DESIGN IN WEAK ROCK MASSES: 

NEVADA UNDERGROUND MINING OPERATIONS 

Tom Brady, 

Spokane Research Laboratory, National Institute for Occupational Safety and Health, 

Spokane, WA USA 

Rimas Pakalnis 

University of British Columbia, Vancouver, BC, Canada 

Lyndon Clark 

Barrick Goldstrike, Elko, NV USA 

Abstract 

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 data on weak rock masses into 

existing design relationships, with an emphasis on updating 

the span design curve for manned entries and the overbreak 

curve for longhole entries. Both curves were originally 

developed at the University of British Columbia, Vancouver, 

BC. The original database has been augmented by information 

from mines throughout the United States, Canada, Australia, 

and Europe. The common factor in all these mines is the 

presence of a weak back and/or walls. In most cases, the ore 

zone is the weakest rock unit and must be stabilized so that 

the mineral-bearing rock can be extracted safety. The current 

NIOSH research attempts to provide rock mechanics tools to 

assist a mine operator in making economic decisions that will 

also ensure a safe working environment. This paper 

documents the Nevada database with a special emphasis on 

Nevada underground gold mines. 

Introduction 

Many of the underground gold mines in Nevada are 

found in very weak ground that creates difficult and 

hazardous mining conditions. A comparative analysis by the 

Mine Safety and Health Administration (MSHA) for the years 

1990 through 2004 (figure 1) shows that the number of 

injuries from roof falls in 13 Nevada underground gold mines 

has varied from a low of eight in 1990 to a high of 28 in both 

1995 and 1997 (3). 

Figure 1.—Injuries from rockfalls in Nevada underground mining 

operations (after MSHA [3]) 

This high injury rate was the prime 

motivator for initiation of the present study by NIOSH. The 

goal is to address the extremely difficult ground conditions 

associated with mining in a weak rock mass and provide mine 

operators with a database that could lead to a better 

understanding of the mechanics of these conditions. 

1
 
Figure 2 shows schematically that existing databases are 

based largely on stronger, less-fractured or altered rock 

masses than those found in Nevada mines.

Figure 2.—Schematic showing jointed rock masses associated with 

Nevada 

 While design 

methods for excavating in a weak rock mass do exist (2), they 

are geared to caving operations. 

The primary mining method used in Nevada is a form of 

mechanized cut-and-fill. This method is amenable to 

achieving a high rate of recovery when mining irregular ore 

geometries. Variations on this method are practiced at 

different mines and within different areas of the same 

operation, depending upon rock quality. Underhand cut-andfill mining (3) is employed and larger spans are required 

under a weak back where the costs of installing roof support 

would be high. For purposes of this study, a rock mass is 

considered to be weak if it has a rock mass rating (RMR) (4) 

of less than 45% and/or a rock mass quality rating (Q) under 

1.0. 

In areas where rock quality allows, the mines employ a 

version of longhole open stoping (5). Where spans from the 

footwall to the hanging wall exceed 7.6 m, transverse open 

stoping is employed with the use of primaries and secondaries 

adjacent to previously cemented rockfill panels. 

Underground mining methods as practiced in Nevada 

dictated which specific databases and stope design curves 

NIOSH would focus upon. Rock mass values were calculated 

during mine visits and varied from an RMR high of 70% and 

a low of 16% in gold-bearing fault gouge. 

Several rock mass design curves developed by the rock 

mechanics group at the University of British Columbia (6) are 

available, but they were not thought to be relevant to the 

mining methods employed within the weak ground of Nevada 

gold mines and therefore were not updated for weak ground. 

Research commenced with visits to Nevada operators in 

June 1999 to address concerns and determine where NIOSH 

would be able to assist. The first technical site visit was on 

June 12, 2002, and initial data were collected (figure 1). The 

major objectives were to obtain information on weak rock 

masses and incorporate this information into existing design 

curves (7) for back spans of manned entries and a stability 

graph (8) for longhole wall design. 

The distribution of the original databases was based on 

Canadian mining data, as summarized in figure 3, and shows 

how few data exist for weaker rock masses. 

Figure 3.— Distribution of original database for back span design 

(top) (7) and stability graph (bottom) (8) 

Span Design, Man Entry 

2
 

The initial span curve was developed by the 

geomechanics group at the University of British Columbia to 

evaluate back stability in cut-and-fill mines. It consists of two 

straight lines that divide a graph into three zones: stable, 

potentially unstable, and unstable. The database for this graph 

initially consisted of 172 data points from the Detour Lake 

Mine of Placer Dome, Inc., in Ontario, with most of the points 

having RMR values in excess of 60% (7). The database was 

expanded to 292 observations in the year 2000 with case 

histories from an additional six mines (9). 

The successful use of empirical design techniques is 

based upon interpolation rather than extrapolation. Thus a 

decision was made to develop a database for a critical span 

curve in weak rock masses. The term “critical span” refers to 
the largest circle that can be drawn within the boundaries of 

the excavation when seen in plan view (figure 4). 

Figure 4.—Critical span curve adjusted to Nevada operations 

The term 

“design span” refers to spans that have no support and/or 

spans incorporating a limited amount of local support (for 

example, pattern bolting in which 1.8-m-long mechanical 

bolts are installed on a on a 1.2- by 1.2-m pattern). 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. The new database included RMR 

values from 24% to 87% with 63% of the cases over 60%. 

Less than 10% of the RMR values fell below 45%, and less 

than 20% fell below 55% (10). An additional 44 observations 

were added to the critical span curve from Nevada operations 

(figure 4; table 1), of which 35 had an RMR less than 45%. 

Table 1.—Nevada mines database: Back spans in weak rock 

RMR (%) Span (m) Condition Other 

Mine 1 

45 

45 

40 

5.5 

9.0 

6.0 

S 

S 

S 

Stable with support 

Stable with support 

Stable with support 

Mine 2 

40 

45 

30 

4.0 

4.3 

3.7 

S 

U 

S 

Stable with support 

Caved with support 

Stable with support 

Mine 3 

40 

45 

26 

25 

55 

45 

7.0 

2.1 

2.1 

4.6 

7.6 

3.0 

S 

S 

S 

U 

S 

S 

Stable with support 

Stable with support 

Stable with support 

Caved with support 

Stable with support 

Stable with support 

Mine 4 

70 

40 

25 

55 

30 

30 

45 

50 

70 

25 

30 

30 

50 

55 

55 

4.6 

4.6 

4.6 

5.5 

6.1 

6.1 

4.6 

6.1 

11.3 

7.3 

3.0 

1.8 

6.1 

7.6 

6.1 

S 

S 

S 

S 

S 

S 

S 

S 

S 

S 

U 

S 

S 

S 

S 

Stable with support 

Stable with support 

Stable with support 

Stable with support 

Caved upon longhole 

Caved upon longhole 

Stable with support 

Stable with support 

Stable with support 

Stable with support 

Prior to support placement 

Prior to support placement 

Stable with support 

Stable with support 

Stale with support 

Mine 5

30 

30 

20 

15 

3.0 

4.3 

5.8 

3.7 

S 

U 

U 

S 

Stable with support 

Caved with support 

Caved with support 

Stable with support 

Mine 6 

45 

40 

40 

4.3 

6.1 

4.9 

S 

U 

S 

Stable with support 

Caved with support 

Stable with support 

Mine 7 

40 

35 

4.6 

4.6 

S 

S 

Stable with support 

Stable with support 

Mine 8 

25 

20 

25 

35 

55 

35 

20 

45 

5.0 

1.2 

2.4 

3.1 

3.7 

4.6 

7.6 

6.0 

U 

S 

S 

S 

S 

S 

U 

S 

Caved. Had to spile 

No. support. Maximum 

round possible

No support. Maximum 

round possible

No support. Maximum 

round possible 

No support. Typical round. 

No problems 

No support. Typical round. 

No spile/shotcrete

Caved. Had to spile. 

Stable with Split-Sets only 

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 (6). 

Excavation stability is classified into three categories; 

each category is further divided into three subcategories. 

1.	 Stable excavation (S) 

a. No uncontrolled falls of ground have occurred. 

b.	 No movement of the back has been observed. 

c. No extraordinary support measures have been 

employed. 

2.	 Potentially unstable excavation. 

a. Extra ground support has been installed to prevent 

falls of ground. 

b.	 Movement has occurred in the back. 

c. Increased frequency of ground movement has been 

observed. 

3. Unstable excavation (U) 

a. Area has collapsed. 

b.	 Depth of failure of the back is 0.5 times the span (in 

the absence of major structures). Within a weak 

rock mass, the depth of failure has been noted as one 

times the span and sometimes even greater. 

c. Limited local support was not effective in 

maintaining stability. 

3
 
A minus-10 correction factor is applied to the final RMR 

when evaluating rock with shallow-dipping or flat joints. 

However, the applicability of this factor in weak ground is 

being reassessed because of its amorphous nature. Where 

discrete ground wedges have been 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-hour period has also been defined as a 

critical amount of movement for safe access (6). This value is 

also being addressed for weak rock masses as it applies to the 

initial database identified in figure 2. This critical value may 

be much greater than 1 mm. 

Stability Graph Method–Nonentry 

 The original stability method for open stope design was 

based largely on Canadian operations and was proposed in 

1981 by Mathews (11), modified in 1988 by Potvin (12), and 

updated in 1992 by Nickson (13). In all instances, stability 

was qualitatively assessed as either being stable, potentially 

unstable, or caved. Recent research at the University of 

British Columbia has augmented the stability graph by using 

stope surveys in which cavity monitoring systems were employed (8). This research has enabled the amount of dilution 

to be quantified. A parameter termed the "equivalent linear 

overbreak/slough" (ELOS) was introduced by Clark (8) and 

was used to express volumetric measurements of overbreak as 

an average depth over an entire stope surface. This has 

resulted in a design curve as shown in figure 5. 

Figure 5.—Stability graph (after 8) 

A limited number of observations existed for RMR 

values under 45% (figure 3-bottom). An additional 45 data 

points were added on the stability graph–nonentry from 

Nevada operations having an RMR under 45% (figure 6; table 

2). 

Figure 6.—Wall stability graph as developed for Nevada operations 

4
 

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 

is defined as the surface area of an opening divided by 

perimeter of the exposed wall being analyzed. 

The following equation was employed for calculation of 

parameters for the database shown in figure 5. 

 N’ = Q’ * A * B * C 

where N’ = modified stability number, 

 Q’ = modified NGI rock quality index (14) 

where the stress reduction factor and 

joint water reduction factor are equal to 

1, as they are accounted for separately 

within the analysis, 

 A = stress factor equal to 1.0 due to relaxed 

hanging wall, 

 B = rock defect factor. This value results 

from parallel jointing and the amorphous 

state of the weak rock mass being set to 

0.2 and 0.3, respectively, as in table 3. 

and C = stope orientation factor as defined in 

figure 5, that is, C = 8 - 6 × cos φ (dip of 

hanging wall). 
Table 2.—Nevada mines database: Stope spans and walls in weak rock. All values of A = 1. 

5 

RMR (%) Dimensions, height by 

length (m) 

Dip B C N HR (m) ELOS (m) Comments 

Mine 1: 

45 

40 

55 

39 

25 

34 

42 

20 by 17 

20 by 16 

49 by 18 

34 by 34 

90 

90 

90 

90 

90 

90 

90 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

8 

8 

8 

8 

8 

8 

8 

2.7 

1.5 

8.1 

1.4 

0.3 

0.3 

0.3 

4.6 

4.4 

6.6 

8.5 

1.8 

3.4 

2.8 

<1.0 

2.0 

1.0 

4.6 

<1.0 

<1.0 

<1.0 

<m of ELOS 

<1 m of ELOS stable (estimated) 

<1 m of ELOS stable (estimated) 

<1 m of ELOS stable (estimated) 

Mine 2: 

40 

50 

11 by 21 

11 by 21 

90 

90 

0.3 

0.3 

8 

8 

1.5 

4.7 

3.4 

3.4 

<1.0 

<0.5 

<1 m of ELOS 

<0.5 m of ELOS 

Mine 3: 

55 

26 

18 by 18 

12 by 18 

70 

70 

0.2 

0.3 

5.9 

5.9 

4.0 

0.2 

4.6 

3.7 

0.6 

>2.0 >2 m of ELOS 

Mine 4: 

25 

25 

25 

25 

55 

45 

45 

45 

45 

25 

25 

45 

25 

25 

45 

25 

25 

35 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

25 

6 by 29 

8 by 36 

17 by 12 

21 by 15 

30 by 26 

6 by 25 

18 by 12 

19 by 16 

6 by 22 

16 by 5 

6 by 26 

6 by 22 

16 by 27 

6 by 20 

6 by 20 

6 by 20 

6 by 24 

6 by 25 

15 by 13 

20 by 15 

21 by 15 

15 by 13 

20 by 15 

21 by 15 

21 by 10 

22 by 14 

23 by 12 

21 by 10 

22 by 14 

23 by 12 

19 by 13 

19 by 13 

22 by 15 

19 by 13 

19 by 13 

22 by 15 

55 

90 

90 

90 

90 

90 

90 

90 

90 

 

 

 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

90 

 

 

 

 

 

0.2 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.2 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

0.3 

4.5 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

4.8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

8 

0.2 

0.5 

0.5 

0.5 

7.2 

2.4 

2.4 

2.4 

2.4 

0.5 

0.5 

2.4 

0.5 

0.5 

2.4 

0.5 

0.5 

2.4 

0.5 

0.5 

1.0 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

0.5 

2.5 

2.4 

3.5 

4.5 

7.0 

2.5 

3.5 

4.4 

2.4 

3.2 

2.5 

2.4 

2.5 

2.3 

2.3 

2.3 

2.4 

2.4 

3.5 

44 

4.4 

3.5 

4.4 

3.8 

3.5 

4.4 

4.4 

3.5 

4.3 

3.8 

3. 

3.8 

4.4 

3.9 

3.8 

4.4 

0.3 

0.1 

0.1 

0.1 

0.1 

0.1 

0.1 

0.1 

0.5 

0.5 

0.5 

0.5 

0.5 

0.6 

0.6 

0.6 

0.9 

1.0 

>2.0 

>2.0 

>2.0 

>2.0 

>2.0 

>2.0 

1.5 

1.5 

1.5 

1.5 

1.5 

1.5 

1.5 

1.5 

1.5 

1.5 

1.5 

1.5 

Cluster average, height/width 

Cluster average, rib 

Cluster average, rib 

Cluster average, rib 

Cluster average, rib 

Cluster average, height/width 

Cluster average, rib 

Cluster average, rib 

Moderate 

Moderate 

Moderate 

Moderate 

Moderate 

Moderate 

Moderate 

Moderate 

Moderate 

Moderate 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Caved visually, estimated < 2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Failed visually, estimated 1-2 m 

Mine 6 

45 

45 

2.6 

2.6 

4.4 

6.2 

<0.5 

1.8 

Typical stope 

Caved stope 
 

 

Table 3.—Nevada mines database: Underhand cut-and-fill 

Mine Amount of cement Span (m) Sill thickness Unconfined compressive Comments 

(%) (m) strength (MPa) 

1 10.0 6.1 3 2.0 Paste 

2a 6.5 7.6 4.6 5.5 Cemented rockfill 

2b 8.0 9.1 4.6 6.9 Cemented rockfill design 

2c 8.0 21.0 4.6 6.9 Mined remotely. No cave. 

3 70 3.0 3. 4-12 Cemented rockfill 

4a 9.0 13.7 4.0 8.3 Cemented rockfill test 

panel 

4b 9.0 3.7 3.0 8.3 Cemented rockfill drift and 

fill 

4c 9.0 7.3 3.0 8.3 Cemented rockfill panel 

5 *7.0 2.7 3.0 3.4 Cemented rockfill 

6 6.75 4.9 4.3 3.4 Cemented rockfill 

7a 10.0 1.8 2.7 0.3 Paste (factor of safety = 

1.5) 

7b 2.4 2.7 0.5 

7c 3.0 2.7 0.7 

7d 3.7 2.7 1.0 

7e 4.3 2.7 1.4 7-day sample 

7f 4.9 2.7 1.8 

7g 5.5 2.7 2.3 

7h 6.1 2.7 2.9 

8 7.0 4.6-6.1 4.6 5.5 Cemented rockfill 

9 10.0 5.0 5.0 4.45 Cemented rockfill 

10 12.8 6-9 6.0 2.0 High-density slurry (78% 

solids by weight) 

11 10.0 3.0 3 (includes 0.9- 2.5 (after 7 days) 10% cemented hydraulic 

m air gap) fill (73%-75% solids by 

weight) 

12 8.0 2.4-4.6 3 (includes 0.6-

m air gap) 

4.8 8% paste (no free water) 

* Includes fly ash binder. 

An initial observation from figure 6 is that the classical 

design curves (ELOS) as shown in figure 5 are inaccurate at 

low N' and hydraulic radius values. If hydraulic radius is kept 

below 3.5 m in a weak rock mass, the ELOS value should 

remain under 1 m. It appears a hydraulic radius under 3 m 

would not result in ELOS values much greater than 1 m. This 

result is being further evaluated. 

Underhand Mine Design 

As noted earlier, most mines use a form of cut-and-fill 

mining for a major portion of their production. As ground 

conditions become weaker, the primary method is mechanized 

underhand cut-and-fill. These methods assure a high degree of 

gold recovery under an engineered back of cemented rockfill 

and/or cemented paste fill (3). To achieve the most production 

at the lowest cost while maintaining a safe work environment, 

mine personnel were interested in developing support 

methods for the back that would best utilize either cemented 

rockfill or cemented paste fill. The result was the development 

of a new empirical database of successful underhand mining 

scenarios for the weak ground of Nevada. 

A major study is also underway at the University of 

British Columbia to develop design guidelines for mining 

under paste backfill. Figure 7 is adapted from Stone (15) and 

is largely based on fixed-beam bending as the critical failure 

mechanism. The guidelines relate material properties to the 

thickness of the beam, the exposed span, and the resultant 

unconfined compressive strength required. A factor of safety 

in excess of 2.0 is incorporated. The analytical relationship 

was augmented by underhand cut-and-fill and shows that 

actual mine design parameters generally exceed those 

required by simple beam theory. The database is shown in 

figure 7 and summarized in table 3. 

6
 
Figure 7.—Underhand mine design curve relating span, sill thickness, 

and fill strength 

Figure 8.—Structurally controlled wedge 

Support Capacity 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 figure 2). Ground support 

in weak rock presents special challenges. Underdesign can 

lead to costly failures, whereas overdesign can lead to high 

costs for unneeded ground support. Figure 8 depicts a classic 

wedge failure controlled by structure. 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 (10). 

Over 400,000 Split-Set (10) friction bolts are used in 

Nevada mines as 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 broken rock 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 Split-Set bolts); mines in Canada, 

however, use 33-mm Split-Set bolts. Canadian mines 

generally 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 9. 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 figure 10. 

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 

they can 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 

enable where they lie on the design curve to be determined. 

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 

of 15% to 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 characterized and analyzed in terms of prevailing 

type of ground support, potential failure mechanisms, and 

rock behavior. 

Variability in field conditions show 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 analyzed. 

7
 
 

 

 

 

 

 

Table 4.—Nevada mines database: Back spans in weak rock 

Rock properties, tonnes Screen Bag strength, tonnes 

Bolt strength Yield strength Breaking strength 4- by 4-in welded mesh, 4 gauge 3.6

5/8-in mechanical 6.1 10.2 4- by 4-in welded mesh, 6 gauge 3.3 

Split-Set (SS 33) 8.5 10.6 4- by 4-in welded mesh, 9 gauge 1.9 

Split Set (SS 39) 12.7 14.0 4- by 2-in welded mesh, 12 gauge 1.4 

Standard Swellex NA 11.0 2-in chain link, 11 gauge, bare metal 2.9 

Yielding Swellex NA 9.5 2-in chain link, 11 gauge, galvanized 1.7 

Super Swellex NA 22.o 2-in chain link, 9 gauge, bare metal 3.7 

*20-mm rebar, No. 6 12.4 18.5 2-in chain link, 9 gauge, galvanized 3.2 

*22-mm rebar, No. 7 16.o 23 

*25-mm rebar, No. 8 20.5 30.8 Note: 4 gauge = 0.23-in diameter; 6 gauge = 0.20-in diameter; 

No. 6 Dywidag 11.9 18.0 9 gauge = 0.16-in diameter; 11 gauge = 0.125-in diameter; 12 

No. 7 Dywidag 16.3 24.5 gauge = 0.11-in diameter 

No. 8 Dywidag 21.5 32.3 Shotcrete shear strength = 2 MPa (200 t/m2

) 

No. 9 Dywidag 27.2 40.9 Bond strength 

No. 10 Dywidag 34.6 52.0 Split-Set, hard rock 0.75-1.5 mt per 0.3 m 

1/2-in cable bolt 15.9 18.8 Split-Set, weak ground 0.25-1.2 mt per 0.3 m 

5/8-in cable bolt 21.6 25.5 Swellex, hard rock 2.70-4.6 mt per 0.3 m 

1/4 by 4-in strap 25.o 39.0 Swellex, weak rock 3-3.5 mt per 0.3 m 

Super Swellex, weak rock >4 mt per 0.3 m 

5/8-in cable bolt, hard rock 26 mt per 1 m 

No. 6 rebar, hard rock 18 mt per 0.3 m, ~12-in 

granite 

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

in diameter.; No. 8 gauge = 1-in diameter. 

NA = Not applicable. 

Figure 9.— Pull-out load versus RMR for SS39. Figure 10.— Pull-out load versus RMR at mine 4 

The results from augmented design curves and pull-out 

tests are presented in the hope that they will aid mine 

professionals in their 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 Nevada gold mining company 

personnel. This continued partnership is critical to the 

development of safe and cost-effective mine strategies. Figure 

1 shows that since the inception of the team approach and 

resultant collaboration, injury statistics have declined dramatically. This decline may be a result of many factors; however, 

it is clear that this approach is important and relevant to mine 

operations. 

8
 
Acknowledgments 

Specific thanks go to Patrick Carroll, chief engineer, 

Newmont Midas Mine; J.J. Hunter, chief rock mechanics 

engineer, Newmont East Carlin and Deep Post Mine; Monica 

Dodd, chief engineer, Newmont Deep Post Mine; Simon 

Jackson, chief engineer, Placer Dome, Turquoise Ridge Joint 

Venture; Shawn Stickler, chief ground control engineer, 

Queenstake Murry Mine; and Rad Langston, senior rock 

mechanics engineer, Stillwater Mining Company. 

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