A Case Study Examination of Two Blast Rounds at a Nevada Gold Mine 

Ed Mchugh and Joel Warneke, National Institute of Occupational Safety and Health 

and 

Cristian Caceres, University of British Columbia 

ABSTRACT 

NIOSH researchers collected basic data from two blast rounds at the SSX-Steer Mine as a component of 

a larger study on controlled blasting, the goal of which is to reduce injuries from falling rocks in 

underground mines. Drilling and blasting procedures at the mine were observed. Rock mass property 

data were collected, a program of seismic monitoring and analysis was conducted, and 3-D laser survey 

scans of the workings before and after each blast were performed. The geologic data showed that the 

rock mass quality was characterized as poor to very poor. It was essentially uniform within the panels 

examined. The recorded seismic data indicated that fill material acted as an effective damper to seismic 

energy. The pre- and post-blast laser scans showed both over-break and under-break conditions, 

highlighting the problems associated with blasting a weak rock mass. By comparing the achieved 

results to the intended design, potential improvements in the drilling and blasting practices can be 

identified. Comparison of these data with similar data from other mines will help define how various 

combinations of rock and fill behave during blasting. Through studies such as these, data collection 

procedures and analysis techniques can be refined. An ultimate goal is to develop blast procedures tied 

to rock mass characteristics that can minimize damage to rock in the perimeter of the opening, limiting 

the amount of loose rock and improving the safety of miners. 

INTRODUCTION 

Falls of ground are a significant hazard for underground miners and a leading cause of fatalities. A 

majority of metal and non-metal mines use drill and blast techniques for drifting and primary 

development. After blasting, the broken rock is mucked out. This is followed by scaling or “barring 

down” to remove any loose rock from roof, walls and face. The stability of the workplace is then 

secured using rock bolts, mesh, shotcrete or other appropriate support systems. The entire process is then 

repeated. All of these activities are potentially hazardous to the miners that perform them. This hazard 

level is increased if care is not exercised in protecting the opening perimeter through the use of careful 

blasting practices. In particular, unwanted blast damage to the perimeter rock increases the hazard to 

those involved in the scaling and reinforcement tasks. It has been found that a majority of ground fall 

injuries involve blocks formed within the half-meter of rock immediately surrounding an excavation 

(Bauer et al., 1999; Mark and Iannachione, 2001). Whyatt et al (2003) summarized methods of 

assessing blast-induced damage in underground mines and discussed the feasibility of improving ground 

control results through the use of controlled blasting. Iverson et al (2007) outlined an approach for 

assessing perimeter blast damage based on rock mass quality, seismic response, and laser scanning. 

To assist industry in adapting good perimeter control blasting practices, the Spokane Research 

Laboratory (SRL), of the National Institute for Occupational Safety and Health (NIOSH), has 

implemented a program focused on developing practical, yet technically sound, perimeter blasting 

design procedures. The objective is to provide techniques so that the as-built openings closely resemble 

those designed. When applied, these will assist mine operators to create stable underground openings 
with a minimum amount of loose or damaged surrounding rock, thereby enhancing the safety of miners. 

The overall research and development program involves theoretical, laboratory, field and modeling 

studies. The results are being integrated into a user-friendly blast design software package. 

Field studies are an important component of the program. The purpose of the field studies conducted by 

NIOSH is to fully document the results of a number of blast rounds carried out under a wide range of 

rock conditions. The documentation includes: the blasting patterns, the actual versus the as-designed 

drillhole locations and orientations, the final excavation profile, the seismic records from the blast, the 

rock mass quality, the ground support, and a number of other parameters. The results of this study will 

provide valuable baseline information on fragmentation practices and add to data compiled on mining 

conditions in different geologic settings. This paper presents some preliminary results of studies carried 

out at the SSX-Steer Mine. 

STUDY SITE AND METHODS 

Background 

The SSX-Steer Mine is located about 60 km (40 mi) north of Elko, Nevada in the Jerritt Canyon mining 

district. Queenstake Resources Ltd., the sole owner of the mine during November 2006 when the 

NIOSH study was conducted, merged with YGC Resources, Ltd. in May 2007 to form the YukonNevada Gold Corp. (Yukon-Nevada Gold, 2007), the current owner and operator. 

Disseminated gold deposits were discovered at Jerritt Canyon in 1972. Open pit mining operations 

involving several small open pits began in 1981 (Queenstake, 2006). Underground mining began in 

1992 as the surface operations began to reach their economic limits (Ash, 2006). The SSX deposit 

(figure 1) was discovered in the early 1990s by following structural trends from adjacent properties. 

Accessed from a portal, it lies at depths ranging from 140 to 300 m (450 to 1000 ft) below surface. 

Mining commenced in 1997. 

 Figure 1. The SSX-Steer Complex showing 

generalized underground workings.

In 2004, the Steer mine portal was collared from 

the old Steer pit. A drift connecting the Steer

mine to the SSX deposit was completed in the 

latter half of 2005. Since that time, the two mines 

have been operated as a single unit referred to as 

the SSX-Steer Complex. Gold mineralization at 

the Steer deposit has been identified in an area 

stretching eastward from the old Steer pit to

halfway along the connection drift to the SSX

deposit, a length of approximately 910 m (3000 

ft). The connection drift facilitated mining in the 

Steer deposit by providing ventilation, secondary 

access, and shared infrastructure. The present

study was conducted in the Steer portion of the 

complex known as Zone 7. Specifically,

information was collected from one blast round 

each in crosscuts XC11 (November 7, 2006) and 

XC7 (November 9, 2006) on the 7170 level

(figure 2). The XC11 and XC7 drifts were driven 
to remove ore remaining between 

primary extraction drifts. After 

mining, these primary drifts were 

filled using cemented fill. Access for 

this study was by way of the Steer 

portal. 

Figure 2. Plan map of Zone 7-7170 level, SSX-Steer Mine 

showing the locations of XC11 and XC7, the seismic monitoring 

stations and the rock classification sections. 

Geology and Rock Mass 

Classification 

The host rock for the gold 

mineralization is the Hanson Creek 

Formation (hc3, Silurian and 

Ordovician), which consists of 

black, fine-grained, carbonaceous 

limestone (Muntean and Henry, 

2006). The limestone has thin, wavy 

laminations and commonly contains 

pods and lenses of black chert. 

Brecciation and stringers of white 

calcite, realgar, and orpiment are 

typical. 

The strength of the rock mass is a 

critical factor in determining the 

stability of the openings. Two 

empirical methods, the Q system and 

the rock mass rating (RMR) system, 

were used to classify the integrity of 

the rock mass. Both systems use 

geologic, geometric and engineering 

parameters for classification. The 

RMR system involves five 

parameters. These include intact 

rock strength, drill core quality 

(RQD), spacing of discontinuities, 

condition of discontinuities, and 

ground water conditions. There is an 

adjustment for discontinuity orientation. The RMR value based upon these factors is expressed as a 

score that ranges from less than 25 for the worst rock conditions to 100 for the best possible rock 

(Bieniawski, 1976). The Q system involves six parameters to represent rock block size, interblock shear 

strength, and the active stress. The Q system ratings are logarithmic-scaled values that range from 0.001 

for exceptionally poor ground to 1000 for exceptionally good, practically un-jointed ground (Barton et 

al, 1974). For this study, the rock mass was classified using both methods and was applied to the two 

ribs and the back at intervals of approximately 3.7 m (12 ft) along the drift. 
Seismic Monitoring 

The seismic signals generated by each blast were recorded using two Instantel MiniMate Plus seismic 

monitors. Each monitor was connected to one or two triaxial geophones with channels to collect vertical, 

longitudinal, and transverse vibrations. Thus, each monitor collected and stored up to six channels of 

information. The monitors, which were set to a sampling rate of 8192/s, would begin recording when 

triggered by a motion exceeding 10 mm/s (0.39 

in/s). The geophones were mounted on 

aluminum brackets and anchored to the ribs of 

the entries (figure 3) with steel masonry bolts. 

Figure 3. Installing a triaxial transducer and 

bracket on the rib in XC 11. 

Each geophone was then connected to a monitor 

with shielded copper cables. The monitors 

themselves were hung from the steel mesh along 

the ribs to avoid damage from blast effects and 

from mining machinery during the mucking and 

bolting cycles. Where blast effects were difficult 

to avoid, a steel-plate box that could protect one 

monitor and one geophone (figure 4) was used. 

Figure 4. Seismic monitor C installed in an 

armored steel box in XC7 

Seismic data were analyzed using Instantel 

Blastware software (version 4.37). The software 

generates waveforms, selects peak values, and 

calculates statistical parameters. One key value 

is the peak particle velocity (PPV). Particle 

velocity was recorded for each of the three axes 

at each point in time sampled. The PPV is taken 

to be the largest of the three values for each 

vibration event. The vector sum is calculated by 

combining the velocities for each of the three 

axes for each data sample. The peak vector sum 

(PVS) is the maximum resultant particle 

velocity. The PVS may or may not coincide 

with any of the peak particle velocities for the 

individual channels. 

Laser Scanning 

Three-dimensional survey scans for the two 

blast rounds in the XC7 and XC11 ore panels 

were conducted using a Leica Geosystems HDS 

3000 laser system. A laser beam reflected from 

any surface in the beam path (rock, fill, 

equipment, etc.) is used to calculate the distance 

and direction to that surface. The result is a 

three-dimensional point cloud representing the 

size and shape of the excavation. It has an 

accuracy of 6 mm over a range of 50 m (0.23 in 
at 164 ft). The scan density was set to a 

sample grid of 7.5 mm vertical x 5 mm 

horizontal (0.3 in x 0.2 in) at a range of 6 m 

(20 ft). Each scan was completed within 

approximately 45 minutes. 

Two scans were required to assess blast 

round results. The first scan provided the 

pre-blast configuration of the drilled face 

including the back, ribs, and blast-hole 

collar locations. The second scan provided a 

survey of the excavation after the round was 

Figure 5. Plan view of XC 11, showing the original face shot and mucked (figure 5).

and the area extracted by the shot on 11/7/06. 

 Successive laser 

scans were merged into a single threedimensional survey using the associated 

software. Point cloud accuracy was improved by placing survey targets within the scanning area. 

Several 15-cm (6-inch) diameter spherical plastic targets designed by NIOSH were mounted along the 

ribs. The targets were screwed into reusable holes that had been drilled and tapped in rock bolt plates. 

The targets were removed during blasting to prevent damage and remounted for subsequent scans. 

Short lengths of PVC pipe were inserted into the blast holes as scan aids so that the laser scanner could 

accurately capture collar locations as well as hole orientations. Software would later be used to estimate 

the 3D location of each drill hole in space to its full depth. 

XC11 BLAST ROUND 

The XC11 entry was the end ore panel on the 7170 level. Panel number 10 containing cemented fill was 

located along the left side while the right side was formed by rock that will not be mined (see figure 2). 

The design for the November 7 blast is shown in figure 6. The blast holes were 2.2 cm (1-7/8 in) in 

diameter and the relief holes were 7.6 cm (3 in) in diameter. All holes were drilled to a depth of 3.7 m 

(12 ft). Perimeter holes were drilled along the back and upper left rib on nominal 0.3 m (1 ft) centers in 

an attempt to limit the extent of breakage. These were not charged with explosives. Except for the lifters, 

all blast holes were charged with ANFO. The lifter holes were charged with sticks of Unigel. 

The rock mass quality was evaluated for six sectors in the XC11 ore panel. Due to the presence of 

bedding-planes and crosscutting fractures, the exposed bedrock appeared highly fractured. The fractures 

were predominately north-trending with shallow inclinations. In addition, west-trending fractures with 

shallow dips to the south, and steeper (75-81° NE) fractures with northwest strikes were exposed. The 

south wall and back had a uniform RMR rating of 35 for all six sectors evaluated in XC11. Q ratings 

were a consistent 0.55 in the back and in sectors 1 through 4 in the south wall. The rock in sectors 5 and 

6 was even more highly fractured than the rest of the crosscut and Q values in the south wall were rated 

at 0.34. Fill in the north wall was not rated. Based upon the Q values and RMR ratings, the rock in the 

XC11 ore panel was classified as very poor. 
Figure 6. XC11 blast pattern diagram, November 7, 2006. 

expected, the peak 

velocities recorded at 

geophone A1 located 

9.8 m (32 ft) from the 

blast face, were 

significantly higher than 

those recorded at 

geophone A2, which 

was 16.2 m (53 ft) from 

the face. The seismic 

waveforms recorded by 

monitor B revealed 

velocities that were 

clearly attenuated by the 

presence of the fill. For 

geophone B1, located 

14.0 m (46 ft) from the 

blast face, the peaks for 

several blast delay 

intervals were muchreduced or essentially 

missing when compared 

to the signals from 

geophone B2 (19.5 m or 

64 ft from the face) 

attached to rock. 

The PPVs for each 

channel of geophone A1 

were near or exceeded 

100 mm/s (3.9 in/s) as 

can be seen in table 1. 

The peak velocities for 

the three channels 

occurred at three 

different blast delay 

intervals, suggesting a

A seismic monitor serving two geophones was installed on each wall of the entry for the November 7 

round in the XC11 ore panel (see figure 2). Both geophones (A1 and A2) located along the right hand 

wall were attached to solid rock. On the left wall, one geophone (B1) was attached to the fill wall of 

panel 10 and the second geophone (B2) was attached to rock (panel 9 between fill panels 8 and 10). 

The Minimate monitors were triggered by the blast at 17:27 and began sampling. The seismic 

waveforms from monitor A showed peaks for each delay in the blast pattern over a duration of about 7 

seconds (figure 7). As 
relatively uniform response to the blast 

pattern. Farther from the blast face at A2, 

the velocities were under 100 mm/s (3.9 

in/s) and concentrated at one time 

interval. On the left wall, all but one PPV 

was associated with the first hole to be 

shot in the burn. The exception was the 

transverse sensor, where the PPV was 

associated with the fifth timed hole (delay 

#4). The PVS value for the first hole shot 

was 74% higher in the rock panel 

geophone than in the fill panel geophone, 

although the fill panel was nearer to the 

blast face.The pre-blast laser scan 

(November 7) showed that the overall 

panel width was 1.8 m (6 ft) or more 

wider than the planned drift blast. The 

remaining ore between the left drift wall 

and the fill had been removed during 

previous blast/muck cycles (figure 8). The 

designed face would be about 18.2 m2 

(196 ft2), but the pre-blast face was about 

27.9 m2 (300 ft2). 

Figure 7. Seismic waveforms from XC 11. a- particle velocity vs time for the duration of the blast 

(detonator delays 0-15), Monitor A1, Transverse channel; b- particle velocity vs time (detonator 

delays 0-9) showing higher particle velocities in the A1 Transverse channel compared to A2 

Transverse channel; c- particle velocity sum (three channels) vs time (detonator delays 0-9) showing 

relatively higher velocities in rock (B2) compared to fill (B1). 

Table 1 Seismic data summary for the blast in the XC11 ore 

panel (November 7, 2006), showing PPV for each channel, 

their PVS at each geophone, and the timing delay of peak 

velocities. 

Geophone 

(Distance from 

face, ft) 

Channel PPV/PVS 

(mm/s) 

Time 

(s) 

A1 (32) Transverse 102 0.015 

 Vertical 99.1 3.201 

 Longitudinal 171 2.051 

Vector Sum (PVS) 190 2.052 

A2 (53) Transverse 41.9 0.332 

 Vertical 39.4 0.321 

 Longitudinal 88.9 0.321 

Vector Sum (PVS) 192 2.052 

B1 (46) Transverse 24.1 0.031 

 Vertical 17.8 0.004 

 Longitudinal 39.4 0.005 

Vector Sum (PVS) 39.6 0.005 

B2 (64) Transverse 57.1 1.440 

 Vertical 35.6 0.069 

 Longitudinal 61.0 0.044 

Vector Sum (PVS) 69.0 0.044 
The plan-view projections of the blast 

holes are shown in figure 9 based on 

the combined pre-blast and post-blast 

scans. The drill pattern had been 

positioned to minimize over-break on 

the left rib with uncharged perimeter 

holes intended to limit over-break 

along the fill as had occurred in 

previous rounds. The line of 

uncharged holes placed along the left 

rib were still present after the blast. 

The post-blast configuration indicates 

that rock broke well beyond the ends 

of the blast holes on the center-left 

side. 

Figure 8. Laser scan collected November 7, 2006 (face view) in 

the XC11 ore panel showing old rock bolts hanging from the 

fill as well as new mesh (upper left side), the painted blast 

hole pattern, and the scan-aid spheres (right side). 

Figure 9. Plan view of the XC 11 ore panel blast round 

(November 7, 2006), showing pre- and post-blast face 

configurations and blast holes as projected from collar 

orientations. Uncharged holes are shown outside the 

resulting excavation 

Figure 10a is a vertical, longitudinal 

section taken through the post-blast 

scan. There is considerable breakage 

above the roof holes near the blasted 

face. The cross-section areas as 

determined using slices taken parallel 

to the face on one foot intervals are 

indicated in table 2. At the 1.5-m (5-ft) 

distance, the area is about what was 

desired. Figure 10c is vertical section 

at taken at mid-round. The projected 

positions of the holes are shown with 

distances from the charged holes to the 

final profile. Along the back, the 

distances from the uncharged holes to 

the final perimeter have also been 

determined. The line of uncharged roof 

holes was apparently not effective in 

controlling the profile. For both walls, 

the final position seemed to correspond 

quite closely to the positions of the 

charged holes. 
Figure 10. A. Vertical longitudinal section through XC 11 from before and after laser scans; showing 

projected drill pattern and position of the mid-round cross-section. B. Outline of the workings showing 

location of sections. C. Vertical slice at mid-round showing drill hole intercepts with distances to the ribs 

and back, solid points were loaded holes and open points were unloaded perimeter holes. 

Table 2. Area of the round taken at 1-ft slices from the face. 

Distance from Face (ft) 1 2 3 4 5 6 7 8 9 10 11 12 13 14 

Area (sq ft) 279 252 229 204 183 166 156 138 137 125 113 88 55 31 

XC7 BLAST ROUND 

As indicated earlier, XC7 was intended to extract the ore pillar remaining between previously mined and 

filled panels 6 and 8. The November 9 blast round was not a standard blast round but was intended to 

turn a corner between the two pillars of fill. Hence, the blast holes were angled to the right from the 

collars shown in the drill plan (figure 11). The hole lengths indicated for each vertical row of holes 

decreased from 3.7 m (12 ft) near the left rib to 1.8 m (6 ft) on the right. The initial opening is a 

combination fan-and-burn cut. As with the XC11 blast round, the blast and relief hole diameters were 

2.2.cm (1-7/8 in) and (3 in), respectively and the same explosive types were used. A third seismic 

monitor (C) equipped with one geophone was attached to the rock on the right rib of the entry into the 

XC7 ore panel (see figure 2). One monitor (B) remained in the left rib of the XC11 ore panel with one 

geophone (B1) in fill (panel 10) and the second geophone (B2) in solid rock (panel 9 between fill panels 

8 and 10). 
Figure 11. XC7 ore panel blast pattern, November 9, 2006 

The rock mass was characterized based on observations made at a single sector in the XC7 ore panel. 

Since cemented fill composed both ribs at the drill face, rock exposure was limited. Steeply dipping (81° 

NE), northwest striking fractures were present. Rock exposed in the back had an RMR of 45 and Q of 

2.42. Both the Q and RMR values 

for the XC7 ore panel indicated 

that the rock mass strength was 

notably better than that in the 

XC11 ore panel and was classified 

as poor to fair. 

The blast triggered seismic 

sampling at 17:21. The seismic 

waveforms from monitor C 

showed peaks for each delay in the 

blast pattern over a period of about 

7 seconds (figure 12). Peak 

particle velocities at geophone C1, 

6.7 m (22 ft) from the blast face, 

were sharply defined and well 

dispersed throughout the blast 

sequence. Waveforms from 

monitor B demonstrated velocities 

that were again attenuated in the 

fill (B1) but somewhat better 

defined in the rock panel (B2). For 

the monitor nearest to the blast 

face (C), both PPV and PVS 

approached or exceeded 100 mm/s 

(3.9 in/s) as shown in table 3. 

Monitor B (located on the opposite 

side of the rock panel) recorded 

particle velocities that barely 

reached the 10 mm/s (0.39 in/s) 

trigger threshold. 

The pre-blast laser scan in the XC7 

ore panel included drill hole collar 

locations and hole orientations 

(figure 13). The post-blast scan 

survey showed that the drilling 

pattern and loading plan was 

effective in excavating the corner 

near the right rib fill with minimal 

over-break into the fill. Some 

over-break did occur along the left 

rib rock/fill contact. 
Figure 12. Seismic waveforms from XC 7. a- particle velocity vs time for the duration of the blast 

(detonator delays 0-15), Monitor C1, Transverse channel; b- particle velocity vs time (detonator 

delays 0-9) showing higher particle velocities in the C1 Transverse channel compared to B2 

Transverse channel; c- particle velocity sum (three channels) vs time (detonator delays 0-9) showing 

relatively higher velocities in rock (B2) compared to fill (B1). 

Table 3. Seismic data summary for the blast in the XC7 ore panel, 

November 9, 2006, showing PPV for each channel, their PVS at 

each geophone, and the timing delay of peak velocities. 

Geophone 

(Distance from face, ft) 

Channel PPV/PVS 

(mm/s) 

Time 

(s) 

B1 (46) Transverse 10.2 1.951 

 Vertical 7.62 1.041 

 Longitudinal 12.7 0.001 

Vector Sum (PVS) 14.0 0.010 

B2 (41) Transverse 15.2 0.038 

 Vertical 8.89 0.033 

 Longitudinal 12.7 0.033 

Vector Sum (PVS) 18.9 0.037 

C1 (22) Transverse 90.2 6.394 

 Vertical 102 3.881 

 Longitudinal 92.7 2.556 

Vector Sum (PVS) 154 3.881 

DISCUSSION AND CONCLUSIONS 

The two blast rounds monitored at the SSX-Steer mine were part of their secondary recovery program. 

XC11 was located on the southeast side of the mining section. The panel immediately to the northwest 

of this panel had been previously mined and filled. The 4.3-m (14-ft) wide drift was being driven
 

leaving a 1.8-m (6-ft) wide rib pillar on the left hand side. The function of this rib was to limit damage 
to the fill and dilution during the

mucking operation. In previous

rounds, both ore and fill had broken

out so that the final overall room width 

was about 6 m (20 ft). 

Figure 13. Plan view of XC7 ore panel blast round with 

projected drill holes and pre- and post-blast face 

configurations. 

XC7 was being driven to extract the 

3.7-m (12-ft) wide pillar remaining

between ore panels 6 and 8. Again, the 

idea was to leave an initial skin of rock 

near the fill walls to limit the potential 

for dilution. Since the round was one 

of the first in the overall extraction 

process, it was not a standard design. 

In both cases, the rock mass was

classified as poor to very poor. 

Although the rock quality was

essentially uniform within the panels 

examined, the rock newly exposed by 

both blast rounds showed signs of

larger block sizes and potentially

greater hazards from rock falls. The 

fill material was relatively friable but 

stood up well where exposed in

secondary drifts. Seismic data indicated that the fill material acted as an effective damper to seismic 

energy. This may have helped to limit blast damage to the remaining rock panels. 

A more detailed look at the results of the standard blast round in XC11 can provide some insight 

regarding drill-blast performance. The post-blast longitudinal and plan sections show that the excavated 

opening tapered down from the drilled face, which was 53% larger than designed, to about the designed 

area at the midpoint of the round, and to much less than designed (about 45%) at the drilled depth of 3.7 

m (12 ft). 

Figure 14 is a cross-section of the round generated by slicing a 0.30-m (1-ft) section of the laser scan. 

The drill hole locations projected to mid-round from the hole collars at the drilled face were used to 

align the designed blast round with the resulting excavation. Deviations from ideal positions for these 

locations reflect both the actual drilled deviations and errors in projecting the drill holes from the scan 

aids used to collect hole orientation. 
Figure 14. Cross section of the XC11 round at 

midpoint, showing over-break and under-break in 

relation to drill holes. 

Although the total cross-sectional area at 

mid-round was approximately as

designed, the laser section at that point 

shows that over-break in the back

accounted for 15.4% of the designed 

area. Over-break ranged from about 18 

to 78 cm (7 to 31 in) in the back and 

averaged 59 cm (23 in) for a total of 2.8 

m2

 (30.2 ft2). Less significant overbreak along the right rib, based on 

perimeter hole locations, amounted to 

about 0.9 m2

 (9.3 ft2), with a range of 20 

to 28 cm (8 to 11 in), and an average of 

24 cm (9 in) or 4.7% of the designed 

area. At the same time, under-break 

occurred along the left rib as shown by 

the retained, unloaded perimeter holes. 

The under-break amounted to about

2.1m2

 (22.4 ft2), with a range of 32 to 55 

cm (13 to 22 in), and an average of 46 

cm (18 in) or 11.4% of design. The 

remaining apparent under-break, about 

10%, was in mostly broken material on 

the floor of the mucked opening. 

The unloaded perimeter holes across the 

back of XC11 failed to prevent overbreak in the blast round examined. The 

closely spaced line of holes across the 

back appear to have been accurately 

drilled, but an arched configuration for 

these holes might have helped to retain more of the back. Gravity certainly played a role in the failure 

as it acted on crosscutting fractures in the very poor quality rock mass where scaling of loose rock is 

likely to extend beyond design limits. At the same time, the unloaded holes along the left rib were still 

standing after the blast, indicating that they were also ineffective in controlling the final configuration. 

The blast round in XC7, with a relatively complex pattern in a tight space, was successful in continuing 

the excavation with only small amounts of over-break. 

SUMMARY 

Two blast round surveys were completed at the XC11 and XC7 ore panels in the Zone 7–7170 level. 

Rock quality was classified as very poor and poor for the XC11 and XC7 ore panels, respectively. PPV 

measurements near the blasts were much lower in the fill than in the rock. The laser surveys showed 

both over-break and under-break conditions in the XC11 ore panel. Generally, the laser surveys illustrate 

the problems associated with blasting a weak rock mass. Study of additional blast rounds can be used to 

document changes in rock mass characteristics or to verify correlations between blasting methods and 

wall rock damage. Data from successive blast round surveys would provide a basis for the design and 
testing of new blast designs. Expanding on the use of laser data will allow more detailed comparison 

between blast designs and final configuration of the resulting excavations. 

Seismic data and blasting outcome case studies will be used to calibrate new models for predicting blast 

damage to the perimeter of underground openings. Drilling and blasting methods that minimize blast 

damage will reduce the hazards of loose, broken rock underground and enhance the safety of miners. 

ACKNOWLEDGEMENTS 

Sam Ash, mine manager, and Jennifer Haas, mine engineer, arranged for access to the mine workings 

and assisted with data collection during the drilling and blasting cycles. Russ Kinsman, trainer and EMT 

coordinator, provided site-specific safety training that allowed entry to the test sites and the use of a 

NIOSH-owned vehicle to transport personnel and equipment. Drillers and powder crews contributed to 

documentation of drilling and loading procedures. 

The findings and conclusions in this report have not been formally disseminated by the National Institute for 

Occupational Safety and Health and should not be construed to represent any agency determination or policy. 

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