RI 9661 

Report of Investigations/2004 

Geophysical Methods to Detect 

Stress in Underground Mines 

Department of Health and Human Services

Centers for Disease Control and Prevention 

National Institute for Occupational Safety and Health 
Report of Investigations 9661 

Geophysical Methods to Detect Stress 

in Underground Mines 

Douglas F. Scott, Theodore J. Williams, Douglas Tesarik, 

David K. Denton, Steven J. Knoll, and John Jordan 

U.S. DEPARTMENT OF HEALTH AND HUMAN SERVICES
 Centers for Disease Control and Prevention
 

National Institute for Occupational Safety and Health
 

Spokane Research Laboratory
 

Spokane, WA
 

March 2004 
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DHHS (NIOSH) Publication No. 2004-133 
CONTENTS
 

Abstract . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 

Page 

1 
 

Introduction ................................................................................... 2 
 

Validation of seismic tomography . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2 
 

Site geology . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2 
 

Ultrasonic core velocity measurements . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3 
 

Finite-difference analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3 
 

Seismic wave analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5 
 

Seismic refraction survey . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5 
 

Discussion . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6 
 

Investigation of electromagnetic emissions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8 
 

Background ................................................................................. 8 
 

Previous studies . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8 
 

EM source mechanisms . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 9 
 

Earthquakes .............................................................................. 9 
 

Laboratory tests . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 9 
 

Theoretical work . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 10 
 

Underground . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 10 
 

Technical approach . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11 
 

Results . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12 
 

Waveform identification . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12 
 

Downhole antenna without filters . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13 
 

Downhole antenna with filters . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13 
 

Suspended antenna without filters . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13 
 

Suspended antenna with filters . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13 
 

Discussion . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13 
 

Conclusions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 14 
 

Acknowledgments . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 14 
 

References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 15 
 

Appendix A: Pressure cell installation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 17 
 

Appendix B: Slot development . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 18 
 

ILLUSTRATIONS 

1. Location of Edgar Experimental Mine . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2 
 

2. Plan view of Edgar Mine pillar ............................................................... 3 
 

3. Geology of Edgar Mine .................................................................... 4 
 

4. P-wave velocities .......................................................................... 4 
 

5 Creating a source of seismic waves . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5 
 

6. Recording seismic waves . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5 
 

7. Source locations . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6 
 

8. Receiver locations ......................................................................... 6 
 

9. Tomogram taken after flatjack installation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 7 
 

10. Tomogram taken at pressure cell reading of 20.7 MPa . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 7 
 

11. Tomogram comparing differences between figure 12 and figure 11 . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8 
 

12. Location of Galena Mine . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 9 
 

13. Data acquisition system setup . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11 
 

14. Plan view of downhole antenna location . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11 
 

15. Waveform generated by air door . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12 
 

16. Waveform generated by barbeque lighter . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12 
 

17. Possible EM waveform . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13 
 

18. Summary of waveforms collected . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 14 
 
ILLUSTRATIONS–Continued 

A1. Installed pressure cell . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 

Page 

17 
 

A2. Hydraulic jack connected to pressure cell . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 17 
 

TABLES 

1. P-wave velocities from core samples ........................................................... 5 
 

2. Material properties used in finite-difference analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6 
 

3. Summary of 1998 P-wave seismic data ......................................................... 6 
 

4. Attenuation of EM emissions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12 
 

UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT 

cm centimeter MPa megapascal 

cm3 cubic centimeter mV millivolt 

cm2 square centimeter V volt 

g/cm3 gram per cubic centimeter in inch 

Hz hertz in2 square inch 

kHz kilohertz in3 cubic inch 

kg kilogram ft foot 

km kilometer ft/s foot per second 

km/s kilometer per second lb pound 

m meter lb/ft3 pound per cubic foot 

MHz megahertz mi mile 

m/sec meter per second psi pound per square inch 

mm millimeter 
GEOPHYSICAL METHODS TO DETECT STRESS IN UNDERGROUND MINES 

By Douglas F. Scott,1

1

Geologist, Spokane Research Laboratory, National Institute for Occupational Safety and Health, Spokane, WA.
 

Theodore J. Williams,2 David K. Denton,2
 

2

Mining engineer, Spokane Research Laboratory, National Institute for Occupational Safety and Health, Spokane, WA.
 

Douglas Tesarik,3

3

Mechanical engineer, Spokane Research Laboratory, National Institute for Occupational Safety and Health, Spokane, WA.
 

Steven J. Knoll,4

4

Rock burst technician, Galena Mine, Wallace, ID.
 

and John Jordan5
 

5

Mining engineer, Stillwater Mine, Nye, MT.
 

ABSTRACT

 Highly stressed rock in stopes continues to be a primary safety risk for miners in underground mines 

because this condition can result in failures of ground that lead to both injuries and death. Personnel from the 

Spokane Research Laboratory of the National Institute for Occupational Safety and Health studied two methods 

for identifying stress in rock. A seismic tomographic survey, finite-difference analysis, laboratory 

measurements of compression wave (ultrasonic) velocities in rock cores, and site geology were integrated to 

evaluate the use of seismic tomography for identifying induced pressures in an underground pillar at the Edgar 

Mine, Idaho Springs, CO. Electromagnetic (EM) emissions were also investigated in the Galena Mine, a deep 

underground mine in Idaho, in an effort to determine if these emissions could be used as indicators of 

impending catastrophic ground failure. 

Results of this research indicated that (1) seismic tomography appears to be a useful tool for determining 

relative stress in underground pillars, while (2) EM emissions do not appear to be significant precursors of 

impending catastrophic ground failure. 
2 

INTRODUCTION
 

This Report of Investigations describes two types of 

geophysical studies conducted by researchers from the Spokane 

Research Laboratory (SRL) of the National Institute for 

Occupational Safety and Health (NIOSH). In the first study, a 

series of measurements was collected from the Edgar Mine, Idaho 

Springs, CO, to determine if 

changes in seismic velocity correlated with stress changes in a rock 

mass subjected to known induced pressures. In the second study, 

measurements of electromagnetic (EM) emissions were collected 

at the Galena Mine, Wallace, ID, to determine whether these emissions were valid precursors to imminent ground failure. 

VALIDATION OF SEISMIC TOMOGRAPHY
 

Previous seismic tomographic studies (Friedel and others, 

1995a; Friedel and others, 1996, 1997; Jackson and others, 

1995a; Jackson and others, 1995b; Scott and others, 1997a; 

Scott and others, 1997b; Scott and others, 1998) have shown 

that seismic tomography can be used to determine relative 

stress. Friedel and others (1995b) explained the use of seismic 

tomography as follows: 

The increase in velocity is related to the closure of 

void space, e.g., pores and cracks. In general, the rate 

of velocity increase is nonlinear and greatest with an 

early incremental increase in stress. As stress increases 

further, the rate at which velocity increases is reduced 

in response to the formation of new cracks (yield 

point) parallel to loading. These observations suggest 

that regions of high velocity are likely to indicate 

zones of high stress concentration, whereas lowvelocity regions indicate zones of stress relief. 

However, to date, no researchers have attempted to validate 

the use of seismic tomography to map stress in an underground 

mine by comparing pressure changes (induced pressure) in a 

pillar to velocity changes in settings where geologic features, 

such as differing rock types, fractures, and stress, can influence 

velocity. Neither have induced stress, predictive models, ultrasonic core velocity measurements, and geology been integrated 

with each other to quantify the relationship between seismic 

tomography and stress. 

Researchers from the Spokane Research Laboratory (SRL) 

of the National Institute for Occupational Safety and Health 

(NIOSH) used different techniques to collect a series of 

measurements at the Edgar Mine, Idaho Springs, CO (figure 1), 

to see if changes in seismic velocity correlated with stress 

changes in a rock mass subjected to known induced pressures. 

Figure 1.–Location of Edgar Experimental Mine. 

The Edgar Mine is located about 55 km (34 miles) west of 

Denver, CO, and is owned and operated by the Colorado School 

of Mines (CSM), Golden, CO. The mine portal (Miami Tunnel) 

is at an elevation of 2,405 m (7,890 ft) (CSM, 2003). Workings 

consist of 305 m (1,000 ft) of crosscuts and drifts that access 

several silver-gold veins. Widths of crosscuts and drifts average 

about 3 m (10 ft), and overburden above the pillar tested is 

estimated to be 120 m (400 ft). The pillar studied is 43 m (140 ft) 

long and 23 m (75 ft) wide (figure 2). Access around the pillar 

is nearly complete, except for 8 m (26 ft) in the northeast part. 

The slot cut for the pressure cell (figure 2) is 0.9 m (3 ft) high 

and 0.2 m (0.5 ft) wide and is cut 3 m (10 ft) into the pillar. 

SITE GEOLOGY 

The mine is developed in Precambrian metamorphic and 

granitic rocks of the Colorado Front Range. Specifically, the 

rocks are assigned to the Idaho Springs Formation. Rocks in the 

pillar (figure 3) include biotite schist, biotite microcline pegmatite, biotite-hornblende schist, quartz-feldspar-biotite gneiss, 

and migmatized gneiss. The biotite schist is black to dark gray 

and well bedded and contains numerous pegmatite dikes up to 

5 mm (0.2 in) thick. The biotite microcline pegmatite is composed mostly of large, white feldspar crystals up to 3 cm (1.2 in) 

long with large pieces of biotite as much as 3 cm (1.2 in) across. 

The pegmatite is blocky in appearance and lacks fractures. The 

biotite-hornblende schist lacks visible bedding and is dark black, 

very hard, and massive. The quartz-feldspar-biotite gneiss is 

white-and-black banded with distinct layers of biotite, feldspar, 

gneiss, and quartz. Migmatized gneiss is dark black and has a 

mottled appearance; it is found adjacent to the pegmatite. Based 

on geologic mapping done during this study, the volume of 

rocks is estimated as quartz-feldspar-biotite gneiss, 35%; biotite 

microcline pegmatite, 30%; biotite schist, 25%; biotite-hornblende schist, 7%; and migmatized gneiss, 3%. From softest to 

hardest, rock hardnesses range from biotite schist, quartzfeldspar-biotite gneiss, biotite microcline pegmatite, biotitehornblende schist, and migmatized gneiss. The slot for the 

pressure cell was cut in the migmatized gneiss. 
3 

Figure 2.–Plan view of Edgar Mine pillar. 

The geology of the interior of the pillar is inferred on the 

basis of geologic mapping along openings around the pillar. The 

pillar is cut by two right-lateral, strike-slip faults that have 

displaced rocks laterally; these faults are oriented roughly N50°E 

and dip 55° northwest. They are filled with quartz, sulfides, 

breccia, and gouge. Fractures in the pillar were also mapped. 

ULTRASONIC CORE VELOCITY MEASUREMENTS 

Two sets of ultrasonic velocity tests were done in the laboratory 

on core samples EM-1, EM-2, EM-3, and EM-4. EM-1 is a darkto black-banded, biotite-hornblende schist containing several quartz 

veins less than 5 mm (0.2 in) long. EM-2 consists of a white, 

medium-grained pegmatite with feldspar clasts less than 5 mm 

(0.2 in) long. The sample lacks biotite, but contains about 5% 

epidote and less than 2% small pyrite clasts. EM-3 is a white pegmatite containing coarse feldspar clasts longer than 5 mm 

(0.2 in) and thin seams of biotite. EM-4 contains a dark, finegrained, biotite-hornblende schist with about 2% fine-grained pyrite. 

The first test was done by CSM personnel on all four 

samples under a condition of no load. Table 1 summarizes the 

P-wave velocities obtained. 

Table 1.–P-wave velocities from core samples. 

Sample 

number P-wave velocity Rock type 

EM-1 

EM-2 

EM-3 

EM-4 

. . . . 

. . . . 

. . . . 

. . . . 

5.42 km/s (17,782 ft/s) 

4.75 km/s (15,584 ft/s) 

4.72 km/s (15,485 ft/s) 

5.00 km/s (16,404 ft/s) 

Biotite-hornblende schist 

Pegmatite 

Pegmatite 

Biotite-hornblende schist 

The second test was completed by Terra-Tek,6

6

The mention of specific products and manufacturers does not imply 

endorsement by the National Institute for Occupational Safety and Health. 

Salt Lake 

City, UT, and consisted of compression tests on samples EM-3 

and EM-4. The cores were 5 cm (2 in) in diameter and 10 cm 

(4 in) long. Axial loads ranged from 1.4 to 52 MPa (200 to 

7,500 psi). Figure 4 shows axial stress differences and P-wave 

velocities. The higher velocities in the pegmatites could be 

explained by the fact that the mica in the schist may allow the 

rock to deform elastically when pressure is exerted. 

FINITE-DIFFERENCE ANALYSIS 

The test pillar was analyzed with the three-dimensional, 

finite-difference program Fast Lagrangian Analysis of Continua 

in Three Dimensions (FLAC3D) (Itasca, 1998) and a MohrCoulomb failure criterion. The purpose of this analysis was to 

compare the extent of the numerically calculated stress zone in 

the pillar to actual seismic velocity contours when horizontal 

stress was applied to the underground pillar by a pressure cell. 

The finite-difference mesh consisted of cubes with 0.3-m (1-ft-) 

sides. Material properties for the model were based on a limited 

number of unconfined compression tests on the biotite microcline pegmatite and biotite-hornblende schist, underground 
4 

observations, and textbook values (table 2) (Farmer, 1968; 

Goodman, 1980). The vein was given low material property 

values to simulate a discontinuity in the continuum (Goodman, 

1980). The modeling sequence consisted of applying initial 

stresses using gravity, running the model to equilibrium, fixing the 

vertical velocities at the top of the pillar and horizontal velocities 

at the northeast corner of the pillar, excavating the slot for the 

pressure cell, grouting the bottom and collar of the slot, applying 

20.68 MPa (3,000 psi) of pressure to the slot walls, and finally 

running the model to equilibrium again. 

Figure 3.—Geology of Edgar Mine. 

Figure 4.—P-wave velocities. 

Results from the finite-difference program indicate that horizontal stress in the rock at mid-height to the pressure cell decays 

to the in situ stress value at approximately 4.6 m (15 ft) on a line 

perpendicular to the longitudinal axis of the pressure cell. Similar 

results are obtained using Bouussinesq’s contours of equal stress 

produced by applying a load on a 0.9- by 0.9-m (3- by 3-ft) square 

foundation resting on a semi-infinite, homogeneous, isotropic, 

elastic solid (Sowers, 1979). These contours indicate that only 2% 

of the applied load was present 4.6 m (15 ft) from the surface. 
5 

SEISMIC WAVE ANALYSIS 

The method requires striking a source (figure 5) with a 4.5-kg 

(9.9-lb) sledgehammer fitted with a trigger connected to a 

seismograph and then recording the first arrival of the seismic 

wave at stations where receivers (geophones) have been installed. 

Figure 5.—Creating a source of seismic waves. 

Geophones are mounted to a rib using rock bolt plates that 

are drilled, tapped, and surveyed into the mine coordinate 

system. Each geophone is hooked to a seismic cable connected 

to the seismograph, and a two-pair shielded cable is used both 

for communication and as a trigger for the seismograph. A 

signal-stacking seismograph (figure 6) is used to record P-wave 

arrivals, and all seismic data are transferred from the 

seismograph to a personal computer (PC). 

Figure 6.–Recording seismic waves. 

 Travel times (first 

arrivals) are “picked” or selected from the seismic records. All 

receiver and source location coordinates (x,y,z) are input into a 

spreadsheet, and travel times are entered into a software 

program (GeoTomAn) for reconstruction. Finally, contouring 

software is used to smooth the tomograms and add text for final 

presentation and interpretation. 

Two seismic surveys were completed in November of 1998. 

Table 3 summarizes the pressures exerted on the pressure cell 

and the mean P-wave velocity for each survey. 

SEISMIC REFRACTION SURVEY 

Researchers from CSM (unpublished report) reported on a 

seismic refraction survey in the floor of the Edgar Mine 

workings that showed the floor to be disturbed to a depth of 

about 0.6 m (2 ft). Velocity equaled 0.83 km/s (2,727 ft/s). 

As depth increased, velocities increased to about 3.7 km/s 

(12,000 ft/s). 

Seismic data collected from boreholes by Jessop and others 

(1992) indicated the maximum velocity of shear waves was 3.5 

km/s (11,600 ft/s). Using a ratio of shear to compressional wave 

velocity of not less than 0.58 gave an upper limit of 6.1 km/s 

(20,000 ft/s). The upper limit was verified by measuring the 

velocity in an intact core cut as a sphere, which also indicated 

6.1 km/s (20,000 ft/s). 

Seismic P-waves in intact granite range from 5.5 to 6.0 km/s 

(18,044 to 19,685 ft/s) (Goodman, 1980). Based on these data 

and the work of Jessop and others (1992), any reading of a Pwave velocity greater than 6.1 km/s (20,000 ft/s) would be 

questionable. Global constraints placed on the 1998 seismic 

inversion were a minimum P-wave velocity of 3.1 km/s (10,000 

ft/s) and a maximum of 6.1 km/s (20,000 ft/s). 

Forty-two sources (figure 7) and 48 receivers (figure 8) 

were used to obtain a calculated wavelength of 4.9 m (16 ft) 

from an average P-wave velocity of 5.9 km/s (18,684 ft/s) from 

the two surveys. Source and receiver locations were spaced 

about 3 m (10 ft) apart, and global constraints for reconstruction 

were 3,000 to 6,000 km/s (~10,000 to ~20,000 ft/s). 

Figure 9 is a seismic P-wave tomogram taken after the slot 

had been cut and the pressure cell installed (X =110; Y = 257). 

It is obvious that the area defined by lower velocities had been 

de-stressed because of blasting and drilling. From the southwest 

part (X = 95; Y = 255) to the northeast part (X = 120; Y = 280) 

of the tomogram there appears to be a rough “break” in 

velocities, which coincides with a fault mapped along the 

southwest edge of the pillar and again along the northeast edge. 

Furthermore, there appears to be an area of high velocity from 

X = 110; Y = 260 to X = 100; Y = 275. Why the high-velocity 

area exists is not known; however, it could be a continuation of 

the very hard migmatized gneiss (the harder gneiss should have 

higher velocities). 
 

 

 

 

 

 

6 

Table 2.—Material properties used in finite-difference analysis 

Rock type 

Modulus of 

deformation, 

MPa (psi) 

Poisson’s 

ratio 

Angle of internal

friction, deg 

Cohesion, 

MPa (psi) 

Tensile 

strength,

MPa (psi) 

Density, g/cm3 

(lb/ft )3 

Biotite-hornblende 

schist . . . . . . . . . . . 

72,395 

(10,500,000) 0.14 35 50 (7250) 35 (5,076) 2.803 (175) 

Biotite microcline 

pegmatite . . . . . . . . 

48,667 

(7,058,600) 0.19 30 30 (4350) 20 (2900) 2.594 (162) 

Migmatite gneiss . . . . 43,115 

(6,253,300) 0.14 30 30 (4350) 20 (2900) 2.803 (175) 

Quartz-feldspar biotite

gneiss . . . . . . . . . . . 

35,365 

(5,129,300) 0.25 30 7 (1015) 5 (725) 2.947 (1840) 

Biotite schist . . . . . . . 22,063 

(3,200,000) 0.25 30 2.5 (363) 0.5 (973) 2.947 (184) 

Vein . . . . . . . . . . . . . 134 

(19,500) 0.30 30 0 0 2.082 (1300) 

Grout . . . . . . . . . . . . . 11,721 

(1,700,000) 0.25 35 11.5 (1670) 4.4 (640) 2.403 (150) 

Table 3.–Summary of 1998 P-wave seismic data 

Survey Pressure from pressure cell Mean P-wave velocity 

0k 0 5.6 km/s (18,373 ft/s) 

3k 20.7 Mpa (3,000 psi) 5.79 km/s (18,996 ft/s) 

Figure 7.—Source locations. 

Figure 8.—Receiver locations. 

Figure 10 is a seismic P-wave tomogram taken when the 

pressure cell was reading 20.7 MPa (3,000 psi). An area of very 

high velocity has developed near the northwest tip of the slot 

(X~107; Y~258), which, according to the finite-difference 

model, should be in tension and actually have lower velocities. 

Higher velocities can be seen extending to the northwest part of 

the pillar. Other areas of change in the tomogram would not be 

associated with the pressure cell because the finite-difference 

model predicted that the pressure exerted by the pressure cell 

would return to the 0k values 4.6 m (15 ft) from the cell. 

Figure 11 is a difference tomogram that compares the tomogram shown in figure 9 to the tomogram in figure 10. Two areas of 

velocity change can be seen, one just west of the pressure cell (X 

= 110;Y = 255), which shows an increase of as much as 2.4 km/s 

(8,000 ft/s), and one just east of the pressure cell (X =115;Y = 257), 

which shows a decrease of as much as 2.4 km/s (8,000 ft/s). 

DISCUSSION 

Stresses west and east of the pressure cell returned to 

background levels of 1.38 MPa (200 psi) within 4.6 m (15 ft) of 

the cell when the cell was loaded to 20.7 MPa (3,000 psi), indicating that the effects of induced stress were limited. However, velocities west of the cell increased by about 2.4 km/s 

(8,000 ft/s), while velocities east of the pressure cell decreased 

by 2.4 km/s (8,000 ft/s). It is hypothesized that the harder rock 

west of the pressure cell (a massive, unfractured biotite microcline pegmatite) retained stress, while the softer rock east of the 

cell (a massive, biotite schist with numerous fractures) was able 

to squeeze and absorb stress. However, this scenario is inconsistent with the work of Friedel and others (1997), who found 

that as pressure was exerted on fractures, the fractures closed 

and velocities increased. 
7 

Figure 9.—Tomogram taken after pressure cell installation. 

Figure 10.—Tomogram taken at pressure cell reading of 20.7 MPa. 
8 

Figure 11.—Tomogram comparing differences between figure 9 and figure 10. 

In summary, geologic structure (faults and some rock types) 

was reflected relatively well in the tomograms. At a small scale 

(9.2 m [30 ft]), stress changes were seen both in the tomograms 

and the finite-difference model. Ultrasonic velocity measurements on core specimens were similar to the velocity measurements seen in the tomograms.
 

Based on this research, the use of seismic tomography to 

identify stress gradients in an underground rock pillar is 

validated. Furthermore, the use of seismic tomography to detect
 

mine workings could be valuable in locating abandoned mine
 

openings in coal or hard rock.
 

INVESTIGATION OF ELECTROMAGNETIC EMISSIONS 

BACKGROUND 

Field studies to collect data on electromagnetic (EM) 

emissions at the Galena Mine, Wallace, ID (figure 12), were 

conducted to determine if such emissions were valid precursors 

to imminent ground failure. Identifying such an association 

could be useful as an indicator of potential rock failure in underground mines. 

Acoustic (including microseismic and seismic) monitoring 

of a rock mass to detect ground movement in deep underground 

mines has been done successfully for several years. Various 

earthquake researchers have noted that occasionally the number 

of EM emissions would increase prior to a large earthquake. 

EM emissions have also been proven to be associated with rock 

failure in controlled laboratory research; however, to date, EM 

emissions have not been used successfully to detect imminent 

ground failure in deep underground mines. 

PREVIOUS STUDIES 

Light emissions and low-frequency electrical phenomena 

associated with seismicity in underground mines and earthquakes 

have been reported for hundreds of years (Brady and Rowell, 

1986). Most of this work involved the identification of EM 

emissions as an indicator of earthquakes. EM frequencies in both 

earthquakes and laboratory tests range from 0.01 Hz to 30 MHz. 
9 

Figure 12.—Location of Galena Mine. 

EM Source Mechanisms 

•	 Brady and Rowell (1986) summarized four mechanisms 

that cause light to be emitted from fracturing rock: (1) rock 

fragments frictionally heated to incandescence, (2) electrostatic discharge produced by the deformation of piezoelectric minerals or charge separation on fractured surfaces, 

(3) plasmas produced by rapid and intense heating of rock 

material, and (4) excitation of the ambient atmosphere by 

particle (electrons or positive or negative ions) bombardment. Brady and Rowell concluded that the light emitted 

from test rocks in the laboratory was caused by excitation 

of the ambient atmosphere by particle bombardment. 

•	 Zi-qiang and others (1988) examined three sources of light: 

(1) heat radiation from friction, (2) electrostatic discharges 

produced by piezoelectric effects or charge separation on 

fractured surfaces, and (3) excitation of the ambient atmosphere by particle bombardment. Because light emissions 

were observed only at the moment when electrons struck air 

molecules, the authors concluded that the most likely source 

of EM emissions was excitation of the ambient atmosphere 

by particle bombardment. 

•	 Brady (1996) detected electrical signals in the frequency 

band of 900 to 5,000 Hz using both coil and monopole antennas. EM emissions coincided with the final failure of an 

unconfined rock sample. Brady then concluded that the lowfrequency signals (900 to 5,000 Hz) recorded after rock 

failure were caused by the rotation and vibration of charged 

rock fragments. His observation was consistent with the 

hypothesis that no low-frequency EM emissions should 

occur if fracturing was confined (thus making particle 

motion impossible). He noted that no emissions were evident at frequencies greater than 10 kHz and that emissions 

were evident only in the near field and not the far field. 

Earthquakes 

Martner and Sparks (1959) noted electrical potential prior 

to the arrival of seismic waves at the surface of the ground. 

About 30 minutes prior to the arrival of main earthquake shocks, 

Gokhberg and Morgounov (1982) recorded EM emissions at 

frequencies of 27, 81, and 1.5 kHz and 1.63 MHz. Later, 

Migunov and others (1984) documented EM emissions in the 

frequency range of 0.5 to 50 kHz that were associated with 

seismicity from earthquakes. Fujinawa and Takahashi (1990) 

observed EM emissions in the 0.01- to 12-Hz and 1- to 9-kHz 

frequency bands hours before earthquake activity in Ito, Japan. 

Fujinawa and Kumagai (1992) observed ultralow-frequency 

(0.01 to 0.6 Hz) to very low-frequency (1 to 3 kHz) electrical 

emissions before, during, and after volcanic eruptions. 

Laboratory Tests 

•	 Tuck and others (1976) tested a cube of quartzite coupled 

with a quartz crystal to determine piezoelectric emissions 

when a 0.5-kg (1.1-lb) hammer was used as a seismic 

source. They concluded that no piezoelectric fabric was 

found; therefore it would be difficult to use EM emissions 

for the exploration of ore bodies. 

•	 Nitsan (1977) fractured quartz crystals, tourmaline crystals, 

and quartz-bearing rocks and recorded EM emissions in the 

frequency range of 1 to 10 MHz. His interpretation of the 

source of the emissions was piezoelectricity. 

•	 Goncharov and others (1980) tested several large (0.55 by 

0.55 by 0.65 m [19.4 by 19.4 by 22.9 ft]) blocks of concrete 

containing pieces of granite by applying load and recording both EM and acoustic emissions as the concrete failed. 

They recognized the fundamental problem of 

simultaneously recording both EM and seismic emissions 

and concluded that the number of EM emissions decreased 

as their amplitude increased. They also found that the ratio 

of EM to acoustic emissions post-fracturing was 20:1. Prior 

to fracturing (initial loading), the ratio had been 7:1. 

•	 In 1981, Bishop studied piezoelectric effects in quartz-rich 

rocks. Using a laboratory-designed system, he attempted to 

prove that the axis of the quartz crystals was a factor in EM 

emissions. He found that a relationship existed between 

EM emissions and predictions of the c-axis orientation in 

quartz crystals. 

•	 Hanson and Rowell (1982) tested quartzite from the Galena 

Mine, Wallace, ID. EM emissions peaked sharply below 40 

kHz on three antennas, leading them to conclude that (1) 

fracture formation coincided with EM emissions, (2) EM 

emissions fell into a frequency range of less than 40 kHz, 

(3) EM emissions seemed to be directional, and (4) the 

amplitude of EM emissions seemed independent of stress, 

but not independent of stress drop. 

•	 Khatiashvili (1984) showed that as the size of fractured 

crystals increased, electrical potential also increased. 

•	 Ogawa and others (1985) used ball antennas covering the 

frequency range of 10 Hz to 100 kHz to measure EM 

emissions from crustal rocks broken in the laboratory. 

They found that sedimentary rocks containing less silica 

emitted less electromagnetism and concluded that the 

source mechanism for EM emissions was either contact or 

separating electrification and piezoelectrification. 
10 

•	 Zi-qiang and others (1988) fractured granite in the laboratory and found that the most intensive light pulse and 

acoustic emissions were recorded simultaneously at the 

moment of rock fracture. 

•	 Yamada and others (1989) also fractured granite in the 

laboratory and recorded EM emissions in the frequency 

range of 80 kHz to 1.2 MHz. They concluded that, based 

on their work, the source of EM emissions was not a 

piezoelectric effect, but was related to new surfaces created 

by cracks. 

•	 Weimin and others (1991) fractured quartz, limestone, and 

granite samples and reported that recorded EM emissions 

were a result of rock fractures. 

•	 Rabinovitch and others (1995) tested granite and recorded 

EM emissions in the 100-kHz range. They also documented EM frequencies of as much as 10 MHz in quartz 

porphyry. Two types of EM pulses were noted, “short” 

pulses of 1-3 microseconds and “lengthy” pulses of more 

than 400 microseconds. 

Theoretical Work 

•	 Rabinovitch and others (2000) attempted to explain the 

mechanism for EM emissions and concluded that following 

early pore closure, microcracking and possibly coalescence 

occurred, while just before peak stress was reached, the 

rock collapsed. A summary of information about the frequency and wavelength of EM emissions showed that their 

frequency range was 1 kHz (with a wavelength of 300 km 

[186 mi]) to 10 MHz (with a wavelength of 30 m [98 ft]). 

•	 Goldbaum and others (2001) identified four distinct EM 

emissions waveforms: short single pulses, a short chain of 

single pulses, an extended chain of pulses, and a new group, 

pulses along baseline voltage changes. Significant to their 

work were EM frequencies reaching 25 MHz (formerly 

believed to be only up to 10 MHz). 

•	 Rabinovitch and others (2001) continued investigating 

mechanisms for EM emissions and concluded that the 

mechanisms for earthquake EM emissions were the same as 

for microfracturing in laboratory tests. They studied the 

Gutenburg-Richter type and Benioff strain-release relationship for earthquakes and found the relationship extended to 

the microlevel. 

Underground 

•	 Sobolev and others (1984) tested the value of collecting 

EM emission data as a method of prospecting for quartz 

veins and base-metal sulfides. They detonated explosive 

charges and measured the EM emissions generated by the 

excited minerals. Their tests showed EM signals generated 

by quartz veins at the Giant Yellowknife Mine (Canada) 

were in the range of about 8 kHz, which was similar to 

emissions from quartz broken in the laboratory. Further 

tests in a sulfide vein at the Sullivan Mine (Canada) also 

produced EM emissions as high as 350 kHz. Their conclusions were that quartz and sulfides such as galena, 

sphalerite, and pyrrhotite emit EM waves along their grain 

boundaries. 

•	 Nesbitt and Austin (1988) recorded seismic and EM 

emissions at a depth of 2.5 km in an underground mine. 

They found that EM emissions preceded the seismic wave. 

•	 O’Keefe and Thiel (1991) recorded EM emissions associated with blasting in rock quarries in Australia and 

recorded signals in the 20-Hz to 20-kHz range. 

•	 Russell and Barker (1991) investigated expected EM emission amplitudes in exploration and found that the identification of true piezoelectric responses was difficult because their data acquisition system recorded both acoustic 

and piezoelectric signals as part of the same waveform. At 

best, they presumed that a portion of the signal collected 

was piezoelectric. 

•	 Butler and others (1994) successfully mapped stratigraphic 

boundaries using emissions responses. They found they 

could map a boundary between glacial till and organic-rich 

fill by collecting emissions waveforms generated by either 

a sledgehammer or blasting caps as a seismic source. Their 

work showed that it was the boundary or interface of the 

glacial till and the organic-rich fill that was responsible for 

the emissions conversion and not the water table. 

•	 Wolfe and others (1996) used emissions studies in an attempt to identify the depth of the water table. Using seismic refraction surveys and dc resistivity surveys in two drill 

holes as a baseline, they showed a consistent depth to the 

water table as compared to the baseline data. Thus their 

study demonstrated that emissions data could be acquired 

in an outwash plain. 

•	 Russell and others (1997) used emissions techniques to 

identify quartz and sulfide veins in three underground 

mines. They were successful in identifying quartz veins, 

sulfide veins, and the boundaries between formations with 

differing permeabilities using data from EM emissions. 

•	 Frid (1997b) concluded that EM emissions in coal mines 

lay in a narrow band from 30 to 150 kHz. He used 100 kHz 

as the most convenient frequency while examining EM 

emissions (1997a). He also concluded that the higher stress 

associated with rock near mine workings increased natural 

EM emissions. 

•	 Frid (1999) used EM emissions to delineate stress in coal 

seams. He measured EM activity during borehole drilling 

and found that a hole nearing a stress peak excited a sharp 

increase in EM activity. 

•	 Frid and others (2000) continued their work in the laboratory and attempted to correlate EM emissions with crack 

dimensions. They found that the amplitudes of EM emissions and their changes with loading were independent of 

both tensile and shear failure and that they were dependent 

only on the area of the entire crack. 

•	 Frid (2001) recognized the value of using EM emission criteria to forecast rock burst hazards in coal mines by using 

the limiting value of broken coal volume, mine working 

width, coal seam thickness, and coal elastic properties. 

•	 Sines and Knoll (unpublished oral conversation, 2000) used 

a data acquisition system to collect both seismic and EM 
11 

emissions on the 4600 level of the Galena Mine. They 

sampled at a rate of about 7,200 samples per second (a 

Nyquist frequency of 3,500 Hz) using two monopole 

antennas 12.5 and 15.2 m (41 and 49.9 ft) long. They used 

no filters to eliminate low-frequency emissions and found 

numerous triggers on the EM antenna, which were initially 

thought to be coincident with seismic activity. However, 

when the EM and seismic waveforms were analyzed, they 

found that most of the seismic emissions had actually 

preceded the EM emissions, which is physically impossible. 

Further evaluation of the collected waveforms showed that 

most of the EM emissions were caused by mine cultural 

noise, which included the opening and closing of air doors 

(60-Hz solenoid), locomotive activity, and chute loading. 

•	 Butler and others (2001) conducted field studies at the 

Brunswick No. 12 Mine in Canada in an attempt to link EM 

emissions with seismic activity and also to delineate sulfide 

ore. They used various antennas covering a range of 

frequencies from 1 Hz to 4.5 MHz. They found that 

broadband EM emissions with frequencies up to 800 kHz 

could be induced by seismicity and blasting. However, 

results did not confirm that EM emissions preceded 

seismicity. 

•	 Vozoff (2002) attempted to demonstrate the use of EM 

monitoring as a warning system for roof failure in a large 

coal seam in Australia. He collected three complete data 

sets and concluded that of the three, one set coincided with 

a roof fall and was correlated with EM activity, one set 

might have had a “weak correlation at best,” and one set 

had no EM correlations with roof falls.

 TECHNICAL APPROACH 

As noted above, many researchers have attempted to 

capture EM emissions before, during, or after ground failure 

(i.e., rock bursts) in underground mines. However, to date, none 

have conclusively linked rock breaking underground with EM 

emissions. The following work describes the methods and results of a study at the Galena Mine, Wallace, ID. It builds on 

the work of previous researchers, but uses new methods in an 

attempt to capture EM emissions from either blasting or rock 

bursting. 

The equipment used included an ESG data acquisition 

system, ULTRAQ, capable of sampling up to 10 million samples 

per second (Nyquist frequency of 5 million samples per second) 

on four channels and a Pentium 166 computer. The system was 

enclosed in a box to keep it as clean and dry as possible; fans 

were installed to keep air moving in the box. Voltages for 

triggering the system could be set as low as 1 mV. Figure 13 

shows the setup for data acquisition. 

Figure 13.—Data acquisition system setup. 

Two monopole antennas constructed of solid copper wire 

and enclosed in plastic pipe sealed at both ends were used to 

collect EM emissions data. The first, 91 m (300 ft) long and 

having a resonance frequency of 821 kHz, was inserted into a 

drill hole extending from the 5500 level down toward an active 

stope (figure 14). 

Figure 14.—Plan view of downhole antenna location. 

The second, 3.8 m (12.5 ft) long and having 

a resonance frequency of 19,737 kHz, was suspended from the 

back above the data acquisition system about 91 m (300 ft) from 

an active stope. 

Based on the work by Hanson and Rowell (1982), the ideal 

antenna length for EM data collection (considering EM 

frequencies of about 107 kHz) would be about 700 m (2,297 ft), 

which compared closely with EM frequencies (100 kHz) 

obtained in laboratory experiments from breaking rock. A 

“wound” antenna was also developed in the laboratory; 

however, this antenna failed to pick up EM waveforms. 

Initially, filters were not installed; however, both high-pass 

(2.56 kHz) and low-pass (102.4 kHz) filters were used after the 

first tests to trap a “range” of frequencies and eliminate any 

triggers that might be related to common electrical interference 

in the mine in the 60- to 120-Hz range. After several data sets 

were collected, use of the low-pass filter was discontinued, and 

only waveforms with frequencies above 2.56 kHz were collected 

and analyzed. 

In an effort to eliminate grounding problems in the mine 

stemming from the mine's ac power source, an independent dc 

power source was tried. However, the noise generated by the dc 

power source was too intense and automatically triggered the 

system. Therefore, the mine's ac power was used to operate the 

data acquisition system. 
12 

EM emissions decay at a rapid rate in rock (table 4), so 

attenuation was also a concern.

Table 4.—Attenuation of EM emissions (resistivity of the rock at 125 ohm/m) 

Frequency (kHz) Skin depth, m (ft) Attenuation (% of energy left) 

5 . . . . . . . . . . . . . . . . . 79.1 (259.4) 64.20 

10 . . . . . . . . . . . . . . . . 55.9 (183) 53.50 

20 . . . . . . . . . . . . . . . . 39.5 (130) 41.30 

40 . . . . . . . . . . . . . . . . 27.9 (92) 28.60 

80 . . . . . . . . . . . . . . . . 19.8 (65) 17.00 

160 . . . . . . . . . . . . . . . 13.9 (45.9) 8.20 

320 . . . . . . . . . . . . . . . 9.8 (32.4) 2.90 

640 . . . . . . . . . . . . . . . 6.9 (22.9) 0.70 

1280 (1.28 MHz) . . . . . 4.9 (16.2) 0.08 

2560 (2.56 MHz) . . . . . 3.5 (11,2) 0.00 

 As frequency increases, the 

distance the waveform can travel from the EM source to the 

antenna decreases. Therefore, if the frequency of an EM 

emission is high (2.56 MHz), the distance from the antenna to 

the EM emission source would be 3.5 m (11.5 ft), with virtually 

no energy left. However, if the EM emissions were at a much 

lower frequency (500 Hz), the distance between the antenna and 

an emission source would have to be nearly 79 m (260 ft), with 

about 64% of the energy left. 

Different settings were configured in an attempt to record EM 

emissions. The EM source was a striker commonly used for 

igniting barbecue grills. The striker gives off a short EM emission. 

RESULTS 

Waveform Identification 

Figures 15 and 16 show waveforms collected with the system. Figure 15 is a waveform collected from the air door solenoid (60 Hz), and figure 16 shows an EM waveform where the 

striker was used as a source. 

Figure 15.–Waveform generated by air door. 

Figure 16.—Waveform generated by barbeque lighter. 

The two waveforms are distinctly 

different. The EM waveform has a high-amplitude spike 

followed by smaller spikes originating from the striker. The air 

door waveform has a large spike that is followed by closely 

spaced decaying spikes. The air door waveform (and all 

electrical noise and grounding effects) is always characterized 

by four to five closely spaced spikes following an initial spike. 
13 

Downhole Antenna Without Filters 

Several different voltages, ranging from 10 mV to as much as 

1 V, were tested as triggers; 50 mV was selected as the final 

triggering threshold at a sampling rate of 1 MHz. Data were 

collected for 5 days. The number of events triggered ranged from 

900 per 24-hour period to as many as 1,589 per 24-hour period, 

with an average of 1,236. All the events were identified as 

electrical mine noise; no EM activity occurred during blasting. 

Downhole Antenna with Filters 

Test voltages ranged from 10 mV to 1 V. All voltages from 

400 mV or less automatically triggered the data acquisition 

system; therefore, the trigger threshold was set at 400 mV at a 

sampling rate of 1 MHz. Data were collected for 21 days. The 

number of events recorded per 24-hour period decreased when 

the filters were in place and ranged from 281 per 24-hour period 

to as many as 581 per 24-hour period, with the average being 

332 events per day. As with the unfiltered antenna, the events 

collected were identified as electrical mine noise, and no EM 

activity was recorded during blasting. 

Suspended Antenna Without Filters 

We concluded that mine electrical grounding in the rock 

created voltage readings up to 1 V that were not EM related. 

We then tried a second approach, which was to suspend an 

antenna from the back where the antenna could not be grounded. 

Trigger voltages ranging from 10 to 30 mV were tested at 

a sampling rate of 3 MHz and a trigger threshold of 30 mV. 

Data were collected for 5 days. The number of events ranged 

from two per 24-hour period to as many as 24 per 24-hour 

period, with a total number of 52 and an average number of 10 

events per 24-hour period. As before, these events were 

identified as electrical mine noise. 

System settings were then changed to a sampling rate of 4 

MHz and a trigger threshold of 25 mV. Seventy-three events 

ranging from one per 24-hour period to as many as 43 per 24­

hour period were recorded over a 9-day period, with an average 

number of eight events recorded per 24-hour period. These 

events again were classified as electrical noise. 

Using a sampling rate of 10 MHz and a trigger threshold of 

25 mV, the system recorded 107 events. Over a 7-day period, 

the number ranged from two per 24-hour period to as many as 

24 per 24-hour period, with an average per day of 15. Again, 

waveforms collected were classified as electrical noise. No EM 

activity occurred during blasting. However, the result was a 

marked decrease in the amount of mine electrical noise recorded 

by the data acquisition system compared to the 24-hour period 

of sampling (with or without filters) in the downhole. 

Suspended Antenna with Filters 

Using information from work by Hanson and Rowell (1982), 

we installed a high-pass (2.56 kHz) filter that allowed EM 

waveforms with frequencies above 2.56 kHz to be collected. EM 

emissions were recorded for 17 days at a sampling rate of 10 MHz 

and a 25-mV trigger threshold. Twenty-eight waveforms were 

collected, of which nine occurred during blasting. Figure 17 is 

one of the waveforms and is possibly an EM emission associated 

with seismicity. 

Figure 17.—Possible EM waveform. 

This was important because only waveforms associated with blasting were possibly associated with seismicity or 

rock breaking. Figure 18 summarizes the number of events 

collected using different methods. 

DISCUSSION 

Cultural noise associated with EM emissions in a deep 

underground mine is caused by, but is not limited to, blasting, 

drilling, motors, air doors, ventilation fans, shaft noise, chute 

activity, power tools, welding, power surges, various types of 

power tools, and water pumps. Seismic activity can also be a 

source of EM emissions. A grounding effect caused by the 

mine's power source creates as much as 1 V of electrical 

interference in the mine rock. The effect of the interference on 

an antenna installed in a drill hole was enough to trigger the data 

acqusition system; therefore, nearly all mine electrical noise was 

recorded and mixed with possible “true” seismic-generated EM 

emissions. Various trigger voltages and the use and nonuse of 

the low-pass and high-pass filters provided a wide range of data. 

The best EM data came from the suspended antenna using a 

high-pass filter. 
14 

Figure 18.–Summary of waveforms collected. 

However, results to date suggest that (1) the number of EM 

emissions prior to recorded seismic activity does not increase, 

(2) some EM signals are generated during blasting, (3) 

interference from mine electrical sources mask true EM signals, 

(4) EM emissions do not give enough warning (compared to 

seismic monitoring) to permit miners to leave a stope, (5) the 

distance an EM signal can travel in the rock is between 18 and 

40 m (58 and 130 ft), and (6) current data acquisition systems do 

not differentiate between EM signals generated from seismic 

activity and random mine electrical noise. In summary, these 

results preclude monitoring EM emissions as precursors of 

impending catastrophic ground failure. 

CONCLUSIONS
 

This Report of Investigations describes two geophysical 

methods examined by NIOSH researchers to identify and 

characterize conditions that might lead to ground failure in 

highly stressed rock in underground mines. Such studies are 

important in that identification of precursors to rock failure 

could lead to measures that could reduce or prevent injuries and 

deaths among miners. 

Key findings concerning the use of seismic tomography 

were that (1) seismic tomograms showed that seismic velocities 

in rock adjacent to mine openings were low, (2) a difference 

tomogram in which in situ stresses on the east and west sides of 

the slot were compared showed that velocities increased west of 

the slot, but decreased east of the slot, (3) geologic features 

(rock types and a fault) identified through geologic mapping 

were recognizable in the seismic tomograms, (4) ultrasonic 

velocity measurements on the rock cores agreed with seismic 

velocity measurements in the tomograms, and (5) results from 

the finite-difference analysis compared well to the seismic 

tomograms west of the slot, but not east of the slot. 

Results from studies of the use of EM emissions as 

precursors to seismic activity indicated that (1) the number of 

EM emissions does not increase prior to recorded seismic 

activity, (2) some EM signals are generated during blasting, 

(3) interference from mine electrical sources mask seismicgenerated EM signals, (4) EM emissions do not give enough 

warning to permit miners to leave a stope, (5) the distance an 

EM signal can travel in rock is between 18 and 40 m (58 and 

130 ft), and (6) current data acquisition systems do not 

differentiate between EM signals generated from seismic 

activity and random mine electrical noise. 

ACKNOWLEDGMENTS
 

The authors wish to thank Harry Cougher (general manager, 

Galena Mine) for access to the Galena Mine. The authors also 

wish to thank Mr. Bill Matthews of Golden Drilling, Canon 

City, CO, for providing diamond drill experience during 

installation of the pressure cell. CSM provided mine access, 

utilities, geologic and mine maps, and engineering students who 

mapped the coordinates for the geophone receiver locations. 

Dr. Michael Friedel, USGS, Lakewood, CO, provided valuable 

editing input for the final manuscript. 
15 

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Brady, B.T., and G.A. Rowell. 1986. Laboratory Investigation of the 

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17 

APPENDIX A: PRESSURE CELL INSTALLATION
 

The Department of Mining Engineering at CSM was 

contacted by researchers from NIOSH to assist in validating 

seismic tomographic methods by conducting several seismic 

tomographic surveys using controlled pressure. The approach 

was to place a pressure cell in a slot drilled into an existing 

pillar at the Edgar Experimental Mine. The pressure cell measured 81.3 cm2 by 1 cm (32 in2 by 0.38 in) and was used to 

induce pressure against the walls of the slot. 

To maximize the area affected by the pressure cell and to 

reduce the effects of fractures created during blasting, a slot was 

drilled 3 m (10 ft) into the pillar, and the pressure cell was 

installed to a depth of 2.1 m (7 ft). Grout was pumped around 

the pressure cell to allow it to press against the walls of the slot 

without expanding excessively. Figure A1 shows the installed 

pressure cell, and figure A2 shows the hydraulic jack attached 

to the pressure cell. 

Figure A1.—Installed pressure cell. Figure A2.—Hydraulic jack connected to pressure cell. 
18 

APPENDIX B: SLOT DEVELOPMENT
 

Several methods for cutting the slot were investigated. The 

first approach was to use a diamond saw. However, even the 

smallest commercial saw available proved to be too large to fit 

through the Edgar Mine portal. The second approach was to 

create the slot by using a diamond drill and overlapping the 

holes. To create adequate clearance in the slot for the pressure 

cell and a grout capsule, it was decided to drill NX-sized holes 

7.6 cm (2.98 in) in diameter on 6.4-cm (2.5-in) centers. 

DRILL HOLE DEVIATION 

The primary concern was that drill hole deviation could 

create areas where a web of rock would be left in the slot, which 

could cause protruding points in the slot to puncture the pressure 

cells as they expanded under loading. 

To minimize hole deviation, a two-post and single-bar 

mount for the diamond drill was used. Precollaring the drill 

holes was accomplished with a jumbo-mounted percussion drill. 

A wooden template was built with 2.5-cm (1-in) holes on 6.4-cm 

(2.5-in) centers. This template was then wedged into place as 

close to the face as possible and used to control collar placement 

of the 2.5-cm (1-in) pilot holes, which were drilled to a depth of 

45.7 cm (18 in). 

Reaming 

A percussion drill fitted with a button-bit reamer having a 

30.5-cm (12-in) long, 2.5-cm (1-in) in diameter pilot lead was 

used to ream the holes. The pilot lead was inserted into the hole 

and the feed shell leveled and aligned using plumb bobs. The 

pilot holes were then reamed to a depth of about 15.2 cm (6 in) 

to create a full slot 15.2 cm (6 in) deep. 

This technique worked so well that we considered creating 

the entire slot using the percussion drill. This was attractive for 

two reasons: first, it would be much faster than diamond drilling, 

and second, the overlapping holes had a tendency to crumble at 

the points of overlap so that there were no sharp points that might 

puncture the pressure cell. To test this approach, the top four 

pilot holes were extended to a depth of 1.2 m (4 ft) and reamed to 

a diameter of 7.6 cm (3 in) in 15.2-cm (6-in) stages. However, 

this approach did not work because when the hole approached a 

depth of 30 cm (12 in), it became obvious that the pilot holes had 

diverged from each other and would never maintain alignment 

well enough to be able to complete the slot to the desired depth of 

3 m (10 ft). This approach was then abandoned. 

DRILLING PROBLEMS 

Unfortunately, the 1.2-m (4-ft) pilot holes previously drilled 

became a problem. Possibly due to the left-hand rotation of the 

percussion drill, the pilot holes had a tendency to drift off to the 

right side of the slot. The diamond drill holes were drilled with 

a right-hand rotation and appeared to have a tendency to drift to 

the left. Thus, the pilot holes wandered out of the drill path as 

they were extended. This caused irregular pieces of rock to 

break into the core barrel and plug the barrel as the slot was 

extended. This was overcome by constantly pulling the barrel 

and cleaning and replacing it. 

When using the diamond drill and starting at the bottom of the 

slot, the bit used was found to be poorly suited to the rock, which 

caused the drill to bind and allowed the hole to wander off-line. 

Alternating holes were then drilled from the bottom to the 

top of the slot. Drilling then proceeded between the intervening 

webs. The major problem encountered was that the alternating 

layers of mica and quartz in the gneiss caused enough deviation 

for the hole to wander off-line, in some cases leaving a web of 

intact rock in the slot. In two cases, an additional hole was 

targeted at the webs, and they were removed with the diamond 

drill. However, in two other cases, the loss of rock at the top 

and bottom of the hole would not allow development of enough 

of a shoulder to keep the drill string on-line, and it became 

apparent that the diamond drill would not be successful in 

removing them. At this point, the slot was 3 m (10 ft) deep, but 

had two webs of rock in the middle that prevented insertion of 

the pressure cell more than 1.7 m (5.5 ft) into the slot. 

Finally, a 12.7-cm (5-in) in diameter reamer for the 

percussion drill with a 7.3-cm (2.8-in) in diameter pilot lead was 

built. Using this reamer, the webs of rock were removed to a 

depth of 2.3 m (7.5 ft). However, a second pass with the 

diamond drill was needed to remove the web, and there was not 

enough shoulder in the hole to allow the reamer to stay on 

course. It jumped off track at 2.3 m (7.5 ft), and no efforts to 

hold it on course were successful. At this point, it was apparent 

the pressure cell could not be successfully installed to a depth of 

3 m (10 ft) in the slot; therefore, the pressure cell would be 

installed at a depth of 2.1 m (7 ft). 

GROUTING 

The remaining work involved grouting the pressure cell in 

place. The pressure cell was inserted into the slot, wedged into 

place with wooden wedges to maintain alignment while the 

grout was pumped, and surveyed to determine its exact position. 

Two support beams were bolted to the face using Hilti-style 

expansion bolts, and a retaining wall was built to hold the grout 

in the slot. Rags were used adjacent to the wall to seal the 

cracks, and wet sand was packed against the rags to prevent the 

grout from getting to the rags. However, the sand was 

ineffective, and grout reached the wall. Leaks were sealed with 

minimal grout loss. The wall was initially 0.6 m (2 ft) high and 

was raised in stages as the grout pour filled the slot. A concrete 

vibrator was used to prevent air pockets from forming in the 

grout, and the wooden wedges were removed as the slot filled. 

The pressure cell was then hooked to a hand pump fitted with a 

gauge that could induce pressures ranging from 0 to 68.95 MPa 

(0 to 10,000 psi).