The Stability Graph Method for Qpen-Stope Design (1)

8/17/2019 The Stability Graph Method for Qpen-Stope Design (1)

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C H A P T E R 5 9

The Stability Graph Method for Qpen-Stope Design

Yves Po tvi rt * and John H ad j i geo rg io i s t

60* 1 INTRODUCTION

In the late 1970s to early 1980s, the underground metal-mining

industry shifted its extraction strategy from highly selective

ent ry methods, such as cut-and-fill, to non -entry methods

such as open stoping. A review of Canadian practice has shown

that 90% of the total production of underground metal mines,

based on reported tonnage, rely on open-stope mining methods

(Poulin et al. 1995). The popularity of open-stope operations can

be attributed to the higher productions levels achieved by

employing larger excavations and using mechanised equipment.

Considering the high cost of developing each stope, there is

a significant incentive to produce a smaller number of large open

stopes. The consequences, however, of exceeding the maximum

critical stable dimensions of a stope can be disastrous. Instability

around open stopes may require large remedial costs for ground

rehabilitation, production delays, mining equipment loss, ore

reserves loss, and, at the extreme, injuries or fatalities to mine

workers.

Pakalnis (1986) reports that in a survey of 15 Canadian

mines, almost half (47%) of the open-stope mines had more than

20% dilution with one fifth suffering excessive dilution of over

35%. Based on field data from 34 Canadian mines, Potvin (1988)

demonstrated that open-stope design was based on past experi-

ence of mine operators in similar mining conditions and on trial

and error. Consequently, it can be argued that the reported high

dilution rates in the early 1980s could be attributed to the

absence of comprehensive engineering design tools. It follows

that there are significant economic gains to be made by

improving open-stope stability.

@Q

d

2 EXC AVATIO N STABIL ITY

Evaluating the stability of a non -entry excav ation such as a stope

can be subjective. Unlike entry excavations in which mine

workers have access, isolated rock falls in stopes are generally of

no consequence, providing that they can be handled by mucking

units. Therefo re, a stope can be considered to be stable if it yields

low dilution (less than 5%) and if there are no ground-fall-

related operational problems. It has been argued (Pakalnis,

Poulin, and Hadjigeorgiou 1995) that there is a unique accept-

able dilution rate for every mine operation. This is defined as a

function of ore gra de, costs, grade of dilution material, and metal

prices. Consequently, provided the operation remains safe and

economical, it is possible to tolerate a level of dilution and a

degree of instability for every stope. Open stopes that display

excessive dilution and/or unmanageable stability problems are

often referred to as caved. In this context, the term cave d indi-

cates major stability problems and should not be confused with

the cave mining interpretation where it refers to orebody failure

(cavings) after undercutting. This overlap of terminology has,

from time to time, been the source of confusion amongst people

not familiar with open-stope mining.

There are multiple and interrelated factors that potentially

contribute to the instability of excavations. For convenience, they

can be divided into two groups: the ones related to the in-situ

conditions prevailing before mining, and the factors related to

the disturbance of these cond itions induced by min ing.

The premining conditions can be characterised by rock-mass

classification schemes and supplemented with structural geology

data and an estimation of the in-situ stress field. The major

factors related to mining are the size, shape, and orientation of

excavations as well as the ground support used (including back-

fill). Blast damage and the effect of time in highly convergent

rock may also affect the stability of excavations.

S 0 a 3 D E S C R I P T I O N O F T H E S TA B S L I T Y -

m A P U ¡1 E T O 0 O

The stability-graph method is an empirical m ethod fo r open-stope

design (non-entry ex cavations). It aims to account for and quantify

the major factors influencing the stability of open stopes. A stability

index for each stope surface is subsequently traced against its

dimensions. A series of empirically derived guidelines allow for

predictions on the overall stability of a stope. Since its introduction

(Mathew s et al. 1981), it has gained wide acceptance and is used

world wide as a design tool. There are documented case studies of

the method being used in Africa, Europe, and the United States,

and extensive databases of case studies in Canada and Au stralia. In

practice, the stability graph can be employed during three distinct

mining stages. Its primary use is during the feasibility stage but it

has also been found useful during individual stope planning.

Finally, through the use of back analysis, it provides an index of

stope performance and allows the mine operation to develop reme-

dial strategies where warranted.

The method traces its origin to the recognition that tradi-

tional rock-mass classification and design tools were based on

tunnelling case studies. A review of some case studies and engi-

neering judgement resulted in the first version of the method,

whereby a stability number (N) was traced against the hydraulic

radius of a stope surface.

The stability-graph method uses the NGI tunnelling index Q

(Barton, Lien, and Lunde 1974) as a basis for estimating rock-

mass quality.

where:

Q = NGI tunnelling index with

RQ D = rock quality designation

* Australian Center for Geomechanics, Nedlands, WA, Australia,

t Laval University, Quebec City, Quebec, Canada.

5 1 3


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5 1 4

F o u n d a t i o n s f o r D e s i g n i

1000

1 0 1 5

H y d r a u l i c r a d i o s ( m )

1 0 1 5

H y d r a u l i c r a d i u s ( m )

F IG U R E S O 1 S ta b i l i t y g r a p h , a f t e r M a t h e w s e t a l . (1 9 8 1 )

F I GU R E 6 0 . 2 S t a b i l i t y g r a p h a f t e r P o t v i n (1 9 9 8 )

J'

r

= joint roughness number

J

w

= joint water reduction number

J

n

= joint set number

J

a

= joint alteration number

SR F = stress reduction factor

Using S RF equal to 1 is a departure fro m the original system

(Barton, Lien, and Lunde 1974). This modified tunnelling index,

Q, is further adjusted to account for stress, rock defect orienta-

tion, and design-surface orientation factors to arrive at a stability

number N. The stability number was plotted against the hydraulic

radius (surface area/perimeter) of the studied surface of an exca-

vation (Mathews et al. 1981). Three zones of potentially stable,

unstable, and caving were proposed with reference to the

predicted stability of an excavation (see Figure 6 0.1).

In its early days, a major shortcoming of the method was

that it was backed by limited field data—26 case studies from

three mines. Once the database was expand ed to 175 cases from

34 mines and the stability graph modified (Potvin 1988), the

method rapidly gained wide acceptance in the Canadian mining

industry. The transition z one fr om stable to unstable was reduced

significantly, thus removing some of the subjectivity in using the

design chart (see Figure 60.2). It should be noted that in the

Potvin database, the adjustment factors were differen t than those

proposed by Mathews et al. (1981). This resulted in what is

commonly referred to as the modified stability-graph method

using a stab ility index AT.

N' = Q'xAxBxC

where:

AT = stability num ber

Q ' = mod ified tunnelling quality index (NG I)

A = stress factor

B = joint orientation factor

C = gravity factor

A Factor

The A-factor is used to account for the resulting

induced stress in the investigated stope surface. A series of charts

that provided preliminary estimates of induced stresses for

1.0

0 .8

0 .6

42

i

0 .4

W

0 .2

0.1

0

1

Gc/Oi


3/8

T h e S t a b i l i t y G r a p h M e t h o d f o r © p e s i -S t o p e D e s i g n

5 1 5

1 0 2 0 3 0 4 0 5 0 6 0 7 0 8 0 9 0

R e l a t iv e d i f f e r e n c e i n d i p b e t w e e n

t h e c r i t i c a l Jo i n t a n d s l o p e s u r f a c e

F I G U R E 6 0 A

( 1 9 8 8 )

D e t e r m i n a t i o n o f t h e O r i e n t a t i o n F a c t o r , a f t e r P o t v i n

1 1 l

30 40 50

i n c l i n a t i o n o f

:

F I GU R E 6 0 . 5 I n f l u e n c e o f g r a v i t y f o r s l a b b i n g a n d g r a v i t y f a ll m o d e s

o f f a i l u r e

© 0 = 4 D I S C U S S I O N O F T H E I N P U T F A C T O R S

As a result of wide dissemination in a number of textbooks

(Hoek, Kaiser, and Bawden 1995, Hutchinson and Diederichs

1996) the input factors for the calculation of N' described above

have now gained broad acceptance from practitioners and

researchers. The applicability of the input methodology on a

case-by-case analysis was reviewed and, with the exception of the

minor modifications to the C factor shown in Figure 60.6, were

found to be appropriate (Hadjigeorgiou, Leclair, and Potvin

1995). On the other hand, some authors (Stewart and Forsyth

1995; Trueman et al. 2000) have indicated their preference for

the formulation of the input factors as originally proposed

(Mathew s et al. 1981).

critical joint > F W

i 1 1 — n r

10 20 30 40 50

I n c l i n a t i o n o f c r i t i c a l j o i n t

F I G U R E 6 0 I n f l u e n c e o f g r a v i t y f o r s l i d i n g m o d e o f f a i l u r e , a f t e r

H a d j i g e o r g i o u L e c l a i r a n d P o t v i n ( 1 9 9 5 )

Several other modifications to the stability graph have been

proposed during the last decade. The following offers a brief

historical review. It should be noted that most of these proposals

have not yet been extensively tested by case studies nor are they

wide ly employed by practitioners.

Scoble and Moss (1994) suggested that there was merit in

adding tw o further adjustment factors, D for blasting and E for

sublevel interval rating with some tentative factors proposed. A

fault factor was been deve loped that can be incorporated into the

stability factor (Suorineni, Tannant, and Kaiser 1999). This fault

factor accounts for the angles between fault and stope surface

and the position where the fault intersects the stope surface. The

fault factor was derived based on modelling and demonstrated

that it could be critical for a series of documented case studies in

Canada and Africa. At the G olden Giant Mine in Ontario, Canada,

it was shown that under high-stress environments the introduc-

tion of a stress-damage factor merited attention (Sprott et al.

1999). Based on 3-D numerical modelling, they used the extra

stress deviato r, the uniaxial resistance of the rock, and the

hydraulic radius to define a stress-damage factor. It has been

argued that the stability predictions of the stability-graph method

may prove inaccurate due to the influence of rock-mass degrada-

tion and relaxation (Kaiser et al. 1997). It was recommend ed that

stope sequencing be used as a tool to minimise stress-induced

rock-mass degradation and to minimise stress relaxation. In their

work, they defined rock-mass relaxation as stress reduction

parallel to the excavation wall—not to stress reductions in the

radial or a reduction in confinement. Rock-mass degradation was

quan tified as loss of rock-mass strength.

S 0

o

S H Y D R A U L I C R A D I U S

Th e term hydraulic radius has been used in the past to charac-

terise the size and shape of stope surfaces (Laubscher 1977). This

is the area over the perimeter of a given stope surface. It has also

been demonstrated that, despite the advantages of hydraulic

radius over span, it still has important limitations (Milne,

Pakalnis, and Felderer 1996). In particular, when applied to

irregularly shaped stope surfaces, it is possible to arrive at the


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5 1 6


F I G UR E 6 0 . 7 D e t e r m i n a t i o n o f t h e r a d i u s f a c t o r , a f t e r M i l n e e t a t .

( 1 9 9 6 )

same hydraulic radius. It has been put forward that a better way

to describe the geometry of an irregularly shaped excavation is

the radius factor (see F igure 60.7 ). This is determined by identi-

fying the centre of any excavation and by taking distance

measurements to abutments at small regular increments:

RF = °

5

I y n = i i

n ^ Q r

0

where

re = distance fro m the surface centre to the abutm ents at

angle q

n = number of rays measured to the surface edge

In principle, the radius factor can be determined at any

point on a surface. If the centre cannot be determined, a series of

calculations are possible with the maximum value assumed to be

the radius factor. Despite its somewhat cumbersome definition,

the radius factor can easily be calculated by a routine integrated

into a computerised design package.

8©oS B E B i m c h a r t s

In reviewing the proposed chart, Figure 60.1 (Mathews et al.

1981), it can be seen that the developed guidelines were some-

what vague for design purposes. This was because there was

insufficient data to provide more accurate zone definition. As

more case studies became available, a narrower transition zone

and a support requirement zone were defined (Potvin 1988).

This has allowed for a calibrated and m ore versatile design tool

(Figure 60.2). A more comprehensive statistical analysis further

mod ified the support zones by introducing lines indicating wh ere

cable bolting could be used (see Figure 60.8) (Nickson 1992). A

review of a larger database (Hadjigeorgiou, Leclair and Potvin

1995) confirmed the general validity of previous work (Potvin

1988, Nickson 1992) within statistical limits.

It should be noted that the work of Hadjigeorgiou et al.

(1995) demonstrated that, for larger stopes with a hydraulic

radius greater than 15, the design curve was in fact flatter (see

Figure 60.9). M ore recent work in the United Kingdom by Pascoe

et al. (1998) and in Australia by Trueman et al. (2000) has

confirmed the same trends.

A series of design guidelines were proposed (Stewart and

Forsyth 1995) that allowed for a finer definition of the types of

F I G UR E 6 0 . 8 S t a b i l i t y g r a p h , a f t e r N i c k s o n ( 1 9 9 2 )

F I GU R E 6 0 . 9 S t a b i l i t y g r a p h d e s i g n li n e s a s d e v e l o p e d b y

H a d j i g e o r g i o u s e t a l . ( 1 9 9 5 )

stope failure, distinguishing between potentially unstable, poten-

tially major, and caving failure separated by transition zones (see

Figure 60.10). In their experience, the boundary between stable

and unstable is clear cut, while the transition between unstable

and major failure is not as well defined. It is of interest that the

transition between a potentially stable zon e and a potentially

unstable zon e is identical to Potvin's transition zone. In practice,

it could be argued that, for open-stope design purposes, it is

somewh at irrelevant to subdivide the area de fining stope failure

into three zo nes as the objective is to design stable stopes.

Cavity monitoring laser surveys have been employed to back-

analyse the resulting volumetric measurements of overbreak/


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T h e S t a b i l i t y G r a p h M e t h o d f o r © p e s i -S t o p e D e s i g n

5 1 7

1000

1

5 1 0 1 5

H y a r a u l i c r d i u ( m )

1 0 1 5

H y d r a u l i c r a d i u s ( m )

F I GU R E 6 0 = 1 0 S t a b i l i t y g r a p h a f t e r S t e w r t & F o r s yt h ( 1 9 9 3 )

F I GU R E 6 0 . 1 2 E s t i m a t i o n o f o v e r b r e a k / s l o u g h f o r n o n s u p p o r t e d

h a n g i n g w a l l s a n d f o o t w a l l s , a f t e r Cl a r k a n d P a k a l n i s 1 9 9 7

S t o p e

„ W i d t h

s?

s*

¡ 4 I C r o s s - S e c t i o n s

G e n e r a t e d f r o m

I M S S u r v e y

L e n g t h j . y . . G ì

l/\ \

E q u i v a l e n t L i n e a r

O v e r b re a k / S l o u g h

( E x p r e s s e d i n M e t e r s

E l

S l o u g h f r o m S t o p e W a i l s

I E q u i v a l e n t l i n e a r o v e r b r e a k / S l o u g h

F I GU R E 6 0 . 1 1 S c h e m a t i c d e f i n i t i o n o f t h e E L OS p a r a m e t e r , a f t e r

C l a r k a n d P a k a l n i s 1 9 9 7

slough and underbreak, and a new index has been prop osed (Clark

and Pakalnis 1997) (see Figure 60.11):

ELOS

equivalent linear overbreak

slough

volume of slough from stope surface

stope height x wa ll strike length

In Figure 60.12, ELOS has been integrated in the stability

graph, providing a series of design zones (Clark and Pakalnis

1997). Although this data presentation does not account for the

influence of support, it provides a useful back-analysis tool for

hanging walls and footwalls in a low- or relaxed-stress state and

with parallel geological structure being present.

All of the above graphs rely on arbitrarily drawn design

curves. The first comprehensive statistical analysis of the then-

available field data (Nickson 1992) clearly demonstrated the

applicability of the modified stability graph (Potvin 1988) and

laid the foundations for further statistical work (Hadjigeorgiou,

Leclair, and Potvin 1995, Pascoe et al. 1998, and Suorineni

1998).

Successful applications of the stability graph recognise that

the method remains subjective. Despite using quantifiable values,

the precise degree of inherent conservatism is not known. It

reflects current and past practice, which m ay have been influ-

enced by legislation, local practices, or geologic al peculiarities, and

does not necessarily constitute an optimum design meth odology.

8 0 . 7 ¡R IS K A N A L Y S I S

It has been argued that the design of non-entry excavations lends

itself to risk analysis much more than the design of access ways

where worker safety is the major concern (Pine et al. 1996,

Pascoe et al. 1998, Diederichs and Kaiser 1996). There are two

basic elements in risk analysis. The first one deals with input vari-

ability and the second deals with calibration uncertainty. For

practical purposes, the major challenge lies in defin ing ho w much

risk is acceptable for design purposes. Using risk probability

procedures for fine-tuning or calibrating site-specific design

guidelines, while attractive, is hindered in that site-specific cali-

brated guidelines will be validated only towards the end of mine

life. At that time, their impact will be limited to p roviding a b etter

understanding of particular field conditions, but it may be too

late to implement m ajor design changes.

6 0 . 8 S U P P O R T R E C OI V I I VS E N D A TS O N S

Potvin (1988) first addressed the influence of support on the

stability of open stopes. The area of the graph that could success-

fully be supported by cable bolts was refined , and a series of design

recommendations made on cable-bolting patterns (Nickson 1992).

The basic concept is that there is a zone where support cannot be

effectively used to stabilise the excavation. It has been shown

(Hadjigeorgiou, Leclair, and Potvin 1995) that the actual support-

able zone was smaller than predicted (Nickson 1992).

A design chart is available to select a suitable cable-bolt

density as a function of relative block size (RQD/J

n

) and the

hydraulic radius (Potvin and Milne 1992). This graph, slightly

modified in Figure 60.13, is most appropriate for stope backs. It

has also been employed for hanging-wall reinforcement design,

provided a systematic and regular cable-bolt pattern is used.


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5 1 8

Foyo i ida tBOin is fo r Des ign

( R Q D / J n )

H y d r a u l i c r a d i u s

F I GU R E 6 0 . 1 3 D e t e r m i n a t i o n o f c a b l e b o l t d e n s i t y

At the time when the design curves for cable reinforcement

were developed, support often consisted of single plain-strand

cables. Recent years have seen a shift towards double-strand and

modified-geometry cables as they provide higher support

capacity. This is achieved in the presence of sufficiently high

ground deformation whereby the strength of the steel is being

mobilized. In other words, pattern reinforcement is designed to

help the rock support itself and not necessarily to support the

dead weight of the rock. A series of semi-empirical design charts

to determine the design spacing for both single- and double-

strand cable bolts has been proposed (Diederichs, Hutchinson,

and Kaiser 1999). It should be noted, how ever, that there are no

documented case studies confirming their application.

SQ 3

L I M I T A T I O N S © F T i H E S T A B I L I T Y

QiRAPU

All empirical methods are limited in their application to cases

that are similar to the one used in the developmental database.

Therefore, the stability graph is inappropriate in severe rock-

bursting conditions, in highly deformable (creeping) rock mass,

and for entry methods. Since its introduction, the stability graph

method has been the subject of extensive efforts to expand its

applicability to better account for the presence of faults, blast

damage, and stress damag e. Unfortunately, some of the proposed

modifications are not supported by field data. Furthermore, when

merging databases from diverse sources, it is necessary to verify

the quality of collected data. In particular, the practice of using

empirical correlations to convert from one rock-mass classifica-

tion system to another should only be used as a last resort and

even then with great caution.

For all practical purposes, the stability graph can be used

during the feasibility stage, during individual stope planning, and

for stope reconciliation or back analyses.

e ®

D

1 0 D E S D G N C O N S I D E R A T I O N S

T H E F E A S I B I L I T Y S T A © E

The determination of adequate stope dimension is one of the

most critical decisions to be made at the feasibility study stage of

a mine. The profitability of an operation is directly linked to

productivity, which in turn, is influenced by stope dimensions.

Validation of stope-design methodology can begin once the first

stope is extracted. By this time, however, the mine infrastructure

is already in place, allowing for no or only minor m odifications to

design stope dimensions. This emphasizes the importance of

developing a reliable stope-design methodology at the earliest

possible stage.

Many practitioners have reported on the reliability of the

stability-graph method during the last 12 years (Reschke and

Romanowski 1993, Bawden 1993, Pascoe et al. 1998, Dunne et al.

1996, and Goel and Wezenberg 1999). When properly used, the

meth od provides a goo d ball park estimate of stable stope dimen-

sions under different conditions. The major limitation at the feasi-

bility study stage is the availability of quality geotechnical data.

This is a concern fo r all design m ethods. C onsequently, it is essen-

tial to optimise all available data while being fully aware of any

limitations. The following guidelines can facilitate the estimation

of realistic stability numbers during the feasibility stage.

An integral part of the stability method is the quantification of

rock-mass quality based on the Q system. At the green f ield stage,

the majority of geomechanical data are derived from boreholes.

Consequently, it is possible to develop a comprehensive database

of RQ D readings, which can easily be integrated into geological

models easily accessed by both planning and rock mechanics. It is

strongly suggested that the number of joint sets be determined by

using oriented diamond-drill cores in the orebody.

There are several case studies where core data were used to

derive representative Q readings for underground mines (Milne,

Germain and Potvin 1992, Germain, Hadjigeorgiou, and Lessard

1996). This has included a simplified approach to determine joint

alteration, J

a

. If the joint cannot be scratched with a knife, J

a

is

assumed to be equal to 0.75, and if it is possible to scratch, it

varies from 1.0 to 1.5. Whe n a joint feels slippery to touch and

can scratched with a fingernail, J

a

is equal to 2; and when it is

possible to indent with a fingernail, J

a

is equal to 4. The joint

roughness parameter (J

r

) is more difficult to assess on a small

exposed surface of a core. How ever, in most cases, it is possible to

estimate whether the surface is smooth or rough . In the absence

of reliable data, joints are assumed to be planar. This allows for J

r

values of 0.5 for slickenslide planar, 1.0 for smooth planar, and

1.5 for rough planar joints.

Factor A can generally be assumed to be equal to 1 for all

stope walls, unless mining is to proceed very deep (say 1,000 m

and deeper). As a first-pass estimation, the stress induced in

stope backs could be assumed to be around 1.5 times the pre-

mining horizontal stress for transversal mining (mining across

the strike of the orebody). In longitudinal mining (mining along

strike), a rough estimate of the induced stress in the back can be

obtained by doubling the premining horizontal stress perpendic-

ular to the orebod y strike. The prem ining stress can be measured

if underground access is available or otherw ise based on regional

data. The uniaxial compressive strength of rock is easily obtained

by standard laboratory tests on cored rock. A larger database of

UCS values can also be gathered at low cost using a standard

point load test. When at least some oriented core is available,

Factor B can be estimated. In the absence of joint orientation

data, a minimum value of 0.2 is assumed. The estimation of

factor C is independent of ground conditions and is, therefore,

straightforward to determine, eve n at the feasibility stage.

A good methodology for the construction of a geomechan-

ical model and the application of the stability graph method for

mine feasibility assessment exists (Nickson et al. 1995). Stability

numbers are calculated for back and walls and displayed on m ine

sections. For each stability number (N ), a hydraulic radius (S ) is

determined from the stability graph in Figure 60.14, using the

upper section of the transition zone. The option of increasing

stope dimensions exists if a systematic pattern of cable-bolt

support is to be used. This can be assessed using Figure 6 0.8.

Unless the ground conditions are consistent throughout the

orebody, a number of stable hydraulic radii will be produced,

which can be grouped into domains and displayed on a longitu-

dinal section. Because most mines employ systematic layouts, a


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T h e St a b i l i t y G r a p h M e t h o d f o r © p e s i -S t o p e D e s i g n

5 1 9

3CQQ EL

F I G UR E 6 0 . 1 4 A c a s e s t u d y d u r in g t h e f e a s i b i l i t y s t a g e , a f t e r

N i c k s o n e t a i . 1 9 9 5

unique stope dimension is usually determined for each mine

domain. Engineering judgement must be used in selecting the

appropriate hydraulic radius. Selecting the smallest hydraulic

radius would ensure that all stopes would be stable, but would

not likely be the most economical option. Notwithstanding the

value of ore, the impact of dilution, acceptable risks, and the

operating philosophy, a good starting point for selecting a mean

hydraulic radius for an entire domain would be to ensure that

approximately 80% of the stopes are stable. The remaining 20%

or so can then be dealt with individually using specific ground

support or different extraction strategies.

The stope height, length, and width within each domain can

be determined from the mean hydraulic radii (roof and walls).

The orebody geometry obviously has an important influence on

the determination of the stope geometry. In many cases, there

will be an economic advantage to maximising the stope height as

it has a major influence on the sublevel interval, and therefore,

on the mine infrastructure cost.

As more and more operations integrate backfill in the

extraction process, its impact on stope stability must be

accounted for. The main function of mine backfill is to limit the

exposure of stope surfaces during extraction by filling adjacent

mined-out stopes. Provided a good quality-control program is

followed, it can reasonably be assumed that backfill provides

adequate support of adjacent mined out stopes. Consequently,

the stability-graph method treats backfill as a rock material when

calculating stope-wall dimensions. In practice, however, it is rare

that a tight fill can be established against a stope back. As a

result, in stope-back analysis, the influence of backfill is assumed

to be minimal and ignored.

@© „ 1 ± D E S I G N C O N S I D E R A T I O N S F O R

I N D I V I D U A L S T O P E P L A N N I N G

It is good engineering practice to employ the stability graph at the

plannin g stage to evaluate the stability of each stop e. At this stage

of development, there is usually underground access that allows

for a revaluation of the rock-mass data collected during the feasi-

bil ity study. Direct underground mapping can provide more

reliable information than diamond-drill hole data. Another

advantage of underground observations is that it can reveal early

signs of stress. This can be complemented by stress measure-

ments allowing for a better assessment of stress influence on the

stopes. At this point, it is possible to integrate numerical model-

ling to investigate optimum sequencing.

Access to more quality data can allow fo r greater confidence

in stope stability estimates than allowed during the feasibility

study. Consequently, it is possible to consider modifications or

fine-tuning to the ground support and extraction strategies. At

this stage, it is also important to assess the influence on stope

stability of some of the factors not well accounted-for in the

stability-graph method. These can include faulting, shear zones,

or areas susceptible to rock bursts.

One o f the great bene fits of using such a meth od at the plan-

ning stage is that it brings geomechanical considerations into

stope design and increases the awareness of mine planners to

ground-control issues. Modern stope designs require the close

cooperation of geology, mine p lanning, and rock mechanics.

6 0 1 2 S T O P E R E C O N C I L I A T I O N

Using the stability graph to assess and document stope perfor-

mance is useful to build site-specific empirical knowledge that

can be used in future design. Once a sufficient number of case

histories have been collected, it may even be possible to refine the

stability graph for a given site or extend its predictive capability

to dilution (Pakalnis, Poulin, and Hadjigeorgiou 1995) or to

quantify the probability of failure (D iederichs and Kaiser 1996).

A major aid in stope reconciliation has been the introduction of

cavity-monitoring systems.

These refinements are interesting and contribute to a better

understanding of stope behaviour. However, the value added to

operations from this effort remains limited in many cases because

the initial stages of mining are completed, the mine infrastruc-

ture is in place, and opportunities for modifying stopes layout are

restricted.

6 0 , 1 3 D I S C U S S I O N A N D C O N C L U S I O N

By definition, empirical design is based on observation and expe-

rience. The stability-graph method owes its popularity in its ease

of use its application at early stages of minin g, and the fac t that it

can provide a reference for stope performance. Invariably, it

cannot provide a successful prediction for every stope at every

operation because the complexity of ground conditions and oper-

ating practices can influence stope performance.

It has been argued in the past that the method only reaches

its full potentia l wh en it is site-calibrated . Th e basic assumption

is that, as more data become available, the design recommenda-

tions can be modified through back analysis. Obviously, this is an

important step in better understanding the site conditions. If the

reconciliation exercise is done rigorously, it can reveal important

information on the efficiency of mine practices such as blasting,

prereinforcement, and sequencing.

However, this should not detract from the main ob jective of

the method as a design tool at the feasibility stage when no such

data is available, but when the critical decision must be made.

The extensive calibration of the method worldwide makes it

very robust and ideal for designing open-stope dimension at the

z

o

Q

HW ROD

Q

:

= 1 0

A = 1

B = 0 .3

C = 3 .5

N' = 10 5

3 = 6

> = 0 %

S > = 20%

l i > = 40%

El

>

= 60%

a > = 7 5 %


8/8

5 2 0


feasibility-assessment stage. The added value of a very refined

site-specific graph towards the end of a mine life can only be

limited.

Over the years, there has been a proliferation of design

charts aiming to refine the m ethod or expand on its applicability.

Design charts that are not backed by field data have limited use.

Similarly, modifications that rely on limited sites should be

viewed with caution.

Complex charts bring new and interesting ideas, but one

should keep in mind that the method can be no better than the

quality of input data available. This is particularly true at the

feasibility stage where data are limited by access.

Introducing many zones to the graph has limited application

at the design stage because the designer has to come up with a

hydraulic radius number for each domain. The transition

between stable and unstable and the potential for increasing

dimensions b y using pattern cable bolts rema in the basis for stope

design

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Documents

The Stability Graph Method for Qpen-Stope Design (1)