© 2005 Institute of Materials, Minerals and Mining and

Australasian Institute of Mining and Metallurgy. Published

by Maney on behalf of the Institutes. Manuscript accepted in

final form 24 February 2005.

DOI 10.1179/037178405X44494 Mining Technology (Trans. Inst. Min. Metall. A) March 2005 Vol. 114 A13

Empirical approaches for opening design in weak rock masses 

Tom Brady, Lewis Martin and Rimas Pakalnis 

A major focus of ground control research presently 

being conducted by the Spokane Research Laboratory 

of the National Institute for Occupational Safety and 

Health (NIOSH) is to incorporate weak rock masses 

(such as are associated with operations in the Carlin 

Trend in Nevada) into existing design relationships. 

The original database that led to most of the empirical 

design relationships presently employed in hard-rock 

mining was derived from fair-to-good-quality rock. In 

this study, the relationship between weak rock quality 

and opening design (non-entry/entry methods) is being 

investigated. The common factor in all mines is a weak 

back or wall. This work attempts to provide tools that 

will enable a mine operator to make economic decisions 

that will also ensure a safe working environment. 

Tom Brady (for correspondence – E-mail: thb6@cdc.gov)and 

Lewis Martin are at the Spokane Research Laboratory, National 

Institute for Occupational Safety and Health, 315 East 

Montgomery Avenue, Spokane, WA 99207-2291, USA (E-mail: 

ljm8@cdc.gov); Rimas Pakalnis is at the Mining Department, 

University of British Columbia, Vancouver, British Columbia, 

Canada (E-mail: rcp@mining.ubc.ca). 

Keywords: Weak rock masses, empirical design, opening 

design 

INTRODUCTION 

Researchers at the Spokane Research Laboratory of 

the National Institute for Occupational Safety and 

Health, Spokane, Washington, USA, and the 

University of British Columbia, Vancouver, Canada, 

have assembled a team to develop underground design 

guidelines for safe and cost-effective mining within a 

weak rock mass. Such work also includes developing 

novel support methods, such as the use of synthetic 

fibre reinforced shotcrete, ways to undermine underhand-and-fill backfilled stopes, and assess supports 

presently in place in weak rock masses. In the present 

study, rock mass interaction with grouted bolt supports 

was investigated in three mines in Nevada and backfill, 

pillar, and bolt support were studied in one. 

Many Nevada gold deposits are found in intensely 

fractured, faulted, and argillised host rock. As a result, 

underground mining is often difficult and hazardous, 

as indicated by the number of injuries and fatalities 

from uncontrolled falls of ground (Table 1). 

Table 1 Ground control injuries and fatalities in underground 

ground gold mines in Nevada, 1985–2000 

Fatalities 7 

Permanent disabilities 4 

Lost-time injuries 49 

Restricted-activity injuries 46 

Other injuries 110 

Total 216 

Reported rock falls with no injuries* 69 

*Includes MSHA data for non-injury incidents where a
 

reportable fall of ground occurred but did not cause injuries
 

because the mining area was unoccupied.
 

A comparative analysis by the Mine Safety and 

Health Administration (MSHA)9 for the years 

1990–1999 indicated that the number of roof-fall 

injuries in Nevada has varied from a low of 8 in 1991 

to a high of 28 in 1995 and 1997. As late as 1999, the 

number of injuries was still in double figures (Fig. 1). 

Analysis of the MSHA data shows 76·7% of the roof 

falls were from the back, 18·6% were rock falls from 

face or ribs with the remaining 4·7% of an unknown 

nature. The mines are required to report all injuries to 

MSHA by law. 

1 Injuries in Nevada, 1990–2004 

Mining is a dynamic process, and ground conditions 

can change over a short distance. A mine opening must 

perform in a predictable manner over its expected life. 
A14 Mining Technology (Trans. Inst. Min. Metall. A) March 2005 Vol. 114

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Brady, Martin, Pakalnis Empirical approaches for opening design in weak rock masses 

Empirical design methods have been used successfully 

over the past 30 years largely because they permit the 

overall behaviour of a rock mass to be predicted easily 

and accurately. The basis for the success of the 

empirical method is a strong foundation of field data 

coupled with on-going field observations that allow 

changing rock conditions to be evaluated as mining 

progresses. Two systems were initially developed from 

case studies and databases originally derived from 

civil engineering applications and augmented by mine 

studies – the rock mass rating (RMR) system and the 

Q system. As stated by the author, ‘the Q system is 

specifically the permanent lining estimation system 

for tunnels and caverns in rock and mainly for civil 

engineering projects’.1 

There are several advantages for utilising the RMR 

system over the Q system. The RMR system has a 

relatively straight-forward scoring system for each 

parameter on a 0–10 scale and consequently is easier 

to learn.10,11 

One of the original extensions of the rock mass rating 

system from civil into mining was conducted under a 

US Bureau of Mines Spokane Research Laboratory 

research grant which developed empirical methods for 

block caving operations for US copper mines.7 

It is important to remember that any method of 

designing an opening must be easy to assess, 

understand, apply, modify if necessary, and reproduce 

for the next application if accepted as an on-going 

operational tool for design. A critical factor is that the 

design incorporates the degree of stability required for 

any mine entry. 

SPAN DESIGN MAN-ENTRY METHODS 

The ‘critical span curve’ was developed in 1994 to 

evaluate back stability in cut-and-fill mines.13 In 2000, 

the span curve database of 172 observations 

developed by the University of British Columbia was 

expanded to include a total of 292 case histories from 

mines primarily in Canada.21 The information from 

these case histories provides the basis for the span 

design curve shown in Figure 2. 

Design span curve 

Table 2 Nevada mines database: back spans in weak rock 

RMR Span 

(%) (m) Condition Other 

Mine 1 

45 5·5 S Stable with support 

45 9·0 S Stable with support 

40 6·0 S Stable with support 

Mine 2 

40 4·0 S Stable with support 

45 4·3 U Caved with support 

30 3·7 S Stable with support 

Mine 3 

40 7·0 S Stable with support 

45 2·1 S Stable with support 

26 2·1 S Stable with support 

25 4·6 U Caved with support 

55 7·6 S Stable with support 

45 3·0 S Stable with support 

Mine 4 

70 4·6 S Stable with support 

40 4·6 S Stable with support 

25 4·6 S Stable with support 

55 5·5 S Stable with support 

30 6·1 S Caved upon longhole 

30 6·1 S Caved upon longhole 

45 4·6 S Stable with support 

50 6·1 S Stable with support 

70 11·3 S Stable with support 

25 7·3 S Stable with support 

30 3·0 U Prior to support placement 

30 1·8 S Prior to support placement 

50 6·1 S Stable with support 

55 7·6 S Stable with support 

55 6·1 S Stable with support 

Mine 5 

30 3·0 S Stable with support 

30 4·3 U Caved with support 

20 5·8 U Caved with support 

15 3·7 S Stable with support 

Mine 6 

45 4·3 S Stable with support 

40 6·1 U Caved with support 

40 4·9 S Stable with support 

Mine 7 

40 4·6 S Stable with support 

35 4·6 S Stable with support 

Mine 8 

25 5·0 U Caved; had to spile 

20 1·2 S No support; 

maximum round possible 

25 2·4 S No support; 

maximum round possible 

35 3·1 S No support; maximum 

round possible 

55 3·7 S No support; typical round; 

no problems 

35 4·6 S No support; typical round; 

no spile/shotcrete 

20 7·6 U Caved; had to spile 

45 6·0 S Stable with split-sets only 
Mining Technology (Trans. Inst. Min. Metall. A) March 2005 Vol. 114 A15

Brady, Martin, Pakalnis Empirical approaches for opening design in weak rock masses 

A ‘critical span’ is defined as the diameter of the 

largest circle that can be drawn within the boundaries 

of the exposed back. The stability of this exposed span 

is related to the type of rock in the immediate back. 

The ‘design span’ refers to backs that have no support 

and/or spans that are supported with localised pattern 

bolting (1·8-m long mechanical bolts on 1·2 × 1·2-m 

spacings). Local support is deemed as support used to 

confine blocks that may be loose or that might open or 

fall because of subsequent mining in surrounding areas. 

Excavation stability is classified into three categories: 

(i)	 Stable excavation: (a) no uncontrolled falls of 

ground; (b) no movement of the back is observed; 

and (c) no extraordinary support measures have 

been employed. 

(ii)	 Potentially unstable excavation: (a) extra ground 

support has been installed to prevent potential 

falls of ground; (b) movement has occurred in 

the back; and (c) increased frequency of ground 

movement has been observed. 

(iii)	 Unstable excavation: (a) area has collapsed; (b) 

depth of failure of the back is 0·5 times the span (in 

the absence of major structures); and (c) support 

was not effective in maintaining stability. 

A –10 correction factor is applied to the final RMR76 

value when evaluating rock with shallow-dipping or flat 

joints. However, the applicability of this factor is being 

re-assessed for weak ground because of its amorphous 

nature and because joint direction is expected to play a 

minor role. Where discrete ground wedges have been 

observed and identified, they must be supported prior to 

employing the critical span curve. 

Stability is generally defined in terms of short-term 

stability because the database is based largely on 

stoping methods that, by their nature, are of short 

duration. Movement of the back greater than 1 mm 

within a 24-h period has been defined as a critical 

amount of movement for safe access.18 

Some 44 case histories from five different mines 

with RMR76 values varying between 20 and 85 were 

added to the information base for the critical span 

curve (Table 2). Several values were less than 55% 

RMR76; the lowest RMR76 value calculated for any 

location was 25%. This information was used to 

augment the original ‘span design curve for man-entry 

mining13 as shown in Figure 3A. The span curve 

enables an operator to assess back stability with 

respect to the rock mass. The information has been 

used successfully to predict the stability of weak rock 

masses and has provided operators with an additional 

design tool for making decisions concerning the 

stability of mine openings. The data are being coupled 

with depth of failure to define the amount of support 

required to arrive at a safe, cost-effective man-entry 

design. A brief description of the use of the critical 

span curve is presented; however, the reader is referred 

to the detailed reference as outlined by Pakalnis.18 

3 Distribution of RMR. (A) Span database;13 (B) 

stability graph database (ELOS)6 

4	 Augmented span curve. Numbers in key correspond 

to mine numbers in Table 2. Letters indicate 

location on the span curve 

STABILITY GRAPH METHOD, 

NON-ENTRY MINING 

The original stability method for open stope design was 

based largely on Canadian operations from 55 case 

histories which included 48 back studies and 7 wall data 

points.15 The original database of case histories 

exhibited a rock mass rating (RMR76) in excess of 50% 

or Q values of 2·0 or greater.4 The method was 

extended and modified by Potvin19 based upon data 

from 34 mines with 175 open stope case histories and 

67 cases of supported stopes. Nickson17 expanded the 

existing database of supported stopes by Potvin, by 

collecting 46 cases histories while visiting 13 mines. In 
A16 Mining Technology (Trans. Inst. Min. Metall. A) March 2005 Vol. 114

Brady, Martin, Pakalnis Empirical approaches for opening design in weak rock masses 

Table 3 Nevada mines database: stope spans and walls in weak rock (all values of A = 1) 

RMR Dimensions, 

(%) height × length (m) Dip B C N HR (m) ELOS (m) Comments 

Mine 1 

45 20 × 17 90 0·3 8 2·7 4·6 <1·0 < 1 m of ELOS 

40 20 × 16 90 0·3 8 1·5 4·4 2·0 

55 49 × 18 90 0·3 8 8·1 6·6 1·0 

39 34 × 34 90 0·3 8 1·4 8·5 4·6 

25 90 0·3 8 0·3 1·8 < 1·0 < 1 m of ELOS stable (estimated) 

34 90 0·3 8 0·3 3·4 < 1·0 < 1 m of ELOS stable (estimated) 

42 90 0·3 8 0·3 2·8 < 1·0 < 1 m of ELOS stable (estimated) 

Mine 2 

40 11 × 21 90 0·3 8 1·5 3·4 < 1·0 < 1 m of ELOS 

50 11 × 21 90 0·3 8 4·7 3·4 < 0·5 < 0·5 m of ELOS 

Mine 3 

55 18 × 18 70 0·2 5·9 4·0 4·6 0·6 

26 12 × 18 70 0·3 5·9 0·2 3·7 > 2·0 > 2 m of ELOS 

Mine 4 

25 6 × 29 55 0·2 4·5 0·2 2·5 0·3 Cluster average, height/width 

25 8 × 36 90 0·3 8 0·5 2·4 0·1 Cluster average, rib 

25 17 × 12 90 0·3 8 0·5 3·5 0·1 Cluster average, rib 

25 21 × 15 90 0·3 8 0·5 4·5 0·1 Cluster average, rib 

55 30 × 26 90 0·3 8 7·2 7·0 0·1 Cluster average, rib 

45 6 × 25 90 0·3 8 2·4 2·5 0·1 Cluster average, height/width 

45 18 × 12 90 0·3 8 2·4 3·5 0·1 Cluster average, rib 

45 19 × 16 90 0·3 8 2·4 4·4 0·1 Cluster average, rib 

45 6 × 22 90 0·3 8 2·4 2·4 0·5 Moderate 

25 16 × 5 0·5 3·2 0·5 Moderate 

25 6 × 26 0·5 2·5 0·5 Moderate 

45 6 × 22 0·3 8 2·4 2·4 0·5 Moderate 

25 16 × 27 90 0·3 8 0·5 2·5 0·5 Moderate 

25 6 × 20 90 0·3 8 0·5 2·3 0·6 Moderate 

45 6 × 20 90 0·3 8 2·4 2·3 0·6 Moderate 

25 6 × 20 90 0·3 8 0·5 2·3 0·6 Moderate 

25 6 × 24 90 0·3 8 0·5 2·4 0·9 Moderate 

35 6 × 25 90 0·2 4·8 2·4 2·4 1·0 Moderate 

25 15 × 13 90 0·3 8 0·5 3·5 > 2·0 Caved visually, estimated < 2 m 

25 20 × 15 90 0·3 8 0·5 44 > 2·0 Caved visually, estimated < 2 m 

25 21 × 15 90 0·3 8 1·0 4·4 > 2·0 Caved visually, estimated < 2 m 

25 15 × 13 90 0·3 8 0·5 3·5 > 2·0 Caved visually, estimated < 2 m 

25 20 × 15 90 0·3 8 0·5 4·4 > 2·0 Caved visually, estimated < 2 m 

25 21 × 15 90 0·3 8 0·5 3·8 > 2·0 Caved visually, estimated < 2 m 

25 21 × 10 90 0·3 8 0·5 3·5 1·5 Caved visually, estimated < 2 m 

25 22 × 14 90 0·3 8 0·5 4·4 1·5 Caved visually, estimated < 2 m 

25 23 × 12 90 0·3 8 0·5 4·4 1·5 Caved visually, estimated < 2 m 

25 21 × 10 90 0·3 8 0·5 3·5 1·5 Failed visually, estimated 1–2 m 

25 22 × 14 90 0·3 8 0·5 4·3 1·5 Failed visually, estimated 1–2 m 

25 23 × 12 90 0·3 8 0·5 3·8 1·5 Failed visually, estimated 1–2 m 

25 19 × 13 90 0·3 8 0·5 3· 1·5 Failed visually, estimated 1–2 m 

25 19 × 13 0·3 8 0·5 3·8 1·5 Failed visually, estimated 1–2 m 

25 22 × 15 0·3 8 0·5 4·4 1·5 Failed visually, estimated 1–2 m 

25 19 × 13 0·3 8 0·5 3·9 1·5 Failed visually, estimated 1–2 m 

25 19 × 13 0·3 8 0·5 3·8 1·5 Failed visually, estimated 1–2 m 

25 22 × 15 0·3 8 0·5 4·4 1·5 Failed visually, estimated 1–2 m 

Mine 6 

45 2·6 4·4 < 0·5 Typical stope 

45 2·6 6·2 1·8 Caved stope 

all instances, stability was qualitatively assessed as 

either being stable, potentially unstable, or caved. 

Research by Mah14 and Clark and Pakalnis6 at the 

University of British Columbia augmented the 

stability graph by using stope surveys in which cavity 

monitoring systems were employed. Mah’s work 

added 96 data points onto Matthews’s stability graph 

under mining conditions whereas Clark5 added an 

additional 88 data points. This research has enabled 

quantification of the amount of wall slough. A 

parameter defined by Clark5 as the equivalent linear 

overbreak/slough (ELOS; Fig. 5) was used to express 

volumetric measurements of overbreak as an average 

depth over the entire stope surface. ELOS is defined as 
Mining Technology (Trans. Inst. Min. Metall. A) March 2005 Vol. 114 A17

Brady, Martin, Pakalnis Empirical approaches for opening design in weak rock masses 

5 Stability graph 

the volume of slough from the stope surface divided 

by the product of stope height times wall strike length 

known as the hydraulic radius (HR). 

volum� o� Slough ELOS = (1) 

The stability graph relates hydraulic radius (HR) of 

the stope wall to empirical estimates of overbreak 

slough. Additionally, the database for the Matthews’ 

method has been augmented by the addition of 400 

case histories which includes open stoping experiences 

for a broad range of rock mass conditions in Australia 

specifically at the Mount Charlotte mine. The 

additional case histories include much larger stopes 

and extend the modified stability graph to a hydraulic 

radius of 55 m as compared to the previous maximum 

value of 20 m.16 The drawback of this study was that it 

was largely qualitative because surveys were not 

available to measure the wall slough. 

A limited number of observations existed for 

RMR76 values under 45% (Fig. 3B). An additional 45 

data points were added on the stability graph–nonentry from Nevada operations having an RMR76 

under 45% (Fig. 3B and Table 3). In addition, mine 4 

reflects over 338 observations that have been averaged 

to reflect the design points shown in Table 2. The 

stability graph relates hydraulic radius of the stope 

wall to empirical estimates of overbreak slough. 

Hydraulic radius (HR) is defined as the surface area 

of an opening divided by the perimeter of the exposed 

wall, analysed as shown in Figure 5. 

N′ = Q′ × A × B × C (2) 

where N′ is the modified stability number; Q′ is the 

Norwegian Geological Institute (NGI) rock quality 

index (after Barton et al.2

), with stress reduction 

factor (SRF) and Jw = 1·0); A is high stress reduction; 

B is orientation of discontinuities; and C is the 

orientation of surface. 

The stress reduction factor and joint water 

reduction factor are equal to 1, as they are accounted 

for separately within the analysis. For the ELOS graph 

points, the database for the stability graph was derived 

from mining operations that are generally dry. 

The following relationship was used to convert 

RMR to Q′ (from Bieniawski3

): 

RMR = 9LnQ′ + 44 (3) 

The A factor accounts for the influence of high 

stresses that reduce rock mass stability and is 

determined by the ratio of unconfined compressive 

strength of intact rock to maximum induced stress 

parallel to the opening surface. It is set to 1·0 if intact 

rock strength is 10 or more times induced stress, which 

indicates that high stress is not a problem. It is set to 

0·1 if rock strength is two times induced stress or less, 

which indicates that high stress significantly reduces 

opening stability. In the mines visited for this study, 

the value of A was equal to 1·0 because the hanging 

wall was largely in a relaxed state. 

The B factor looks at the influence of the orientation 

of discontinuities with respect to the surface analysed and 

states that joints oriented 90° to a surface do not create 

stability problems and the B factor is assessed a value of 

1·0. Discontinuities dipping up to 20° to the surface are 

the least stable and represent geological structures that 

can topple. In this case, the B factor is equal to 0·2, which 

was the value used for the Nevada database. In extremely 

weak rock masses (RMR76 = 25%), the material largely 

resembles an amorphous mass with geological structures 

throughout; therefore, reduction due to jointing is 

suspect. The authors are presently analysing the data to 

augment this factor. 

The C factor considers orientation of the surface 

being analysed and is assigned a value of 8 for the design 

of vertical walls and a value of 2 for horizontal backs. 

The C factor reflects the inherently more stable nature of 

vertical walls compared to a horizontal back. In this 

paper, the ELOS curves employ a value for C of 8·0 for 

the footwalls. For a more complete explanation of 

factors A, B, and C refer to papers by Potvin19 and Clark 

and Pakalnis.6 

SUPPORT GUIDELINES 

The development of support capacity guidelines is 

critical to the overall success of the mining method 

selected in terms of ensuring a safe work place (see Fig. 2). 

Ground support in weak rock presents special 

challenges. Under-design can lead to costly failures, 

whereas over-design can lead to high costs for 

unnecessary ground support. Figure 6 depicts a classic 

wedge failure controlled by structure if the ground 

support has been under-designed. It is critical to design 

for the dead weight of the wedge in terms of the breaking 

load of the support, as well as the bond strength 

associated with embedment length.8 Split-Set and 

Swellex have been defined as continuous friction coupled 

(CFC) non-grouted bolts by CSIRO.24 The load transfer 

mechanism between the rock and borehole depends on 

the borehole characteristics (hole diameter and 
A18 Mining Technology (Trans. Inst. Min. Metall. A) March 2005 Vol. 114

6 

Brady, Martin, Pakalnis Empirical approaches for opening design in weak rock masses 

Classic wedge failures 

roughness) compared to the outside diameter of the bolt 

element. Independent tests conducted by NIOSH,12 

Tomory 20 25 et al., and Villaescusa et al. confirm that 

relationship for all diameter of bolts – 33-mm, 39-mm 

and 46-mm Split-Sets. 

Over 400 000 Split-Set8 friction bolts are used in 

Nevada mines for primary support. Friction bolts are 

particularly useful in fissile, buckling, or sheared 

ground where it is difficult to secure a point anchor. 

Caution must be used when using this method of 

primary support because of the low bond strength 

between the weak rock mass and the bolt and because 

of the susceptibility of the bolt to corrosion. In mine 

4, Split-Set bolts had a life of 6 months because of 

corrosion resulting from acidic ground conditions. An 

analysis of the performance of friction bolts in mines 

with weak rock (as determined by RMR) needed to be 

addressed. With one exception, Nevada mines use 39­

mm Split-Set bolts (the exception uses 46-mm SplitSet bolts); however, mines in Canada, generally use 

33-mm Split-Set bolts. Canadian mines normally use 

these bolts only in the walls and not in the back. Table 

4 shows an updated support capacity chart as 

augmented by this study. 

Data points gathered from several pull tests in 

weak rock were plotted as shown in Figure 7. A neural 

Table 4 Nevada mines database: back spans in weak rock 

Rock properties (t)
 

Bolt strength Yield strength Breaking strength
 

5/8-inch mechanical 6·1 10·2 

Split-Set (SS 33) 8·5 10·6 

Split Set (SS 39) 12·7 14·0 

Standard Swellex NA 11·0 

Yielding Swellex NA 9·5 

Super Swellex NA 22·0 

*20-mm rebar, No. 6 12·4 18·5 

*22-mm rebar, No. 7 16·0 23 

*25-mm rebar, No. 8 20·5 30·8 

No. 6 Dywidag 11·9 18·0 

No. 7 Dywidag 16·3 24·5 

No. 8 Dywidag 21·5 32·3 

No. 9 Dywidag 27·2 40·9 

No. 10 Dywidag 34·6 52·0 

1/2-inch cable bolt 15·9 18·8 

5/8-inch cable bolt 21·6 25·5 

1/4 × 4-inch strap 25·0 39·0 

Note: No. 6 gauge = 6/8-inch diameter; No. 7 gauge = 7/8-inch 

diameter; No. 8 gauge = 1-inch diameter; NA = Not applicable. 

Screen Bag strength (t) 

4 × 4-inch welded mesh, 4 gauge 3·6 

4 × 4-inch welded mesh, 6 gauge 3·3 

4 × 4-inch welded mesh, 9 gauge 1·9 

4 × 2-inch welded mesh, 12 gauge 1·4 

2-inch chain link, 11 gauge, bare metal 2·9 

2-inch chain link, 11 gauge, galvanised 1·7 

2-inch chain link, 9 gauge, bare metal 3·7 

2-inch chain link, 9 gauge, galvanised 3·2 

Note: 4 gauge = 0·23-inch diameter; 6 gauge = 0·20-inch
 

diameter; 9 gauge = 0·16-inch diameter; 11 gauge = 0·125-inch
 

diameter; 12 gauge = 0·11-inch diameter. 


Shotcrete shear strength = 2 MPa (200 t m–2).
 

Bond strength 

Split-Set, hard rock 0·75–1·5 Mt per 0·3 m 

Split-Set, weak ground 0·25–1·2 Mt per 0·3 m 

Swellex, hard rock 2·70–4·6 Mt per 0·3 m 

Swellex, weak rock 3-3·5 Mt per 0·3 m 

Super Swellex, weak rock > 4 Mt per 0·3 m 

5/8-inch cable bolt, hard rock 26 Mt per 1 m 

No. 6 rebar, hard rock 18 Mt per 0·3 m, ~12-inch 

granite 

net23 was superimposed on the mine data to determine 

if trends or predictions could be made. The neural net 

methodology has been used in establishing design 

curves for span and stope design by Wang et al.

22 The 

graph shows a strong trend between RMR and bond 

strength; this relationship is being assessed as part of 

on-going research. Preliminary results are shown in 

Figures 7 and 8. 

7 Rock mass rating versus pull-out strength. A neural 

trend line is superimposed 

Variability in test results shows the difficulty in 

assessing overall support for a given heading. Thus, it 

is important that mines develop a database with 

respect to the support used so as to design for variable 

ground conditions. Factors critical to design, such as 

bond strength, hole size, support type, bond length, 

and RMR, should be recorded so as to determine 

their influence on the design curve to be determined. 
Mining Technology (Trans. Inst. Min. Metall. A) March 2005 Vol. 114 A19

Brady, Martin, Pakalnis Empirical approaches for opening design in weak rock masses 

8 Pull-out load versus RMR at mine 4 

CONCLUSIONS 

The Spokane Research Laboratory and the University 

of British Columbia Geomechanics group are 

focusing on the development of safe and cost-effective 

underground design guidelines for weak rock masses 

having an RMR in the range 15–45%. Weak ground 

conditions, ground support, and mining methods used 

in several Nevada underground mines were observed. 

The RMR values were calculated to update both span 

design calculations and stability graphs, and a 

database on underhand mining methods was developed 

to reflect existing Nevada mining conditions. The 

immediate rock mass was also characterised and 

analysed in terms of prevailing type of ground support, 

potential failure mechanisms, and rock behaviour. 

Variability in field conditions shows the difficulty in 

assessing overall support for a given heading. It is 

imperative that mines develop their own databases 

based on the type of support used in their mines so 

unexpected ground conditions can be analysed. The 

results from augmented design curves and pull-out tests 

are presented in the hope that they will aid mine 

professionals in the task of designing a safe workplace. 

A systematic approach allows an operator to understand overall failure mechanisms and resultant loads 

that could affect the system. This approach would allow 

an engineer to develop an optimal support strategy for 

the mining method employed. 

The work would not have been possible without the 

partnership between NIOSH, the University of British 

Columbia Geomechanics Group and the Nevada gold 

mining company personnel. This continued partnership 

is critical to the development of safe and cost effective 

mine strategies. Figure 2 shows that since the inception 

of the ‘team’ approach and resultant collaboration that 

injury statistics have declined dramatically. The cause 

and effect may be a result of many factors; however, it is 

clear that this approach is important and relevant to 

mine operations. 

REFERENCES 

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site characterization and tunnel design’, Int. J Rock 

Mech. Min. Sci., 2002, 39, 185–216. 

2.	 N. BARTON, R. LIEN and J. LUNDE: ‘Engineering 

classifications of rock masses for the design of tunnel 

support’, Rock Mech., 1974, 6, 189–236. 

3.	 Z. T. BIENIAWSKI: ‘Rock nass classifications in rock 

engineering’, Proc. Symposium on Exploration for Rock 

Engineering, Johannesburg, 1976. 97–106. 

4.	 Z. T. BIENIAWSKI: ‘The aeromechanics classification in 

rock engineering applications’, Proc. 4th International 

Congress on Rock Mechanics, Vol. 2, Rotterdam A. A. 

Balkema, 1979, 41–48. 

5.	 L. CLARK: ‘Minimizing dilution in open stope mining 

with focus on stope design and narrow vein longhole 

blasting’, Unpublished MSc thesis, University of British 

Columbia, Vancouver, 1998. 

6.	 L. CLARK and R. PAKALNIS: ‘An empirical design 

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stope hangingwalls and footwalls’, 99th Annual 

AGM–CIM conference, Vancouver, 1997. 

7.	 R. A. CUMMINGS, F. S. KENDORSKI and Z. T.BIENIAWSKI: 

‘Caving mine rock mass classification and support 

estimation’, US Bureau of Mines Contract Report, 

J0100103, Washington DC. 

8.	 J. M. GORIS: Personal communication, Thiessen Team 

USA Inc., February 20, 2003. 

9.	 T. HOCH: Ground control seminar for underground gold 

mines, Elko, NV, September 20, Sponsored by MSHA 

Ground Control Division, Pittsburgh, PA, 2000. 

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Authors 

Tom Brady is a Senior Scientist with the Spokane Research 

Laboratory, formerly of the U.S. Bureau of Mines and now 

part of National Institute for Occupational Safety and 

Health (NIOSH). He works on health and safety issues for 

underground mining with his main focus or empirical mine 

design for metal and nonmetal mines. Recently, Tom Brady 

was selected as a Henry Crumb lecturer by SME. 

Lewis Martin is a Mechanical Engineer with Spokane Research 

Laboratory. He works on safety issues for under-ground 

mining with emphasis on development of equipment. Lewis 

holds several patents for instruments which monitor 

underground loads and movements. 

Rimas Pakalnis is the Rock Mechanics Professor at the 

Department of Mining Engineering at the University of 

British Columbia.