to Good, and is relatively broken with RQD’s from drill core logging confirming this. Due to the broken nature of the ground, high levels of stress are unlikely, even if the mine was extended to greater depths.

The well-known cases of mine induced seismicity, such as Beaconsfield and Mt Charlotte examples are mines that extend 1000m below surface level and are excavated within ground conditions which are considered strong and competent. The greater mining depth and strong rock mean that higher stresses can develop around the mine openings, in turn leading to mine induced seismicity.

P:\2241_G \4819_V4 74



8 RISK ASSESSMENT CONSIDERATIONS

As part of a formal risk assessment process planned to be undertaken by Terramin at a later date for the BIH project, Terramin requested that Mining One document an initial risk assessment for the geotechnical aspects of the project. This is intended to be used as a starting point for the formal risk assessment workshop at a later date which Mining One is expected to be attending.

Terramin provided Mining One with a risk assessment profile spreadsheet and this was completed. The potential impact event descriptions itemised in the risk assessment are shown below:

1. Depth of Clay deeper in Boxcut Location;

2. Depth of Clay deeper in Portal/Decline Design;

3. Collapse of BEBO structure during backfilling;

4. Collapse of BEBO structure while using area above as laydown area;

5. Surface instability of dumps and cuttings, including erosion;

6. Subsidence above initial spiral decline close to the surface, Laydown area;

7. Surface vibration due to mine induced seismicity;

8. Impact of mine development if the clay/rock surface is not accurate;

9. Impact of ground water pressures on ground support design;

10. Corrosion of ground support;

11. Stability of vent shafts effected by ground conditions and water;

12. No-dewatering and grouting ineffective/failure causing significant water inrush and flooding;

13. Fibrecrete ineffective surface support due to water pressure and inflow;

14. Large water inflow/pressure once HW fault structure intercepted in shafts and flatbacking drives;

15. CRF sill pillar design not strong enough to mine underneath the sill pillar below;

16. Mining subsidence during and after completion of cut and fill mining;

17. Central access into flatback levels need to be offset from each other; and

The detailed risk assessment ratings are shown in Appendix J. It highlights that the water inflow/pressure and grouting effectiveness are classified as the higher risks. No information to date has been supplied to Mining One regarding the grouting effectiveness for the project, hence this has been identified as a higher risk. It is understood that Terramin is currently investigating this using a specialist grouting consultant. These risk assessment ratings may not reflect Terramins’s current internal studies.

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9 SUMMARY, CONCLUSIONS AND RECOMMENDATIONS

Mining One conducted a geotechnical assessment for the Bird-In-Hand Gold Project during 2016. The assessment was based primarily on drill hole data collected by Terramin and mining plans, also provided by Terramin. The geotechnical assessment found that the clay material thickness at the surface varies significantly across the site. The rock conditions vary between Poor to Good and bedding/foliation is dominant in a generally blocky rockmass. Major structures will have a very important influence on stability of underground openings. Systematic grouting is expected to be required to control water inflows and this will improve the rockmass in terms of rock quality with the grout filling structures and making the rockmass more competent, when compared to the ungrouted state.

The mine was separated into Geotechnical Domains for the geotechnical assessment and are known as the Clay/Highly Weathered, Upper Level Decline, Middle Level and Lower Level Domains, with part of the Middle Level Domain and the Lower Level Domain including the ore body. Ground support derived for each Domain is dependent on the expected water inflows/pressures and ground conditions, but fibrecrete and either a resin bar bolt or a grouted splitset are the recommended ground support choices for development and production mine areas. Cablebolts will be required for wide spans in excess of 6m span.

The boxcut position allows access to rock conditions at shallow depth below surface, but some blasting may be required to break the rock at the base of the boxcut. When underground mining commences, the rock conditions are expected to be too hard for a roadheader, meaning drill and blast methods will be required. The alternative of moving the decline to position it within the Clay/Highly Weathered Domain will not allow roadheader excavation methods as the material is expected to be too soft to cut effectively.

Underground infrastructure is positioned to consider the mining method, induced stress conditions, ground conditions, and proximity to surface infrastructure. Shafts to surface will be influenced by the thickness of the surficial clay materials, with all current shaft position options likely to require lengthy pre-sinks with precast concrete liner within the clay materials.

Induced stress conditions are expected to be manageable for the current mining method and depth of mining. Recommended ground support systems utilise methods that are commonly used at many underground mines in Australia and are widely accepted for their success in controlling rock fall risk.

The surface stability of excavations was assessed for two designs provided by Terramin (2016 design and a May 2017 design). For the 2016 design, all the excavations assessed were found to be stable for the purpose designed for. This assessment is also seen for the May 2017 design for the peak production and closure design. However, for the construction phase of the IML infrastructure, a 5m vertical wall is designed as part of the stripping of the surface prior to the mullock being introduced as fill. This 5m high vertical wall in clay/highly weathered material is close to failure (Table 6-2). It is recommended that this batter be designed at approximately 40o to 45o, as recommended as a guideline in Table 6-1.

The most significant excavation assessed by Mining One was the boxcut and this is intended to be a temporary excavation with a structural tunnel liner inserted into the boxcut, with the boxcut then backfilled to ground surface level.

Surface stability due to underground void collapse and mining induced vibrations were assessed and both found to be unlikely. The mining method utilising backfill in the excavated

P:\2241_G \4819_V4 76



voids has a large influence on maintaining surface stability as there are no large underground voids either during or following mining. The shallow depth of the mine, low induced stress change, small mining area, small blast sizes and blocky rockmass will mean the sudden release of energy due to mining is unlikely.

Mining One recommends the following be completed to verify the geotechnical model and design prior to commencement of mining:

Further work on the major structure model is required to gain as much confidence in the position and description of the major structures as reasonably possible. The major structures are very important to both the hydrogeology study (underground water inflows) and the stability of underground openings. Without depressurisation of the hangingwall faults and other prominent water bearing structures, Mining One consider that there would be a significant risk to the operation and to underground employees from ground failure (due to high water pressures) and flooding.

Confirmation of the grouting plan and expected water inflows/pressure into the underground openings and applying this information to check the influence of final assessment water inflows and grouting extents on the ground support design.

Dedicated investigation drill holes and geotechnical assessment is required for each shaft to ground surface.

The structural tunnel liner to be placed in the boxcut will require final engineering confirmation of the liner load bearing capacity and construction specifications for the footings and backfill types/compaction requirements.

CRF sill pillars will need proper design for each sill pillar location and width. The design work will be required to commence when there is access to the mullock on the IML stockpile, to allow for mixing of samples using mullock from the BIH mine.

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REFERENCES

1. Terramin, Oct 2013, “Bird-in-Hand Scoping Study”, Internal report.

2. Grimstad and Barton, N.R, 1993, “Updating the Q-System for NMT”.

3. Ground Support 2016 Conference Paper, “Selection of ground support for mining drives based on the Q-System”. Y. Potvin, Australian Centre for Geomechanics, The University of Western Australia, Australia. J. Hadjigeorgiou, University of Toronto, Canada.

4. MAP3D Software – Mine Modelling Pty Ltd. Version 65. Rock Mechanics and Stress Analysis for the Mining, Civil and Geomechanics Industries.

5. Letter Report by Mining One, “Angas Zinc Mine LOM MAP3D Modelling Assessment”, 17th December 2012, 1605_G\3194v2.

6. Peck and Lee, May 2008, ”Stability of Raised Bored Shafts in Australian Mines”, 13th Australian Tunnelling Conference.

7. Wiles, T.D. 2006: “Reliability of Numerical Modelling Predictions”, International Journal of Rock Mechanics and Mining Sciences, 43 (2006), 454-472.

8. Jefferies, M., Lorig, L, and Alvarez, C. 2008: “Influence of Rock Strength Spatial Variability in Slope Stability”, in Continuum and Distinct Element Numerical Modeling in Geo-Engineering (Proceedings, 1st International FLAC/DEM Symposium, Minneapolis, August 2008), Paper No. 05-01, R. Hart et al, Eds. Minneapolis, Itasca Consulting Group Inc.

9. AMC, 2002, “Cracow Project Cemented Rockfill Crown Pillar Specifications”, report number 202058.

10. D. Finn & M. Dorricott, 2002, “Cemented Rock Fill for Extraction of High-Grade Crown Pillars at Crusader Gold Mine”, Underground Operators Conference 2002, AUSIMM.

11. AMC, 2011, “CRF Mix Recipe Review for Kilkenny Ore Block at Cracow”, report number 111095.

12. Mitchell, R.J., Olsen, R.S. & Smith, J.D., 1982, “Model Studies on cemented tailings using mine backfill”, Canadian Geotechnical Journal, vol.19, pp.14-28.

13. K. Guilfoyle & T. Gready, “Stawell’s Cemented Rock Fill System”, Underground Operators Conference 2002, AUSIMM.

14. Sainsbury, D., & Sainsbury, B., “Design and implementation of cemented rockfill at the Ballarat Gold Project”, Minefill 2014 ACG.

15. Mitchell 1983, Mitchell R.J., 1983. “Earth structures engineering”, London, Allen & Unwin Chapter 6.

16. Carter T.G., Cottrell B.E., Carvalho J.L. and Steed C.M., 2008, “Logistic Regression improvements to the Scaled Span Method for dimensioning Surface Crown Pillars over civil or mining openings”, 42nd US Rock Mechanics Symposium and 2nd U.S. – Canada Rock Mechanics Symposium, San Francisco.

17. Lanzano G, Bilotta E, Russo G., “Tunnels under seismic loading: a review of damage case histories and protection methods” Meeting Project, Strategies for reduction of the seismic risk, Laboratorio di Dinamica strutturale e Geotecnica, Termoli, Italy.

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18. ASCE, 1974, “Earthquake damage evaluation and design considerations for underground structures”, American Society of Civil Engineers, LA.

19. Dowding C.H., Rozen, A., 1978, “Damage to rock tunnels from earthquake shaking”, American Society of Civil Engineers, Journal of Geotechnical Engineering Division, Vol 104, p 175-191.

20. Owen G.N., Scholl R.E., 1981, “Earthquake engineering of large underground structures”, Report no. FHWA/RD 80/195, Federal Highway Administration and National Science Foundation.

21. Sharma S.W.R., Judd, 1991, “Underground opening damage from earthquakes”, Engineering Geology, vol.30, p263-276.

22. Power M.S., Rosidi D., Kaneshiro J.Y., 1998, “Seismic vulnerability of tunnels and underground structures revisited”, Proceedings of North American Tunnelling ’98, Newport Beach, CA, Balkema Rotterdam, p 243-250.

23. Corigliano M., “Seismic response of deep tunnel in near-fault conditions”, Tesi di Dottorato in Ingegneria Geotecnica, Politenico di Torino, 2006.

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Appendix A Geotechnical Cross Sections

Leapfrog Model

Geotechnical Cross Section Locations

1

2

3

4

5

6

7

8

9

10

W11

3

BH00

8

BH04

5

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime0.01 – 0.10.1 – 1.01.0 – 4.0

4.0 – 10.010.0 – 40.040.0 – 100.0

100.0 – 200.0> 200.0

SurfacesBIH Old WorkingsFaultProposed Underground Workings

Q Prime Cross Section 6,129,440 North

Scale: 1:2,500

Vertical exaggeration: 1x

0m 200m

broache

Stamp

broache

Rectangle

W11

4

W11

5

BH00

4

BH04

4

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH00

3

BH02

4

BH02

5

BH03

1

BH04

2

BH04

3

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

W04

1

W04

2

W04

4

W04

5

BH00

5

BH02

4

BH02

5

BH02

6

BH02

7

BH02

8

BH02

8W

BH02

9

BH03

1

BH05

5308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH01

8

BH02

1BH

024

BH02

6

BH02

7

BH02

9

BH03

0BH

038

BH04

1

BH05

5

BH05

7

BH05

9

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH01

6BH03

2

BH03

4

BH03

9

BH05

0

BH05

4

BH05

6BH

058

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH02

2

BH03

3

BH03

5BH

036

BH04

6

BH05

0

BH05

1

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH046

BH047BH048

BH052

BH053

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH047BH048BH052

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0



Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH047

BH049

BH052

308500 308700 308900 309100

0

90

180

270

360

450

W E

Q Prime< 0.010.01 – 0.10.1 – 1.0

1.0 – 4.04.0 – 10.010.0 – 40.0

40.0 – 100.0100.0 – 200.0> 200.0

SurfacesBIH Old Underground WorkingsFaultPlanned Underground Workings


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

W113

BH008

BH045

308600 308800 309000 3092000

90

180

270

360

W E


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0

RQD Cross Section 6,129,440 North

Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

308600 308800 309000 3092000

90

180

270

360

W E


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH003

BH025 BH028BH028W

BH031

BH042

BH043

308600 308800 309000 3092000

90

180

270

360

W E


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

W042

W044

BH005

BH024

BH026

BH027BH029

BH031

BH055

308600 308800 309000 3092000

90

180

270

360

W E


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0> 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH021

BH026

BH027

BH029BH030

BH041BH055

BH057

BH059

308600 308800 309000 3092000

90

180

270

360

W E


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0> 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH016BH032

BH034

BH039

BH050

BH054

BH056

BH058

308600 308800 309000 3092000

90

180

270

360

W E


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0> 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH022

BH033BH035BH036

BH046

BH050

BH051

308600 308800 309000 3092000

90

180

270

360

W E


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0> 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH04

7

BH04

8

BH05

3

308600 308800 309000 309200

0

90

180

270

360

W


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0> 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH04

7

BH04

8

308600 308800 309000 309200

0

90

180

270

360

W


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0> 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

BH04

9

308600 308800 309000 309200

0

90

180

270

360

W


RQD< 10.010.0 – 20.020.0 – 40.0

40.0 – 60.060.0 – 80.0> 80.0


Scale: 1:2,500


0m 200m

broache

Stamp

broache

Rectangle

P:\2241_G\4819_V4

Appendix B Geotechnical Rock Mass Domains

Histogram Charts

Figure B1: Upper Decline Domain – Q’ Distribution

Figure B2: Middle Decline Domain – Q’ Distribution

0

20

40

60

80

100

120

0.001 ‐0.01

0.01 ‐0.1

0.1 ‐ 1 1 ‐ 4 4 ‐ 10 10 ‐ 40 40 ‐100

100 ‐400

400 ‐1000

Length of Core (m)

Upper Decline ‐ Q' Distribution

0

10

20

30

40

50

60

70

0.001 ‐0.01

0.01 ‐0.1

0.1 ‐ 1 1 ‐ 4 4 ‐ 10 10 ‐ 40 40 ‐100

100 ‐400

400 ‐1000

Length of Core (m)

Middle Decline ‐ Q' Distribution

Figure B3: Lower Decline Domain – Q’ Distribution

Figure B4: Upper Decline Domain – RQD Distribution

Figure B5: Middle Decline Domain – RQD Distribution

0

10

20

30

40

50

60

0.001 ‐0.01

0.01 ‐0.1

0.1 ‐ 1 1 ‐ 4 4 ‐ 10 10 ‐ 40 40 ‐ 100 100 ‐400

400 ‐1000

Length of Core (m)

Lower Decline ‐ Q' Distribution

0

20

40

60

80

100

120

140

160

0 ‐ 10 10 ‐ 25 25 ‐ 50 50 ‐ 75 75 ‐ 90 90 ‐ 100

Length of Core (m)

Upper Decline ‐ RQD Distribution

0

10

20

30

40

50

60

70

80

90

100

0 ‐ 10 10 ‐ 25 25 ‐ 50 50 ‐ 75 75 ‐ 90 90 ‐ 100

Length of Core (m)

Middle Decline ‐ RQD Distribution

Figure B6: Lower Decline Domain – RQD Distribution

Figure B7: Clay Domain – Weathering Distribution

Figure B1: Upper Decline Domain – Weathering Distribution

0

5

10

15

20

25

30

35

40

45

50

0 ‐ 10 10 ‐ 25 25 ‐ 50 50 ‐ 75 75 ‐ 90 90 ‐ 100

Length of Core (m)

Lower Decline ‐ RQD Distribution

0

100

200

300

400

500

600

RS CW HW MW SW FR

Length of Core (m)

Completely Weathered/Clay ‐Weathering

0

20

40

60

80

100

120

140

160

CW HW MW SW FR

Length of Core (m)

Upper Decline ‐Weathering

Figure B2: Middle Decline Domain – Weathering Distribution

Figure B3: Lower Decline Domain – Weathering Distribution

Figure B4: Clay Domain – Strength Distribution

0

50

100

150

200

250

CW HW MW SW FR

Length of Core (m)

Middle Decline ‐Weathering

0

20

40

60

80

100

120

CW HW MW SW FR

Length of Core (m)

Lower Decline ‐Weathering

0

50

100

150

200

250

300

350

400

450

500

S1 S2 S3 S4 S5 R0 R1 R2 R3 R4 R5

Length of Core (m)

Completely Weathered/Clay ‐ Logged Stength

Figure B5: Upper Decline Domain – Strength Distribution

Figure B6: Middle Decline Domain – Strength Distribution

Figure B7: Lower Decline Domain – Strength Distribution

0

50

100

150

200

250

S R0 R1 R2 R3 R4 R5 R6

Length of Core (m)

Upper Decline ‐ Logged Strength

0

20

40

60

80

100

120

140

S R0 R1 R2 R3 R4 R5 R6

Length of Core (m)

Middle Decline ‐ Logged Strength

0

20

40

60

80

100

120

S R0 R1 R2 R3 R4 R5 R6

Length of Core (m)

Lower Decline ‐ Logged Strength

Figure B8: Upper Decline Domain – Controlling Infill Type Distribution

Figure B9: Middle Decline Domain – Controlling Infill Type Distribution

Figure B10: Lower Decline Domain – Controlling Infill Type Distribution

0

20

40

60

80

100

120

140

CA CH CY GR NN QZ

Length of Core (m)

Upper Decline ‐ Controlling Infill Type Logged Interval

0

20

40

60

80

100

120

CA CB CH CY LI NN PY QZ ZE

Length of Core (m)

Middle Decline ‐ Controlling Infill Type Logged Interval

0

20

40

60

80

100

CA CH LI NN

Length of Core (m)

Lower Decline ‐ Controlling Infill Type Logged Interval

Figure B11: Upper Decline Domain – Controlling Defect Type Distribution

Figure B12: Middle Decline Domain – Controlling Defect Type Distribution

Figure B13: Lower Decline Domain – Controlling Defect Type Distribution

0

20

40

60

80

100

120

140

BG FL JN SH XX

Length of Core (m)

Upper Decline ‐ Controlling Defect Type Logged Interval

0

50

100

150

200

BG CN FL JN SH XX

Length of Core (m)

Middle Decline ‐ Controlling Defect Type Logged Interval

0

20

40

60

80

100

120

140

160

BG FL JN

Length of Core (m)

Lower Decline ‐ Controlling Defect Type Logged Interval

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Appendix C Ore Body Rock Mass Quality

Modified Tunnelling Index - Q’

Figure C1: Main Ore Body Reef Showing DH Pierce Points

Figure C2: Deeper Hangingwall Conditions - Q’

Figure C3: Immediate Hangingwall Conditions (0 to 3m) - Q’

Figure C4: Ore Body Conditions - Q’

Figure C5: Pillar Conditions - Q’

Figure C6: Immediate Footwall Conditions (0 to 3m) - Q’

Figure C7: Deeper Footwall Conditions - Q’

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Appendix D Raw Water Dam and IML Stability

Figure D1Raw Water Dam Stability Cross Section

Figure D2IML Stability Cross Section

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Appendix E Boxcut Stability Analysis and Design

Kinematics BH047

1

Kinematics BH048

2

Kinematics BH052

3

Kinematics All

4

Kinematics Analysis

5

Eastern Slope

Southern Slope

Western Slope

Kinematic Assessment Assumed:

Friction Angle of 30°

Lateral Limit of 20°

Slope Dip in Fresh Rock 80°

Kinematics Analysis

6

East Slope - Sliding

Kinematics Analysis

7

East Slope - Wedge

Kinematics Analysis

8

East Slope – Flexural Topple

Kinematics Analysis

9

West Slope - Planar

Kinematics Analysis

10

West Slope - Wedge

Kinematics Analysis

11

West Slope – Flexural Topple

Kinematics Analysis

12

Southern Slope - Sliding

Kinematics Analysis

13

Southern Slope - Wedge

Kinematics Analysis

14

Southern Slope – Flexural Topple

Kinematic Analysis

Summary Table

15

Low Low to Moderate Moderate Moderate to High High

<5% 5 – 15% 15 – 30% 30 – 50% >50%

Failure ModeCritical

StructuresEastern Slope Southern Slope Western Slope

60° 70° 80° 60° 70° 80° 60° 70° 80°

PlanarAll 0.41% 1.64% 3.28% 4.30% 4.71% 6.15% 5.53% 5.74% 5.74%

Joint Sets Nil Nil Nil 5.56% (JS1) 5.56% (JS1) 5.56% (JS1) 26.09% (Vein) 26.09% (Vein) 26.09% (Vein)

Wedge All 2.53% 5.35% 9.45% 7.82% 10.72% 14.55% 23.84% 29.15% 33.04%

Joint Sets Nil Nil Nil Nil Nil Nil Nil Nil Nil

Flexural TopplingAll 1.84% 2.66% 3.89% 1.43% 2.05% 2.87% 4.10% 4.30% 4.51%

Joint Sets Nil Nil Nil Nil Nil Nil Nil Nil Nil

Box Cut Stability Analysis

Completely weathered (CW) material parameters are based on Lab test results.

Moderately weathered material parameters were derived by reducing the UCS of the fresh rock.

Fresh rock material parameters based on residual tri-axial test results. GSI was reduced from 100 to 70 to replicate rock slopes as opposed to laboratory test samples.

16


Material Properties and Results Tables

17

Profile AssessedCW Batter

Angle

CW Material Properties

Recommended Lower Limit

CS2 – Western Slope 45° 4.31 2.32

CS2 – Eastern Slope 50° 5.20 2.75

CS1 – Southern Slope 50° 3.66 2.08


Eastern Slope

18

Recommended CW material props. Lowest Observed CW material props.


Southern Slope

19



Western Slope

20



Completely Weathered Design Chart

21

1.01.21.41.61.82.02.22.42.62.83.03.2

0 10 20 30 40 50

Factor of Safety

Slope Height (m)

Completely Weathered (Recommended)Design Chart

45deg

50 deg

60 deg

0.6

0.8

1.0

1.2

1.4

1.6

1.8

0 10 20 30 40 50

Factor of Safety

Slope Height (m)

Completely Weathered (Lower Limit)Design Chart

45deg

50 deg

60 deg

P:\2241_G\4819_V4

Appendix F Site Stabilisation and Erosion Letter Report

By Landloch, December 2016

1

25 July 2017 Attention: Ben Roache Geotechnical Manager Mining One Pty Ltd Level 9, 50 Market Street MELBOURNE VIC 3000

1255.16b

Site stabilisation and erosion control, Bird In Hand (BIH) Gold Project

Dear Ben,

I have inspected the BIH site and reviewed associated plans and reports. I note that there has been some slight revision of plans for the surface layout of the site since my initial report, but that the general patterns of stormwater movement and management on the site have not been changed significantly. I note that a contaminated land investigation by Golder Associates (2017) has carried out an intensive program of soil sampling and analysis. However, although comprehensively dealing with potential contamination, it has significant limitations as a source of data for soil management, in that:

Sample depth increments are not stated; and Very few analyses relevant to soil management were made.

Nonetheless, it may still be possible to gather soil management information from that work, provided:

i. The drill logs can be assessed to derive reasonable estimates of topsoil depth; and ii. A range of appropriate samples could be obtained – either from retained drill cores or

from the laboratory to which samples were sent – and analyses appropriate to soil management then carried out. Involvement of an experienced soil scientist would be advisable.

The soil management information will be needed to guide the initial establishment of the site, and although probably not critical at this stage, is something that will need to be addressed in the not too distant future.

2

1. Site setting and overview

The site is located within an area of significant agricultural/horticultural development, as shown in Figure 1 (taken from Tonkin (2016a). As such, the broad environment surrounding the site - although far from degraded or unstable - is not pristine, with some areas of bare soil bordering roads, farm tracks, stock water points, etc. Equally, some major flow lines in the general area show points of instability and riparian vegetation is generally present but not always extensive.

Consequently, it can be expected that runoff flows in the area of the project site will carry small to moderate loads of sediment, with possibly some low concentrations of nutrients from animal wastes, and traces of herbicides/pesticides from horticultural operations.

It is reasonable to expect that the BIH mining operation would aim to ensure that the quality of water discharged from its site would be of equivalent or better standard than runoff from surrounding areas. This can be achieved by:

Maintaining high levels of vegetative cover on all non-work (undisturbed) areas; Managing potential areas of sediment or pollutant sources to minimise impacts on

runoff water quality; Containing or treating low-quality water to prevent its uncontrolled discharge; and Managing flow paths to both ensure their stability and to optimise runoff water

quality.

2. Stormwater plan

The stormwater plan developed by Tonkin (catchments and planned drainage pathways and runoff treatment – Figure 2) provides a useful and practical basis for management of water quality on and leaving the site, and will be used to support some following suggestions and comments.

It quite usefully separates flows from sources of differing quality and directs them to appropriate areas for treatment. There are aspects of the plan that may need to be adapted or considered in greater detail when being installed, but the overall concept is sound.

A more recent plan by Tonkin for site layout is shown in Figure 3. Key differences are movement of a Hillside Screening Bund closer to the site’s western boundary (to be downslope of the Raw Water Dam, rather than to be upslope of the dam), and some additional earthworks to the south south-east of the raw water dam. Overall, however, these changes are not greatly significant in terms of the site’s functionality

3

Figure 1: Site catchments (from Tonkin (2016a)

4

Figure 2: Conceptual stormwater plan (Tonkin 2016a)

5

3. Key source areas for runoff and sediment

3.1 Integrated Mullock Landform (IML)

This is anticipated to be the largest area of bare material generating sediment in runoff on the site.

3.1.1 Construction

The current design requires the excavation of an area under the footprint of the IML. This is planned to reduce the overall height of the final IML and utilise a smaller overall footprint than previously designed.

Landloch is advised that topsoil and subsoil from the excavated material will be used during rehabilitation of the area once mining is completed and the post-closure land use has been agreed upon.

Eventual closure/rehabilitation of this area will require careful planning and use of topsoil stockpiled for this purpose, and parts of the “amenity bund/landscape mound” could be used as a source of capping material if required. That approach would be appropriate at closure, as a visual barrier would no longer be required.

It is assumed that – to maximise storage area, the IML landform will have the cross section shown below. That landform does create some risks with respect to erosion, with (potentially) a significant volume of runoff developing on the low-gradient top area, and then triggering scour as it discharges onto the steeper outer batter. (A sharp increase in gradient will trigger scour at the point of transition.)

3.1.2 Avoidance of mass failures on IML

The stormwater plan indicates that runoff on the IML will move across the landform in a northerly direction and discharge to the waste runoff drain that will be constructed along the northern side of the IML.

6

The IML should be constructed so that drainage from south to north is not impeded or ponded. Where dumping creates a barrier to drainage movement (generally at the outer edge of the landform), there is potential for water to pond against that barrier during major rainfall/runoff events, and result in mass failures and relatively large-scale mud flows that can overtop drainage networks. The following Photo 1 shows the impact of such a failure in a much larger stockpile on a Queensland coal mine.

Photo 1: Failure point (arrowed)

3.1.3 Minimising erosion and sediment loads

Generally, the dumping process will provide strong compaction of the relatively flat top of the dumped landform. However, the outer batter slopes – especially if left at angle of repose – will be uncompacted and strongly susceptible to erosion. Some thought will need to be given to ways of managing that erosion risk, as high rates of sediment movement off the IML could block the constructed drain designed to convey runoff from the IML to the mine water dam.

One option would be to grade the batters down to – for example – 1:2 V:H, and to dozer track the batters to compact them. If, following that treatment, the erosion rate was found to still be such that the constructed drain did not have capacity to carry the sediment-laden flows, then a soil stabilising compound could be sprayed onto the batters to provide a high level of erosion resistance. (Use of vegetation to stabilise an active waste dump structure is not a practicable option.)

There is a range of products on the market, with, e.g., either a relatively viscous product (such as Stonewall from Vital Chemicals) that would also reduce infiltration and drainage, or

7

a less viscous product (such as Envirobinder from Global Road Technology) that would possibly be more effective in preventing scour. (It is recommended that Terramin investigate options and make decisions on products. Landloch does not specifically recommend the products mentioned.)

Generally, it would be advisable to avoid having runoff flows concentrate and then move along some steep pathway to exit the IML, as the potential for scour along that flow line would be quite high. A situation where a large sediment load entered the drain at a defined point would have considerable risk of a large deposited fan causing the drain to overtop at that point and discharge low-quality runoff to the adjoining landscape.

3.1.4 IML seepage

AMD assessment (Tonkin 2016 b) indicates that a quite small quantity of PAF material will be excavated, and the availability of large quantities of limestone waste means that neutralisation of that PAF material should not be difficult if co-disposed and/or encapsulated with limestone. Seepage from the IML may be mildly saline (due to neutralisation of PAF material), but – given that the base of the IML will be initially excavated below ground level – it is less likely that seepage from the IML will discharge onto the soil surface and enter the drain bordering the IML which reports to the mine water dam.

Nonetheless, if saline seepage mixes with runoff from the IML, then it may increase sediment deposition in the mine water dam, and there is potential that that structure may need to be cleaned of sediment periodically if deposition rates are high. Regular monitoring of water quality in the mine water dam will be required.

Both compaction and surface sprays of a relatively viscous soil stabiliser/dust control product applied to the flat top of the IML could be used to minimise seepage if seepage quality is found to be undesirable.

3.1.5 Dust

Dust from the IML may well become a concern if works are being carried out during prolonged dry periods.

The same compounds used for erosion control could also be used to control dust.

Operationally, it would be desirable for segments of the IML to be constructed and stabilised progressively, so that the working area with potential for dust generation is kept as small as possible.

8

3.2 Landscape mounds

The landscape mounds or amenity bunds being constructed to manage views of the site from neighbouring properties will need to be managed properly to avoid them becoming visually unattractive sources of sediment.

Although I am fully aware that the recently-constructed bund shown below (Photo 2) has a number of reasons for its lack of vegetation and visible erosion, it nonetheless highlights that successful and appropriate vegetation of those bunds is quite important, and that poorly vegetated and eroding structures will be visually unattractive.

Specifically:

Stripping and management of topsoil should be planned as part of the construction of the landscape mounds; and

Although establishment of trees on the bunds is undoubtedly a priority, establishment of at least 50% surface cover of grass to stabilise slopes and to cover the bare soil should be the initial target for revegetation works;

Hydromulching or hydroseeding may be needed to ensure adequate grass establishment together with surface soil stability to erosion during establishment;

Irrespective of grass species used, rapid establishment of a dense shrub/tree cover will ensure that – over most of the life of the landscape mound - grass biomass on the landscape mounds is prevented from becoming particularly high

Lower batter gradients may be needed for the higher bunds.

9

Photo 3: Recently constructed landscape mound prior to spray seeding (October 2016)

3.3 Existing points of erosion

Although erosion rates on the site are visibly not high, there are some points where localised flows have caused on-going scour. These include:

The line of flow concentration draining the western catchment, and A large rill or very small gully along the tree line in the southern creek catchment

(Photo 4), possibly caused by a cattle track along the fence line.

10

Photo 4: Large rill along tree line in southern creek catchment

In general, as site works are carried out, the areas that will remain grass covered should be inspected, and any points of scour should be cross ripped and smoothed to prevent these scour points from becoming larger and more active.

Consistent with this, it would be desirable for the grassed areas in the various catchments to be managed to maximise surface contact vegetation cover. (Pasture management to maximise cover whilst also controlling levels of biomass is recommended.) Control of recognised agricultural weeds would also be desirable. Likely actions include:

Reseeding any bare areas, and/or patch-burning to reinvigorate growth; Cutting and baling the grassed areas, or intensive cell grazing, at least once each

winter season to encourage greater basal area cover; and Use of herbicides to control weeds (if possible, considering the need to establish

trees on landscape mounds and the possibility that some herbicides may injure trees).

11

Management of these areas of minor scour will be important as some of them are close to discharge points and, therefore, any erosion that develops will be more likely to impact on the quality of water discharged from the site.

3.4 Flow pathways

The flow pathways outlined in the stormwater design provided by Tonkin (2016) appear to be appropriately sited, and include suitable controls to manage water quality.

The flow pathway most likely to require care in its construction is the stormwater swale planned to be installed on the southern side of the existing creek. It will be important that provision be made for the swale to overtop during large runoff events in a stable manner (or at designed, stabilised, locations), as instability adjacent to the creek will not only immediately deliver sediment to flows moving off-site, but could also develop rapidly into significant points of scour.

4. Forward work

Items to be considered as forward work include:

Development of a topsoil management plan covering: o Soil stripping; o Construction of the IML and landscape mounds; and o Planning for eventual site closure;

Development and refinement of guidelines for establishment of vegetation on the landscape mounds;

Consideration of landform designs for the IML; Development of erosion and dust management guidelines for the IML; and Development of a management plan for grassed areas on the site.

Landloch is advised that a range of bushfire risk management practices will be an integral part of the forward work, and control of fire risk on minesites is strongly endorsed.

Yours faithfully

Dr Rob Loch

PRINCIPAL CONSULTANT

12

REFERENCES

Golder Associates (2017). 192 Pfeiffer Road, Woodside, South Australia. Baseline Contamination Assessment. Report No. 1659870-001-R-Rev 1.

Terramin Australia (2016). Surface layout V8 Summary. BIH Gold Project.

Tonkin Consulting (2016 a). Bird in Hand Gold Mine Stormwater Management Plan, Terramin Australia Limited. Ref No. 20155706R002C.

Tonkin Consulting (2016 b). Acid and Metalliferous Drainage Assessment, Bird in Hand Mine, Woodside, SA. Terramin Exploration Pty Ltd. Ref No. 20161077R002.

P:\2241_G\4819_V4

Appendix G Shaft Location Assessment

Stress Modelling

Shaft Option 1 Q Prime Shaft Option 1 RQD

112m

Shaft Option 2 Q Prime Shaft Option 2 RQD

69m

Soil strength materials




Figure G1: Q Prime for Shaft Options 1, 2 and 3

Shaft Option 1 –Sigma 3 change due to mining




Figure G2: Sig.1 and Sig.3 for Shaft Options 1, 2 and 3

Figure G3: Sig.1 and Sig.3 for Lower Decline and Vent Return

P:\2241_G\4819_V4

Appendix H Wedge Analysis of Mine Design

Table H1: Decline 8o Plunge – No Support

Table H2: Decline 8o Plunge – 2.4m Split Sets @ 1.5m Spacing


0 0.45 0.55 0 0.3 1.4 10

20 0.45 205 0 2.7

40 0 0.5 1 159

60 0 0.1 1 147

80 0 0.05 1 50

100 1.4 0.5 0 0.02 1 3.6

120 1.4 17 0 0.01

140 1.4 45

160 1.4 69 1 0.04

180 1.4 143 1 1

200 1 4.3 0.45 26

220 1 62 0.45 0.4

240 1 230

260 1 628

280 1 29 0 0.08 1.4 0.04

300 0 0.014 1.4 147

320 0 0.2 1.4 334

340 0 0.2 1.4 179

Trend Drive

(o)


Decline with 8 Degree Plunge ‐ No Support


0

20 1.1 205

40

60

80 1.4 0.08

100

120 1 0.13

140

160 1 0.04

180

200

220

240

260 1.1 628 0 0.04

280

300

320

340

Trend

Drive

(o)

Decline with 8 Degree Plunge ‐ 2.4m Split Sets


Table H3: Decline 8o Plunge – 2.4m Resin Bolts @ 1.5m Spacing

Table H4: Decline 0o Plunge (stockpile) – No Support


0

20

40

60

80 1.4 0.08

100

120 1 0.13

140

160 1 0.04

180

200

220

240

260 1.42 628 0 0.04

280

300

320

340

Trend

Drive

(o)

Decline with 8 Degree Plunge ‐ 2.4m Resin Bolts



0 0 0.8 1.4 74

20 0.45 45 0 0.5 1 0.5

40 0.45 0.2 1 80

60 0 0.08 1 314

80 0 0.05 1 139

100 1.4 0.5 0 0.04 1 7

120 1.4 39 1 0.06

140 1.4 101 0 0.08

160 1.4 170 0 0.3

180 1.4 73 0 0.8

200 1 0.5 0 0.5 0.45 46

220 1 80 0.45 0.2

240 1 314 0 0.08

260 1 138 0 0.05

280 1 7 0 0.04 1.4 0.5

300 1 0.07 1.4 40

320 0 0.08 1.4 102

340 0 0.3 1.4 171

Right WallTrend Drive

(o)

Decline with 0 Degree Plunge ‐ No Support

Left Wall Back

Table H5: Decline 0o Plunge – 2.4m Split Sets @ 1.5m Spacing

Table H6: Decline 0o Plunge – 2.4m Resin Bolts @ 1.5m Spacing


0

20

40 0.45 0.2

60 0 0.08 1.22 314

80 0 0.05

100

120 1 0.06

140

160

180

200

220 0.45 0.2

240 1.15 314 0 0.08

260 0 0.05

280

300 1 0.07

320

340

Trend

Drive

(o)




0

20

40 0.45 0.2

60 0 0.08

80 0 0.05

100

120 1 0.06

140

160

180

200

220 0.45 0.2

240 0 0.08

260 0 0.05

280

300 1 0.07

320

340

Trend

Drive

(o)



Table H7: FW Access Infrastructure 0o Plunge – No Support

Table H8: FW Access Infrastructure 0o Plunge – 2.4m Split Sets @ 1.5m Spacing

Table H9: FW Access Infrastructure 0o Plunge – 2.4m Resin Bolts @ 1.5m Spacing


0.6 204 0 0.04 0.6 36

1.07 1136

0 0.24 0 1.75 1.07 890

0.39 338

0.4 3.2 0.39 21 0.8 305

1.07 3606

1.07 1575 0 5.4 0.43 1118

0.9 77 0.4 871stockpile 126

stope access 110

cross cut

40

0


(o)

Infrastructure with 0 Degree Plunge ‐ No Support



stope access 110 1.12 1136 0.39 0.22 1.01 11.2

0.62 0.12 0.6 0.19 1.07 890

0.52 338

0.43 0.26 1.3 21 1.36 305

1.07 3606

1.11 1575 0.39 0.12 0.62 1118

0.69 871

cross cut

40

0

stockpile 126


(o)

Infrastructure with 0 Degree Plunge ‐ Split Sets 1.5m Spacing



stope access 110 1.22 1136 0.39 0.22

0.62 0.12 0.6 0.19 1.07 890

1.04 338

0.43 0.26 0.4 0.09 1.1 3606

0.39 1.8

1.2 1575 0.39 0.12 1.19 1118

1.47 871

cross cut

40

0

stockpile 126


(o)

Infrastructure with 0 Degree Plunge ‐ Resin Bolts 1.5m Spacing


P:\2241_G\4819_V4

Appendix I Roadheader Assessment Option

By McMillen Jacobs Associates, December 2016

Rev. No. B/Dec 2016 1 McMillen Jacobs Associates

Technical Memorandum

Technical Memorandum

To: Ben Roache Project: Bird-in-Hand Mine

From: Peter Campi cc: Peter Kottke Mike McRae

Date: 22/12/2016 Job No.: 5590.0

Subject: Roadheader assessment for mine decline in clay

Revision Log

Revision No. Date Revision Description A 16/12/2016 Initial Draft

B 22/12/2016 Final – issued to client

1.0 Introduction

Mining One Consultants (M1) has engaged McMillen Jacobs Associates (MJA) to provide an initial evaluation of the proposed Bird-in-Hand Mine decline. The intent of this memorandum is to evaluate the feasibility of excavating the main decline of Terramin’s Bird-in-Hand Mine (Adelaide Hills, South Australia) with methods other than drill & blast. Two common alternatives to drill & blast are roadheaders and excavators with attachments. This memorandum outlines the limitations of these methods and includes a summary of the shotcrete lining requirements. The planned excavated dimensions of the decline are 5.5 m wide and 5.5 m high. A number of options have been developed siting the mine decline. MJA has been informed that there a two main options:

OPT_v5, majority of the decline positioned in rock. “clay” option where the portal is located away from sensitive sites and the upper reaches of the

decline pass through a significant thick overlying clay stratum before encountering the underlying rock unit.

1.1 Information Provided

M1 provided a number of documents and a leapfrog model. The documents provided via Dropbox included:

Borehole logs and core photos for BH47, BH48, BH52 and BH53. SOIL classification file. Summary of Lab Testing – BIH. UCS and Triaxial Test Results.

1.2 Geology

The general site geology consists of weathered and fresh rock of medium-strong to very-strong rock (40 to 190 MPa) overlain by a bowl of stiff, mostly low plastic (CL) clayey soils (see Figure 1).

Bird-in-Hand Mine Roadheader assessment for mine decline in clay


Figure 1: General configuration of main decline sited in rock with overlying clays shown in brown.

The overlying clays, which have typically greater than 80% clay fraction, generally increase in strength with depth and the consolidated-undrained triaxial tests report strengths starting at 193 kPa with 151 kPa of confinement and increasing to over 1555 kPa at 1500 kPa of confinement. Each of the five depths sampled was tested at three different confining pressures. The results of the triaxial testing indicate a stiff to hard clay with cohesion increasing linearly with depth (see Figure 2). It is unsure if the tested confinement of the clay is an accurate representation of the expected confinement within the clay stratum. Further work will need to be undertaken to validate the clay strength and structure and calculate the expected overburden loads.

Figure 2: Clay shear vs confinement chart

The rock is medium-strong to very strong with unconfined compressive strengths ranging from 40 to 190 MPa (see Figure 3). Generally, the rock strength has been underestimated in the field logs in comparison



to the laboratory UCS strengths. The rock structure and blockiness is important in the cuttability assessment for roadheaders. If the rock is too blocky and very strong then damage to the picks is an increased likelihood. A review of the borehole cores indicates that there will be sections of the rock decline where pick damage is likely.

Figure 3: Rock strength vs depth chart

2.0 Roadheader Evaluation

The rock cutting ability of roadheaders is heavily controlled by the machine weight and the installed power as well as the intact rock strength and the rock structure. Larger machines generally have more power delivered to the cutting drum and are more effective at cutting stronger materials. The upper limit rock strength that can be efficiently cut by road headers is typically around 130 MPa (for relatively intact rock) for the largest machines.

Model

Weight (tons)

Cutting Power

(kW)

Machine

Width (m)

Machine

Height (m

)

Cutting Width

(m)

Cutting Height

(m)

Rock Strength

Applicab

ility

(1)

Benching

Required

Can

Excavate

>= 150 M

Pa

SLB‐350S 120 350 3.4 4.8 8.8 8.8 Soft to Medium‐Hard Rock No No SLB‐300S 90 300 3.4 4.8 8.3 8.8 Up to 100 MPa No No MRH‐S300 95 300 4 4.7 7.5 6.5 Up to 130 MPa No No MRH‐S200 49 200 2.9 3.85 6.4 6 Up to 80 MPa No No MRH‐S100 27 100 2.8 1.8 5.1 4.5 Up to 80 MPa Yes No MRH‐S65 20 65 2.8 1.5 4.2 3.8 Up to 60 MPa Yes No (1) http://www.mitsuimiike.co.jp/english/product/excavator/rh/

As summarised in the above table, even the largest roadheaders available from Mitsui Miike, cannot efficiently cut the high strength rock (up to 190 MPa) that may be encountered along OPT_v5 decline. In addition, such a machine may not be available in Australia within a reasonable timeframe.



If the decline is relocated into the clay stratum, and the clay is suitably hard to stiff, a roadheader could be used to excavate the clay. A machine with a cutting area greater than the required decline section would be required to prevent benching. A significant risk of using roadheaders in clay is the clogging of the cutting drum teeth. Most of the clay sampled is low plasticity but high plastic clays are also present. Additionally, roadheaders are typically difficult to relocate quickly and the 2.9 m width of the MRH-S200 (the small machine capable of excavating a full face) would prevent most ground support equipment from accessing the face to install ground support between advances. These factors would significantly reduce the excavation rate. The stability of the invert would also be a concern in the clay materials and an aggregate road base would have to be continuously advanced (and maintained) up to the cutting face.

3.0 Excavator Evaluation

An alternative to using a roadheader in the “clay” decline option is to use a specialised short boom excavators fitted with different excavation tools which are common for soft ground tunnels and for short tunnels in weak or highly fractured rock. For a tunnel of this size an excavator around the 20 tonne size would seem feasible. The short boom, although reducing the reach, helps prevent the boom articulation from hitting the crown above the excavator. These machines generally have quick hitches that allow efficient change between excavating tools such as drum cutters and rock hammers. The rate of excavation is significantly slower than a roadheader in rock formations but these machines are significantly cheaper to operate. The excavator has the flexibility to cut a full heading where the roadheader cannot and are very good in profile control. The availability of these machines is good and a smaller excavator is generally a standard piece of kit when a roadheader is used, to help with clean up and excavating selective structures like sumps and hitches.



4.0 Tunnel Support

For the OPT_v5 decline, a combination of bolts and shotcrete or bolts and mesh would be the likely ground support measures. Such support systems are standard for mine designs and can be prepared by most geotechnical engineers experienced in underground design. For the “clay” decline option a change in support philosophy is required and would likely require a reinforced shotcrete shell and possibly pre-support measures such as spiling and/or face dowels. Design of such a support system will require a more detailed structural lining design. The sequencing of the excavation and support is very important. Initial support may require some form of canopy support in spiles and face dowels to increase face stability. The liner will most likely include a sprayed concrete (shotcrete) generally fibre reinforced or may require increased reinforcement in the form of mesh embedded in the shotcrete. If straight walls are required to prevent excess excavation volume, then a shear moment connection and wall reinforcement will be required between the base of the wall and the invert slab. Notwithstanding, until further analysis is needed to be undertaken on the expected loads in the liner which may require a closed invert. It is assumed the tunnels are drained, however, sufficient water management processes will need to be incorporated in the form of strip drains behind the shotcrete lining and weep holes to prevent any water loads on the lining. If an undrained (tanked) tunnel is proposed then a complete rethink on the decline shape, shotcrete lining and waterproofing is required. The shape below the groundwater table would be circular/oval with a closed reinforced invert and depending on the watertightness criteria a membrane maybe required. The most significant time/cost factor will be the sequencing of the works. For both excavation methods in clay a shotcrete liner will be required. Smaller roadheaders will not be able to cut a full face in the clay and will require benching. This makes the installation of the concrete sprayed liner difficult as the load comes onto the liner it will need to be directed into ground that is capable to taking these temporary loads. A sequencing issue that will effect both the roadheader and excavator excavation methods is the strength gains in the concrete sprayed liner required prior to taking the next advance. It is likely the full face advance will be limited to 1 to 2 m prior to installing the shotcrete. Depending on the design thickness the shotcrete will need to be installed in a number of layers, and this can be done in stages/advances. The concrete will require a certain strength prior to allowing the next cut taken. This can be addressed in using accelerators in the shotcrete but will need further assessment.

5.0 Conclusion

Roadheaders will likely be unable to efficiently mine the very strong rock at the site and would experience slow cutting rates and high pick use. Additionally, roadheaders will likely encounter difficulties associated with clogging of the cutter head when excavating in clays.

If it is decided to locate the decline entirely in the rock, then drill and blast is the only practical option for excavating the decline as the rock is too strong for even the largest roadheaders.

Alternatively, if the decision is made to move the decline into the clay area then a possible approach for mining the decline would involve:

● Excavate the section of decline in clay with an excavator. ● Construct an underground magazine at an appropriate depth and location along the decline. ● Complete the decline utilising drill and blast methods.

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Appendix J Risk Assessment Appraisal

Description of Potential

Impact EventMine Life Phase

Potential

Impact Event

ID

Source Pathway Receptor

Outcome

required?

i.e. is receptor

reasonably

expected to be

adversely

impacted by the

source?

Evidence for linkage or lack of linkageDescription of Grouped

Environmental ImpactImpact ID

Factors that limit / mitigate impact

(control measures)

Significance of

expected impact

1 = Negligible

2 = Low, 3 =

Med

4 = High

Understanding of

source, pathway,

receptor (given

limitations in

information used in

assessment)

Validation of

modelling

Effectiveness of design

measures

Effectiveness of

management measures

Sensitivity to change of any

assumptionsActions to reduce uncertainty

Consequence of

changeLikelihood

Risk Ranking

(low, medium,

high, extreme)

Impact covered by

alternative Act or

Limit (Yes/No)

Act / Limit

ReferenceProposed Outcome Outcome measurement criteria

Leading indicator

criteria (where

required)

Depth of Clay deeper in

Boxcut LocationSurface Works Geotech_1 Weathering Depth

Lower Rockmass

Strength

Mine Workers

Mine designYes

Boxcut designed predominantly for rock

conditions. Will need to change design if clay is

deeper.

Injury/fatality due to collapse of

boxcut during construction, and

operation.

Test pit and geotech drilling to

identify the sub‐surface

conditions. Development of a 3D

model

1

Medium: high

Test Pit and drilling

completed

Low: Data to validate

model will only be

available after

construction

commences

Medium:high

Clay/Rock boundary

supported by Test pits and

drilling

High: Drilling and test pit to

define weathering is

common

Medium: Change to clay depth will

require change to design

Could do further drilling or test pits

in vicinity of proposed Boxcut.Moderate Possible Low:Medium No N/A

Depth of Clay deeper in

Portal/Decline Design

U/G

Development

Works

Geotech_2 Weathering DepthLower Rockmass

Strength

Mine Workers

Mine designYes

Clay strength much weaker than rock, hence

design needs to alter.


decline during construction, and

operation.

Test pit and geotech drilling to

identify the sub‐surface

conditions. Development of a 3D

model

1

Medium:

Geotechnical drilling

completed


model will only be

available after

construction

commences

Medium:

Clay/Rock boundary

supported by drilling results

High: Drilling to define

weathering boundary is

common

Medium: Significant change to

decline design if its in clay rather

than rock.

Drill In‐fill holes to better definfe

rock/clay boundaryModerate/Major Possible Low:Medium No N/A

Collapse of BEBO

structure during

backfilling

Surface Works

during U/G

development

Geotech_3

Manufacturers

procedures and

weight limits not

followed

Machinary

Weights

exceeded.

Procedure not

followed

Mine Workers

Mine designYes

Backfilling required to manufacturers

specifications based on an engineered

structure.


decline during construction, and

operation.

Follow manufacturers quidelines

regarding backfilling consolidation

methods and maximum machinary

loads allowed.

1

High: Engineered

structure with

guidelines from

Supplier

NAHigh: Engineered structure

with guidelines from Supplier

Medium:

Not commonly used in

mining industry but is an

engineered product

Low: Engineered structureFollow QA/QC recommendations

from manufacturerModerate Unlikely Low No N/A

Collapse of BEBO

structure while using

area above as laydown

area

Surface Works

during U/G

development

Geotech_4

Surface Weight

limits exceeded or

BEBO strength not

adequate

Surface Weight

limits exceeded.

BEBO strength

lower due to

installation

Mine Workers

Mine designYes

Engineered structures have strength limits and

these must not be exceeded


BEBO during construction,

operation and closure.

Follow manufacturers quidelines

regarding backfilling maximum

machinary loads allowed on the

surface above BEBO

1

High: Engineered

structure with

guidelines from

Supplier

NAHigh: Engineered structure

with guidelines from Supplier

Medium:

Not commonly used in

mining industry but is an

engineered product

Low: Engineered structure, with

loadings testes by supplier

Follow QA/QC recommendations

from manufacturerModerate Unlikely Low No N/A

Subsidence above initial

spiral decline close to the

surface, Laydown area?

Surface Works

during U/G

development

Geotech_5Weathering depth

and surface loads

Lower Rockmass

Strength, Surface

Loads

Mine Workers

Mine designYes

Upper decline first spiral section close to

surface. Hence strength of cover above decline

very important.


Decline and surface subsidence

during construction, operation and

closure.

Cover depth above decline and

also the strength of the cover

materia (i.e.

clay/weathered/fresh). Surface

Loads crossing decline path

1

Medium:High

Dillhole through

spiral shows ground

conditions. Stress

Modelling assessed

surface loading

aspects

Medium: Data to

validate model will

only be available after

construction

commences

Medium:High

Dillhole through spiral shows

ground conditions. Stress

Modelling assessed surface

loading aspects

Medium:High

Decline spiral located in rock

and 2D modelling completed

to determine surface

loading effects

Medium: If rock surface was lower

than decline may be formed in

weak material and assumptions

would be changed.

Support the decline opening with

appropriate ground support once

the ground conditions are exposed.

Moderate/Major Unlikely Medium No N/A

Surface vibration due to

mine induced seismicity

Construction,

operation,

closure

Geotech_6mine induced

seismicity

mine induced

seismicity

Mine workers

Local communityYes

Vibrations can be felt by local community and

can cause concern and structural damage if

significant

Damage to local community

infrastructure. Potential of

surface damage to lease

infrastructure.

Ground conditions dictate if mine

induced seismicity is possible

together with the mine design

within the orebody and mining

method adopted

2

Medium:High

Ground conditions

favourable i.e. not

massive. Mining

method cut and fill

and not generating

large stress changes

quickly

Medium: Ground

conditions and mining

method not

attributable to mining

induced seismicity

Medium:High

Ground conditions

favourable i.e. not massive.

Mining method cut and fill

and not generating large

stress changes quickly

Medium:High

No modelling completed but

ground conditions and

mining method are not

favorable for mining

induced seismicity

Low:Medium

Not a deep mine, mining method

likely to be favourable and ground

conditions also likely to be

favourable.

Follow the cut and fill mining

method and and this will mininmise

the potential.

Moderate Unlikely Medium No N/A

Impact of mine

development if the

clay/rock surface is not

accurate

Construction,

operation,

closure

Geotech_7 weathering depthLower Rockmass

Strength

Mine Workers

Mine designYes

Strength of rockmass is very dependant on

ground suppot design for declines.


Decline during construction,

operation.

Geotech drilling to identify the sub‐

surface conditions. Development

of a 3D geotechnical model

1

Medium:

Geotechnical drilling

completed

Medium: Data to

validate model will

only be available after

construction

commences

Medium:

Coverage of drilling and test

pit work is reasonable to

have medium confidence

Medium:

Coverage of drilling data is

reasonable along decline

path to determine the

weathering boundary

Medium:

Drilling indicates complete

weathering is deep in places.

Model based on drill spacing.

Ground support required would be

significantly changed

Could do further infill drilling to test

the clay/rock boundary.Moderate Possible Medium No N/A

Impact of ground water

pressures on ground

support design

Construction,

operation.Geotech_8 ground water

Higher Loads on

ground support

Mine Workers

Mine designYes

Water pressures add significant extra loading

to gravity and mining induced stress


Decline/Headings due to added

pressure on ground support during

construction, operation. Water in‐

rush potential.

Development of a comprehensive

structural and groundwater model

critical

1

Medium: Regional

groundwater model

results provided.

Localised influences

not highly understood

Low: M1 not aware of

status/confidence of

groundwater

modelling

Low:Medium Ground water

pressure model confidence

not provided to M1

Low:

No dewatering of mine

considered unconventional

and requires specialized

expertise in grouting and

groundwater experts

Medium:

Analysis/Design has assed for Dry,

Medium and Large

inflows/pressures.

Definitive groundwater modelling Major Possible/Likely Medium/High No N/A

Corrosion of ground

support

Construction,

operation.Geotech_9

groundwater and

support type

Reduced strength

capacity of

ground support

Mine Workers

Mine designYes

Ground support type chosen must be able to

resist corrosion due to water in rockmass for

purpose of support. i.e. decline life of mine.


Decline/Headings due to

innefective ground support during

construction, operation.

Development of a groundwater

model and understanding the

chemistry of the water

encountered in the rockmass i.e

salinity, acidity

1

Medium:High

Likely to be seepage

in rockmass due to

not dewatering

Low/Medium

M1 not aware of


groundwater

modelling



not provided to M1

Medium:High

Quite common at sites with

water issues. Support type

required to survive the life

of infrastructure type.

Medium:

Assumption has been made that

wet ground conditions will occur

and appropriate support required

to resist corrosion.

Testing of composition of water at

depth.Moderate Possible Low/Medium No N/A

Grout Curtain to reduce

water inflows/pressures

ineffective

Construction,

operation.Geotech_10

groundwater and

grouting

Increased

pressure on

ground support

from water

Mine Workers

Mine design

Local community

Yes

Expected to be significant water pressures in

areas of the mine. No‐watering allowed, hence

grout curtains are critical to the mine design.





rush potential.



critical. Curtain grouting study to

understand if water targets can be

met through grouting

2

Low

No grouting study

results provided to

M1 to ascertain

whether water

targets are achievable



groundwater and

grouting modelling

Low:

Grouting study effectiveness

not completed yet.

Low:





groundwater experts

High:

No details regarding curtan

grouting known as of yet.

Definitive assessment of grouting

effectiveness.Moderate/Major Possible/Likely Medium/High No N/A

Stability of vent shafts

effected by ground

conditions and water

Construction,


faults and

groundwater

localised lower

rockmass strength

and higher water

pressure

Mine Workers

Mine designYes

Shafts can become unstable in poor ground and

were high water pressures or inflow is

occuring.

Lack of ventilation in mine

workings forcing mine design and

operations to be effected,

Development of 3D geotechnical

and structural model. Specific

geotech diamond hole along vent

path recommended also

1

Low:Medim

Shafts should also

have specific geotech

hole to determine

ground conditions


model will only be

available after

construction

commences



not provided to M1. Specific

ground condition along shaft

should be investigated with

diamond hole.

Medium:

Will require specific geotech

hole. Common practice for

designing of shafts.

Medium:

No detailed design completed yet.

Location assessed by stress and

ground conditions.

Development of a structural model

and groundwater model.Major Unlikely/Possible Medium No N/A

No‐dewatering and

grouting ineefective

causing signficant water

inrush and flooding

Construction,


groundwater and

groutingWater inflow

Mine Workers

Mine design

Local community

YesWater In‐rush potential on the HW fault

structure significant, based on modelling





rush potential.



critical. Curtain grouting in

advance to reduce/eliminate

inflows

2

Low

No grouting study

results provided to

M1 to ascertain

whether water

targets are achievable



groundwater and

grouting modelling

Low:

Ground water pressure

model confidence not

provided to M1. Also

grouting effectoveness not

assessed.

Low:





groundwater experts

High:

No details regarding curtan

grouting known as of yet.

Definitive groundawter and

grouting model developed.Major Possible/Likely Medium/High No N/A

Fibrecrete innefective

surface support due to

water pressure and

inflow

Construction,

operation.Geotech_13 groundwater

Loss of Surface

Support strength

Mine Workers

Mine designYes

Surface support can be lost if fibrecte is not

sticking to the wall or high water pressures are

cracking it.


Decline/Headings due to

innefective surface support for

ground control.



critical. Grouting to reduce inflow

and improve F/C strength

1

Medium

No detailed

groundwater model

completed. More

regional



groundwater and

grouting modelling

Low:Medium





assessed. Inter‐related with

fibrecrete.

Medium:High

Quite common at sites with

water issues. May require

mesh if F/C not working in

places.

Medium:

Ground water pressure model

confidence not provided to M1.

Also grouting effectoveness not

assessed. Inter‐related with

fibrecrete.

Develop a comprehensive

groundwater model.Minor/Moderate Possible Low:Medium No N/A

Large water

inflow/pressure once HW

fault structure

intercepted in shafts and

flatbacking drives

Construction,


groundwater and

faulting

Loss of stability of

infrastructure and

drives

Mine Workers

Mine design

Local community

YesWater In‐rush potential on the HW fault

structure significant, based on modelling

Injury/fatality due to water inrush

due to innefective grounting

during construction, while

intercepting water bearing

structures



critical. Curtain grouting in

advance to reduce/eliminate

inflows

2

Medium

No detailed

groundwater model

completed. More

regional



groundwater and

grouting modelling

Low:





assessed.

Low:





groundwater experts

Medium:

Ground water pressure model

confidence not provided to M1.

Also grouting effectoveness not

assessed.

Develop a comprehensive

groundwater model.Major Possible/Likely Medium/High No N/A

CRF sill pillar design not

strong enough to mine

underneath the sill pillar

below

Construction,

operation.Geotech_15 CRF pillar design

Loss of pillar

strength and

mining block

Mine Workers

Mine designYes

CRF design critical for the strength of pillar.

Cement quantity critical.


CRF sill pillar during construction

and operation..

CRF design mix and testing in

combination with stress analysis

for deformation analysis

1

Low:Medim

No CRF design or

testing has been

completed to date

Low:Medim

No CRF testing has

been completed to

date

Low:Medim

No CRF testing has been

completed to date

Medium:High

Using CRF sill pillars at other

sites is quite common

Medium

No CRF testing has been completed

to date.

CRf testing to be completed once

mine is at that stageMajor Possible Medium No N/A

mining subsidence during

and after completion of

cut and fill mining

Construction,

operation,

closure

Geotech_16ground conditions,

mining method

Loss of stability of

ground conditions

Mine Workers

Mine design

Local community

YesLoss of ground control movement from

orebody to surface

Injury/fatality and surface

infrastructure damage due to

large scale subsidence during

construction, operation and post

closure

Mining and backifilling method

has a significant role in subsidence

together with undrstand the

geology and geotechnical domains

3

Medium:

Backfilling of most of

the orebody will

minimise potential of

subsidence


model will only be

available after

construction

commences

Medium:

Mining method of backfilling

of most of the orebody will

minimise potential of

subsidence

Medium:High

Subsidence in cut and fill

type mines is un‐common

due to lack of voids

Low:Medium

No seinsitivy analysis completed.

Based on Mining method of

backfilling of most of the orebody

will minimise potential of

subsidence

Could complete high level 3D

numerical modelling to support

evidence that it is a low risk.

Major Unliekly High No N/A

Central access into

flatback levels need to be

offset from each other

Construction,

operation.Geotech_17 mine access design Mine design

Mine Workers

Mine designYes

mine design. Can cause thin wedge of rock in

backs unless accesses are offset.


back rock mass during

construction and operation..

The stope access mine design 1

High:

Final design to

incorporate offset

between adjacent

Stope access drives.

High

Adjust mine design

High

Adjust mine design to have

accesses offest to each other.

High

Low:

Mine Design of central access

would be required.

Change to current mine design. Minor/Moderate Unlikely Medium No N/A

Mining under wide CRF

sill pillars will be needed

and may not be strong

enough to support span.

Construction,


CRF pillar design

and mine design

Mine design and

CRF strength

Mine Workers

Mine designYes

mining spans need to be able to be supportable

with rock mass strength or CRF strength


CRF sill pillar during construction

and operation..

CRF design mix and testing in

combination with stress analysis

for deformation analysis

1

Low:Medim

No CRF design or

testing has been

completed to date

Low:Medim

No CRF testing has

been completed to

date

Low:Medim

No CRF design or testing has

been completed to date

Medium:High

Using CRF sill pillars at other

sites is quite common

Medium:

No CRF testing has been completed

to date.

CRF testing and modeeling Major Possible Medium No N/A

P:\2241_G\4819_V4

DOCUMENT INFORMATION

Status Final

Version 4

Print Date 26/7/2017

Author(s) M. Bijelac, B. Roache

Reviewed By Dr Peter Fuller

Pathname P:\2241_G Terramin Bird In Hand Feasibility Report\WPO\Final\June 2017\4819_V4.docx

File Name 4819_V4

Job No 2241_G

Distribution PDF emailed to client

DOCUMENT CHANGE CONTROL

Version Description of changes/amendments Author (s) Date

1 Draft Report M. Bijelac,

B. Roache

18/01/2017

2 Final Report M. Bijelac,

B. Roache

06/02/2017

3 Final Report – updated for new surface infrastructure layout

M. Bijelac,

B. Roache

09/06/2017

4 Final Report – updated for new version of erosion report

M. Bijelac,

B. Roache

27/07/2017

DOCUMENT REVIEW AND SIGN OFF

Version Reviewer Position Signature Date

3 Dr. P. Fuller Consultant Geotechnical Engineer

09/06/17

BIRD IN HAND GOLD PROJECT · 2019-06-19 · FINAL REPORT BIRD-IN-HAND GOLD PROJECT GEOTECHNICAL...

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Transcript of BIRD IN HAND GOLD PROJECT · 2019-06-19 · FINAL REPORT BIRD-IN-HAND GOLD PROJECT GEOTECHNICAL...