Department of Health and Human Services

Centers for Disease Control and Prevention

National Institute for Occupational Safety and Health

RI REPORT OF INVESTIGATIONS/2001

9655

Geomechanics of Reinforced Cemented

Backfill in an Underground Stope at the

Lucky Friday M ine, M ullan, Idaho
Report of Investigations 9655

Geomechanics of Reinforced Cemented Backfill 

in an Underhand Stope at the Lucky Friday Mine,

Mullan, Idaho

T.J. Williams, D.K. Denton, M.K. Larson, R.L. Rains, J.B. Seymour, and

D.R. Tesarik

U.S. DEPARTMENT OF HEALTH AND HUMAN SERVICES

Public Health Service

Centers for Disease Control and Prevention

National Institute for Occupational Safety and Health

Pittsburgh Research Laboratory

Pittsburgh, PA

July 2001
This document is the public domain and may be freely copied or reprinted.

Disclaimer: Mention of any company or product does not constitute endorsement by NIOSH.

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DHHS (NIOSH) Publication No. 2001-138
CONTENTS

Page

Abstract . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1

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

Mining method . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2

Description and placement of instruments ............................................................ 3

Mining and monitoring procedures . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6

In-mine observations and data analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8

Closure readings . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8

Tests of cemented backfill specimens ............................................................. 9

Backfill pressure readings . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11

Instrumented rock bolts . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11

Rock bolt load cells . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12

Backfill temperature . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13

Truss loads . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 13

Conclusions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 16

Acknowledgments . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 16

References . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 16

Appendix A: Results of unconfined compression tests of backfill . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 18

ILLUSTRATIONS

1. Location of Lucky Friday Mine, Coeur d’Alene Mining District, Idaho . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2

2. Intersection truss system . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3

3. Rock bolts instrumented with load cells . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 4

4. Wall closure instrument locations . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5

5. Location of instruments in 5660-05-level intersection ............................................... 6

6. Longitudinal projection of Lucky Friday Mine . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 7

7. Cut sequence for 5660 and 5750 sublevels showing location of instrumented intersection ................... 8

8. Gap and backfill closure versus time ............................................................ 9

9. Thrust-fault-type fill failure, November 4, 1997 ................................................... 9

10. Backfill closure as a function of time . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 10

11. Compressive test of cemented sandfill, stress/strain curve . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 10

12. Backfill closure versus backfill pressure, West 2 . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11

13. Calibration curve showing relationship between frequency and load on instrumented rock bolts . . . . . . . . . . . . . . 12

14. Backfill closure versus rock bolt load, East 2 . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12

15. Closure sequence in backfill . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 14

16. Temperatures recorded on rock bolts in backfill and in mine atmosphere . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 15

17. Load cells at truss four-way bracket . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 15

18. Backfill failure at four-way bracket . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 15

19. Data from instrumented bolts in truss . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 15

TABLES

1. Original stope width . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 9

2. Closure readings . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 10

3. Loads on instrumented bolts and load cells for three cuts of mining . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 12

4. Load readings on trusses and rock bolts during intersection failure . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 16
UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT

cm centimeter ft foot

hr hour in inch

Hz hertz lb pound

kg kilogram psi pound per square inch

km kilometer °F degree Fahrenheit

kN kilonewton

kPa kilopascal ° degree

MPa megapascal % percent

m meter

N newton

°C degree Celsius
GEOMECHANICS OF REINFORCED CEMENTED BACKFILL IN AN UNDERHAND

STOPE AT THE LUCKY FRIDAY MINE, MULLAN, IDAHO

By T.J. Williams,1 D.K. Denton,1 M.K. Larson,1 R.L. Rains,2 J.B. Seymour,1 and D.R. Tesarik3

ABSTRACT

Because backfill has occasionally collapsed into an active working area, posing a hazard to miners, engineers

from the Spokane Research Laboratory of the National Institute for Occupational Safety and Health and Hecla

Mining Co. installed instruments in a cemented, backfilled, stope-ramp intersection at Hecla’s Lucky Friday

Mine, Mullan, ID. The purpose was to measure stress and strain changes in the backfill and reinforcing

members during undercut mining. The instruments were monitored for 6 months while three successive cuts

were mined below the intersection. Readings showed induced loads up to 3450 kPa (500 psi) in the backfill

as stope walls converged 2.5 to 12 cm (1 to 5 in). The backfill then deformed against the top and bottom plates

of the 2-m- (6-ft-) long vertical rock bolts installed as reinforcement, producing loads to 177 kN (40,000 lb)

on the rock bolts. 

We hypothesize that a compressive zone was created in the backfill that allowed the backfill to remain stable

as long as the compressive zones from adjacent rock bolts overlapped. This hypothesis is presented in graphical

form. 

Of particular interest was the effect of loading on trusses installed to augment the vertical rock bolts and wire

mesh typically installed in backfill. Data from the instruments indicate that wall closure perpendicular to the

vein induced loads in truss legs parallel to the vein and in the rock bolt driven through the center of the truss,

but, because they are designed to function under tension, truss legs perpendicular to the vein supplied

insignificant support as a result of compressional forces from wall closure. Based on this study, use of trusses

was discontinued, and an alternative support system of wood beams and posts was installed as needed to ensure

the safety of miners working beneath the backfill in stope-ramp intersections.

1

Mining engineer.

2

Geologist.

3

Mechanical engineer, Spokane Research Laboratory, National Institute for Occupational Safety and Health, Spokane, WA.
2

 Figure 1.—Location of Lucky Friday Mine, Coeur d’ Alene

Mining District, Idaho.

INTRODUCTION

Mining-induced wall closure in cemented, backfilled underhand stopes at Hecla Mining Co.’s Lucky Friday Mine in

Mullan, ID, can cause the backfill to fracture. These fractures

generally do not pose a hazard to miners working in the stopes

beneath the backfill because any broken material is contained

by wire mesh and Dywidag4

 rock bolts used for backfill

reinforcement. Still, some collapses of backfill have occurred

in stope-ramp intersections as the backfill was undercut.

Options for improving miner safety while miners are working

under backfill include placement of a layer of low-modulus

backfill above a reinforced cemented sill, installation of

deformable plastic sheets along the centerline of the stope

(Fredericksen et al. 1993; Krauland and Stille 1993), or

placement of additional support to reinforce the backfill. The

latter option was chosen by Hecla, and in addition to vertical

rock bolts, trusses manufactured by Western Support Systems,

Salt Lake City, UT, were installed in the four-way ramp-stope

intersections.

To monitor the effectiveness of the trusses and to compare

stresses in the truss-supported backfill with stresses previously

recorded in stopes without truss systems, engineers from the

Spokane Research Laboratory (SRL) of the National Institute

for Occupational Safety and Health (NIOSH) and Hecla Mining

Co. mounted instruments on both bolts and trusses. The

instruments included vibrating-wire strain gauges and load cells

on the rock bolts and the legs of the trusses, earth pressure cells

in the cemented backfill, and string potentiometers across the

stopes. Redundancy was designed into the instrumentation plan

to overcome known instrument survival problems associated

with deformation of the fill and to provide sufficient data to

interpret the interaction of the individual support components.

In addition, cylinders of wet backfill material were collected

during filling of the instrumented stope. The cylinders were

cured in a room where humidity and temperature were

controlled. They were then tested for 7- and 28-day unconfined

compressive and splitting tensile strengths.

MINING METHOD

The Lucky Friday Mine (figure 1) uses a mechanized

underhand cut-and-fill mining method to mine lead-silver ore

from a steeply dipping, 2.4-m- (8-ft-) wide vein at a depth

exceeding 1.6 km (1 mile) (Scott 1990). In the underhand

mining method, the mined-out stope is backfilled with reinforced, cemented mill tailings following each cut, which provides a safe stope back or roof for the following cut. Approximately 10% cement by weight is added to the mill tailings to

strengthen the backfill rapidly so that mining of the following

cut can resume under the backfill without a long wait. This

amount of cement has been selected after years of experience in

balancing miner safety, cost of the cement, and the need to start

mining the following cut soon after backfill placement.

Approximately 70% of the stope height is backfilled, leaving

a 1-m (3-ft) gap between the bottom of the previous fill and the

top of the new fill. The backfill is delivered from the surface in

a pastelike consistency, which lowers water content, thereby

increasing the final strength of the backfill (Brackebusch 1994).

Backfill reinforcement consists of 2- or 2.4-m- (6- or 8-ft-) long

Dywidag rock bolts driven vertically on 1.2-m (4-ft) centers into

loose muck on the floor. When the end of the rock bolt driven

into the loose muck is exposed during mining of the following

cut, chain link fencing, a bearing plate, and a nut are installed

for ground support.5

 The high horizontal in situ stresses at the

Lucky Friday Mine (Whyatt and Beus 1995; Whyatt et al. 1995)

result in rapid closure of the wall rock in the mined-out portion

 4

Mention of specific products and manufacturers does not imply endorsement by the National Institute for Occupational Safety and Health.

 5The Garpenberg Mine in Sweden (Krauland and Stille 1993) and the Henty

Mine in Western Tasmania, Australia (Henderson et al. 1998), use similar

reinforcement with their undercut-and-fill mining methods. 

 of the vein, and this wall closure is the main factor affecting the

stability of the backfill.

Each mining cut is 3.3 m (11 ft) high and 2 to 3 m (6 to 10 ft)

wide and extends approximately 75 m (250 ft) along the vein to

each side of an access ramp (slot). The broken ore is stored in

a muck bay on the opposite side of the vein from the access slot.

This creates a four-way intersection. Thus the backfill may span

a distance of up to 9 m (30 ft) diagonally.
3

Figure 2.—Intersection truss system.

DESCRIPTION AND PLACEMENT OF INSTRUMENTS

Although stope wall closure and fill pressure had been

monitored previously at the mine (Williams et al. 1992; Hedley

1993), loads on backfill reinforcement had not been. Thus, the

instrumentation plan was designed to determine closure of the

backfilled mine openings, pressure in the backfill, strength of

the backfill, and induced loads on the truss and rock bolts placed

in the backfill. Because support is provided by mechanical

interaction among these elements, a complete evaluation of the

engineering parameters of each element was needed to

determine how the whole system functioned. Failures of the

instruments from deformation and breakup of the backfill were

anticipated; therefore, redundant systems were used.

A four-way truss was installed at the center of the

intersection in addition to the vertical rock bolts typically used

as reinforcement. The truss was constructed using No. 7

Dywidag rock bolts having a minimum yield strength of 160 kN

(36,000 lb). It consisted of one vertical rock bolt, four legs

angled 15° to 20° from the horizontal, one four-way horn

bracket, and four stirrup U-bolts bent upward at 25° from the

horizontal. The horizontal legs consisted of two or three rock

bolts joined together with couplings so that the ends of the legs

lay outside the intersection into the east and west sides of the

stope along the vein and into the slot and the muck bay (figure

2).

Instrumented No. 7 Dywidag rock bolts were obtained from

Roctest, Plattsburgh, NY, to measure induced load on the rock

bolts used for backfill reinforcement. Roctest installed a

vibrating-wire strain gauge and thermister in one end of each

rock bolt. These instrumented rock bolts were placed at the

ends of all the truss legs. They were also used to replace some

of the vertical rock bolts throughout the stope. All instrumented

rock bolts were installed so that the instrumented end was in the

backfill to protect the instruments from blasting as the next cut

was mined. 

ALC10 rock bolt load cells were also obtained from Roctest

to provide redundant readings on the vertical rock bolts and the

truss. Load cells were the only instruments that could be used

to obtain load readings at the junction of the stirrup and the

horizontal legs at the truss bracket because blasting at this

location would have destroyed the instrumented rock bolt wires.

The load cells were installed on the lower end of instrumented

rock bolts (figure 3) and on the truss legs after mining had

passed. The load cells survived better than the instrumented

rock bolts because the wires were not subjected to backfill

deformation. 

 At various locations, wall-to-wall closure was measured in

three different vertical positions (figure 4). A closure plane in

the 1-m (3-ft) gap above the backfill was measured with string

potentiometers. Another closure plane across the backfill was

measured with string potentiometers inside collapsible steel

casings, and a closure plane in the new stope cut was measured

with a tape extensometer. All closure readings were taken

between 15- by 15-cm (6- by 6-in) bearing plates on 1.2-m-

(4-ft-) long rock bolts to achieve as much accuracy as possible.

Compressional loads in the backfill were measured with

690 kPa, 23-cm- (1000 psi, 9-in-) diam Model 3500 earth

pressure cells from Geokon, Inc., Lebanon, NH. Each earth

pressure cell had a backfill wall closure instrument, a gap wall

closure instrument, and an instrumented vertical rock bolt

installed at the same location. 
4

Figure 3.—Rock bolts instrumented with load cells.
5

Figure 4.—Wall closure instrument locations.

Figure 5 shows the location and types of instruments

installed in the intersection.

A CR10 datalogger manufactured by Campbell Scientific,

Inc., Logan, UT, was chosen because of its ability to record

information from a number of instrument types (Seymour et al.

1998; Larson et al. 1995; Larson and Maleki 1996), including

vibrating-wire strain gauges and thermistors, load cells, pressure

transducers, and voltage potentiometers. An Excel spreadsheet

was used for data analysis. 
6

Figure 5.—Location of instruments in 5660-05-level intersection.

MINING AND MONITORING PROCEDURES

After the instruments were calibrated at SRL, they were

installed in the mined-out cut 8 of the 5660-05 stope prior to

placement of the backfill. Figure 6 is a longitudinal projection

of the Lucky Friday Mine with the test stope location shown in

the lower west corner. Figure 7 is an expanded view showing

the cut sequence for the 5660 and 5750 sublevels and the

location of the instrument site. On October 15, 1997, the

intersection was backfilled with 2.3 m (7.5 ft) of cemented mill

tailings so that the truss legs were completely covered. Because

the stope was only filled for 2.3 m (7.5 ft) of its 3.3-m (11-ft)

height, additional string potentiometers were installed wall-towall in the gap on October 16. The datalogger was then moved

to a position in the gap between the fills so it would not be

damaged by blasting during mining of cut 9. 

The instruments were initially monitored hourly, then every

2 hr, and by the end of the project, every 12 hr. The monitoring

rate was changed to lengthen the time the system would operate

without having to retrieve the data storage canister. 

Mining the next cut (5660-05 stope, cut 9) began by blasting

the bottom of the cut 8 ramp on October 20. The load cells were

installed on the exposed ends of the rock bolts on October 29

after mining had proceeded far enough so that the instrument

wires would not be damaged by blasting. The mine installed a

3-m- (10-ft-) long cap across the west side of the intersection in
7

Figure 6.—Longitudinal projection of Lucky Friday Mine.
8

Figure 7.—Cut sequence for 5660 and 5750 sublevels showing location of instrumented intersection.

cut 9 to protect miners from a possible fill collapse following

the collapse of a 3-m (10-ft) section of the wall in the northwest

intersection corner that exposed 1.5 m (5 ft) of the side of the

fill. This was 3.7 m (12 ft) from the West 1 instrument location,

so the collapse did not affect instrument readings. Later,

another cap was installed across the muck bay side of the

intersection when a crack in the backfill was noticed there.

Monitoring the instruments continued to April 2, 1998, during

mining of three successive cuts below the instrumented backfill

location.

Changes in the truss leg readings on March 1, 1998,

indicated that the intersection had failed. Visual inspection

revealed that the northwest corner of the intersection had

collapsed on to the top of the cut 9 fill in the same area where

the stope wall had failed during mining of cut 9. At this time,

the active mining face was 9 m (30 ft) below with two backfill

horizons between it and the failed backfill, so the failure posed

no hazard to the miners.

IN-MINE OBSERVATIONS AND DATA ANALYSIS

CLOSURE READINGS

Stope wall closure readings were taken in the mining cut,

across the backfill, and in the gap above the backfill between

rock bolts to determine horizontal convergence of the stope

walls as mining progressed. The closure instruments showed

that the walls of the slot and vein were converging, but that the

walls of the muck bay were not moving.

String pots placed in the backfill showed the vein walls had

converged an average of 7.9 cm (3.1 in) during mining of cut

9 on the 5660 level, 8.9 cm (3.5 in) during mining of cut 1 on

the 5750 level, and 8.0 cm (3.2 in) during mining of cut 2 on the

5750 level. At the same time, vein closure measured by string

pots in the gap area averaged 14.0 cm (5.5 in) for cut 9 on the

5660 level, 11.5 cm (4.5 in) for cut 1 on the 5750 level, and

9.7 cm (3.8 in) for cut 2 on the 5750 level. Increased amounts

of closure across the gap were caused by a lack of support for

the walls in this area. Gap closure also began 1 or more days

before closure in the backfill (figure 8), a further indication that

the backfill was supplying wall support. The amount of overall

wall closure would probably be the same, but the backfill

supplied enough support so that fractures in the stope walls

closed before the backfill started to yield. There were no

borehole closure extensometers in the walls, so this hypothesis

could not be confirmed.

In the active mining area, a tape extensometer recorded an

average of 9.4 cm (3.7 in) of convergence across the vein and

7.6 cm (3.0 in) across the slot. Closure measurements in the

active mining area began after mining was at least 9 m (30 ft)

past the instrument locations; therefore, these measurements do

not represent total closure related to mining of this cut.

A thrust-fault-type of fracture (figure 9) was noticed at

the top of the backfill while retrieving the CR10's data canister

on November 4, 1997. The failure went across the intersection

and into the east and west headings along the vein. Both 2- and

2.4-m- (6- and 8-ft-) long vertical rock bolts had been placed in
9

Figure 8.—Gap and backfill closure versus time.

Figure 9.—Thrust-fault-type failure, November 4, 1997.

the area. Failure appeared to be above the 2-m (6-ft) rock bolts,

but blocked by the 2.4-m (8-ft) rock bolts, the ends of which

were just visible at the top of the fill.

This failure was caused by convergence of the vein walls on

the backfill. At this time, there had been 2.5 to 7.6 cm (1 to

 3 in) of horizontal closure across the vein in the backfill and 7.6

to 15 cm (3 to 6 in) in the gap above the instrumented backfill.

 Fracturing of the backfill caused upward buckling that

gradually reduced the gap between the backfills over time.

Buckling continued until the initial 1-m (3.5-ft) vertical gap was

reduced to less than 0.3 m (1 ft) before the backfill from cut 7

collapsed onto the instrumented cut 8 backfill. Table 1 is a

summary of original stope widths, and table 2 shows closure

readings for

the three mining cuts. Figure 10 shows backfill closure as a

function of time for three mining cuts. Total closure includes

some rapid closure as the cut was mined past the instruments

and more gradual time-dependent closure resulting from all

previous mining in the area.

Table 1.—Original stope widths.

 Location m in

West 1 . . . . . . . . . . . 34.5 136

West 2 . . . . . . . . . . . 25.4 100

East 1 . . . . . . . . . . . . 27.4 108

East 2 . . . . . . . . . . . . 36.6 144

Muck bay . . . . . . . . . 39.6 156

Slot 2 . . . . . . . . . . . . 30.0 118

West 1 . . . . . . . . . . . 52.1 205

West 2 . . . . . . . . . . . 35.6 140

East 1 . . . . . . . . . . . . 37.1 146

TESTS OF CEMENTED BACKFILL SPECIMENS

At the time the stope was backfilled, samples of the fill were

collected for compressive and tensile strength tests. The 7-day

unconfined compressive strengths ranged from 1703 to

2744 kPa (247 to 398 psi), with an average of 2082 kPa (302

psi), while 28-day unconfined compressive strengths ranged

from 2654 to 3902 kPa (385 to 566 psi) and averaged 3254 kPa

(472 psi). The laboratory tests agreed with the 2757-kPa

(400-psi) ultimate strength recorded across the vein at approximately 15 days.

Movement of the platten head of the compression test

machine was also recorded to determine an apparent modulus

for the 7- and 28-day compressive tests. Figure 11 is an example of the stress-strain curve for the tests. The modulus for

the samples is calculated between the 20% and 50% strength

values because this is the straight line portion of the curve and

most representative of the true response of the backfill. The

equation used is —

20-50 modulus = (50% strength - 20% strength) 

÷ (50% strain - 20% strain). 

The five samples tested at 7 days had a range of apparent

modulus from 593 to 2013 MPa (86,000 to 292,000 psi) with an

average of 1041 MPa (151,000 psi). Apparent modulus for the

28-day tests ranged from 1172 to 1641 MPa (170,000 to

238,000 psi) and averaged 1370 MPa (198,700 psi).

Seven-day tensile strengths ranged from 441 to 551 kPa

(64 to 80 psi) and averaged 496 kPa (72 psi), while the 28-day

tests ranged from 537 to 579 kPa (78 to 84 psi) and averaged

551 kPa (80 psi). These strengths are consistent with other tests

recently conducted on samples of cemented fill from the Lucky

Friday Mine. Appendix A provides a summary of recent tests.
10

Figure 10.—Backfill closure as a function of time.

 Figure 11.—Compressive test of cemented sandfill, stress/

strain curve.

Table 2.—Closure readings.

Instrument and 5660 level, cut 9 5750 level, cut 1 Cuts 9 + 1 5750 level, cut 2 Cuts 9+1+2

location cm in cm in cm in cm in cm in

Muck bay . . . . . . . . . . . . . . 0.10 0.04 0 0 0.10 0.04 0.10 0.04

Backfill string pots:

West 1 . . . . . . . . . . . . . . 7.77 3.06 11.28 4.44 19.05 7.5 12.19 4.8 3.24 12.3

West 2 . . . . . . . . . . . . . . +4.65 +1.83

East 1 . . . . . . . . . . . . . . 12.14 4.78

East 2 . . . . . . . . . . . . . . 3.78 1.49 6.63 2.61 10.16 4.1 3.81 1.5 14.22 5.6

Muck bay . . . . . . . . . . . . 0.08 0.03 0.10 0.04 0.18 0.07

Gap string pots:

West 1 . . . . . . . . . . . . . . 12.4 4.88 11.05 4.35 23.44 9.23

West 2 . . . . . . . . . . . . . . 11.33 4.46

East 1 . . . . . . . . . . . . . . 16.36 6.44 11.86 4.67 28.22 11.11 9.68 3.81 37.90 14.92

East 2 . . . . . . . . . . . . . . 15.80 6.22

Slot 1 . . . . . . . . . . . . . . . 8.71 3.43

Tape extensometer:

Slot 2 . . . . . . . . . . . . . . . 6.55 2.58

West 1 . . . . . . . . . . . . . . 9.12 3.59

West 2 . . . . . . . . . . . . . . 12.47 4.89

East 1 . . . . . . . . . . . . . . 6.76 +2.66

+ = Quit working.
11

Figure 12.—Backfill closure versus backfill pressure, West 2.

BACKFILL PRESSURE READINGS

Model 3500 earth pressure cells from Geokon, Inc., were

placed in the cemented backfill to monitor pressure as the walls

closed. The backfill pressure increased rapidly as wall closure

commenced (figure 12). The readings peaked with mining of

cut 9, after which backfill pressure dropped as each subsequent

cut was completed. Peak pressures ranged from a low of 793

kPa (115 psi) in the intersection to 4178 kPa (606 psi) at West

2; the average across the vein was 2757 kPa (400 psi) at 15

days. The data collected in the mine were within the range of

the data collected in the laboratory (7-day unconfined compressive strength of 2082 kPa [302 psi] and 28-day unconfined compressive of 3254 kPa [472 psi]). 

The 20-50 modului determined at locations E1, E2, W1,

and W2 were 681, 1513, 2361, and 5095 MPa (98,760, 219,

400, 342,400, and 738,800 psi), respectively, and averaged 2412

MPa (349,800 psi). Strain was determined by dividing

measured wall closure by the original opening width. These

modulus values were documented 8 to12 days after the backfill

had been poured and as mining of the following cut passed the

instrument locations. These data show that the ultimate strength

of the cemented backfill was surpassed by loading induced by

the large amounts of wall closure experienced in the stope.

Thus, rock bolt reinforcement is needed to maintain the integrity

of the backfill so that it will be safe for miners to work under.

Broken backfill was observed in the chain link below the

backfill, and backfill heaving was seen in the gap at the top.

The backfill eventually broke up to where it could no longer

carry load and support itself. It then collapsed onto the backfill

below it. When these collapses occurred, they presented no

danger to miners because mining was usually two or more cuts

below, and two or more intact backfill horizons were between

the miners and the collapsed fill remained.

INSTRUMENTED ROCK BOLTS

The vibrating-wire instrumented rock bolts and load cells

were initially calibrated in a Tinius-Olsen testing machine at

SRL to 53 kN (12,000 lb); however, readings recorded at the

mine with the instrumented rock bolts using the 27-N/Hz

(6-lb/Hz) calibration indicated that the yield and ultimate

strengths of the rock bolts had been exceeded. 

To determine if the initial readings were accurate or if this

were a calibration problem, two instrumented rock bolts were

tested in tension to failure. The instrumented rock bolts yielded

at 160 kN (36,000 lb) and failed between 209 and 221 kN

(47,000 and 49,900 lb) at the gauge location. 

These tests showed that the hole drilled for the vibrating wire

did not affect the yield strength of the bolt, but that it did reduce

ultimate strength from 240 to 215 kN (54,000 to 48,450 lb).

The slope of gauge response was linear between 8.9 and 124 kN

(2000 and 28,000 lb) at 27 N/Hz (6 lb/Hz). Gauge response was

then rapidly reduced to 2 N/Hz (0.45 lb/Hz). At 142 kN (32,000

lb), the gauges failed before the yield point of the rock bolts was

reached. Figure 13 shows calibration load versus gauge

response frequency for the bolts tested to failure. These tests

showed that the vibrating-wire gauges installed in the end of the

bolts were inadequate for this application and any future tests

should use a gauge with a higher load range. The data were

reanalyzed using this calibration curve.

Load on six of the nine instrumented rock bolts placed

vertically in the backfill exceeded 142 kN (32,000 lb), which

was the limit of the vibrating-wire gauge. Information from

most of the instrumented rock bolts was eventually lost because

deformation of the backfill apparently broke many of the signal

wires in the backfill. 

The data showed that the vertical reinforcing rock bolts in

the backfill did a good job of resisting backfill deformation

and provided a safe back for miners to work under. Initial loads

began at zero and increased over time, indicating a direct

relationship to wall closure.
12

 Figure 13.—Calibration curve showing relationship between frequency and load on instrumented rock bolts.

Figure 14.—Backfill closure versus rock bolt load, East 2.

ROCK BOLT LOAD CELLS

Data from the rock bolt load cells indicated that the vertical

rock bolts were under loads from 41 to 179 kN (9300 to 40,300

lb) after cut 9 was mined except in the muck bay, where loads

ranged from 2.9 to 19.1 kN (660 to 4300 lb). Loads on the

vertical rock bolts along the vein dropped as cuts were subsequently mined. This drop was probably a result of the continued breaking up of the backfill as the walls closed. Figure 14

shows the relationship of rock bolt load to backfill closure

at East 2. The data indicate rock bolt load leveling off and

decreasing as closure continued following each mining cut.

This drop was probably caused by backfill failure around the

rock bolt bearing plates as the walls converged. Inspection of

the bottom of the backfill revealed bags of broken backfill in the

chain link fencing between the rock bolt plates. Visual

observation of the top of the 2.4-m- (8-ft-) long vertical rock

bolts also confirmed that the 15- by 15-cm (6- by 6-in) bearing

plates resisted backfill deformation and transferred load to the

rock bolt.

Table 3 is a summary of the loads recorded on the instrumented rock bolts and load cells. The readings of 164 and

179 kN (36,900 and 40,300 lb) are the only ones exceeding the

yield strength of the rock bolt at 160 kN (40,900 lb); the rest

were below the yield strength. This is important because it

Table 3.—Loads on instrumented bolts and load cells for three cuts of mining.

5660-05, cut 9 5750-05, cut 1 5750-05, cut 2

Location Bolt Load cell Bolt Load cell Bolt Load cell

kN lb kN lb kN lb kN lb kN lb kN lb

Vertical bolts:

Slot . . . . . . . . . +142 +32,000 102 22,900 44.3 9,960 35.1 7,890

Muck bay . . . . . 8.9 1,990 0.9 194 6.7 1,510 1.8 400 2.9 655 10.1 2,270

West 1 . . . . . . . 63.1 14,200 42.5 9,340 503 11,300 34.1 7,680 25.3 5,700

West 2 . . . . . . . +142 +32,000 116 26,200 19.4 4,370 9.9 2,230

East 1 . . . . . . . +142 +32,000 179 40,300 157 35,400 133 30,000

East 2 . . . . . . . +142 +32,000 157 35,500 +142 +32,000 122 27,600 104 23,400

Intersection A . +142 +32,000 +142 +32,000 +142 +32,000

Intersection B . 103 23,200 164 36,900 53.3 12,000 1.8 408 43.3 9,740

Truss bolts:

Slot . . . . . . . . . 31.4 7,060 35.8 8,060 5.4 1,220

Muck bay . . . . . -2.5 -570 12.5 2,810 8.7 1,960 19.1 4,300 5,170 2.8 660

West . . . . . . . . 19.9 4,480 23.7 5,330 6.5 1,460 41.7 9,370 6,320 159 35,800

East . . . . . . . . . 36.0 8,100 4.3 960 20.1 4,530 48.3 10,860 63.1 14,200

Center . . . . . . . +142 +32,000 113 25,600 +142 +32,000 104 23,500 +142 +32,000

+ = Out of range or quit working.
13

indicates that the yield strength of the rock bolts was not being

exceeded.

Figure 15 is an idealized drawing of how the instrumented

rock bolts interacted with the cemented backfill in the vein

portion of the stope as mining progressed. This interpretation is

based on conventional theories of rock bolt reinforcement that

state that tensioned rock bolts create a self-supporting

compressive arch across the opening (Lang 1961; Hoek and

Brown 1980; Stillborg 1986; Brady and Brown 1993). The

illustration is also based on visual observations, closure

measurements, and load readings on the vertical rock bolts

gathered during this project. Initially, in stage 1, there is no wall

closure, pressure in the backfill, or load on the reinforcing rock

bolts. Then forces in the backfill, created as the bearing plates

resist backfill deformation caused by wall closure, form a cone

of compression in the backfill (stage 2). The backfill is selfsupporting until the backfill breakup progresses to the point

where the overlap of the cones of compression between adjacent

vertical rock bolts is eliminated (stage 3). The backfill then

collapses because of gravity. 

BACKFILL TEMPERATURE

The instrumented bolts and two of the earth pressure cells

had temperature sensors attached in case large temperature

fluctuations required that calculations be made to compensate

for these fluctuations. The sensors on the bolts were on the end

of the bolt with the strain gauge, and the temperature readings

reflected the position of the gauge with respect to the surface of

the sandfill. The readings from the 2.4-m- (8-ft-) long rock bolts

in the intersection were lower than the other readings because

the ends of these bolts were exposed to the air. Figure 16 shows

average temperatures for the bolts in the backfill and those in

the intersection. The 47.2 °C (117 °F) in-fill temperature at day

2 was the highest temperature recorded in the backfill and

stemmed from the heat of hydration of the cemented backfill.

The 32.2 °C (90 °F) recorded on the IA, IB, and IC rock bolts

was air temperature in the unventilated area above the backfill.

The data show that no temperature compensation was required

for the strain gauges because loads were significantly higher

than the temperature compensation. The data also give an

indication of the heat load to the ventilation system from the

heat of hydration of the cemented backfill.

TRUSS LOADS

Immediately after installation on day 14 and during mining of

cut 9 on the 5660 level, the load cell on the vertical rock bolt in

the center of the truss took on loads to 116 kN (26,000 lb), while

the load cells on the horizontal leg showed loads from 0.4 to

33.4 kN (100 to 7500 lb) (figure 17). Loads on the horizontal

legs remained fairly constant until day 87, at which time cut 1

from the 5750 level was being mined past the instrument

locations 6 m (20 ft) below.

On February 15, 1998 (day 122), load on the horizontal legs

increased suddenly while load on the vertical leg decreased.

Visual inspection on February 18 showed that the backfilled

intersection of cut 7 had slumped onto the top of the

instrumented backfill. It was noticed that the yoke of the slot leg

of the cut 7 truss had come off the horn bracket (figure 18) and

supplied no support at all. Then, between 6:00 p.m. on February

28 (day 136) and 6:00 a.m. on March 1, the load cells on the

four horizontal legs of the truss showed sudden increases of 6.7,

7.6, 9.3, and 19.1 kN (1500, 1700, 2100 and 4300 lb). The next

readings (12 hr later at 6:00 p.m. on March 1) showed that loads

on the east and west legs of the truss had increased from 111

and 80 kN (25,000 and 18,000 lb) to 178 and 165 kN (40,000

and 37,000 lb), respectively, while loads on the slot and muck

bay legs had dropped from 53 and 49 kN (12,000 and 11,000 lb)

to 13.3 and 2.7 kN (3000 and 600 lb), respectively.

Load on the vertical instrumented rock bolt in the truss also

showed an increase from 35.6 kN (8000 lb) to the 142-kN

(32,000-lb) limit of the vibrating-wire strain gauge on day 12 as

cut 9 on the 5660-05 level was mined under the intersection.

At approximately the same time, the instrumented rock bolts on

the east and west truss leg ends reached loads of 60.9 and 27.7

kN (13,687 and 6236 lb), respectively. These loads then

gradually leveled off at 36.5 and 19.6 kN (8200 and 4400 lb). 

After day 87, the instrumented rock bolts showed decreases in

load on the east and west legs and an increase on the muck bay

leg, while the load cells all showed an increase in load.

Readings from the rock bolt on the east leg of the truss stopped

shortly after, probably as a result of backfill deformation cutting

the signal wire. 

At day 136, the two remaining instrumented rock bolts on

the muck bay leg and the west leg also recorded sudden

increases in load similar to, but of a lower magnitude than, those

of the load cells at the truss bracket. Readings from the instrumented rock bolt on the west truss leg stopped on day 142

after cut 2 from the 5750 level had mined under. Figure 19 is a

graph of data for all three mining cuts.

Data from both the load cells and instrumented bolts indicate

that a major redistribution of load was taking place in the

intersection during this 18-hr period. Table 4 is a summary of

the readings over the 18-hr period for the rock bolt load cells

and the instrumented rock bolts. The truss legs along the vein

functioned as expected, but the legs extending into the muck bay

and slot failed to hold load and were ineffective.

Visual inspection of the intersection on March 3 showed that

the backfill in the northwest corner of the intersection had

collapsed onto the backfill in cut 9, but that the truss was still

above the backfill. It was not possible to determine how far the

failure extended along the west side of the stope, but it is

thought that, based on instrument response, failure was limited

to the first 3 m (10 ft) of the west side. The 3-m- (10-ft-) long

cap placed here probably stopped further collapse, but the area

was not accessible to confirm this belief. The weight of the

backfill from cut 7 on top of the instrumented backfill and

continued wall closure from mining in cut 2 on the 5750-05

level caused the intersection to collapse. The collapse was not

hazardous to miners because there were two intact backfilled

cuts between the collapse and the active mining area.
14

Figure 15.—Closure sequence in backfill.
15

 Figure 16.—Temperatures recorded on rock bolts in backfill and

in mine atmosphere.

Figure 17.—Load cells at truss four-way bracket.

Figure 18.—Backfill failure at four-way bracket.

Figure 19.—Data from instrumented bolts in truss.
16

Table 4.—Pressure readings on truss and rock bolts during intersection failure.

Instrument February 28, 6:00 p.m. March 1, 6:00 a.m. March 1, 6:00 p.m.

kN lb kN lb kN lb

Load cells:

IS truss . . . . . . . . . . . . . . . . . 46.7 10,489 53.7 12,075 13.6 3,064

IM truss . . . . . . . . . . . . . . . . . 42.4 9,533 50.2 11,288 2.9 644

IW truss . . . . . . . . . . . . . . . . . 62.7 14,095 81.8 18,398 165 37,067

IE truss . . . . . . . . . . . . . . . . . 104 23,485 114 25,584 180 40,539

IC truss, vertical . . . . . . . . . . . 89.6 20,144 75.4 16,943 2.9 650

Instrumented bolt:

IM truss . . . . . . . . . . . . . . . . . 10.0 2,242 12.4 2,783 17.1 3,847

IW truss . . . . . . . . . . . . . . . . . 6.9 1,555 17.4 3,911 29.2 6,558

Load cell on vertical bolt:

W1 . . . . . . . . . . . . . . . . . . . . . 33.7 7,578 29.2 6,560 23.7 5,321

W2 . . . . . . . . . . . . . . . . . . . . . 13.0 2,920 8.3 1,875 8.5 1,908

CONCLUSIONS

An extensive instrumentation project carried out in the intersection area of cut 8 of the 05 stope on the 5660 level of the

Lucky Friday Mine showed that the intersection truss installed

to provide additional support did not function fully because

closure across the vein reduced the effectiveness of the horizontal truss legs in the slot and muck bay. Data from the

instruments indicated that wall closure induced loads in the truss

legs parallel to the vein and in the vertical rock bolts, but that

insignificant support was supplied by the truss legs perpendicular to the vein. Therefore, the mine staff decided to use an

alternative support system of wood beams and posts to ensure

the safety of miners working beneath the backfill.

Project data also showed that some rock bolts placed vertically

in the backfill for reinforcement were taking loads past their

yield strength of 160 kN (36,000 lb). This is the first documentation of mining-induced loads on rock bolts in backfill

at the Lucky Friday Mine. The instruments also documented

for the first time significant closure across the slot and an

almost total lack of wall rock movement in the muck bay. All

instruments recorded changes as mining of subsequent cuts

passed by the instrument locations.

An interpretation of the interaction among wall closure,

backfill deformation, and induced loads in the vertical rock bolts

in the cemented backfill is presented in figure 15 and indicates

how the reinforced backfill support system may work. This

knowledge is important for designing backfill support systems

for other mines to ensure the safety of miners working in

underhand stopes.

ACKNOWLEDGMENTS

Clyde Peppin, chief engineer at the Lucky Friday Mine,

purchased instruments and coordinated activities at the mine.

Doug Bayer, rock mechanics engineer, helped install the instruments and supplied information on mining sequences. Frank

Reed, mine production foreman, was very patient and helpful

when researchers were working around the miners in the stope.

This work could not have been completed without the help

of a number of support staff at the Spokane Research Laboratory. In particular, Paul Pierce and Mike Jones helped with

building the closure instruments; their work is greatly

appreciated. 

REFERENCES

Brackebusch, F. W. Basics of Paste Backfill Systems. Min. Eng., Oct. 1994,

pp. 1175-1178.

Brady, B. H. G., and E. T. Brown. Rock Mechanics for Underground Mining,

2nd ed. Chapman and Hall, 1993, 571 pp.

Fredericksen, A., N. Krauland, H. Stille, and S. Stromberg. Improved

Undercut-and-Fill Mining at the Garpenberg Mine. In Minefill 93, ed. by H. W.

Glen (Johannesburg, S. Afr., Sept. 1993). Symp. Ser. S13, S. Afr. Instit. of Min.

and Metall., Johannesburg, S. Afr., 1993, pp. 17-23.

Hedley, D. G. F. Properties of Cemented Paste Backfill at Hecla’s Lucky

Friday Mine. Report on Stiff Backfill Project of Canadian Rockburst Research

Project. April 1993, 13 pp. Available from author.

Henderson, A., G. Jardine, and C. Woodall. The Implementation of Paste Fill

at the Henty Gold Mine. In Minefill 98, Proceedings of the Sixth International

Symposium on Mining with Backfill, ed. by M. Bloss (Brisbane, Australia, April

14-16, 1998). Aust. Inst. of Min. and Metall. Pub. Series 1/98, 1998, pp. 299-

304.

Hoek, E., and E. T. Brown. 1980. Underground Excavations in Rock. Inst.

Min. and Metall., London, UK.

Krauland, N., and H. Stille. Rock-Mechanics Investigations of Undercut-andFill Mining at the Garpenberg Mine. In Minefill 93, ed. by H. W. Glen

(Johannesburg, S. Afr., Sept. 1993). Symp. Ser. S13, S. Afr. Instit. of Min. and

Metall., Johannesburg, S. Afr., 1993, pp. 47-54.

Lang, T. A. Theory and Practice of Rock Bolting. Trans. 1961, vol. 220,

pp. 333-348.

Larson, M. K., and H. Maleki. Geotechnical Factors Influencing a TimeDependent Deformation Mechanism Around an Entry in a Dipping Seam. In

Proceedings, 15th International Conference on Ground Control in Mining, ed.

by L. Ozdemir, K. Hanna, K. Y. Haramy, and S. Peng (Golden, CO, Aug. 13-15,
17

1996). CO School of Mines, 1996, pp. 699-710.

Larson, M., B. Stewart, M. Stevenson, M. King, and S. Signer. A Case Study

of a Deformation Mechanism Around a Two-Entry Gateroad System Involving Probable Time-Dependent Deformation. In Proceedings of 14th

International Conference on Ground Control in Mining, ed. by S. S. Peng

(Morgantown, WV, Aug. 1-3, 1995). Dept. of Mining Engineering, WV Univ,

1995, pp. 295-304.

Scott, D. F. Relationship of Geologic Features to Rock Bursts, Lucky Friday

Mine, Mullan, Idaho. In Rockbursts and Seismicity in Mines. Proceedings

 of the 2nd International Symposium on Rock Bursts and Seismicity in Mines,

ed. by C. Fairhurst (Minneapolis, MN, June 8-10, 1988). Balkema, 1990,

pp. 401-406. 

Seymour, B., D. Tesarik, M. Larson, and J. Shoemaker. Stability of

Backfilled Cross-Panel Entries During Longwall Mining. In Proceedings of

17th International Conference on Ground Control in Mining, ed. by Syd S.

Peng (Morgantown, WV, Aug. 4-6, 1998). Dept. of Mining Engineering, WV

University, 1998, pp. 11-20.

Stillborg, B. Professional User’s Handbook for Rock Bolting. Trans Tech

Publi. Series on Rock and Soil Mechanics, vol. 15, 1986.

Whyatt, J. K., and M. J. Beus. In Situ Stress at the Lucky Friday Mine (In

Four Parts): 1. Reanalysis of Overcore Measurements From 4250 Level. U.S.

Bur. Mines Rep. of Invest. 9532, 1995, 26 pp.

Whyatt, J. K., T. J. Williams, and W. Blake. In Situ Stress at the Lucky

Friday Mine (In Four Parts): 4. Characterization of Mine In Situ Stress Field.

U.S. Bur. Mines Rep. of Invest. 9582, 1995, 26 pp.

Williams, T. J., J. K. Whyatt, and M. E. Poad. Rock Mechanics Investigations

at the Lucky Friday Mine (In Three Parts): 1. Instrumentation of an

Experimental Underhand Longwall Stope. U.S. Bur. Mines Rep. of Invest.

9432, 1992, 26 pp. 
18

APPENDIX A: RESULTS OF UNCONFINED COMPRESSION TESTS OF BACKFILL

Sample designation Type of sample Date of placement Location

A . . . . . . . . . . . . . . . . . . . . . . . . Cored Jan. 14, 1997 5660-05 ramp-stope intersection cut

B . . . . . . . . . . . . . . . . . . . . . . . . In situ Mar. 11, 1997 5660-05 east (left) side of stope cut

C . . . . . . . . . . . . . . . . . . . . . . . . In situ Oct. 10, 1997 5660-05 ranp-stope intersection cut

D . . . . . . . . . . . . . . . . . . . . . . . . Cored 5500-01 ramp stope intersection cut

Specimen Compressive strength, psi

series Curing time, days Average Range Coefficient of 

Minimum Maximum variation, %

Sample type A:

1 . . . . . . . . . . . . . . . . . . . . . . . 43 628.7 577.5 685.1 5.3

2 . . . . . . . . . . . . . . . . . . . . . . . 45 632.0 571.2 694.9 6.4

Sample type B:

1 . . . . . . . . . . . . . . . . . . . . . . . 14 450.6 414.3 477.4 5.7

2 . . . . . . . . . . . . . . . . . . . . . . . 28 499.8 469.2 530.3 4.8

3 . . . . . . . . . . . . . . . . . . . . . . . 90 613.2 575.4 698.0 7.4

Sample type C:

1 ....................... 7 302.8 247.7 398.8 18.0

2 . . . . . . . . . . . . . . . . . . . . . . . 28 472.4 385.7 556.2 14.7

Sample type D:

1 ....................... ? 249.6 241.3 263.1 3.4
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