The Development and Application of Destressing
Techniques in the Mines of INCO Limited, Sudbury, Ontario.
By J. Denis P. O'Donnell Sr.
Laurentian University
SUDBURY. Ontario.
A thesis submitted to the School of Graduatr Studies in partial fulfilrnent of the
requirements for the Degree of Master of Science.
Apr. Yd. 1999
APPDEMCAoc
O Copyright by J. Denis P. O'Donnell Sr. 1999 National Library Bibliothèque nationale I*I of Canada du Canada Acquisitions and Acquisitions et Bibliographie Services services bibliographiques 395 Wellington Street 395, nie Wellington Ottawa ON KIA ON4 Ottawa ON K 1A ON4 Canada Canada Your fi& Votre réUnmai
Ow fi/e Notre reierence
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The author retains ownership of the L'auteur conserve la propriété du copyright in this thesis. Neither the droit d'auteur qui protège cette thèse. thesis nor substantial extracts from it Ni la thèse ni des extraits substantiels may be printed or otherwise de celle-ci ne doivent être imprimés reproduced without the author's ou autrement reproduits sans son permission. autorisation. iii Acknowledgements
The destressing project. the research. and compilation of this thesis was carrird out with the help and encouragement of a number of people. Sincerr thanks are due to al1 of them. In particular. I wish to acknowledge: Doctors C. D. Martin and D.H. Rousell for their guidance and suggestions for the manuscript: INCO Limited. C. Langille for his encouragement and Cor sreking approval for the project as a thesis topic fiorn Mines
Technical Services management: M. Sylvestre for approvd of the project as a thesis. Thanks are aiso extended to management at Stobie Mine particularly: J. Loring ( Manager): G. Elliott
(Mine Superintendent): H. Buksa and H. Parsons (Chief Mine Engineers): W. Quinn
(Grneral Foreman) for supplying the required support to kerp the project moving. Thanks are also extended to: Colin Mc Anulty (Division supervisor): Micliarl Paventi (plamer): and
Chris Wereley (technical representative with Dyno Nobel): for their direct involvement in the project. The author wishes also to acknowledge Chris Preston. research scientist with
Dyno Nobel and guest Professor of the Advanced Explosives course at Laurentian
University. for his encouragement and guidance at the on-set of the project.
In particular I wish to thank my wife. Susan O'Donnell. and my farnily for their encouragement. and patience during my absence while taking courses and panicularly during the writing of the thesis. iv Table of Contents
Page
Abstract
Acknowledgements
Table of Contents
List of Appendices
List of Figures
List of Tables xii
Chapter 1 Introduction
Chapter 2 Rockbursts in Underground Mining
2.1 Introduction
Factors Which Contribute to Rock bursting
2.1 .1 Lithostatic and Tectonic Stresses
2.1 .a Planar Discontinuities
2.2 Stored Strain Energy and Stress
2.3 Rock Strength v 2.4 StresdStrength an Indicator of Rock Mass
Insta bility 11
2.5 Evidence of Over Stressed Regions in a Mine
2.6 Rockburst Mitigation
2.6.1 Rockburst Research
2.6.2 Support
2.6.3 Controlling the Failure Process
2.6.J Changing Mining Methods
3.63 Destressing
2.8 Summary
Chapter 3 Destressing Case Histories 35
3.1 Introduction 35
3.1.1 Developrnent of Destressing at INCO Limited 35
3.2 A List of Destressing Techniques Developed by
the Author and Associates at INCO Ltd. 39
3 2.1 Drift Destressing 39
3 2.2 Break Through Destresssing 39
3.2.3 Destressing in Stopes. 40
3.2.4 The Destressing Pillars on the Si11 CUI. JO vi 3.2.5 Destressing a VRM Extraction Horizon at 2 Mm.
3.2.6 Destressing Stopes at Mid Height.
3.2.7 Destressing of Drifts Driven in Stressed Pillars
3.2.8 Abutment Destressing
3.2.9 Hanging Wall Destressing in VRM Topsill.
3.3 Cautions and Controls with Destressing
3.4 Problems Destressing RockWith Widely
Spaced Joints
3.5 Destressing in Tabular Ore Zones in South Africa 55
3.6 Examples of Pillar Destressing
3.6.1 Galena Mine 1970
3-62Campbell Mine 1980's
3.6.3 Macassa Mine 1987
3.6.4 Mt. Charlotte Mine 1995
3.6.5 Sigma Mine 1996
3.7 Analysing the Success of Destressing
3.8 Summary
Chapter 4 Destressing a Regional Pillar at Stobie Mine
41.1 Geological Setting
4.1 2 Stress Conditions At Stobie Mine
4.13 Rock Strengths vii 4.1.4 Joints
4-13 Mining Sequence
4.2 Mining Induced Seismicity
4.3 Numerical Modeling
4.3.1 Stages Modelled
4.3 .? Phases Evaluated
4.3 -3 Strength PararnetersIFailure Cntenon
4.3.4 Mode1 Indicating the Effect of the Destress Slot
4.3.1 Summary of Modelling Results
4.4 Explosives and Blast Design
4.4.1 Configuration of Destress Holes
4.4.2 E~plosives
4.4.3 Timing of Holes and Kg of Explosives per Del-
4.5 Instrumentation
45.1 Stress Monitoring
45.2 Microseisrnic System
4.53 Blast Monitor
4.5.4 Observations Before and Afier the Blast
4.6 Interpretation of Results
Chapter 5 Summary and Conclusions viii
Appendices
Appendix A Time line of the Project 120
Appendix B Numerical Analysis
B. 1 Description of the Mining Phases Modelled
8.2 Surnmary of Key Findings
B.3 Conclusion
Appendix C Glossary of Terms
Appendix D Correspondence and Approvals 139
D. 1 Letter recomrnending to management to destrrss 25 pi llar 129
D.2 Lrtter to INCO reqursting approval of destressing as a thesis 130
References 131 List of Figures
Figure Number Description
Exarnples of strain. pillar. and fault-slip rockbursts at Creighton Mine
Measured stress variations with depth
Extraction ratio and pillar stress
Discing of 150-mm core of granite
Spalling in the sidewall of a bored mise
Progressive developmrnt of the notch geometry in a tunnel at URL
Borehole breakout in a 28cm diameter raise bore pilot hole
Borehole breakout in 16.jcm diameter production holes on 55rn lrvel
Shear failure and sidewall spalling in highly stressed drifts in quartzite
Curved failure in drift corners
Progressive failure dur to spalling
Effect of total tidal force on rockbursts and falls of ground
Exarnples of cable lacing installation and hardware
Slender and squat pillars
General mangement of destress holes during the 60's
Photo of a blasted destress hole
Breakthrough destressing
Darnage caused by destressing a drift wall
Destress pattern 2070m mechanized cut & fil1 unit
Destress pattern used on 2 195m level extraction horizon X List of Figures Continued
Figure Num ber Description
Abutment destressing
Hanging wall destressing in a VRM topsill
Types of failurr in different rock masses and stress regimes
Destressing pattern for drifts in rock with widely spaced joints
Instantaneous failure and gradua1 sloughing in pillars
Destressing pattern at the Galena Mine
Layout of destress holes ai the Campbell Mine
Layout of destress holes at the Macassa Minr
Preconditioning at the Mt. Charlotte Minr
Blasthole layout and charging design at Mt. Charlotte Mine
Longitudinal sections of the Sigma Mine
Destress pattern at the Sigma Mine
Geological map of the Sudbury Structure
Frood-Stobie Offset dike and Section of Stobie Mine
Long. Section of Frood-Stobie indicating 85% extraction ratio
Section of 25 pillar indicating mining
Rockbursts per year at Stobie Mine
Rockbursts and seismic events associated with 25 Pillar
Plot of numerical mode1 geometry from Surface to 730m Ievel
Plot of the destress slot area xi List of Figures Continued
Figure Number Description
Plans indicating geometry of nurnerically modelled phases
Plot of (s, on a vertical grid through the 25 pillar
Plot of ai on a vertical grid through the 25 pillar
Safety factor plot on a vertical grid through the 25 pilla
Plan of 55Om level indicating stress change between phases 2 and 3
Plan of 55Om level indicating mining pnor to destress blast
Long. section of destress slot indicating attitude of destress holrs
Destress hole spacing and detonation sequence at Stobie Mine
Photos of the boosters with detonators and their installation
Photos of the destress drift prior to the blast
Plot of the wave form of the destress blast
Photos of the destress slot after the blast
Photos of the back of the destress slot after the blast
Plot of tnces captured of the 2.4 Mn rockburst
Plot of strain change captured by the stress ce11
Photo of a CSiRO hollow inclusion stress ce11 xii
List of Tables
Table Number Description
Influence of q/o, Ratio on Stability
Data from significant pillar destrrssing projects
Formulae from stress measurements at Stobie Mine
Caiculated stresses in the 25 PiIlar area
Uniavial and triaxial compressive strengths of rocks near 35 pillar
Attitudes of joint sets on 420m to 550m Ievel
List of seismic rvents and rockbursts associated with 75 pillar
Propenies of RUS rmulsion manufactured by Dyno Nobel Ltd.
Changes in stresses caused by the drstress blast at Stobie
PPV and peak sound pressure Ievels of the drstress blast
List of runs and descriptions of phases modelled
Abstract
Rockbursts are the result of the sudden release of stored strain energy derived £iom the stresses induced by rnining at depth or to a high extraction ratio. Conml methods range fiom modifiing geometries to altering the rock properties by destressing. The physical expression of the release of stored strain energy includes dog-earing, core disking, slabbing, pillar failure and bursting. A review of the histoncal development and a comprehensive list of the various rnethods of destressing are outlined. Special interest is paid to destressing of pillars, culminating in the description of the destressing of 25 Pillar, a regional pillar at NCO's Stobie Mine (world's Iargest destress blast to date). The destress slot, which was designed to promote failure in the core of the pillar, provides a dot dong which mining can safely proceed. The nsk of bursting has been reduced for the possible extraction of 1.8 million tons of ore between the 550 and 520-m levels.
Resume
Les coupes de troit sont la consequence de soudaine liberation d'energie accumulée. Cet energie denve des tensions provoquées par la profondeur ou un haut rapport d'extraction. Les methodes de control vont de la modification de la geomeüie a l'alteration des proprietés de terrain par des tires de relaxation. L'expression physique de cette liberation d'energie inclus la formation d'elipse dans les trous forages et de disques carrotes, I'ecaillage, et les coups de toit. La revue du developement historique et un appercu detaillé des methodes utilisées pour modifiées les tenains sont exposés. Un interest special est donné aux methodes utilisées pour les pilliers, suivi par la description du cas particulier du pillier # 25, un pillier regional a la mine d'MC0 Stobie. La rainure de tires de relaxation qui a été dessiner dans le but de promouvoir l'échec au noyau du pillier, represente le terrain au long duquel l'opération de minage peut continuer avec securité. Le risk de coups de roche a été reduit pour l'extraction de minerai entre les niveaux 550 et 520-m. Chapter 1 Introduction
Underground mining is subject to a number of risks that have safety and economic ramifications in terms of threat to personnel. dismption of production and expensive reconditioning costs. The cornmon physicai phenomena associated with these risks are an
unexpected in-rush of water. fire. methane explosions. falls of ground and rockbursts. Of these physical phenomena. rockbursts are the most rlusive. difficult to prevent. and have the potential to significantly impact the operational economics. eg.. die cost of rockbursts in the
Coeur d' Alene Mining district has been pegged at 8 to 18% of the total mining costs or $10 million per year (Bolstad 1990).
The cause of rarthquakes associated with major faults. such as the San Andreas. is the sudden release of stored strain enerpy (Scholz. 1990). Slip does not proceed smoothly and uniformly dong the rntire fault surtàce. Instead. it is local and abrupt. A rockburst is an instantaneous. violent failure of rock whereby fragments of rock are ejrcted into mine openings (Hedley 1992). Rockbursts are also associated with the release of stored strain energy and generally occur at great depths. However. rockbursts cmalso occur at shallow depths. in pillars under high mining-induced stresses. and in shallow areas of high tectonic stress.
The classic rnethods of rockburst aileviation are proper mining sequences: reasonablr opening dimensions. geometries and mining rates: stringent filling practices: the use of destressing or preconditioning. md the installation of enhanced support to control the damage. Of these methods destressing and support are the ody ones that cm be used by the excavation personnel. While the role of support has been extensively investigated as a mitigative measure to control rockbursts. destressing has not (Camiro 1995).
Destressing is a ground control technique. whereby explosives are used to release the stored strain energy in a controlled manner by.(l) creating or estending a Fracture zone around openings or (2) fracturing pillars that are likely to burst. A successhl destressing program reduces the potential occurrence of a rockburst and also reducés the need for costly support such as cable king.
This thesis is arranged in five chapters. Chapter 2 introduces the topic of rockbursts as a mining hazard that must be rnanaged. highlights the indicators for increased rockburst potential and summarises rockburst mitigation measures comrnonly used in mines. Chapter
3 is a summary of case studies where destressing. particularly pillars. bas been used as a mitigation tool against rockbursting. Chapter 4 describes the application of destressing to a regional pillar at Stobie Mine. Sudbury. To the author's knowledge. this application of destress blasting. is the largest undenaken by the international mining industry. Finally
Chapterj summarizes the main findings fiom this research topic.
Several appendices are included to accommodate data and figures from numerical modelling and detailed descriptions of various destressing techniques developed at MC0
Lirnited since 1962. with special ernphasis on applications developed by the author and his associates at INCO since 1980. Chapter 2 Rockbursts in Underground Mining
2.1 Introduction: Factors which contribute to rock bursting.
Kaiser (1988) proposed that rock mass failures are of iwo types: falls of ground (non- violent. gradua1 failures) and rockbursts (violent. instantaneous. brinle failures). Hedley
(1991)suggested there are three types of rockbursts: (1) strain bursts. (2)pillar bursts and
(3) fault slip bursts (Fig. 2.1).
Complex mechmisms produce rockbursts. According to Hedley (1 987) the major causes ofrockbursts are: (1) surface instabilitirs at or near a working face characterised by spalling at the free face: (2) propagation of shear fractures in the rock mass ahead of the stope face: (3) sudden collapse of over stressed pillars: and (4) slip dong Faults or bedding planes. Rockbursts are always associated with rrgions of over-stressed rock. Factors which lead to over stressing include: high hydrostatic and tectonic stresses. high extraction ratio and planar discontinuities. These are described in the following sections.
2.1.1 Lithostatic and Tectonic Stresses.
Ground stresses are caused by geological and artificial phenornena. The most comrnon environment where a rock will be subject to high in-situ stress is at depth. Where the in-situ stress is caused by lithostatic and the locked in stresses of tectonic ongin (Franklin &
Dusseault 1989). A mine opening dismpts the local stress field and a new set of stresses is induced in the surrounding rock. A knowledge of the magnitudes and directions of the in-situ and induced stresses is essentiai for the design of mine openings, because if the strength of the rock is exceeded the resulting instability may have serious consequences on the behaviour of the excavations. Lithostatic (vertical) stress is due to the mass of the overly ing rock colurnn (column length x rock density) and has a range of 0.026MPalm to 0.0324MPdm (Fig. 2.2). Brown and Hoek (1 978) use 0.027MPa /mas the average fiorn rnining and civil engineering projects around the world. Accordingly the vertical stress at the 550m level at Stobie is 15MPa. In the Sudbury area the maximum principal tectonic stress is horizontal and in a NE-SW direction (Cochrane 199 1). The ratio of the horizontal to vertical stress decreases with depth from 5 at surface to 1 -7 at the 550111 level (Fig. 2.7b).
Most mining methods require the leaving of pillars between stopes to reduce closure and to afford stability to adjacent openings. In some mines. pillars can be placed in Iow grade ore where as in others. the pillars are within the ore body and may be permanent or trmporary. In some cases the pillars have been designed for recovery at the final phase of mining. As rnining proceeds. the extraction ratio (mined material/original volume) x 100 increases (Fig. 2.3). At 25% extraction the stress acting on a pillar is 1.33 times the stress acting on the orebody plane. Reference to the plane of the orebody is significant because generally the stress acting on a pillar is a vector component of the vertical and the tectonic stresses perpendicular to the plane of the orebody. At Stobie Mine the hangingwall (HW) contact (70°.3 15') best represents the plane of the orebody. Assuming that the pillar stress is caused by the gravitational and tectonic forces (Galbraith 1992), the average pillar stress at Stobie Mine at the 550m level. with a 75% extraction ratio is 4 times the stress normal to the orebody plane, rquivalent to the stress at the 2.200m level of the rockburst-prone
Creighton Mine (Fig. 4.3b). Pillas in a mine with a high extraction ratio usually fail. This failure can be a violent pillar burst or gradua1 sloughing and spalling. O 10m
-SCALE-- -- LONG. SECTION CROSS SECTION DRIFT ON 2195m LNEL
CROSS SECTION OF 2070m LEVEL FEB. 22 1980
C-1 1 25 CROSS SECTION O- 10m SCALE
FLOOR HEAVE -- TRqCK DISPLACED c-4LONG. SECTION - 1585m LEVEL ROCK OISPLACE0 8Y ROCK8URST O 1 Om -SCALE
Fic. 2.1 Exarnples of the types of bunts ai Creighton mine. a) a strain burst in a drift at the 2195m level, b) pilla burst in a rib between two cut and fil1 stopes on the si11 cut and c) a fault-slip burst at the 1585m level caused by slip dong the 6 Shafi shear induced by driving the development drift. Fic. 2.2 Measured stress variations with depth. a) Vertical stress component can usually be predicted from weight of overburden. b) Ratio of horizontal to vertical stress is about 1.5 at depths below 1 Km, but rapidly increases to values greater than 5 in the 100 m nearest to surface. (Data fiom Herget 1986) FIG.2.3 (a) Extraction ratio = mined materidonginal volume x l O0 = 75%. (b) Pillar stress to extraction ratio using vibutary area. Areas with high tectonic stress and areas in close proximity to faults are two other environments in which mine openings are subject to high stress. The stress associated with the movement of the North American plate accounts for the high horizontal stresses measured in the Canadian Shield (Cochrane 199 1).
2.1.2 Planar Discontinuities
Planar discontinuities such as dykes and faults enhance rockburst activity as indicated by mine seismic networks (Ryder 1988). Stress magnitudes and orientations are significantly dfected by geological structures in the Underground Research Laboratory at
Pinawa. Manitoba (Martin and Chandler 1993).. For rxarnple stress magnitudes mrasured above and below a fracture zone ranged from -1 to +43 MPa and stress orientations rotated as much as 90" when this fracture zone was crossed.
2.2 Stored Strain Energy and Stress
Strain rnergy is a function of the stress state and the elastic conditions (Young's modulus and Poisson's ratio) of the rock mas. The stress distribution around an underground opening cm be determined using comrnercially available sobvare such as MAP3D or
EXAMINE 3D. These programs provide the stress distribution in klimensions for irregular shaped geometries.
To determine the snalli energy. the Young's modulus of the rock mass must be known.
This parameter is seldom known with any confidence and hence elastic strain eneqy is not widely used in evaluating burst potential in mines. Pnctitioners instead rely mainiy on the identification of regions in the mine where mining induced stresses exceed stress magnitudes that have ken empincally determined as cntical for a particular mine or rnining camp. A value of a,= 125 MPa has been used for Creighton and other Sudbury Basin mines as a stress level at which bursting is highly likely (P. H. Oliver. retired iNCO Rock Mechanics supervisor. persona1 communication 1986). Particular attention should be paid to areas that the mode1 indicates are in tension or have a reduction in confinement in close proximi~a zone of high stress. Based on cornparisons of numerical mode1 results with field evidence. the author has considers any areas. with o,2 70 MPa and o;ranging fiom a negative value (tension) to half the far field conîinement value. having a hi& potential of bursting.
2.3 Rock Strength
A rock mass is made up of intact blocks separated by discontinuities. The rock mass strength is rmpirically determined using Rock Mass Rating (RMR) and Tunnelling Quality
Index (Q) values. while intact rock strengths or unimial compressive strengths (03 cm be deterrnined in the lab from intact fracture-free specimens. The fiequency and orientation of joints. significantly reduces the rock mass strength rock relative to oc. The presence of geological structures influence stress magnitudes and orientations. Rousell (1984) indicares numerous joint sets exist in the rocks of the Sudbury Basin. with as many as 7 joint sets at a given station. P. H. Oliver (personal communication 1988) recommended the use of a rock mass strength of 62MPa for the strongly jointed rocks of the Sudbury Basin with a a, of 155MPa.
2.4 StresdStrength an Indicator of Rock Mass Instability
Rock must be subjected to stress greater than its compressive strength to fail violently. However. the overail far field in-situ stress regime need not be greater than the strength of the rock to cause failure because the geomeû-y of openings induce local zones of stress concentration which may locally exceed the breaking stren-gth. Martin et al. (1998) have indicated that brittle failure occurs at approximately a stress level of 115 of the (03 of a rock.
Hoek and Brown (1980) and Stacey and Page (1986) suggestrd that the ratio o,/q might be used to mess ground conditions in underground mines. Based on observations in
South Afncan mines. they proposed the classification in Table 2.4.
Table 2.1 Influence of o,/o,Ratio on Stability (Stacey and Page 1986). q/q 1 Description of Condition
I < 0.2 No particular problerns.
0.2-0.4 Spalling from the surface paralle1 to o,.Heavirr support required.
0.4-0.5 Heavy support required. Major Spalling.
0.5-0.67 Very dangerous and di fficult to keep open. Support heaw and costly.
>0.67 Impractical or rxtremeiy difficult to maintain open.
Martin et al. (1998) proposed that brittle failure occurs at a ratio of o,/o, = 0.33 and that this simple index could also be useci to mess the burst potential of a pillar. They suggested that if the stress in the pillar falls below one third of the cc(c& W3)the burst potential in a pillar is significantly reduced.
Using the average pillar stress of 100MPa at mid height of 25 pillar at Stobie Mine and the average ocfor disseminated quartz diorite of 2 12MPa. then o& = 0.47. Burst potential exists at mid height of 25 Pillar (375m)and it would increase towards 55Om level.
2.5 Evidence of Over Stressed Regions in a Mine
As rnining progresses. pillas and abutments are subject to increased stress. which may lead to early gentle failure or later violent failure. Hence it is important to establish which regions in a mine are overstressed and if possible. to determine by what mode the energy wdl likely be released.
Stifiess is best understood by cornpiving a mine to the reactions of a stiffand a soi? testing trame. When a rock specimen yields in a stiff testing framr. the testing frarne looses its load with only small arnounts of yielding by the specimen. Al1 of the energy released from the loading system is used in deforming the yielding rock sample. Thus the stiff testing thme cannot drive the yielding sarnple to failure unless additional energy is supplied by pumping more oil into the hydraulic system. When a rock specirnen yields in a sofi testing frame. the testing frame tends to maintain its load over large amounts of deformation. More energy is released from the loading system than can be used in deforming the yielding rock smple.
Thus the sofi testing framr drives the yielding sample to Failure without any additional rnergy being supplied. This example is directly analogous to a pillar. abutment or fault loaded between the footwall and hangingwall (Wiles et al. 1998).
The ability of an element in a mine to store energy is dependant on the geometry of the element (e.g. pillar = widthheight) and the properties of the surrounding rocks. When a stope panel is blasted. the reduction in confinement dong one or twro sides of a block of rock is accompanied by a release of the stored strain energy as the block expands. This process of energy release is Merdriven by the expansion and energy release of adjoining blocks. This aggregate of energy releasing blocks is the loading system. The loading system has properties that define the tpe of failure, which will occur. It is the enerw of the loading system. which dictates the behaviour subsequent to initial failure. whether yielding is induced by over stressing, loss of confinement. reduction OF the clamping forces on a fault. creep. fatigue. asperity shearing. large hcnire Formation or even blasting (Wiles et al. 1998).
Equally important is the stiffness of the loading system. which is the slope of the Ioad versus the displacement. The total energy released by the loading system. cm be calculated by intrgrating the load through the displacement denvrd by a boundary element nurnencal mode1 (Wiles et al. 1998).
Releases of stored dnenergy by slip dong faults causes earthquakes. Failure releases energy and this released energy cmproduce additional Mure. which in tem releases additional energy. Tlius failure and releasc: of stored strain energy once initiated cmbe self-propagating.
Failure and cnergy relerises occur on scales. ranging fiom earthquakes dong plate margins. to failure in mine oprnings ( fiom spalling to rockbursts).
Diamond drill core. of any diameter. rxtracted hmregions of high stress oRen fractures into thin discs (Fig. 7.4). This discing is the result of high stresses perpendicular to the borehole a.-+&: the higher the stress the thi~erthe discs. Various attempts have been made to relate the direction and magnitude of the principal stresses with the concavity and thickness of the discs (Dyke 1989; Song and Haimson 1997). FIG.2.5 Spalling in the sidewall of a bored raise in massive rock. The direction of the major principal stress is shown. Spalling initiates at points of maximum compressive stress concentration which occur at right angles to the major principal stress direction (fiom Hoek et ai. 1995). Damage or breakout in the wall of a borehole is cornmonly referred to as "dog earing" or -'eggingW.Borehole-breakout simulation tests (Maloney and Kaiser 1989) indicate the following: (1) failure of the borehole wall is due to shearîng in bands parallel to
(perpendicular to O,) (Fig. 2.5); (2) spalling occurs at the intersection of shear surfaces and extensional zones: (3) block failure follows as fngments slough into the borehole. The extent of the borehole elongation increases with increase in the stress parallel to the hole avis and with the increase in magnitude of the difference between the stresses in the plane perpendicular to the hole ais (Fig. 2.5). Because the breakout is formed in a direction perpendicular to a,. observation and documentation of this featwe provides an inexpensive means of determiring variations in the near field stress direction. Because rock hilure is scale invariant (Tschalenko
& Ambraseys 1WO), borehole breakout occurs in holes of al1 siws. Martin and Young ( 1993).
Lee and Hamison ( 1993) and Martin ( 1997) document the formation of borehoie breakout type notches in a 3.511 diameter tmel driven in massive uniform. granite using perimeter lim drilling and mechanical breaking of the rock stub (Fig. 2.6). Boreholr breakout. occuned within a few hours derdrilling a raise bore pilot hole (d = 28cm) at Creighton Mine (Fig. 2.7) The formation of borehole breakouts in production holes (dia = 16.5cm)in the 25 pillar at Stobie mine signalled the advent of failure and the need for destressing (Fig. 2.8).
Another indicator of high stress mapinides is the slabbing and spalling From the roof and walls of underground excavations (Hoek et al 1995). Exarnples of slabbing include the following: (1) shear failure in the corner of a drift (Fig. 2.9~1);(2) sidewall spdling (Fig. 2.9b);
(3) curved slabbing due to shear failure in corners (Figs. 2.10% 3.1 Ob): and (4) progressive levels of failure due to spdling (Figs. 3.1 la, 2.1 I b). Lu et al. (1 989) compare and interpret the brinle failures observed in the excavating of three. 10k long headrace tunnels for the Tianshengqiao hydro povier station in SE China to
mechuiisms of failure obsemed in tests of rock specimens at different confining pressures. The three mechanisms of failure observed were failure by brittle splitting, tensile failure associated
with the shear process and failure by pure shear fracture. These mechanisms occurred at low. moderate and high confining pressures. respectively. The location of the failure in a drift cohedthis. The splitting mechanism ofa rockbwst vas hcturing of high-strength materiai
under low confining stress. The slabbing on the sidewdls occurred in areas of intermediate confinement. The shear failures. in corners of drifts rvere in tightly confined zones. with high.
induced stress and high confining pressures. The strain rnergy released due to spalling is less
than the energy releaxd due to bursting.
2.6 Rockburst Mitigation
As previously mentioned. the classic methods of rockburst alleviation are proper
mining sequencrs: reasonable dimensions. geometrirs and mining rates: and stringent filling practices. These are strategic methods. The "tactical" methods (Morrison and MacDonald
1989) are the use of destressing or preconditioning and the installation of enhanced support to control the damage.
Hedley (1992. p 725) in the Rockburst Handbook for Ontario Hardrock Mines. reconfirms the validity of the conclusions reached by the Rockburst comminee 10 years prior. "The prediction of the time of occurrence of a rockburst is of little actual value to a mine operator and it wodd be far more usefd to delineate areas predisposed to bursting and to apply means to inhibit and control the burst". FIG.2.6 Progressive development of the notch geometry in the roof and floor of the
Underground Research Laboratory test tunnel over a five-month period fiom Martin (1 997). FIG.2.7 Borehole breakout that occurred within a few hours of drilling a raise bore pilot hole (dia. = 28cm) at Creighton Mine. The tricone bit jamrned with debis fiom the breakout. FIG.2.8 The formation of borehole breakouts in production holes (dia = 16.5cm)in the
25 pillar at Stobie mine signalled the advent of failure and the need for destressing. Frc. 2.9. a) Shear failure surfaces in the corner of an underground excavation in highly stressed quartzite. b) Severe sidewall spalling in an excavation in quartzite in a deep level
South African gold mine (both after Hoek & Brown 1980). FE. 2.10. a) Stanrock Mine, Eiiiot Lake, Ontario. Note curved sidewall of service way. b) Lac-nor Mine Elliot Lake, showing a pillar face fomiing the curved sidewall of a stope (both
&er Momson 1976). Fie. 2.11 a) & b) Progressive levels of failure due to spalling (Hoek et al. 1995). Momson (1976) considered the pre-rnining stress distribution. the structure and properties of the rock mass. and the support and geometry of the excavations. as the major variables contributing to rock failure. graduai or sudden. He considered the pre-mining stress distribution and rock mass properties as facts that had to be accrpted and concentrated his rockburst control efforts on support and geometry. Today the practitioner attempts to alter the distribution of the pre-mining stresses and modie the properties of the rock mass in and about the excavations using destressing techniques. The following are a few specific exampies. The methods of stress shdowing developed in South Africa. to suit the geornetry of shallow-dipping. deep. tabular. goid veins include: over stoping (Leih 1990): driving "TM shaped driRs (Jager et al 1990): over slotting: mining of over-reefs first: and rnining hanging walls before footwalls (Cook et al. 1977).
Areas of hipher stress and associated risk of rockbiirsting in the Creighton 1 1 shah transfer drifi. defined by discing of core in pre-development diamond drill holrs. were subjected to additional destressing (Garrood 1982). These "structure" holes were drilled solely for the purpose ofdefining the zones with high stress. Emphasis has been placed recentiy on imaging systems such as Tomography (Fridel et al. 1996) at the Homestake
Mine. Successful identification of high stress zones would natiirally direct one to use destressing techniques to reduce the risks of bursting, and increase the ground support to accommodate any potential bursts.
Back analysis of bursting sites to define the conditions and causes of bursts have entailed numerous new techniques. Over coring detailed stress risee (Whyatt et ai. 1993) and first motion analysis confi~rmedfault slip mechanism (Jenkins et al. 1990) at the Lucky
Friday Mine. Wallace. Idaho. 2.6.1 Research on Rockbursts
A Fault-slip event at the Falconbridge mine in 1984. resulted in four fatditirs (Bharti
& West 1984). The subsequent Stevenson Inquiry. recomended a Canadian Rockburst
Research Program. whose purpose was to investigate rockbursts in Canadian mines. The first five years of the research program. directed by Dr. D. Hedley of CANMET. concentrated from its onset on developing seisrnic systems to; (1) locate the rvents. (2) capture wave forms and (3) determine source mechanisms. and culminated in the publication of the
'Rockburst Handbook for Ontario Hardrock Mines'. The handbook documents fundarnentals. a historical review of bursting in Canadian Mines. and a list of destressing techniques used by Canadian companies. During the subsequent five yrars (1 990-1 995) the Canadian
Rockburst Research Program. which was funded by the Industry through the Mining
Research Directorate (Mm)and both the Federal and Provincial Governments. concentrated its efforts on seismology. mine design. support. backfill and monitoring. The results are contained in a six-volume set. bound in two tomes (Camiro. 1995). Drstressing was not included in any of the research phases of the final 5 years.
An analysis of bursting that occurred during the mid 1980's in the Copper Cliff
North Mine indicated abnormally low confinhg pressures as the cause of bursting (Momson
& Galbraith 1990). The lower confining pressure produced a lower rock strength than expected.
The conditions that cause rockbursts are numerous and their interplay is cornplex. A study by O'Donnell (1995) documented the influence of solar and lunar tides on the frequency of rockbursts and falls -ofground at ten INCO Lirnited mines during the period of
1985 to 1994. A frequency of rockburst occurrence of 1.5 times the average was 0.15 -
014 - 0.13 - Olt - 0.11 -
01 -
009 -
O08 - O07 - 006 -
TOT& Ti%L FORCE -+AVERAGE NUMBER OF EVENTS PER DAY
26-40 41-50 51.60 61-70 71-10 81-00 91-1W 101-113 TOTAL TiDAL FORCE - AVERAGE NUMBER OF FALLS FER DAY
FIG. 2.12 a) Histogram of the fiequency of rockbursts and major seismic events recorded at Creighton Mine between 1985 and 1994 versus total tidal force. b) Histogram of the fiequency of ground falls recorded at al1 MC0 mines between 1985 and 1994 versus total tidal force (O'Donnell 1995). observed on days with the lowest total tidal force. tnterestingly. the frequency of falls of ground was highest when the total tidal force was the highest (Fig. 2
Blake et al. (1998) state that there have been few well instrumented field tests to validate the mechanisms by which destressing reduces the rockburst potential. However. a recent request for submissions on destressing resrarch has some possibility of advancing the state of the art. conditional on the sponsorship fiom indus- and the availability ohsuitable venue for in situ tests.
2.6.2 Support
Standard support systems (e. g. mechanical bolts. grouted rebars. friction bolts and welded wire) are inadequate in areas subject to heavy rockbursts. brcause the. cannot accommodate the larger displacements. Accordingly speciality support systems such as cable
king have been devised.
Cable lacing is a three-tiered support system consisting of grouted bars. chain link mesh and cables. The grouted smooth bars knoiçn as shepherd's crooks. or hairpins. (Fig.
2.13) are installed on a 1m x 1 m to 2m .u 2m pattern. depending on the Irvel of risk. Chain
link mesh is fastened to the rock with tensioned cables that are interlaced through the pins or shepherd's crooks (Hedley 1992). This robust support is commonly used in deep South
African gold mines and occasionally in Sudbury. Swan ( 1989) defines cable king as the final solution at Falconbridge Limited whereby rockfalls. rehabilitation and exposure of employees is mitigated (Fig. 2.13 b).
Other suppon systems include: (1) the Cone bolt. a grouted yielding tendon developed by the Charnber of Mines Research Organisation (COMRO) (Ortlepp 1994); and
(2) shotcrete, which is used as a coupling element with rock bolts and meshing (Cook et al. I I
HAlR PIN
SHEPHERD'S CROOK
14MM 8AR INSIDE OIAMETER OF EYE 37mm
LEGS 1.8 OR 2.4m
EXAMPLES OF A SHEPHERD'S CROOK
a) AND A HAlR PIN USE0 IN CABLE LACING
Fie. 2.13 a) Exarnples of support hardware used in cable lacing. b) Photc cableilacing at Stratcona Mine (Davidge et al. 1988). 1977. Langille & Buss 19%). The energy absorbing capacity of mesh reinforced shotcrete
(5Om thick), is greater than 16mm rebar shepherd's crooks. The performance of steel-fibre reinforced shotcrete is erntic (Stacey et al. 1995). .4t the Lucky Friday Mine. before dnlling the next round. shotcrete is applied to the back and \vails which are reinforced with bolts and chain link rnesh (Blake et al. 1998).
2.6.3 Controlling the Failure Process
If failure in a conf guration or sequence is inevitable. then attempts should be made to control the timing and rate of failure. Controlled bilure may be achieved by one or more of the following methods: designing slender pillars, allowing time in the mining sequence for energy to be released. delaying entry. and installing soA support.
The geornetry of a pillar affects its strength and ability to store strain energy (Parker
1974). A slender pillar will fail under low stress conditions. A squat pillar with a height to width ratio of 0.25 or less will devrlop conf nement in its core and will build up stress. Thus regional or boundq pillllars should be squat and pillars that are required to fail should be slender (Fig. 2.1 4).
Some configurations are prone to bursting. Blasts in these areas cmbe scheduled at times when a natural luIl in activity would follow. during which time energy could safely dissipate by bursting when no one is around. The blasting of footwall slots in the silling process on the 2075111 level at Creighton had a high potential of inducing rockbursts. To ailow the groound to release stress safely, the blasts were detonated at the end of the 4-1 2 shifi on Fridays. giving till Monday rnorning for the ground to release the excess energy.
The decay time of a block of ground affected by a vertical retreat mined 0or crown blast can be detemined by monitoring seismicity over a number of blasts. Once the decay time has been defined, entry to a workplace or a portion of a mine cm be delayed. to reduce the nsk of injury.
The installation of a stiff support element. shortly afier blasting a drift that is under high stress. cm rupture the support. A case in point is the insrallation sequence used in drifts at Creighton. Since the drifts are destressed. considerable deformation is expected in the drift back and walls. Mechanical bolts are installed as the first pass of support. Rebars are only installed a day to a week later. because if they are installed earlier. they will rupture at approximatrly 0.3m from the collar. The installation of mechanical bolts has a second-. safety feature in that they are quicker to install and thus the exposure time for the miner is much less than for a rebar installation. Anempts have bren made to develop macro-delayed resins for placement at the collars of rebar for use in conjunction with the qui& setting cartridge at the toe of the hole (Laneille. persona1 communication. 1997).
2.6.4 Changing Mining Methods
The diminishing pillar. created as a cut and fil1 stope is mined. and the pillas in a stope and pillar bulk mine attract stress and may burst. Some major initiatives that have been implemented to reduce rockbursts are: (1) conversion from the cut and fil1 method to a primary undercut and fil1 method at the Lucky Friday Mine (Noyes et al. 1988): (2) centre out mining with reduced spans in VRM topsills at the Copper Cliff North Mine (O'Donnell
199 1): and (3) designing and excavating a stress relief dot at Creighton Mine (Oliver et al..
1987 and MacDonald et al.. 1988). Changes of this nature bring out other questions: (1) cm a consolidated fil1 be developed that wilI accommodate the mining method (Brackebusch
1988) and (2) will the support potential of the cernented-fil1 influence rockbursts. The fil1 vu w/ / HEIGHT : WlOTH RATIO
PILLAR HEIGHT : WlDTH RATIO IS AROUNO 1:2 or 0.5 - CENTER OF PILLAR IS CONFINED HENCE VERY -STRONG
PlLtAR HEIGHT : WlDTH RATIO IS AROUNO 1:4 or 0.25
DEC1 F4.ûWC
Fic. 2.14 Examples of slender and squat pillas (Parker 1971). a) Height: Width =
1 :1; The edges of the pillar fail in shear then tension cracks open up in the centre. b) H:W
= 1 :2; Gradual deterioration of the outer part of the pillar decreases confinement on the centre and increases the load on the remaining central portion. c) H:W = 1:4; The roof deflects a little and exerts an inward force on the pillar. voids or gaps. in both the under cut and fil1 and VRM mining methods. generally preclude the ability of the fil1 to develop resistance to closure that would re-induce wall stresses of a magnitude. for bursting to develop. Bursting would be more likely to develop due to relaxation and slip associated with reduced clarnping forces (Whyatt et al. 1989).
2.6.5 Destressing
There is considerable uncertainty conceming the mechanics of destressing. The main reason for this uncertainty is that destressing was developed as a practitioner's tool. While there have been several attempts to quantify the rffects of destressing through field rxpenments. to the author's knowledge nonr of these experîments have shed any light on the underlying iùndamentals of destressing. None the less there is a dearth of operational experience that shows desvessing does reduce the rockburst hazard.
Numerous studies have been undertaken to simulate the effect of destressing by numerical modelling. Attempts have been made to compare different destress hole locations
(Mitri et al 1988) and calibrate mode1 results with field observations. Design tools developed in South ficato assess different stratrgies included: ( 1) Energy Rrlease Rate (Cook 1966); and (2) Excess Shear Stress (Ryder 1 988). Wiles et al. ( 1998) used numencal modelling to correlate the occurrence of bursting to an rnergy release density. The Local Energy Release
Density was equated to the area under the loading system stiffness curve divided by the sample volume. Reasonabie correlation has been indicated in his preliminary srudies.
Destressing techniques have been developed. by the author and associates at INCO's
Creighton mine to accommodate various mining methods (O'Do~eii 1992). They essentially use explosives to create and extend Fracture zones or promote pillars to fail.
Before studying the techniques and their development, some of the notions that have been put fonvard as to why destressing works. are reviewed.
Oliver et al. (1987) States that stability requires the establishment and maintenance of a skin of hiled material around al1 openings. and recognising this is the key to minimising the rockburst hazard in development openings at depth. This skin of failed material is produced by destressing.
Rorke and Brurnrner (1988) suggested destressing serves to release strain energy in highly stressed rock whereas preconditioning serves to fracture and weaken large areas of unstressed rocks. to prevent the dangerous acc~imulationsof strain energy as mining progresses. Thry described the desired results of preconditioning of deep tabulu gold vrin in South Afiica. as reduction in horizontal clmping forces thus rncouraging the occurrence of stable shear movements dong fracture planes.
O'Domeii (1 988) stated that destressing in primary development drifts destresses and preconditions the rock. The face holes. generally drilled twice the lrngth of the round. push the stresses ahead of the excavation. where as the shoulder (corner) holes fracture the ground pnor to blasting even the cut. As the excavation is formed. the stress concentration at the shoulder induces failure in the weakenrd rock initiating the progressive failure that promotes wall closure. slippage dong joints and release of energy as the excavation progresses to the desired state of low potential energy. synonymous with Iow risk of bursting. The bolting procedure prornotes this process by the injection of water into the rock mass and along joints. It is indeed at this point that a significant increase in audible rock noise is detected as the transition to stability proceeds.
Blake (et al. 1998) sumrn~sethe goal of destressing as the promotion of fracturing or mobilising movement on pre-existing fractures resulting in a change in the deformation mode from bnttle to plastic.
2.8 Summary
Rockbursts are associated with: deep mines. areas of high tectonic stress. faults. and high extraction ratio. Mining sequence. pillar geometry. destressing. as well as timing of re- entry are steps that cmbe taken to reduce the risk of rockbursts. Chapter 2 reviewed: (1) the various methods by which stored strain energy is released: (2) the causes of rockbursts: and
(!) the solutions available to reduce their risk. The following chapters document significant pillar destressing projects in order to drvelop a Frmework with which to compare the destressing of 25 Pillar at Stobie Mine. Chapter 3 Destressing Case Histories
3.1 Introduction
The previous chapter deds with stress. strength. stored strain energy. rockbursts and destressing as an effective tool in releasing stored strain rnergy to reduce the risk of a rockburst occurring. This chapter documents the application of destressing to drifts. stopes and tabular ore zones and presents a number of case studies on pillar destressing.
3.1.1 Development of Destressing Techniques at INCO Limited.
Destressing up to 1980
Dickhout ( 1962) was one of the fints to elaborate on the use of destressing at INCO
Limited's Creighton Mine. During the earl y sixtirs the cut-and-fil1 method was being introduced to replace the square set mining method. brlow the 1220m Irvel. Expanded rnetal mesh and rock bolts were being used to replace sets as drift support on the 7010rn Ievel. It was ai this transitional stage in the development of mining that destressing mrthods were developed. Ammonium Nitrate and Fuel oil (ANFO). which had been recently introduced as an explosive. made long destress holes frorn a working face possible. Dickhout indicated the destressing at that time consisted of two drill holes (d=57mrn). one 7.6m the other l5.2m in length (Fig. 3.1). The long hole was cornpletely filled with ANFO and collar primed. This arrangement produced the required degree of destressing and produced a relatively clean hole that could be readily exarnined and occasionally was used in the cut (Fig. 3.1). New destress
holes were ddled when 1.8m of the previous destress holes remained. Destress was a
success because the openings were stable and there was a marked reduction in overbreak and
spalling. nie latter was very significant for miner safety since the 2.4m x S.7m drifts were
driven using jack leg drills. The successful use of destressing permitted the continued use of
rock bolts for drift support as opposed to reverting to the more expensive timber or steel sets.
Drilling the long holes and reconditioning after the blast were time consuming and
expensive. Fiirther refinements determined that two face holes, hÿice the length of the round,
would adequately destress a face.
NOT TO SCALE
PREVIOUS DESTRESS HOLE 1 r--- DESTRESS HOLES 1.8m I I 1 15mH I , DRIFT 7 ma PlANNED AOVANCE
FIG.3.1. Plan of a drift showing the general arrangement of destress holes used during the 60's (after Dickhout 1962). FIG.3.2 Enlargement of a blasted destress hole (after Dickhout 1962). The lack of data on destressing techniques in the interval 1962- 1980 is apparentiy because
drillers were responsible for drilling and breaking rounds. If the rounds break cleanly then
no one interferes with the process. In some cases the only information available is anecdotal
rvidence. For euample. the idea of placing destress holes in the corner of working faces is
credited to a crew working on the 2070m level at Creighton mine. The rule of not drilling
within .l 5m of a previously blasted hole (bootleg), and twin destress holes in a face. make
locating of the cut dificult. The crew noticed during the process ofdrilling bolt holes in the
shoulder. that the magnitude and frequency of rock noises (low level seismicity) increased above back-ground. When the holes were completed. the seismicity Fe11 below background.
Apparently. corners trigger seisrnicity. Accordingly. destress holès at the corners produce
beneficial results. However. face holes are still required to push the fracture zone ahead of
the face.
For destressing to be readiiy accepted by rniners the following are required: (1) they
must be aware of the risk of rockbursts: (2) the method has to be easily performed with the tools and supplies that are available: (3) the process rnust not significantly delay the drilling and blasting of a round: (4) there has to be minimum damage and repair work required because of the destress blast; and (5) the results must be convincing. 3.2 A Comprehensive List of Destressing Techniques Developed by the
Author and Associates at INCO Limited.
3.2.1 Drift Destressing
The configuration developed over the years to reduce the risk of strain bursts in dzvelopment drifts is the drilling of. two face holes twirc: the lèngth of the round. and holes in the upper corners the sarne length as the round. Wall holrs and holes in the bottom cornes were required in some cases (Oliver et al. 1987). The method is successful and no reconditioning of the drifts is required. There is generally an increasr in the fracture pattern at the face that causes some problems in drilling the next round.
3.2.2 Break Through Destresssing
When two drifts are being driven towards each other. die strain enerçy and associated seismicity increase as the distance between them decreases. The technique developed by the author in order to reduce the risk of a rockburst is: the headings are stopped four round lrngths apart. and two holes are drilled. twice the Iength of a round. from each heading (Fig.
3.h). Instead of one meter of ANFO at the toe of each hole. the holes are loaded from the toe out to a le& of approximately 75% of the length of the hole. It should be noted that these holes have a strong kick back or shot gun effect. Care should be taken that thesr holes are not in line with any crucial equiprnent or power lines. The rernaining four rounds are usually taken from one side using the customary corner and wall holes.
A similar procedure can be used if a drift is to break into a stope and there is a nsk of bursting (Fip. 3.3 b). The procedure used in the 1290 zone on the 2070m level at Creighton mine when dnving a cil! drift towards a filled cut and fill block was as follows: A diamond drill was used to driil two destress holes L5.2m long towards the stope. Approximately 75% of the hole (from the toe out) was loaded with ANFO and blasted. Access into the area was prevented for a predetermined decay period. usually a minimum of 8 hours.
3.2.3 Destressing in Stopes.
Destressing was used in cut and fill stopes on the 195Om and 20 10m levels during the mid-seventies. A series of holes were drilled in stiff portions of gabbro or granite nb pillars. These holes were used in both traditional high fill cut and fill and in uppers cut and fill stopes. In the latter a ten foot slicr was blasted from hanging wall to footwall at once. using narrow diameter 35mm drill holes. The new back was then bolted off the rnuck pile. which was subsequrntly mucked out. The destress holes were ?.-lm long and on a 3rn spacing. There is no wtitten evidence of the amount of explosives used. Reconditioning of the walls was required after the destress holes cntered the walls and peeled back the screen
(Fig. 3.4).
3.2.4 The Destressing of Pillan on the Sill Cut.
Methods developed in the early eighties to dcstress rib pillars on the si11 cut of mechanized cut and fiIl complexes. and the destressing of extraction drift ribs. are detailed in O'Donnell (1 992). However. to make this document complete the pertinent facts will be out1 ined.
Bursting of pillars on and above the 1950 m level did not occur until removal of the second to fourth cut (Oliver et al. 1987). Pillar bursts started on the 2010m level. In the early eighties a program of destressing was developed by the author and associates in cut and fil1 PLAN VlEW
a) BREAKTHROUGH DESTRESSING FOR A DRIFT BREAKING INTO A STOPE
b) BREAKTHROUGH DESTRESSING FOR A DRIFT BREAKING lNTO A DRIFT
PORTION OF THE HOLE LOADEO AND DESTRESSED
FIG. 33 a) Breakthrough destressing for a drift breaking into a Siope. b) Breakthrough destresshg for a drift breaking into a drift. FRACTURE ZONE
Sm LONG 32 mm OIAMETER OESTRESS HOLE
Fic. 3.4. Damage caused by destressing a drift wall.
PLAN SlLL CUT
SECTION A-A'
FIG.3.5 Destress pattern in the 2070m level Mechanized cut and fil1 units. complexes on the 2073m level (Fig. 3.5). The complexes had 7.9m wide stopes and 4.9m wide pillars with a 9. lm wide central slot. The stope direction was parallel to the least principal stress.
Destressing included the blasting of 1.jm of ANFO at the base of twvo 6m. face holes.
The first 3 m of these holes were ddled with a 5Jmm cut bit to accommodate the coupling which was required to extend the hole to 6m with a 43mm bit. The corner holes were 3m long and J3mm dimeter and contained O.jm of ANFO. These holes were drilled up and out at 30 degrees to the advance direction. Wall holes at 45 degrees to the direction of advance
(3m long. dia43mrn) were drilled 1.8m back from the face and 1.8m above the floor. These holes inclined up a few degrees to allow the drill water to drain out. were destressed with
0.9m of ANFO. The pillars were inspected daily and if they remained competent the destress holes were re-blasted. Records were kept of the pillars that yielded as a result of proper destressing.
The yiclding of the pillars was significant. As the stopes rxpanded. a cntical dimension was reac hrd at which the back and floor converged as indicated by an episode of elevated seismicity. Pillas that were not properly dcstressed would burst.
The level of support that proved adequate to support the pillars through the episode of convergence was the following; the pillars (walls) were screened to the floor with f: 4 gauge welded wire mesh. bolted with 1.7m long ss39 split sets and 1.8m long 19mm diameter resin grouted rebars.
Silling of the cut and fil1 unit on the 21 34m level proceeded with fewer problems due to the implementation of techniques developed on 2073m level. These included: a)
Developing the stopes in a wedge fashion (Le. pushing the stress outward). b) Planning stopr liniits to maintain a smooth outline. c) Blasting of the slots through the pillars on the 16h00 to 24h00 shifis on Fridays. to allow the stresses to redistribute during the weekend, when no one was present in the workplace. d) Increasing the widths of the dots from 3.6 to jm
reduced the hazard of loose ground to the construction crews. and e) al1 destress holes were
tashed out and the collars painted; the? were treated as bootlegs.
The state of the art progressed from a number of uncontrolled bursts in the initial
silling of the 2073m level cut and fill. to the occurrence with the drvelo~mentblast of 10 out
of the eleven bursts recorded during the silling of the 2 134 levrl mining block. During the
development of the 21 95 m level VRM extraction horizon 85% of the bursts were initiated
with the development round.
3.2.5 Destressing of the Vertical Retrert Method Extraction Horizon on the
2 MmLevel
The extraction horizon for the Vertical Rrtreat Mining (VRM) on the 21 95m level
contained 4.9 x 4.9 m drifts with 4.9m wide pillars designed to fail easily/sofly with proper
desvessing (Fig. 3.6). The drstress pattern of the extraction drifts was the most involved of
the time with 10 destress holes per round including; two 7.3m face holes with 2.7 meters of
ANFO and four wall and four comer holes each 3.6m long and Mmm in diameter with 0.6m
of ANFO. The four corner holes were ddled up and out and dom and out at 30" to the
driving direction. The bonom comer holes were drilled about .3m above the si11 to prevent
water fkom entering. The four wail holes were drilled at 45" to the direction of advance and
a few degrees above horizontal to allow drill water to drain. The wall holes were staggered C' C' ' SECTION ' 'C
Fic. 3.6 Destress hole pattern used on the 2195m level extraction Horizon at
Creighton Mine (O'Domell 1992). so they would not intersect holes from adjacent drifts. Typically the holes on the left wall
were ddled at the shoulder and Z.4m above the sill. The holes on the right hand wall were
drilled at 1.2m and 3m above the dl. Improper alignment of the wall holes cmdestroy the
wall of the adjacent drift.
3.2.6 Destressing Stopes at Mid Height
On a number of occasions cut and fil1 stopes at Creighton have developed stress
problêms. as they reached rnid height between levels (33 m). Evidence included spalling at
the shoulder and an increase in seismicity. The abovr occurred in stopes that had been
destressed at the si11 cut and some that had not. The stress problrm \vas solvrd by destressing
the rib pillars using the up and our 30" holes in the corners and holes at 45" to the advance
at mid height of the cut. This procedure was carried out in stopes that showed the need
between the 1 830m and the 2 13 Sm levd.
3.2.7 Destressing of Drifts Being Driven Through Stressed Pillars
At Copper Cliff South Mine during the mid-eighties two drill drifts on the 700m
level were being driven through a stressed pilla. Spalling at the face and shoulden occurred.
This was mitigated by using the standard two face holes twice the length of the round and
two top corner holes up and out at 30".
3.2.8 Abutment Destressing
The abutments of stopes below the i 83Om level at Creighton would be under stress and hence prone to burst. The nsk of bursting fiom these sites was diminished by drillhg and blasting destress holes. In the cut and fiIl stopes. face holes hvice the length of a round were
ddled in the last round on the hanging ivall and footwall of each stope in the rnechmized
unit (Fig. 3.7a). Appro~imately1 meter of ANFO was detonated in these 44mm diameter
holes. If a slice is mined in an under cut and fiIl stope dong the hanging wall or footwall
contact. the destressing is as follows: a ?.-lmlong holr (dia.=32mm) at 45" to the direction
of advance. is drilkd at mid height into the wall with each round. Approximately 0.5 meter
of ANFO is detonated in the holes (Fig. 3.7b). If the under cut is mined transverse to the
contact then face holes are destressed in the last rounds into the hanging wall and foot~vall.
3.2.9 Hanging Wall Destressing to Eliminate Hole Closure, Struck Rods and
Reduce Dilution.
Longitudinal stopes mined by the vertical retreat mined (VRM) rxpenence hole
closure when depths excerd 1800m or whcn stressed pillars are mined. Rods become stuck
causing delays and loess of 'in the hole' (ITH) hammers.
Figure 3.8 indicates the geometry of panel #3 mined between the 1830m level and
the 1890m level at Creighton mine. Holrs were drilled parallel to the hanginy wall to reduce
delays and losses due to hole ciosure. Rings of holes (dia.= 170mm) were drillrd 3 meters
apart with a 3m spacing benveen the toes of the holes on the 1890m level. The holes in ring
13 and holes B and C in ring 14 were dnlled without problems. Holes A. D and E in ring 14 and al1 the holes in ring 15 were closing (egging and plugging) and caused delays due to
stuck rods. A row of holes (dia.=[ 70mm)were dnlled dong the hanging wall. at a 1 meter spacing, parallel to the E holes. The plans were to blast these holes at the end of a 4-12 shifi lm ANFO 44mm DIA, > cHOLES.
- - - - pp - a) A PLAN OF ABUTMENT DESTRESSING CUT & FlLL MlNlNG
O.5m ANFO, 32mm DIA. LONG
SANOFILL
/ SECTION 0-8'
b) A PLAN OF ABUTMENT DESTRESSING UNOERCUT & FlLL MlNlNC
Fic. 3.7 a) Abutment destressing in Cut and Fil1 klining. b) Abutment destressing in
Undercut & Fil1 Mining. FRACTURE e RlNG
LEVEL
ZONE
NOT ~TOSCALE PLAN CREIGHTOM MINE 1830m LEVEL SECTION RlNG 15 RlNG HOLES O O O NOT TO SCALE HW HOLES PANEL SIZE 15m x 7.6m x 60m PANELS 3 &4 HW HOLES DRILLE0 PANELS 5 ON, FRACTURE ZONE EXTENDED BMOND THE NEXT STOPE.
FK. 3.8. Hanging wall destressing used to reduce hole closure, stuck rods and dilution in VRM stopes Mined under Stress at Creighton Mine. on a Fnday evening in order to let the ground settle on the weekend. The hopes were that the
Fracture zone on the hanging wall would reduce drillinç problerns in rings 15 to 17. Prior to these holes being blasted. a crack propagated tiom hole to hole. Drilling of the remaining holes proceeded without problems. Drilling of closely spaced HW holes was required in a subsequent ljmlong panel. Af'ter that the geometry of a 60m by 7.6m by 60m slot induced the fracture zone to extend beyond the fifth ring in the next panel. allowing holes to be drilled without a problem. Hanging wall dilution was also reduced by the clean Fracture produced along the hançing wall .
3.3 Cautions and Controls with Destressing
Destress holes cmpose a risk. for a number of them are located in the side walls that rnust be bolted. There may be non-detonated powder dur to a rnisfire. Destress holes must be considered as "bootlegs" or "sockets" as the South Afncans cal1 them. As such they must be washed out and identified by painting the collar with a circle. This task is assigned to the bolting crew for they are present at this part of the cycle. they are also at risk ofdrillinç into un-cleaned holes when bolting. Even with the availability of stick powder (cartridges) ANFO is preferred because it can be easily washed out.
It is also advisable to have the inspections of destress holes on the ground control personnel's check list when evaluating workplaces. Destressing has on occasion been considered ineffective only to find that either the holes were drilled and not blasted or drilled at an improper angle usually too far in the wall to be effective.
As mentioned elsewhere destress holes shotgun out with enough impact to tear screen. This is usually noticed when a drift is tuming. Another case in point was a destressing request to quiet an active draw point pillar on the 91 jm level at C.C. North Mine.
The crew was instmcted to drill the destress holes but not to blast them until they were inspected. This was a fortunate move because the holes were aiming at the cables canying main mine power. Needless to say the holes were re-drilled with a different orientation before blasting.
Occasionally destressing is required in the walls of drifts or stopes. Destress holes blasted. as the first holes in a round do not produce any wall or shoulder damage. This would be due to the confinement of this configuration. However. 3m long holes (dia.=j?rnm) with
0.3 meters of ANFO will blow back the screen and produce a conical-shaped Crater at the collar (Fig. 3.4). Damage of this type is viewed very negatively by the crews. who have to do the remedial work. It is important to note that reconditioning of ten such holes would take one shifi. whereas a burst along the sarne distance of wall could require considenbly longer to recondition. The damqing desuess blast occurs when the heading is clrared whereas the timing of a rockburst is unpredictable. A difficulty that anses when destressing in a specific location in an a-seismic mine is the effort and time required to convince the group from miners to management the benefit of destressing. For people to do extra work when on bonus there has to be immediate pay back. Since the effects of destressing are not always readily distinguishable. it cm be difficult to convince miners to do the extra work. Low stress levtls High stress levels
Ham jodrak subjaaî r~ lu@ insitrimesdthruRrrocLmur ainouiri;iu Oie opcning fUb by sli& in#wdirwaimuiiiarndc~of rock piau. C?#rr have ud ahdl cbrimrniypurlrrniluof~ryp of f'iin.
FIG. 3.9. Types of failure which occur in different rock masses under low and high in situ stress levels. (Afier Hoek, et al. 1995). Destress blasting to reduce primary bursts in drifts is quite successfbl in heavily jointed rock in high stress levels. The success rate decreases with decreasing number of joints. 3.4 Problems Destressing Rock With Very Widely Spaced Joints.
It is difficult to destress large (4.9mx 4.9m)drifts excavated in hard. brittie rock with a joint pattern which foms orthogonal blocks approximately 7m by Zm (Fig. 3.9). This type of ground has led to bursting during the excavation of the rmp between the 21 35m level and the 2 19511 level at Creighton. despite the drilling and blasting of the standard two face holes and two comer and two wall holes. Since the rock generally lound below the 201 0m level at Creighton cm be successfully destressed. then one could assume that the rock quality. frequency of joints and discontinuities and blast damrigr fractures al1 are significant characteristics that contribute to the destressing process.
It is postulated that. bursting could be climinated if a rock mriss with large blocks were converted by the judicious use of e.uplosives to a mass with small blocks (Le. a wdl fractured yielding drift periphery). The configuration and lenyth of the holes and the amount of explosives would require empincal tests in a controllrd test site. The configuration would have to be altered to suit the stress directions. Based on the succrss of the pattern used to develop the extraction drifts on the 2 19511 level at Creighton (O'Donnell 1997) the following configuration is proposed for destressing a drift in extremely good ground: Four face holes twice the length of the round: four standard comer holes up and out/down and out. at 30" to
* the drive direction: two wall holes on each side and an additional long wall hole on each side of the drift. as well as in the back. These hoies must be angled to locate their destressed portion beyond the bolt range of the drift (greater than 2.h) (Fig. 3.10). The holes. dthough traditionaily timed to go off al1 at once and pnor to blasting the cut. should in this complex pattern be sequenced. using electronic caps if necessary. t-w
c /' SECTION ' 'C
13 HOLES PER ROUND F FOUR 7,3m FACE HOLES W FOUR 3.6m WALL HOLES C FOUR 3.6m CORNER HOLES D THREE Sm DEEP HOLES
FIG. 3.10. Proposed destressing pattern for drifts in rock with widely spaced joints. 55
3.5 Destressing in Tabular Ore Zones in South Africa
Success of a 1957 destress program at East Rand Proprietary Mines Limited. a deep high Suess tabular gold mine. was not acknowledged till 1982 (Ortlepp 1984). However the program had been discontinued due to difficulties drillinç the long holes and loading hem with explosives.
At the Blyvoomitzicht Gold Mine in South Africa. a tabular ore body (depth
1.850rn). the design for the destress blasting in the up dip remnants (pillars). called for several 10rn long holes drilled below the ore seam. in gullies. to accommodate advance of lm per blast. The intent was to promote loosening and slip dong parting planes. Rorkr et al.
(1990) and Adams et al. (1993) summarize the effect of 57 destress blasts which contained
1 .-I km of holes. This intensive study was monitored by a fiill wave form seismic system. convergence points were measured for ride and closure. fractures were mapped using stereo photography. and ground penetrating radar was used before and atier the preconditioning blasts to quanti@ the changes. The optimums detemined for drstressing were: breast panels with holes drilled. 3 to 5 meters ahead of the face. drilled parailel to the dip and parallel to the face and sternmed for 5m using grave1 and clay. Both 76 and 89mm diameter holes were used; no preferred holr size was reported. A 29% increase in production and proof that preconditioning could be done safely were the principal results of this project.
Recent destressing trials at the 2600m level of the Western Deep Levels South Mine were successfd. Over 8000m2of reef was mined without a face burst and a 40% increase in face advance was reported (Blake et al. 1998). 3.6 Exarnples of Pillar Destressing
Rockbursts anse fiom unstable energy changes in the host rock mass of a mine and involve release of seismic energy fiom the zone of influence of the mining (Brady and BCOWTI
1994). Coats (1977) defined three conditions that are required for a burst to occur: (1) the breaking strengrh of the rock must be exceeded by the applied stress: (7) the rock must trmsform the rnajority of the work done on it by stress. into strain energy. rather than shed it by some form of plastic strain: (3) the failure must be violent brittle fracturing accompanied by the sudden release of al1 the stored elastic strain energy.
Pillar failure may affect only the imrnediate area of the pilla or. in some cases. the stability of the whole mine. For example. in a room and pillar mine a domino effect cm be induced by the failure of a pillar. which cm tdce place over a period of months or in a matter of hours (D.G. F. Hrdley formerly of Canrnet. personal communication. 198 1 ).
Individuai pillars may fail instantaneously (Fig. 3. I la) or _gadually slough (Fig. 3.1 1b).
Gradua1 sloughing usually occurs when there are numerous discontinuities at a steep angle to the pillar walls.
Pillars that burst violently may have one or more of the following properties: the rock is massive and brinle; discontinuities. if present. are perpendicular to the pillar walls and do not slide easily; there is a moderate to high level of confinement at the core of the pillar: and the transfer of stress is instantaneous due to blasting in an adjoining stope. When destressing a pillar under hi& stress there is a possibility of triggerîng failure or bursting elsewhere in the mine
Stored strain energy. which is accumuiated in a volume of rock. is dissipated on surfaces by Iiictional sliding or by extensional fracturing. Destressing reduces the stifiess of a rock and releases stored energy (Blake et al. 19%).
In sumary pillars and abutments in mines store strain rnergy as mining progresses.
The openings in rock reiease the stored strain energy as yielding occm. Destressing prornotes the release of stored strain energy by promohg slip dong newly created microhctures as well as dong pre-rxisting discontinuities.
The cut-and-fil1 mining mctthod is versatile and widely used in recovering both irregular. narrow veins and wide orebodies with irregular contacts that require selective mining to reduce dilution. The problem associated with cul-and-fil1 mining at increased depths is the creation of pillars which. as mining progresses. load-up and burst. According to Oliver et al. (1987. p 12) "The key to mining at depth is to rrcognize that failure around openings is both inevitable and desirable." Pi llars designed with littlr laterai confinement are weak and will fail under low levels of stress. Ground modification blasts soften pillar material and allows them to hi1 without bursting.
Table 3.1 lists several examples where pilla desuessing was carried out. The pillar size. quantity of explosives and the criteria used to judge success are listed to indicate the state of the art at the time the Stobie regional pillar was destressed. The principal examples are surnrnarized in this chapter. The Stobie 23 pillar blast is described in detail in Chapter
4. Frc. 3.1 1 a) Individual pillars rnay fail uistantaneousiy or b) through graduai sloughing.
Photos Erom the 2070m level at Creighton Table 3-1 Pillar Destressing
I I I I I I I I I I I I I 1 u idenis\wpfilasWdannelnprildessxls 'Decked with aggregate. "For Stobie the volumes of the destress slot and the piliar are given I~iiiarminm WIVI out a burnp or large seismic event I Mining . no sersmiaty. h-Sbng in desvess gap
Mining wiVi very lime saismiaty
1 Fractured out dia not minad I Pillar panially aesmsseu Adlacent pillar ount t year laier
2 O Mn Ount with blast mined for 2 years, Numemw I 1bunu in Gaps. Large buni al en6 on feuils 1 Inducad ptliar ciosure wiVi out inmase in swss Mining nat completea 1996 Tha nillar was not min& but analvzea efiecnveness of uesmssing 1 8 Mn wim lin on end. Mined with out a bunt destress blast is auesnonaMe Table 3.1 Data fiom significant pillar destressing projects for comparative purposes.
Mine 1 Pilbr IMinlng Yeu Pillar Sire 1 1Stemmina l Commfmb IDapth1Mahod Type 1 LIWI H Iml lmlmlm
C&F 3 6.7 AnFo 1 1 [RIW minsd wim OUI a km(r oc lvps i«anuc avant -- I 7CBF I Star 24 10 C&F 3 28 2200
I 1 1 C&F 4.5 6 90 Air 1.5 AnFo 1 245 1 0.2 I~11wmimi ~urrt
CLF 4.5 7.5 112 Air 1.5 AnFo Fracturd but 6d not mimi -CUF 3 1 20 2.5 UCBF 3 24 1415 6
-Opwi SloQ 35 36 901 '~)ecked' 4.5 Emulsion 1 3480 1 tndocsd ptUw dowe nlh wt incrasse in slrasa Mining I I 1 I I I I I I CIF 1.2 8.5 7.1 1 128 1 8 5 38 1 1 8 1 128 1 Cemenil 3 1 AnFo 1 27 1 O 2 [TM pllw was no1 mnsd but ancrlyzed efîecbvenair of I 111 I I I I I I UCdF 1 1271 1 '70 1 11 8 ~n wim lm on end ~tnedwim out a bural 1 SuMevel 120 J 550 46li2~0000l213l102l12 l11501 Air 1 21 ~Ernulsion1100451 56 A 2 4 Mn 45 diys lalu Mimng ir prograrungwilh lori Cave 1.81 21 11.711777"1 1 1 1 1 1 1 1 1 S~IU~YLTIV - 1 1 111 I I 1 I I I 1 iI\aUd.ra li! ked with ; jgregate. "For Slobi the volumes of the destress slot and the pillar are given . 3.6.1 Galena Mine 1970
Blake (1972) describes the destressing of a cut and fil1 croun pillar in the Galena mine (Coeur doAlene district. Idaho). The pillar. depth of 1.1 3Om. was 3Om long. 3rn wide and 6.7m high. ANFO was used in srnall diarneter holes. A seismic velocity survey was conducted on the pillar both before and after destressing in order to determine if the pillar was tiactured effectively. The project was deemed a success as the pillar was mined without a burnp and only low-level noises were rxperienced (Fig. 5.12).
3.6.2 Camp beil Mine 1980's
Four destress blasts were ini tiated during the 1980's at the Campbell Mine in the Red
Lake district of NW Ontario (Neumann and Makuch. 1984). Destressing of crown and siIl pillars on the 1 8 levrl. 870m below surface was the most successtùl of the four blrists. These destress holrs (d. = 14mm)of genenlly less ihan jm. were drillrd at a 1.8m spacing. to within 1.5 m of the overiy ing drift. in a 4.5m by 45m crown. Six-m long holes were drilled on a 1.8m spacing into the 25m long si11 pillar above the level (Fig. 3.13). To help stabilize the ara during and afrr the blast. the 18 level drift was reinforced with 1.8 and 2.4m long resin-grouted rebars and welded wire mesh. The stope below was filled with 10: 1 çemented tailings.
A rockburst occurred 30 seconds afler the blast and nurnerous small seismic events were recorded dunng the following few hours. There was significant darnage to the back of the 18 level drifi and 30 cm of broken rock covered the drift floor. The crown pillar was mined successfùl1y. A small bunt occurring in the pillar above the level following a Fic. 3.12 Sections illustnting the stop pillar and destressing pattern at the Galena Mine (Blake 1972).
Fic. 3.13 Layout of destress holes in the 1902 and 1802 crown and si11 pillars at the Campbell Mine (Neumann and Makuch, 1984). production blast. The fiactured ground made drilling difficult. ïhe destressed si11 pillar above the level was not mined due to the unconsolidated fil1 above it (Hedley 1992).
3.6.3 Macassa Mine 1987
The Macassa mine (Kirkland Lake. Ontario) has a history of rockbursts some of which have caused fatalities (Hanson et al. 1987). Gold bearing zones. 2 to 6rn widr and dipping at 7j0 are associated with a prominent fault and are mined between depths of 1400 and 2 150m.
Cut and fil1 crown pillars brcome burst pronr at 60% extraction when the pillar width approaches 1 8m.In 1987 the 5840 crown was destressed using O. 15Kg/m3 of ANFO in 16 to ?lm long holes (d. = 64rnm) at a 3m spacing with 2.5m of stemming (Fig. 3.1 4). To monitor the blast, waveforms were recorded. source locations were determined and convergence of 25mm was rrcorded in the stope. Subsequent mining of the crown pillar released a small burst that darnaged raise timbers and increased convergence to 33mm.
Apparently the crown pilla had been only partially destressed. A year later. a series of bursts displaced a 1000 tons and darnaged adjacent drifts. The stope below the destressed pillar was not affected.
Since 1986 the use of unconsolidated waste rock as the sole means of filling. was rrplaced by poured bacWi11. Cut-and-fil1 mining has been replaced by under-cut and fil1 in order to reduce the number of pillars created. The need to destress pilla has been almost eliminated (Hedley 1992). one zome
Fit. 3.14 a) Section showing the layout of destress holes in the 58-40 crown pillar at the Macassa Mine. b) Location of microseismic activity following the destress blast
(Hanson et al. 1987). 3.6.4 Mt. Charlotte Mine 1995
Gold minenlisation in the Mt. Charlotte mine (Kalgoorlie. Australia, Fig. 3.1 5a) is hosted in a strong. stiff and competent dolerite located near major faults (Mikula et al. 1995).
Near vertical ore bodies are 30 to 80 m thick with 200 to 300 m strike lengths (Fig. 3.1 jb).
The ore is recovered using longhole open-stoping mining methods with sequentially mass- fired. associated rib and crown pillars. Rib failures have occurred in three modes. by shearing on steeply dipping structures. violently when they have been formed progressively and gradually when formed at their final size by sloning the next stope against the future ribs. C
A large rock rib piilar. with a high risk of violent failure. was preconditioned by blasting in advancr of mining. The objective was to encourage the pillar to close and thus redistribute loads to the stiffer. stronger abuments. This \vas achieved by blasr-fracturing a thin zone of rock in the pillar. In situ stiffness was reduced. Gcological structures and small individual blocks were more likrly to move in the vicinity of the preconditioned zone than elsewhere.
Numerical modeling of several preconditioning options lrad to the decision to precondition a zone 55m high. 36m long. and 3 to 4 m wide. at the hanging wall contact of the lower section of the pillar. khveen the 900m and 95Om levels. The blast design consisted of a total of seventeen vertical blast holes (d = 14Omm. total length = Sm. Fig. 3.16a). Six pairs of pre-split holes were drilled 3rn apart at a spacing of6.hdong strike. Five decked charged blast holes were drilled between the pairs of pre-split holes. The holes were surveyed Fic. 3.15 Preconditioning at the Mt. Charlotte Mine afier Mikula et al. (1995). a)
Location of Mt. Charlotte Mine. b) Long section of Stoping Blocks. c) Plan and section indicating preconditioning zone. d) Cross section at 1 1 rib pillar. e) Blast design concept. to ensure they were within the 2% deviation typically achieved for 5Om holes at Mt.
Charlotte mine. The pre-split holes were charged with JO kg of bulk emulsion explosives placed at the midpoint of the hole and air decked (rmpty) above and below the explosives.
Four 150kg decks of emulsion. separated by J.jm of crushed rock aggregate as stemming materiai. were loaded in each blast hole. The blast holes were designed to hcture the ground between the pre-split holes in a direction parallel to the dot (perpendicular to the fractures between the pairs of pre-split holes) (Fig. 3.1 je).
The blast was monitored as follows: (1) ground vibrations were measured both underground and on surface: (2) seismic monitoring with a full wave form 33 channel PSS system; (3) stress change monitoring with CSIRO-Hi cells and vibrating wire cells; (4) extensometer monitoring to mesure dilation or shear across knoun geologic structures: and
(5) closure monitoring on the 900m level. During the mining only a few seismic events were recorded within the pilla. in contrast to numerous rvents in the stope tloor and surrounding areas.
The stress changes in an rast-west direction were +9MPa for the vibrating wire ce11 and +2 IMPa for the CSIRO Hi ceil and +2bIPa and 4MPa in the north-south direction for the two cells. respectively. Closure monitors recorded 7mm displacement across the preconditioned zone and extensometers recorded up to 2 1mm displacement dong geologicai structures. The preconditioning appeared to have ken successful as the pillar closed with out a major increase in stress or seismicity (Mikula et al. 1995). DECKEO PRESPLIT SIDEWALL OF PILLAR LINE BLASTHOLE , BLASTHOLE 1 O O O oX O O O O O
4.5m STEMMING 3.0m STEMMING BLASTHOLE PLUG
150 kg POWERGEL
4.5m STEMMING EMPT
150 kg POWERCEL
431 STEMMING PRIMERS 40 kg POWERGEL BLASTHOLE PLUG PRIMER 1 rn / 150 kg POWERGEL PRIMER 1 1
4.5m STEMMING EMPT
PRIMER 1 I1 150 kg POWERCEL PRIMER pJ
Fic. 3.16 a) Idealized blasthole layout. b) Blasthole charging design. Mer Mikula et al. (1995). 3.6.5 Sigma Mine 1996
A test was carried out in a sill pillar of an abandoned stope at a depth of 1500m (Fig.
3.17) at the Sigma Mine (NW Quebec) to determine the efficiency of pillar destressing in reducing the potential for violent failure in loaded mining structures (Labne et al. 1997). The sheared ore zone (dip =5j0). which contains qua-tourmaline veins and secondary sulfide. chloritic and carbonaceous minerais. is hosted by relatively homogeneous porphyritic diorite.
The rock is of good quality with a rock mass ratinç (RMR) of over 80 and a Q Index of 1 -6 to 3.3.
The drilling pattern consisted of fi fteen 8.5m long (d. = 38mrn) holes. dnlled at 55'
(to the horizontal) dong the vein in the drift back up to the siIl. The holes were dnlled on nvo lines 1.2m apart. one dong the footwall contact the othrr at the center of the drift. Tne holes were drilled 1.8m apart dong the line and the holes were staggered 0.9m Crorn one line to the other (Fig. 3.1 8b). The holes were loaded with ANFO For a lrngth of 5.h. and stemmed with 3m of cernent. The cernent was separated from the explosives usine a plastic cover and jute to prevent any reaction between the explosives and the cernent. The cernent sternrning extended beyond the I m deep fracture zone caused by the drift. The powder factor of the blast was 0.2 Kg/m3 .
The holes were initiated al1 at the same tirne and a 1.O Nuttley magnitude event was determined fiom the waveforms analyzed. nie cernent sternming was ejecied from the holes. the screen was peeled back at the collars of the holes and light spalling from the walls was noted. Geotornographic surveys were conducted before and after the blast. Drspite a reduction in the velocity of seismic waves in the pillar after the blast. it was not possible to calibrate the &op to a change in rock properties. Stress variations were recorded in the center and at both ends of the stope using the doorstopper method. The variations in the maximum stress were as follows: the stress at the crnter of the pillar dropped Iiom 104 MPa to 59 MPa: the ce11 on the east end recorded a drop fiom 64 to 19Mpa: whrreas the ce11 on the West end measured m increase from 3JMPa to 42MPa. Only one of three triavia1 cells remained operative after the blast. it indicated a change of I or 3 MPa of stress. Dilatometer tests were conducted in bore holes and the in situ moduius of detomarion average was 55GPa.
A decision on the efficiency of the destress blast could not be made due to the variability of the rock propenies and the difficulties in drtermining results with the instrumentation used.
3.7 Analyzing the Success of Destressing
Blake et al. ( 1998). after an extensive review of international practice. concluded that destressing of developrnent openings can be highly succrssful. However. the effectiveness of pillar destressing is more difficult to determine. It is genrrally accepted that once a pillar has been destressed mining should be able to proceed without the rockbursting hazard.
However there are only limited comparative records of mining "uith" and "without" destressing. Hence success in areas where comparative records do not exist is a very subjective rating . Fie. 3.17 a) Longitudinal section of the P Zone with 3420E Stope b) Longitudinal section of the 3420E Stope and the destressed pillar show hatched (Labrie et al. 1997). A-A' P
@ Tornographic holes 0 Walls @ Blastholes ? Bst view
a)
arnitt of 25 x 2OOmm
Iectric detoaaton
ANFO bulk explosive
Fic. 3.18 a) Destress hole pattern. b) Loading schematics at the Sigma Mine after
Labrie et al. (1 997). 3.8 Sumrnary
In summary. destressing techniques have been in use since the 1960's. These techniques have developed such that they are routinely implemented with the equipment used to drive developrnent headings. Recent tests in South -4frica have demonstrated that destressing cmincrease productivity from 30 to 40°/o. -4lthough pillar destressing requires significant planning and the involvement of men and equipment during implementation. the results are worth it in terms of reduced rockbursts. 73
Chapter 4 Destressing a Regional Pillar at Stobie Mine
The Stobie mine. (Fig. 4.1) a large tonnage. low grade deposit. is the largest producer in iNCO's Ontario Division. The mining methods are sublevel cave in the lower grade. upper portion of the mine and vertical retreat method (VRM) in the higher grade. deeper levels.
The 25 pillar ertends from surface to the 73Om level and is 190m long and 122m widr
(footwall to hanging wall).
4.1.1 Geological Setting
The Sudbury Structure is comprised of three components: the Sudbury Basin. the
Sudbury Ignrous Complex (SIC) which surrounds the basin and an outer zone of shatter-conrd and brecciated rocks (Rousell et al. 1997). The basin consists of four confomable formations that form the Whitewater Group. The main mass of the SIC consists of. from top to bottom. uanophyre. quartz gabbro. norite and contact sublayer (Fig. 1.1). Quartz diorite offset dikes C were traditionally interpreted as a component of the sublayer. but recently were linked by trace element concentrations to the felsic main mass norite. They intrude footwall rocks as ndiating and parallel dikes. (Lightfoot et al 1997). The inclusion rich sublayer. quartz diorite offset dykes and Foowall Breccia host the M-Cu-PGE ores. Archean gneisses. rni&patites.ganites and volcanics of the Superior Structural Province extend from the west. northwest. north and nonheast of the Sudbury Structure. Supercrustal rocks of die Huronian Supergroup and
Nipissing Diabase of the Southem province are found to the south. south-rast and north-east of the SIC. Ftc. 4.1 Geological map of the Sudbury Structure (after Dressler et al. 1992) Sudbury Breccia is found within 80km of the SIC in al1 of these outlying rocks
(Rousell et al 1997). This breccia and the presence of shatter cones led Deitz (1964) to suggest a rneteorite impact origin for the Sudbury Structure.
The tectonic history includes two periods of crusta1 extension and closure known as
Wilson cycles. The latest being the Grenville Orogeny. at 1000 Ma. which caused a NW-SE closure (Rousell et al. 1997).
The Frood-Stobir ore deposit occurs in the Frood-Stobie Offset Dike (Fig. 4.2). The dike consists of pods of the quartz diorite in a zone of mineralized and variably recrystalized
Sudbury Breccia. The offset. lies 2km frorn and parallel to the Main Mass norite contact. extending over a strike Irngth of I 1 km (Grant and Bite 1984).
Lower greenschist to middlr arnphibolite facies regional metmorphic grades are present in the Frood-Stobie area. Thermal metmorphism from the impact event. the SIC intrusion and the ore emplacement event are masked by the regional metarnorphism.
Okell et ai. (1979) describes the faulting in the Frood Stobie area as block style ~ith
NW-SE and NE-SW striking faults intersecting at 90'. The orebody is situated within a discontinuous envelope of breccia which slightly cross cuts the interbedded metavolcanics and metasediments of the Stobie formation. The contact between the quartz diorite body and the Stobie formation dips at 70 to 75" to the NW.
The three principal ore types at Stobie Mine are disseminated sulphide. inclusion massive sulphide and contorted schist inclusion sulphide. The Stobie orebody is made up of a low-grade disseminated ore zone above the 670m level and a hi& &grade inclusion massive ore block below the 670111level. The copper-nickel ratio is relatively constant through out the Fic. 43 a) The Frood-Stobie, Kirkwood and Mccomeil Offset Dikes (after Grant and
Bite 1984) b) Geological section of Stobie Mine (after Stobie geological staff). Stobie deposit at 0.9 to 1 .O (Anom. 1993).
The three principal ore types at Stobie Mine are disseminated sulphide. inclusion massive sulphide and contorted schist inclusion sulphide. The Stobie orebody is made up of a low-grade disseminated ore zone above the 670m level and a hi& grade inclusion massive ore block below the 670m level. The copper-nickel ratio is relativeiy constant through out the
Stobie deposit at 0.9 to 1 .O (Anom. 1993).
4.1.2 Stress Conditions at Stobie Mine
The overcoring method was used to Setermine stresses on the 730m level. and the lO3m level (Galbraith 1992). The values determined indicate that the mine openings influenced somr of the resuits. The parameters in Table 4- 1 were used for numencal modelling purposes and to calculate field stresses on various lrvels in 25 pillar (Table 4.2).
Table 4.1 Formulae from Stress Measurements at Stobie Mine
Stress Magnitude kirnuth Plunge
1 01 ( 10.9MPa + 0.0407MPa/m) 100" O
(3 (8.7Mpa + 0.0326MPa/m) --- O
03 (0.039MPdm) --- 90 Table 4.2 Calculated field stresses (MPa) in the 25 Pilla. area and Pillar stresses at 85%
extraction ratio.
Level Gl "1 '33 0s D~
275111 22MPa 18MPa 8MPa 15MPa 1OOMPa
420111 28MPa 23blPa 12MPa I8MPa 120MPa
I s50rn l 27MPa l '6MPa I
The stress on a mine pillar (q)cm be caiculated as follows.
op=osX1OO1%P
%P=100-%E
The stress perpendicular to the plane of the orebody (0,) is a vector component of o,,
a, and a, and %P is the percent of the original pillar remaining and %E is the percent
extraction.
The percent extraction of the Frood and Stobie mines down to the 55Om level is
85% (Fig. 4.3a). To calculate an average pillar stress at the center of 23 pillar. at the
175m level. the stress perpendicülar to the ore body is 15MPA. at 85% extraction ratio
the pillar stress is 100 MPa. On the j5Om level the stress perpendicular to the svike of the ore is 24.5 MPa thus with 85% extraction the average pilla stress is 163MPa (Fig. 43b). 1 1 a) SECTION
STRESS (MPo) 20 40 60 80 1pO 120 140 160 180 r 1 1 1 1 v I T 1 CALCULATEO AVERAGE Wp STOBIE MINE A 25 PlLLAR m3 GUI 85% EXTRACTION O000 O A m3 ELASTIC MOOEL Ci AVERAGE ep STOBIE MINE 25 PILLAR STRESS 85% EXTRACTION
Fit. 4.3 a) Longitudinal section of Frood and Stobie Mines indicating an extraction ratio of 85% above the 55Om level. b) Plot of stress versus depth. o,,a, & a, = Maximum, intermediate & least principal stresses respectively. a, = average stress acting on a pillar. 4.1.3 Rock strengths
The ore in 25 pillar consists of disseminated sulphidrs in quartz dionte host rock. while the hangingwall and footwall rocks are meta-sediments and meta-gabbros. The strength data determined from core (d=j4mm) specimens fiom a borehole on 27.5 section close to 25 pillar. are tabulated in Table 4.3 (O'Domell 1995).
Table 4.3 Uniaxiai and Triaxial Compressive Test Results of Rocks from BH 90264 ROCK (MW ROCK (MW TYPE TYPE Quartz 206.2 Mean = 21 2.1 220 Mean = 307.9 Diorite Il 212.6 Std-Dev.= 23.9 H 377 Std.Dev.= 80.2 Il 183.6 Maximum = 258.8 H 326.7 Il 215.9 Minirnum=179.8 1 tt 213.7 - 11 228 Metasab- Mean = 373.3 t I t 1 182.3 41 1.8 Std.Dev.= 54.4 , t 1 219.7 II 258.8 Amphib. 414 Mean = 361.5 Il 179.8 11 291 -5 Std-Dev.= 76.7 Il 232.6 1I 300 11 , 440.4 ,
TrEia~Compressive Test Results of Rocks from BH 90264 r 1 ROCK Confining Vertical pressure at failure Deviatoric stress at failure TYpe pressure ~1 (MW (1-1 (MW 1 Quartz 1 226.5 225.5 Diorite
1 Metased. 1 10 1 358 1 348 1 4.1.4 Joints
Diederichs (1997) identiiied three major joint sets in the footwall of 25 pillar on die
120m and 180m levels 89" 333". 85" 242" and "01" 068". Joint sets at deeper levels are set out in Table 4.5. The joints weaken 25 Pillar and they would becorne the loci for rockbursts.
Table 4.4 Attitudes (Dip & Dip Direction) of Joint sets identitird between 430m and
550m levels in 25 PilIar at Stobie Mine.
Level Dip & Dipdirection
J20m 88" 353" 77" 009"
465m 90" 186" 82" 360" 75" 03 1O
390m 90" 006" 88" 180" 03' 10O0
520m 88" 101" 90" 367" 89" 050"
550m 88" 279" 84" 186" Oz0 190"
4.1.5 Mining Sequence
Sublevel cave mining requires the influx of cave rock as cover to maintain a safe system. Othenvise. possible catastrophic hangingwall failure would cause an in-rush of air.
The method chosen to mine the top of the block (420m to 165m levels) and to induce caving in a controlled manner was VRM (Fig. 4.3). This was done by excavating a topsill on the
420m level and a bottom siIl. or extraction horizon on the J65m level. The general configuration was to mine VRM panels dong the hanging wall and proceed diagonally towards the foohvail and towards the south. The panels were generally 33m long, 18m wide and 46m high. with a planned 80% recovery of the broken ore once caving had started. The remaining 20 % was left as cover. Caving of the back proceeded well with only minor encouragement required derthe third stope had been mined. The encouragement consisted of drilling and blasting a few 100mm-diarneter holes in die back above the open VRM stopes. The ore outlined behveen the 420m and 465m sublevels enabled the extraction of the north half of 25 pillar. leaving a segment 90m along strike and Smfrom footwall to hanging wall. The resultant pillar had a height to width ratio of 0.43 on the 420111horizon and 0.76 on the 465m horizon. The average height to width ratio for this portion of the block was 0.6.
Failures observed as mining progressed.
As mining proceeded. cracks 3 to 6 meters apart. transverse to the pillar. drveloprd in the walls and backs of the drifts on 42Om and J65m sublrvels. They were rnapped and monitored regularly. At this time. the height to width ratio of the remaining pillar was 0.5 on 420m and 1 on the 465m level. Initially. the cracks were thought to represent tàilure along a preferred plane parailel to fabric in the rock. but splitting of the pillar as stress increased. sirnilar to the axial splitting observer in uniavial samples of core in a press is the interpretation proposed by the author. The failure and associated cracking progressed from single straight cracks in the ore and shotcrete. to the formation of loose blocks. then failure of the rebar bolts resulting in minor falls of ground. It is important to dari@ that larger falls of ground were experienced on the 420m level close to the caving front.
To ensure continual monitoring of the reaction of the pillar rock mass to mining, a novel monitoring method was developed. A band or strip of shotcrete was applied on the wdls of the drifts. This strip. which was 0.5m wide and 25 to 40mm thick. was applied a rneter above the floor and started at the entrance to the block on 465m level. The shotcrete strips were applied to specific drifts at right angles to one another to monitor movement in different directions. The response of the rock mass to mining was sasily documented by scanning the ~vallsregularly and painting and dating the cracks. A monitoring prognm using extensometers. data loggers and control points would have cost in the $50.000 to $100.000 ranges. As expected any tensile failure resulted in a crack and stress failure resulted in spalling.
The mining method used on the J90m lrvel and subsequent levels was sublevel cave.
The sublevel cave mining from 190m lrvel up to the J65m sublevel kvas started at the 2650 rock stope in the north west corner of the pillar and progressed towards the footwall. The remaining rock pillar is 90m by 90m. giving a height to width ratio of 1.
4.2 Mining Induced Seismicity
Despite the rxpected increase in occurrence of rockbursts as estraction of the deposit progressed. Stobie Mine still has a low level of seismicity and a corresponding low frequency of rockbunts. Between 1970 and 1988 an average of one rockburst per year occurred. During the years 1988 to 1998 an average of 5 rockbursts were recordrd per year with highs of 8 recorded in 1996 and 1997 (Fig. 4.5). Currently three to ten Iow level seismic events (no displacement) due to re-adjustment of the stresses after blasting, occur per day.
One rock burst per year since 1995 was associated with the marcasite filled shear on 91 5m level. Since November of 1995 four 2.4 to 2.8 Nuttli magnitude seismic events. and 4 rockbursts associated with 25 pillar have occurred (Fig. 4.6 and Table 4.5). FIG.4.1 Longitudinal section indicating the mining in 25 pillar between 420m and
550m leveis. Rockbunts per year 1970 to 1989 Year 1965 1970 1975 1980 1985 1990 1995 2000
Fic. 4.5 Histograrn indicating the number of rockbursts at Stobie Mine per year. Table 4.5 List of Seismic Events and Rockbursts Associated with 25 Pillar. See Fig.
4.5 for the location of the events.
1 No. / Diqlaced. / Mag.MNn*l
1 I A Nov. 1 7/95 2 1 :O9 0455m 25 Pillar 2.6
I 0 Nov.2 1 /95 2 I :3 8 0455m 35 Pillar 2.6
ST49 Jan. 19/96 0652 Sudace Cave area 6000Tons 2.8
ST50 Feb.28/96 1 1 :O0 550m Mn S Dr. OSTon
ST62 Nov.28/97 00:03 425111 25 Pil.Gar. 2.5Tons ST64 1 2262 X-C. 1 OSTons I 1 19:39 I 550m ST66 Jun.03/98 15:35 425m 25 Pil.Gar. lTon -
ST67 Jun.24198 07:19 580m 25 Pillar clTon 2.4
1.* Mn indicates a Nunli Magnitude, which is sirnilar to Richter Magnitude but used
for central and castern North America (Hedley 1992).
1 SElSMiC €VENTS ASSOCIATED WlTH 25 PILLAR 1 SCALE 'dm0 SEISMIC MNT OESTRESS SLOf LLTiUAk
Frc. 4.6 Rockbursts and Seismic Events Associated with 25 Pillar of Stobie Mine plotted on a Longitudinal Section This significant Ievel of seismicity associated with the initial stages of mining 25 pilla was used as an indicator of the potential of a very large seismic event or rockburst if
25 pillar should fail. Presurnably. filure would occur while mining on the 55Om level. since it was the first level on which the pillar was to be mined right across from 12 block to 3 1 block (Fig. 4.4).
4.3 Numerical Modelling
Numerical modelling was used to evaluate the various scenarios for mining 25 pillar and assess the consequences of rach mining sequence. Tlie modelling was conductrd using
Map3D, a comprehensive three-dimensional rock stability anal y sis package ( Wilcs 1996) distributed by Mine Modelling Limited. The package was usrd to construct models. analyze and then display displacements. strains. stresses and strength factors of various phases of the mining. The program fonnulation is based on the Indirect Boundary Element Mrthod and incorporates simultanrous use of both fictitious force and displacement discontinuity elements. A version of Map3D with linear elastic solutions was used. However zones with different moduli were used in some of the evaluations. The caved areas on surface and mine openings down to the 73Om level were includrd in the mode1 (Fig.4.7).
4.3.1 Stages Modelled
The rnodelling was carried out in two Stages. In stage 1. numerical modelling was used to evaluate when in the overall mine sequence 25 pillar could be rnined. The models indicated the selective mining frorn the 420m level to the 55Om level would not have adverse effect on the pillar or adjacent rocks. However if 25 pillar were to be kept intact and oniy mined from the 640m level down. the pillar would start to fail before the onset of mining.
Once the rnining on the 640m level would extend across 25 pillar an extensive failure zone is propagated both above and below the mining horizon. Thus it was of paramount importance to mine 23 pillar From the 420m level dom. Another negative impact of mining the pillar later would be the lower mine grade due to the reduction in high grade ore available to balance the mine output. If mined early the low grade 25 Pillar would becomes incremental ore. which could be mixed with a greater portion of the high grade reserves.
4.3.2 Phases Evaluated
In stage 2. a second series of numerical modeis were used to analyze the 25 pilla destressing project. Specific objectives were as follows: ( 1 ) define the level of stress and rxtent of failure induced as mining progressed: (3) evaluate the effect of drstressing the pillar: (3) evaluate the effect of mining without destressing: and (4) determine whether a destress dot could be usrd. and during which phase of the mining it should be implemented
(Fig. 4.8).
Six phases or mining were identified For assessrnent purposes: Fig. 4.9 displays simplified plans of four of them. For each phase the relevant portion of the J90m. 520m. the
55Om and 610m levels were constmcted (indicated as mined) in the model. (See dso Fig.
4.4.)
The destress slot represented in the numericd model by three methods: (1) a 2m thick slot with weakened rock. (a, of 10% of the intact value). (2) a 0.311 thick vertical mined out zone and (3) a 2m thick mined out zone. Phases without the destress slot were modelled for comparative purposes (Fig.4.10). When dealing with mine-wide geometry it is often not practical to look at individual openings For detail. Given that the destress slot is
an extremely narrow dit. it was necessary to incrrase the size of the dot in the numerical
mode1 beyond its physical dimensions to ensure that the numerical results were reasonably
correct. It is well known that thin elements or elements with extreme aspect ratios are
numerically inaccurate. Hence to over corne this problem the destress slot was analyzed as
a O.3m and 2.0m thick dot in the numerical model.
4.3.3 Strength Pararneters/Failure Criterion
Hoek-Brown strength panmeters were used to determine safetp factors (Hoek et al.
1 9%). The modi fied Hoek-Brown failure criterion for intact rock. an empirically derived
relationship may be expressed in the following form:
where m, is the value of the constant m for the rock mass.
s and ïr are constants which depend upon the characteristics of the rock mass
oc is the uniaxial compressive strength of the intact rock piecrs and
o., and o are the axial and confining effective principal stresses
respectively.
The values used were a tende strength of O. a uniaxial compressive strength of 155
MPa. m equal to 1 .J and s = 0.01.
The stress regime used was a, = 10.9MPa+O.O407MPa/m S. o2 = 8.7MPa +
0.0326MPalm S. and 0,= 0.029MPdm S where S = depth below surface in meten. The trend of the principle stress is 100 degrees. The plunge used was O degrees. stop 1 Stobie Mine - 2S PIUAR Eualuation Destressing Progran SSSPDE62/11/9? t I
FIG. 4.7 Plot of the numerical mode1 geometry (MAP3D used) indicating the openings brtwern surface and the 73Om level.
FIG. 4.8 Plot of the destress slot area. 490m LEVEL 52Om LEVEL 550m LEVEL
PHASE 1
b) PHASE 2
C) PHASE 3
d) PHASE 5 MINING IN 25 PILLAR DESTRESS SLOT- amsi eza
Fie. 4.9.Plans indicating the opening included in each phase modelled. a) Phase 1. J90m level66% mined, 520m 25% mined, 550m rock stopes, 610m level30 block 50% mined. b) Phase 2,490m level 100% mined, 520m level50% mined, 550m 25% mined. Conditions prior to blasting the destress dot. c) Phase 3,
49Om level 100% mined, 520m 50% mined, S5Om 35% mined, Destress dot blasted. d) Phase 5,490m level
100% mined, 520m 80% mined, 550m 50% mined, no demess slot. Note Phases 4 and 6 and 6lOm level are omitted in this figure. 4.3.4 Model lndicating the Effect of the Destress Slot.
A list of the models analyzed in stage 2, detailed summaries of individuai models and the key findings are contained in Appendix B. Al1 of the electronic data for stages 1 and 2 of the modelling is archived on the Stobie Mine netsvork. Hard copies of the o,.o, and safety factor plots and summary reports are on file at the mine.
Model S3PD29I indicates the rffect of the destress slot very well. The destress slot is represented in this mode1 as a 1.8m wide excavated slot and results are plotted on a vertical erid through the pillar. The o,plot indicates a band of reduced stress through the pillar. High C stress zones are located above and below the destress slot (Fig 4.10). The o;plot indicates a zone in tension on the HW and FW of the dot. Zones of increased confinement are located above and below the slot (Fig 4.1 1). The safety factor plot indicates the core of the pillar above and below the slot has failed (Fig 4.17).
4.3.4 Summary of the Modeling Results
Modeling indicated that failure would begin with the removal of between 25 and
50% of the 5jOm level. Also. the central portion of the pillar. above and below the destress slot (core of the pillar). would either fail early in the mining history with the blasting of the destress slot. or later in the mining. as the extraction of the 55Om to 52Om level block proceeds. In surnrnary. destressing was required to promote timely failure of the core of 25 pillar. (MPa) al (Psi)
t
Fic. 4.10 A G,plot on a vertical _gid through the pillar in Mode1 S25PD29I in which the destress slot is represented as a 1.8m wide excavated dot. FIG.4.11 A a, plot on a vertical grid through the pilla in Mode1 S25PD29I in which the destress dot is represented as a 1.8m wide excavated slot. Fic. 4.12 A safety factor plot on a vertical grid through the pillar in Mode1 S2jPD29I in which the destress dot is represented as a 1.8m wide excavated dot. 4.4 Explosives and Blast Design
The objective of destressing 25 pillar was to: promotr failure of the pillar dong the dot. induce failure above and below the slot at the time of or shortly afier the destress blast. and to produce a failure zone dong which mining could progress without building up stress. nie following stcps were taken to promote these objectives: (1) Mining wsdirected to induce the highest level of stress dong the destress slot and (7) Two sets of fan holes were located on the ends of the destress slot. to provide more of an opening for the destress blast to break to.
If there were preferences. a destress blast would in itself producr muck. and would be scheduled or designed before stresses are elevated (before the pillar becomcs a pillar) which wouid make it a preconditioning blast. Consideration could not be given to designing a preconditioning blast because in the Stobie case the pillar pre-existed the drcision to recover a portion of it. However. if there is not a high lrvel of stress. the blast will not have the potential to fracture and condition the ground to the same rxtent. Mining sequences require the mining to start at a fault or opening (slot) and progress away from it.
4.4.1 Configuration of the Destress Holes
The destress slot contained two parallel rows of holes. which are also parallel to the strike of the hanging wall contact. Two fans of sublevel cave rings were placed on each end of the destress slot. one against the rock stope on the north and the other against 20 block on the south (Figs. 4.14.4.15 and 4.16). The fans and rings of holes. were designed to: (1) create a positive opening for the first few destress holes to break into; (2) reduce the length of the STRESS CELL & POINT #1 POlNT #2
STRESS FROM NUM. MODEL AT GRID POINT #1 53 30 17 MPa STRESS FROM NUM. MOOEL AT GRlD POlNT #2 56 22 16 MPo
550m LEVEL PHASE 2
STRESS FROM NUM. MODEL AT GUID POlNT #1 14 6 2 MPo STRESS FROM NUM. MOOEL AT GRID POlNT #2 37 23 16 MPo CHANGE IN STRESS FROM NUM. MODEL AT GRID POlNT #1 (STRESS CELL) 0ETWEEN PHASE 2 & 3 -39 -24 -15 MPo POINT #2 BETWEEN PHASE 2 & 3 -8 +3 +1 MPo RECORDED CHANGE IN STRESS CELL -8 -3 -2 MP~ 550m LEVEL PHASE 3 MINING IN 25 PlLLAR DESTRESS SLOT-
Fie. 4.13 Plan of S5Om level with sirnplified geometry of Phases 2 and 3 indicating stress regime fiom numencal mode1 and stress change determined by stress cell. destress dot and (3) concentrate the stress on the slot. The fans (Fig. 4.15) are typical wagon wheel fans with rings inclined above the horizontal at 50". 65". and 70" grading into 80" rings. The fans to the north and south of the destress slot were blasted together on the first blast and the 80" rings were blasted altematively on the subsequent blasts. It is important to note that the rings were not blasted till al1 the destress holrs were loaded with emulsion. This was to ensure that if a stress or seismic problem developed the holes could be primed and wired quickly.
The holes were 2 1.3 rneters long by 102 mm in diameter. The rows were designed to be 1.8 m apart but ranged from 1.5 to 1.8m due to variations in drill capabilities. The rows followed the destress drift and essentially were panllel to the hanging wall contact joining the southeast corner (HW) of the 2340 rock stope and the northeast corner of 20 Block. The holes were inclinrd at 85" from the horizontal. There were differences in the spacing of the holes in the hanging wall row of destress holes. The holes on the north half of the drift were spaced 2.4 meters apart and the holrs on the south half were spaced 1.8m apart.
The holes in the row on the footwall sidr of the drift were drilled 0.61m apart and every second hole was loaded and blasted. These holes were the pnrnary destress holes. The empty holes were buffer holes and were to induce stress and promote fracturing. The holes in the hanging wall row were designed to extend the fracture zone created by the primary destress holes.
4.4.2 Explosives
The emulsion used in the destress blast was the same product used in regular sublevel SCALE
Fic. 1.14. Plan of 550m level indicating the geometry of Phase 3 of the numerical model. This geometry coincides with the position of the mining front when the destress blast was taken. Fic. 4.15 Longitudinal section of the destress slot indicating the fan rings on the
North and South ends. The general attitudes of the destress holes are indicated. Not al1 the rings are plotted as the holes were on a spacing of 0.6rn. cave blast rings at Stobie Mine. The drills. loading equipment and explosives used were the standard used in the mine on a daily basis. Dpo Nobel Ltd. manufactured the ernulsion wd under the trade name of RUS Emulsion and its properties are listed in Table 4.6.
Table 4.6 Properties of RUS Emulsion manufactured by Dyno Nobel Ltd.
Energy 690 cal/gm or 860 cakc
Relative Weight Strength 0.78 - Relative Bulk Strength 1.19
Vrlocity of Detonation 1 5200 m/sec
Detonation Pressure 85 Kbars
Gas Volume 4 1 .7 Moleskg
Water Resistance 1 Excellent
Minimum Diameter
Fume Class t
The holes were loaded with a toe loader. The emulsion was purnped through a wand that was inserted up the hoie to the toe and retracted towards the collar as the emulsion was pumped. The holes were designed to be loaded to 1 .jm and 3m fiom the collar on alternate holes. In the field. most holes were loaded to 3.lm from the collar. A total of 10,045 kilograrns of explosives was loaded in the destress holes. The rings on the south side and north side had 12,575kg and 10.900 kg of emulsions respectively. 4 --. -4.5*I m
#-* 21 70 OESTRESS ORlfT a#-- -- wu- -#* r
40~~t- DISTANCE BETWEEN HOLES
DETONATION SEQUENCE
HOLE LOADED WlTH EMULSION O BUFFER HOLE NOT LOADED
4 DETONATION SEQUENCE
FIG.4.16 Simplified plan of 2170 Destress drift on 55Om Ievel indicating the hole spacing and detonation sequence. FIG.4.17 a) Boosters with Nonel delay detonators which were used to prime the emulsion Ioaded destress holes. b) Blaster and Dyno Nobel Technicd representative placing the boosters at the correct distance up the hole with a fiberglass rod. FIG.4.18 a) Back of the destress dot facing North. Holes are Ioaded and primed. b) Loaded and primed holes in the back of the destress dot, view facing South. 4.4.3 Timing of Holes and Kilograms of Explosives per Delay
Due to the number of holes in the destress blast. both Nonel Unidets (short period detonators) and Nonel long period detonators were usrd to ensure that al1 the cap delays would have been initiated before the first hole would detonate (Werely 1998). The holes were double primed at 1O.7m and 7.2m from the collar (Figs. 4- 1 7a. b and 4.1 8a. b). The short period detonators used were. 1-22.24.26.18. 30.32. 36.40. The timing for the short period detonators is delay# x 2jms (milliseconds). thus a number 32 would be 32x3ms =
800ms (#40 =1000ms). The Long Period (LP) detonators used. ranged fiom a X3LP
(1 100ms) and ended with #17LP (7700ms). The longest delay a number 17 would have extended the duration of the blast to 7.2 seconds. Since two holes on the foohvall and 3 holes on the hanging wall were crushed and were not successfully initiated. the blast only lasted
4.9 seconds. Emulsion was observed after the blast at the collars of the crushed holes on the
South end of the destress slot. Two delays had 564g of explosives: the remainder had 1 87kg of explosives per delay.
3.5 Instrumentation
4.5.1 Stress Monitoring
A CSIRO HI stress ce11 (Council for Scientific & Industrial Research Organization of Australia Hollow Inclusion Cell) was installed February 1 7Lh.1998 in a 10.b long hole ddled in the back of 21 85 X cut (Figs. 4.13.4.23 and 4.24). The hole was inclined at 85" towards the destress slot and the cell was located 14 m from the slot. The ce11 was wired to a multiplexer and data logger on February 25th. Readings were taken every hom. The data logger was dom loaded twice a week.
4.5.2 Microseismic System
An Electrolab MPXO microseismic system has been used at Stobie since 1988 to monitor strength and location of seismic events. This system does not capture waveforms or plot wave traces of the blasts or events recorded. Active areas cmbe identified and the çvent frequency and density (number of events per hour and volume) cmbe determined. The decay tirne aftrr a major event is also an important parameter.
4.5.3 Biast Monitor
Explotech Engineering Ltd. monitored vibrations at three locations with three
Instantel digital seismographs. rquipped with triaxial çeophones (Fig. 4.1 9). The geophones have a Peak Particle velocity range of up to 254mmls and a frequency response of 2-30 Hz. in addition to the digital seismograph a microphone with ri Peak Sound Pressure Level
(PSPL) range of up to 1J2dB was used to record overpressure at the site in a sub-division
1.8km from the mine (Corkery 1998).
4.5.4 Observations Before and After the Blast
Stress had been evident on the 55Om level fiom September of 1997 when the "in the hole" (ITH) holes of 2260 rock stope (fiom 5jOm to 610m level) started dog-earing. The formation of these borehole breakouts in production holes. in the 25 pilla. at Stobie mine signalled the advent of failure and the need for destressing (Fig. 2.8). The presence of egging in drill holes is used as an hdicator that the stress level is sufficient for primary destressing to be effective. Considenble squeezing and sloughing with associated seismicity occurred during the blasting of the sublevel cave rings in the minor pillar between 2340 rock stope
(55Om to 565rn level) and 2530 rock stope (490m to 55Om level) during the months of
Januq and February of 1998. Sloughing occurred in the walls of 2260 rock stope at the
55Om level topsill in late February of 1998. Cracking and shear failure of the shotcrete and egging of the destress holes in the south end of the destress slot were noted on the 1' of April
1998. On May 4Ih additional egging was noted in the destress holes at rings 2066 and 2077.
The last ring on the South side that was double primed. as per the design. was ring 2079.
Rings 2066 and 2077 were collar-primed. On May 4'. Rockburst ST-64 displaced 1 ton from the back of 7262 driti on the 55Om lrvel. On May Yh spalling was noted in the shotcrete at
2080 orepass drift on the 55Om level. The sequence of rvents described above indicates the increase in stress as mining progressed and the release of stored strain energy that occurred wiith the Failure around the openings. The timr to blast the drstress slot was close at hand.
The destress blasr was detonated at l4:Q on May 9Ih1998. Al1 three adjoining mines
(Stobie. Little Stobie and Frood) were cleared for the blast in case a Richter (Nuttli) magnitude rockbunt should occur. Access to 490.520.550 and 580 m levels was resuicted tiil 7:00 the day following the blast. Work was performed on the bels below 590m during the afternoon shift of the 9Ih after the smoke and fumes cleared.
Conditions in the destress slot were reviewed on May 10Ih. There was no damage to any of the mine openings outside of the destress slot (Fig. 4.20). The back of the destress slot was cratered up to twelve feet above the original back. The muck was relatively fine with occasional pieces having the longest dimension up to 0.5 meter.
The stress ce11 rneasured changes induced by blasts taken on the Zorn. 550 and 610m levels. As expected the destress blast had a marked effect compared with the other
blasts (Fig. 4.23). The slope of the plot of the values is an expression of the creep of the
epoxy glue between the rock and the crll.
The changes in stress calculated from the strain changes recorded by the stress ce11
between Fndg May 8Ih the day before the blast and Tuesday May 12Ih. 3 days after the blast
are contained in Table 4.7.
The minor stress drop cm be compared to the values calculated using the numerical
modeling of phases 2 and 3 (Fig. 4.13). While a large drop in stress would be desirable the
failure of the core of the pillar was considered a primary goal of the blast. The mode1
indicated the destress blast would promote failure of the core of the pillar above and below the destress dot. however it is difficult to assess this in the tield. The 1.4 Nuttli Magnitude ment that occurred at 07: 15 on June 24th in the north wall of 3 pillar on 1900 level is viewed as a significant event in the expected continued failure of 25 pillar.
Table 4.7 Changes in stress caused by the destress blast rneasured by the stress ceil in 2185 X-cut on the 550m level.
Principal Stress MPa Di p Bearing
61 -8.1 29 O 234 "
% +3.8 57 O 23 O
b3 +1.1 14 O 135.6 I Figure 4.13 indicates the stress regime in phases 2 and 3. calculated at two grid points, one at the location of the stress cell. and the other 18m to the footwall. at the same elevation as the stress cell and on the same section. The stress drop indicated by the stress ceIl is also plotted.
The destress blast \vas classed as a mild blast with minimal vibrations felt and recorded on surface. A strong reaction was expected on surface due to the confinement of the explosives (lack of free face) in the centrr of the destress dot and the high stress the pillar was subjected to. which would promote better transmission of the shock waves from the blast. A greater that 7 Nunli magnitude (Mn) seismic event was anticipated within 24 hours of the blast as the pillar failrd and released its stored strain energy. Neithrr a large shock wave nor a seismic rvent was associated rvith the blast. The low kilograms of explosives per delay producrd acceptable Peak Particle Vclocity (PPV) on surface and at the garage on the
550m level. 150m fiom the destress dot. A 3.4 hiln seismic event occurred at 07: 19 on June
24Ih on the 58Om level dong the north wall of 75 pillar (Figs. 4.6 and 4.2).
The destress blast was recorded over five "event-windows" by the microseismic system due to its long duration. Thirty-four events were recorded over the two days following the blast. On the third day afier the blast only 3 events were recorded. By comparison 3 1 events were recorded on the 24 Ih of June. the day the 2.4 Mn seismic event occurred on the 380m level.
There were an average of 17 seismic events recorded per day during the month of
April and an average of 12 and 1 1 per day during the months of May and June respectively.
By comparison the month of October had 2.4 events per day. The data recorded during the July vacation shut dom indicated an increase in activity in the areas of the marcasite shear
on the 91jm level but no activity in the 23 pillar area. The events recorded during a
production shut dom are a good indicator of the areas in the mine. which are eidier retuming
to their naturai stress strtte. or are naturally active without mining induced stress. One could
postulate that the destress blast and the 2.4Mn event transferred stress to the abutmrnt dong
the rnarcasite shear. The recorded vibrations and overpressure induced by the blast are
tabulated in Table 4.8.
Table 4.8 Peak Particle Velocities and Peak Sound Pressure Levels recorded of the
May 9'hdestress BIast (Corkrry 1998).
Location PPV (rnm/s) PSPL (dB)
1800 Level Garage 10.03 NIA
Il Eng. Equipment Vault <1 .O0 N/A
100 Village Cr. 0.25 88
1.6 Interpretation of Results
A complete evaluation of the success of the destress blast can only be properly made
once the 55Om level is completely mined. A preliminary evaluation indicates the destress
blast was a success. based on the stress drop recorded by the stress cell. the reduction in daily
seismicity through out the mine and the absence of any new stress related problrrns on the
550m level since the destress blast (January 1999).
The occurrence of a 2.4Mn rockburst on June 24'' 1998 can be viewed as a
significant event in the expected, continued failure of 25 pillar. - I Loi 1 --î
FIG. 4-19 Plot of the wave form of the destress blast recorded by Explotech
Engineering Ltd. (Corkery 1998) FIG.4.20 a) View of the desness dot from the entrance drift. View facing towards the hanging wall. Note there is approximately 6 feet of ore in the slot. b) Close up view of the muck pile in the destress slot. FIG.4.21 a) View of the back of the destress slot facing North. b) Close up of the back facing south where the rings did not detonate. wrnaow 1 - tvenr I 199846-24-1 1 :Il:35.6580
Fit. 4.22 Plot of the traces captured by the Eastern Canada seismological network of the 2.4 Mn event on June 24" 1998 (CANMET 1998). CUMMULATIVE STRAlN CHANGE 25 OESTRESS SLOT
- -- -. . -Gauge 6 -Gauge 3 Gauge 4 I -Gauge 5 2 RINGS SdUW -Ref 120 STAR ' OF DESTRESS ' "ORrn ROT FAN MNGS -Gauge 1 ME0NORTH h MUTH -Gauge 2 8 RINGS SCUTH I -Gauge 7
JUUAN DATES
Fie. 4.23 Plot of the suain change data capnired by the stress ce11 and data logger.
Note the corresponding blasts are indicated. FI~.4.21 Photo of a CSIRO (Australia) hollow inclusion stress ce11 Chapter 5 Summary and Conclusions
Rockbursts are the result of the sudden release of stored strain energy derived from the stresses induced by mining at depth or to a high extraction ratio. Control methods range frorn modi@ing geometries to altering the rock properties by destressing. The physical expression of the release of stored strain energy includes dog-earing. core disking. slabbing, pillar failure and bursting. A review of the historical development and a comprehensive lis1 of the various methods of destressing are contained in this thesis. Special interest is paid to destressing of pil1;ü.s. culminating in the description of the drstressing of the regional pillar at Stobie Mine. which the author designed and directed.
The significant milestones of this project are: (1 ) mining was directed to induce stress along the destress slot and the pillar was succcssfully brought to a point that failure was eminent: (2) over 10.000 kg of rsplosives was used to destress the pillar without damaging the mine infrastructure and (3) the drstress slot. which was designed to promote failure in the core of the pillar above and below the dot provided a zone from which mining could safely proceed.
The risk of bursting has been reduced for the extraction of 1.8 million tons of ore between the 550 and 520-m levels. Another 5.6 million tons of ore is available if mining proceeds io the 64Om level. It is recommended that the seismicity induced as mining proceeds in 25 pillar be reviewed in the final evaluation of this project once the ore on the
550m level has been mined. Appendix A Time Line of the Project.
1992 Silling and developing on 420m level
January to June 1993 Modelling of mining 25 Pillar
June 1994 First VRM Production blast between 420m and 465m levels
August to October 1994 Three reportable falls of ground
Januay to June 1997 Numerical Modelling of destress dot
June 1996 Joint Preliminq Report on Destressing 25 Pillar Preston and O'Do~ell.
February 1997 Approval From MC0 to Use the destressing project as a Thrsis
September 8th 1997 Photographed ITH holrs egging in 2260 rock stopc on 55Om
Nov. 38 1997 Rockburst X 62 on 425m Irvrl in 35 Pillar Garage.
Feb 1 7/98 Stress ce11 installed.
Feb. 25 stress ceIl wired.
March 8th/98 Walls slouging in 1360 rock stope.
May 4th Last 6 rings on the South side fan blasted.
May 4th Rockburst $64 on Som level in 2262 X-cut
May 4th increase in the deformation (egging) of the holes in the south end of the
destress slot at rings 2066 to 2077, the lnst ring primed is 2079.
May 9th Destress Blast
June 24th Major seismic event of 2.4 Mn. Magnitude at 07:l-l on 580m level
elev. at 2490 East & 1880 North
December 1998 No new stress related detenoration noted in 25 pillar on the 520 and
55Om levels. Appendix B Numerical Analysis.
The purpose of the numerical modeling analysis of the 25 pillar destressing project is to 1) determine the stress magnitudes and estent of failure as mining progresses. 2) evaluate the effect of destressing the pillar and 3) evaluate the effect of mining without destressing. Al1 numerical analyses were camed out with the three-dimensional boundary element numerical program MAP3 D (Wiles 1996).
B.l Description of the Mining Phases Modeled:
The stress regime used was G, = IO.OMPa+0.0407MPdm S. 0, = 8.7MPa +
0.0326MPdm S. and flj = 0.029MPdm S where S = drpth below surface in rneten. The trend of the principle stress is 100 degrees. The plunge used was O drgrrrs. Hoek Brown strength parameters were used to determine saîèty factors. The values used were a tensile strength of
O. a unimial compressive strength of 1 5jMPa. m equal to 1.4 and s = 0.0 1 .
Five phases or stages of the mining process have been identified for assessrnent purposes. The first phase represents the mining grorneûy to Janu.; of 1997 at which time the following was rnined: (1) 66% of 25 block on 490m level was mined (sublevel cave): (2)
25% of 520m level: (3) 100% of the 55Om lrvei rock stopes and (4) 50% of 30 block on
6 10m level were mined. Figure 4.9a shows this initial phase of the mining in 25 block.
Phase 1 of the mining was evaluated using models S25PDESl to 4. The efiect of the georneûy on the stress regime and the extent of failure in 25 Pillar are displqed on 4 grids.
Grid # I is a vertical transverse section frorn 38Om level to 6 10m level along 2 140 section. this is referred to as view SXPSTRN. Grid #2 is a vertical longitudinal section along 1060 line fiom 380m level to 610m level and is Iabeled view S2SPSLNG. Grid #3 is a horizontal grid at 5201-11 level, the view is from the bottom up. with the footwall at the top of the page.
This view is referred to as S25PPUDR. Gnd #J is aiso a horizontal grid. it is located at 55Om level and is viewed from the bottom looking up. the same view as used for grid #3.
Phase 2 of the mining represents the geometry of mining at Aprii 1997 with 100% of 490m 1eveI mined. 50% of 52Om level mined. and 25% of SSOm mined. Models
S25PDESj to 8 evaluate the geornetry of phase two.
The mining conditions projected for June 1997 are representrd by Phase 3. which was modeled by the runs S2jPDES9. S2ZPDElO to 12. The geometry of phase 3 represents
100% mined on 490m level. 50% of 52Om levrl. and 25% of 55Om level. Mining in this phase includes a vertical 6' thick soti zone from j20m to 5ZOm level joining 12 block to the sublevel cave front. representing the destress dot.
Phase 4 has 100% of -F9Orn mined. 80% of SXm 50% of 550m level and the 6 Z Om level foot wall drift rnined. A 6' wide softened zone represents the destress slot. The models evaluating phase J are S25PDE 13 to 16.
Phase 5 has the rame mining configurations as phase 4 but without the destress slot. it is evaluated by models S25PDE17 to 20.
Models S25PDE29 to 32 evaluate the geometry of phase 3 with a 1' thick slot to simulate the destress slot. Models 33 to 36 have the geometry of phase 3 without the softened
6' thick slot of models 39 to 32 nor the 1' thick slot of modeis 9 to 12. B.2 Summary of Key Findings
The models SEPDESI to 4 indicate that phase 1 is a stable configuration with an average of 60MPa 0,stress in the core of the pillar. The confining stress sigma 3 is about
13.8MPa. The strength factor used is the most appropriate to use when assessing pillar
failure. it is the factor that gives the smallest value thus it is the most conservative. The strength factor is defined in terms of the o,strength divided by the current stress assuming a constant value of sigma 3. The analysis indicated a short to long term stability in the core of the pillar. a safety factor of 1 -3 to 1.6. Failure is limited to the periphery of 17 block and
30 block.
Phase 2 was evaluated with models S25PDESj to S. They indicate a drop in
confinement in the core of the piilar from an average of 13.8MPa in phase 1 to 9.6MPa in
phase 2. This drop in confinement reduces the stability of the pillar and an increase in the
rxtent of the failure zone is ctvident in the plots cornparrd to phase 1 plots.
Models S25PDE33 to 36 evaluate the state of phase 3 (without destressing). The
plots indicate little if any increase in the principlr stress in the core of the pillar but a further
reduction in confinement to values as low as 4.8MPa (plot S25DUS3). A reduction in pillar
strength in phase 3 follows the reduction in confinement. The edge of the mining on 55Om
level induces a rise in confinement just below 550m level as well as an increase in
confinement to above 20.6MPa in the hanging wall and footwall. This local increase in confinement has an associated increase in strength at the 55Om level horizon.
Models S25PDES9 and S25PDE10 to 12 represent phase 3 with a 1.8m thick destress
slot sirnulated by reduced rock strengths of 25MPa pillar strength and an angle of intemal 122 friction of 25 degrees. tension as usual is 0. Mode1 SEPDE13 to 16 represents phase 4 with a destress slot of similar properties. Modelling the destress zone in this fashion has not shown any change in the results. In comparing models SZSPDE 13 to 16 (Le. phase 4 with a sofiened destress dot) with SZ5PDE 17 to 10 (Phase 4 without destressing) there is no difference in the results. This would indicate the material properties assigned are not low enough to effect the system.
A cornparison of the results of phases 3 and 4 without destressing indicates very linle change in confinement and extent of failure caused by the increase in mining in phase 4.
A destress dot in phase 3 was simulatrd by creating a O.3m thick dot (mined not sofiened) tiom 52Om to SjOm level in models S2jPDE29 to 33. In comparing these rnodels to S25PDE33 to 36 (phase 3 without destressing) the destress slot causes an increase in o, up to 13.8MPa to the footwall of the destress dot. A signiticant reduction in o3and a tensile zone occurs to the hanging wall of the slot. This causes the two hilure zones in mode1 33 to join into one continuous zone.
In cornparhg models 3 1 and 33 with 35 and 36. the destress dot greatly extends the zone failure zone through the pillar.
Models S2jPDE37 to 40 are copies of models 13 to 16. they were analyzed to evaluate whether a 1.8m wide sofiened zone with a unimial compressive strength of 1 SMPa
( 10% of the intact value) and m = 0.0 14 and s = 0.000 1 could simulate a destress slot. There is absolutely no difference between the results of models 13 to 16 and the results of 37 to 40.
To evaluate mining without a destress dot models S25PDE4 1 to 44 were analyzed.
They have 75% of 550m level mined. To evaluate the conditions in the center of the remaining pillar the vertical transverse grid was moved to 3080 section. In comparing the results of models 41 to 44 with models 17 to 20. which have 50% of 5Om level mined. an increase in 0,is noted. The a; plot indicates an increase in confinement in the hanging wall of model 4 1 and a decrease in the confinement in the core of the pillar. A large increase in the extent of the failure zone is indicated by plot S25D4 1SF. The increase in failure zone is not as evident on the long section plot S25D42SF nor on the plans at 520m and 55Om leveis
Plots S25D43SF and S25D44SF respectively. It cmbe surmisrd that the pillar has îàiled when 50 % of 55Om level has been minrd. This is why the destress dot is pianned to bc blasted when the rnining on 5Wrn is confined to hanging wall area and only 35% of j5Om
Ievel bas been mined.
B.3 ConcIusion
The purpose of the destrrss slot was to promotc a tirnely tàilure of 35 pillar and reducr the risk of injury by controlliny the failure with blasting the destress dot. Numerical modeling and comparing the changes in the sxtent of srress and failure zones as rnining progresses provided an indication of early tàilure pnor to destressing. This early failure may result from the safety factor used. There have been no previous cases that could be used to calibrate the model. Assuming that failure will start prior to rnining the destress dot. there is still a need to destress since the core of the pillar does not fail till the destress slot is taken.
In summary. the modeling indicates a need to destress 23 block to promote a tirnely failure of the core of the pillar.
C 'DNIS\WPF~LESLVL~IODC~~I~~~rpt 124
Table B.1 List of Runs and Descriptions of Phases Modelled. Mining Comments I Phase 1 Phase 1 Conditions at January 1997 490m level213 mined, 520m 114 mined, 550 rock stopes, 610m level3O block mined. Phase 2 Conditions at April l997J9Om level 100% mined, 520m 50% mined, 550m I/J mined. Phase 3 Conditions at June 1997490m level 100% rnined. 520m 504'0 mined, 550m 3506 mined, Destress slot 1.8m wide type 3 parameten 25 MPa pillar strength, 25 degree angle of 1 interna1 friction. 1 Phase 4 49Om 100% mined, 530111 80% rnined, 550m 50Y0 mined, 6 10m level footwall mined. Destress slot 1.8m wide type 3 parameters 25 MPa pillar strength, 25 degree angle of internal friction. .. 490m 100% mined. 530m 8096 mined. 55Om 50°h mined. no dest ress dot. 1 *. Copies of models 9- 12. J90m t 00% mined. 520m 509'0 mined. 55Om 25% mined. destress dot rnined 1.8m thick no softening parameters. .. Copies of models 9- 12.4901~1100% mined. 520m 50°h mined. 550m 2596 mined, destress dot mined 0.3m thick no softening parameters. Slot has a twist, is non planar. Runs Copies of models 9- 13. J90m IOO% mined. 52Om 509.o mined. 55Om 25% mined. destress dot mined 0.3m thick no 1 softening parameters. Slot has no twist, is planar. .. Copies of modeIs 9- 12. J90m 10094 mined. Zorn 5090 mined. 55Om 35% mined. no destress slot. .. Copies of models 13 - 16. J90rn level 100°'a mined. 520m 809'0 mined. 550rn 5096 mined. 6 10m level footwall rnined. 1 Destress slot 1 .8m w ide type 3 parameters uniaxial compressive strength of 15.5MPa m = 0,014 and s = 0.000 1. .. Copies of models 17 - 30.490m 10094~mined. SlOm 80% mined. 550m 75% mined, no destress slot. 1
Additional Runs Dec. 22/98
As a result of a revicw of the models by Drs. Derek Martin and D. McCreath
additional models were run to attempt to define or elirninate the stress riser visible in models
S35PDE29 to 3 1.
The problem was partially eliminated by adding joiner plates between the destress slot and the large mining blocks. However it required a wider destress slot to eliminate the problem. With a O.3m \vide slot. the image of the destress slot was transparent when viewed frorn the foohvall. When viewed from the hanging wall it was opaque. Increasing the destress slot to a 1.8m width eliminated both the transparency problem and the stress riser. Models
S25PD32H and S25PD321 were used to plot the stress values at two points. Models
SXPD291 to 321 were developed to replace btodels SD25PDE29 to 3 1.
Model S25PD29I has a vertical gnd through the pillar. The o,plot indicates a band of reduced stress through the pillar. High stress zones are located above and below the destress dot. The ojplot indicates a zone in tension on the HW and FW of the slot. Zones of increased confinement are located above and below the slot. The Safety factor plot indicates the core of the pillar above and below the slot has failed.
Model SXPD321 has a horizontal grid through the destress dot at 533m level. The a,plot indicates the stress level through the pillar is at the background level. The o3plot indicatrs tension dong the destress dot. The Safety factor plot indicates a zone !O the foot~vallof the destress slot which has not failed. This is due to the increase in conilnement in this area.
Model S25PD3 11 has a horizontal grid just above the destress dot at 5ZOm level. At the destress slot the grid cuts the stress riser above the dot. The Strength factor plot indicates failure through the core of the pillar.
Model S25PD3OI has a vertical grid just to the footwall of the destress dot. Failure is indicated across the pillar along the top half of the destress slot. The zone with a strength factor greater than 1.6 that coincides with an increase in confinement is indicated at the bottom half of the destress dot.
In summary models S25PD29I to 321 indicate the destress slot causes a &op in stress in 25 pillar between 520m and 55Om levels. Failure is promoted in the core of the pillar between 460m and 6 10m Ievels. Appendix C Glossary of Terms
Collar The Collar of a hole is the opening or start of the hole. The term collaring or
starting a hole is commonly used. The end of a hole is referred to as the toe.
Cut The cut in a drift round is a groupkonfiguration of 6 to 9 holes of which 3 to
6 are of a larger diarneter and usually are not charged with explosives. The purpose of the
cut is to provide a free Face to which the remriinder of the round may break (Singh !990).
Cut The cut in a stope designates the lifi or slice mined. (i-e. The crew is mining
on the Yd cut.)
Crown Blast The crow blast in a Verticai Retreat Mined (VRM) or VerticaI Crater
Retreat (VCR) stope is the final blast taken through to the topsill. Usually the crow is 8 to
10m in thickness prior to blasting.
Nonel Anodet The trade name oîà non clectric delayed drtonator.
Primary Undercut and fil1 Method. A mining method that uses multiple slices driven from a main dot to recover a block of ore. Each slice is tilled with a tirnber and screen mat and cemented till. Once a cut has been completed the nen slice is taken below the previous cul. This is an underhand stoping method. It is usrd in extremely high grade. where complete recovery is desirable. in ore that cannot be supported conventionally using bolts. and in areas of high stress where the diminishing pillars created by ciit and fil1 mining are likely to burst. The term primary differentiates the method from a pillar or secondary undercut and fil1 stope.
Rebar Rebar bolts are sti ff bolts made of 400 grade steel (same as reinforcement rods used in concrete) which can have a forged head or threaded end. Rebar bolts are grouted in the rock using resin or cernent.
Silling Silling is the term used to denote the preliminary development phase of a stope during which the blasting requires a cul. This is to differentiate between silling the first Lift and down breaking a second lifi. Appendix D Correspondence and Approvals
D. 1 Letter recornmending to management to destress 25 Pillar
INCO LIMITED
WBKR SUsmic Event Stobie
Stobic mine expcrienced a 2.3 Mn (Richter) magnitude rismic event on Damiber 18, 1996. at 8: 13. A signifcant stress buüd up is cxpectcd as miaiDg continues in 25 block. This stress wilî dissipate whcn the geometry of the diminishhg piliar U such tbat fulm taktq place. Failurc may bc acçompanied with bursting. Once mining has pmgrcsd to rbc point that 30 block ud 12 block arc jowd by the mining of 25 block the Iiicclihood of seismic events king inducd wouid k vcry smaii. It might bc desirable to &stress the diminLching pillar to rclease this energy at a prcscnil tiw rathcr than it failhg at an Uhdctenninablc the.
Destrcssing of pülars is rclaiively new, with some data availabfe on prcconditioning blasts at the Mt. Cbarfottc mine in Australia.
Numerid modelling data. devclopment. drilling. ad explosive costs wül be tabulateci during Jaouvy so chat an iaformed decision can k made with respect to appmvhg the desuesshg of 25 block.
X.C. G. Ellioac. M. Sylvestrr, L. Cmhnne, R. Myk, H. Buksa. W. Quinn. C. McAnuity, C. Langille, File. D.2 Letter to Inco management requesting approval of destressing as a Thesis.
INCO LIMITED
m / C. ~angiiie DA^ 4 February 1997 naw: D.P. O'DOM~~
NULCT Thesis approvd
Further to our conversation, 1 request approval to write an MSc thesis on the destressing of 25 Pillar between 1700 and 1800 levels at Stobic Mine. The deswing of this pilla is required to induce failure of this pillar at a desirable lime. Documentation of this project will increase INCO's profile in the mining and academic community. The cost to iNCO will be minimal as the thesis would be written on my own the.
Your assistance in geaing official approval would be appreciated.
Denis O'Donnell Rock Mechanics Specialist Frood, Stobie and Shebandowan mines,
U--N.
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