10

DISCUSSION AND CONCLUSIONS

This analysis of underhand backfill practice in the
Coeur d’Alene Mining District shows that backfill
contributes substantially to reducing rock burst hazards.
Numerical modeling studies indicated that omitting backfill
altogether increased the ERR by 42 pet, whereas modify-
ing backfill placement practice could reduce ERR by as
much as 30 pet. The easiest measure would be to reduce
the unfilled void between cuts; complete elimination of this
space, typically 15 pct of stope volume, could reduce the
ERR by 14 pct. A further, but more costly, measure
would be densification of the backfill. Finally, there was
no indication that the initial strength gained by the
addition of cement to form lean cemented backfill mixes
provided any reduction in the ERR.

In applying these results to a particular mine, special
attention should be paid to the fact that the two-
dimensional numerical models used in this analysis assume

a very long vein that is capable of attaining complete
closure for a significant length without end effects. Also,
the ERR method is aimed primarily at rock bursts caused
by pillar and abutment-crushing rock burst mechanisms,
but does not apply directly to discontinuity slip-type rock
bursts.

Finally, this analysis pointed out the importance of
developing rock burst hazard assessment methods, like
ERR, that are empirically linked to rock burst experience
outside of South African mines. Bureau of Mines research
is examining hazard assessment methods at the Lucky Fri-
day mine for application in the deep mines of the Coeur
@Alene District of northern Idaho.

REFERENCES

1, Barrett, A. J,, H. A. D. Kirsten, and T. R. Stacey. Backfilling in
Deep Level Tabular Stopes. Ch. in Innovations in Backfill Technology
(Proc. 4th Int. Symp. on Mining with Backfill, Montreal, QU, Oct. 2-5,
1989). A. A. Balkema, 1989, pp. 177-186.

2. Board, M. P., and S, L, Crouch. The Physical, Mechanical and
Hydrologic Properties of Backfills From the West Driefontein Mines.
Available from Dep. of Civ. and Min. Eng., Univ. MN, Minneapolis,
MN, 1984, 23 pp.

3. Boldt, C. M. K, P. C. McWilliams, and L. A. Atkins. Backfill
Properties of Total Tailings. BuMines RI 9243, 1989, 21 pp.

4. Corson, D. R. Field Evaluation of Hydraulic Backfill Compaction
at the Lucky Friday Mine, Mullan, Idaho. BuMines RI 7546, 1971,
24 pp.

5. Corson, D.R, and W.R. Wayment. Load-Displacement Measure-
ment in a Backfilled Stope of a Deep Vein Mine. BuMines RI 7038,
1967, 51 pp.

6. Crouch, S. L. Analysis of Stresses and Displacements Around
Underground Excavations: An Application of the Displacement
Discontinuity Method. Univ. MN Geomechanics Rep., Dep. of Civ. and
Miner. Eng., Minneapolis, MN, 1976, 268 pp.

7. Cundall, P., and M. P. Board. A Microcomputer Program for
Modelling Large Strain Plasticity Problems. Ch. in Numerical Methods
in Geomechanics, ed. by G. Swododa (Proc. 6th Int. Conf, Apr. 11-15,
1988, Innsbruck, Austria). A. A. Balkema, 1988, pp. 230-238.

8. Hedley, D. G. F. Utilization of Backfill Support in Longitudinal
Cut-and-Fill Mining. Can. Cent. for Miner. and Energy Technol., Min.
Res. Lab., Elliot Lake, ON, Rep. MRP/MRL 84-76 (TR), 1984, 16 pp.

9. High-Level Committee on Rockbursts and Rockfalls. An Industry
Guide to the Amelioration of the Hazards of Rockbursts and Rockfalls.
Chamber of Mines of S. Afr., Johannesburg, S. Afr., 1977, 123 pp.

10. Jenkins, F. M., T. J. Williams, and C. J. Wideman. Analysis of
Four Rock Bursts in the Lucky Friday Mine, Mullan, Idaho, USA.
Paper in International Deep Mining Conference: Technical Challenges
in Deep Level Mining, ed. by D. A. J. Ross-Watt and P. D. K. Robinson

@ohannesburg, S. Afr., Sept. 17-21, 1990). S. Afr. Inst. of Min. and
Metall., Johannesburg, S. Afr., 1990, v. 2, pp. 1201-1212.

11. McMahon, T. Rock Burst Research and the Coeur d’Alene
District. BuMines IC 9186, 1988, 49 pp.

12, Noyes, R. R, G. R. Johnson, and S. D. Lautenschlaeger.
Underhand Stoping at the Lucky Friday Mine in Idaho. Pres. at 94th
annu, meet. NW Min. Assoc. (Spokane, WA, Dec. 2, 1988). Paper No.
12, 8 pp5 available from J. K. Whyatt, Spokane Research Center,
BuMines, Spokane, WA.

13, Pariseau, W. G., J. K. Whyatt, and T. J. McMahon. Rock
‘Mechanics Investigations at the Lucky Friday Mine (In Three Parts):
3. Calibration and Validation of a Stope-Scale, Finite-Element Model.
BuMines RI 9434, 1992, 16 pp.

14, Patchet, S.J. The Use of Fill for Ground Control Purposes.
Chapter 11 in Rock Mechanics in Mining, ed. by S. Budavari Practice.
S. Afr. Inst. Min. and Metall., Johannesburg, S. Afr., Monograph Series
No. 5, 1983, pp. 241-255.

15, Salamon, M.D. G. Rock Mechanics of Underground Excavations.
Chapter in Proceedings of the 3rd Congress of the International Society
for Rock Mechanics (Denver, CO, 1974). Nat. Acad. Sci., Washington,
DG, pp. 951-1099.

16. Salamon, M. D. G. (CO Sch. Mines). Personal communication,
1988.

17. Wemer, M. A. The Lucky Friday Underhand Longwall Mining
Method. Ph. D. dissertation, Univ. of ID, Moscow, ID, 1990, 132 pp.

18. Whyatt, J. K. Geomechanics of the Caladay Shaft. M. S. thesis,
Univ. of ID, Moscow, ID, 1986, 195 pp.

19. Whyatt, J. K, and M. P. Board. Numerical Exploration of Shear-
Fracture-Related Rock Bursts Using a Strain-Softening Constitutive
Law. BuMines RI 9350, 1991, 14 pp.

20, Williams, T. J., J. K. Whyatt, and M. E. Poad. Rock Mechanics
Investigations at the Lucky Friday Mine (In Three Parts): 1. Instru-
mentation of an Experimental Underhand Longwall Stope. BuMines
RI 9432, 1992, 26 pp.



ROCK MECHANICS INVESTIGATIONS AT THE LUCKY FRIDAY MINE

(In Three Parts)

2. Evaluation of Underhand Backfill Practice
for Rock Burst Control

By J. K. Whyatt,' T. J. Williams,’ and M. P. Board?

ABSTRACT

The U.S. Bureau of Mines has been conducting a series of rock mechanics investigations at the Lucky
Friday Mine in the Coeur d’Alene Mining District of northern Idaho. In the present study, the energy
release rate (ERR) index was used to evaluate three underhand stope backfill parameters—cement
content, density, and placement gap—to determine their influence on rock bursts. This index provides
a relative measure of the likelihood of rock bursting for a given rock mass. It is known that the elastic
idealization of the rock mass inherent in this method causes underestimation of stope closure and hence
underestimation of the influence of backfill on rock bursting. However, this evaluation still showed that
(1) changing from present practice to one in which backfill was omitted altogether would increase the
ERR by 42 pct and (2) perfect placement of backfill would reduce the ERR below present practice
levels by 28 pct. The easiest step toward perfect placement would be elimination of the placement gap,
which accounted for half of the difference in ERR between present practice and ideal placement
conditions. There was no indication that increasing initial fill strength through increased cement content
would affect the ERR.

'Mining engineer, Spokane Research Center, U.S. Bureau of Mines, Spokane, WA.
Project engineer, Itasca Consulting Group, Inc., Minneapolis, MN.


Fixed velocity in x

Fixed in x

Fixed in y

Figure 6.—Original problem mesh (A) and

forming paste fill pillars (8) during numerical mod-

el simulation of stope closure. Closure is simulated with displacement boundary conditions.

and the 1.5 pet strain-to-failure for the laboratory test
model. These discrepancies were caused in part by the
laboratory and field fills being composed of slightly
different materials and in part by mesh dependency prob-
lems in the experimental constitutive law (19). Corson and
Wayment (5) reported lateral pressures of 350 to 500 psi
after closures of 14 to 18 pet in sandfill placed at 85 pet of
maximum density (a 15-pet effective gap in the model) in
overhand stopes at the Lucky Friday Mine. This range of
values lies between the 10- and 20-pct curves shown in
figure 7 for paste fill.

The perfectly placed paste fill model (no gaps and
maximum density) represented conditions very similar to
those established for confined compression laboratory tests
on fully consolidated samples. Because the fill was totally
confined, loss of cohesion with plastic strain was over-
whelmed by the increase in frictional resistance. In
models with gaps of 10 and 20 pet, there was a low-
strength region between the initial peak strength and the
onset of confined compression with contact between fill
pillars.

LATERAL STRESS, 103 psi

Figure 7.—Stress-s'

pillar.

KEY
—— No gap
——-= 10% gap
—"= 20% gap

Sf, 7
|

L141 1

01 #02 08 04 05 06
STRAIN

response at center of model backfill



recent study by Hedley (8) showed that increasing stope
height in an overhand cut-and-fill stope, which is analo-
gous to placing backfill with a temporary gap, significantly
influenced rock bursting.

This report examines the influence of this gap, as well
as the influence of cement content and degree of fill con-
solidation, on rock burst potential as measured by the
ERR. The purpose is to evaluate the relationship between

backfill practice and rock bursting in underhand longwall
stopes. Other Reports of Investigations in this series
describe the overall LFUL stope investigation (20) and
calibration of a stope-scale numerical model (73). Addi-
tional information on stope design and operation were re-
ported by Werner (17) and Noyes, Johnson, and Lauten-
schlaeger (12), respectively.

ACKNOWLEDGMENTS

Close cooperation of the Hecla Mining Co. and the
University of Idaho is gratefully acknowledged. Fred
Brackebush, manager of mining research for Hecla (now
president of Mine Systems Designs, Kellogg, ID), was
instrumental in organizing research efforts. Dave Cuvelier
and Mike Werner, mining engineers, Hecla Mining Co.

coordinated mine access. The contributions of Bureau of
Mines staff, including Doug Scott, geologist, and Mark
Board, mining engineer (now with Itasca Consulting
Group), were invaluable to planning and monitoring the
instrumented stope.

BACKFILL BEHAVIOR

Fundamental to studying the role of backfill practice in
reducing rock burst hazards is an appreciation of the
physical properties of a specific fill. Patchet (14) identifies
fill characteristics that control behavior: particle size
distribution, cement content, and density or void ratio after
placement.

Since the raw material for fill is mill tailings, the
particle size distribution is largely fixed. That is, the mill
is not likely to modify its grinding and processing
operations to satisfy fill specifications. However, it has
been common practice to cyclone off slimes from the total
mine tailings to make a coarser sandfill that dewaters
easily, Original LFUL stope specifications called for paste
fill, defined as low-water-content total mill tailings with
cement added. This paste fill would have an angle of
repose high enough to allow placement tight to irregular
backs. Typical uniaxial compression test results for paste
fill cylinders made from fine and coarse total mine tailings
with 6-pct cement are shown in figure 2. For a detailed
discussion of Coeur d’Alene fill characteristics, see Boldt,
MeWilliams, and Atkins (3).

However, at the Lucky Friday Mine, the high viscosity
of this paste fill presented difficulties when it was pumped
a mile down into the mine and then horizontally across
levels. Viscosity was reduced by increasing water content
and removing some slimes. The resulting fill was a com-
promise between the planned paste fill and a traditional
sandfill with added cement. In initial cuts of the LFUL
stope, cement was added during filling of the bottom half
only, and uncemented fill was placed in the remaining
void.

Stope closure and fill pressure instruments were paired
for placement in both the cemented and uncemented

portions of the fill. These instruments showed that the
initial increase in strength resulting from the addition of
cement quickly disappeared when the backfill fractured
(fig. 34). In this case, the 2 in of stope closure created by
mining the following cut was more than sufficient to
fracture the cemented fill completely, thus nullifying the
fill’s contribution to regional support before attainment of
full confinement. Laboratory confined-compression tests
on Coeur d’Alene backfill showed cement content had very
little influence on load-carrying capacity after fracturing.

2.0 T b=~~
w 7
B ars

ist
w
© 2
Po Zi0

Go
O2
>i
cto
a6
o
N

0 0.01 0.02 0.03
STRAIN

Figure 2.Unconfined compression stress-strain curves for 6-

pet cement fill using fine and coarse total mill tallings.


INTRODUCTION

Rock bursts have a long history in the Coeur d’Alene
Mining District of northern Idaho (17)? and are threat-
ening the future of many district mines. The U.S. Bureau
of Mines has long been involved in developing alternatives
to the traditional overhand cut-and-fill stoping method
(fig. 14) that would reduce rock burst hazards and be
amenable to mechanization. One of these methods, the
underhand longwall cut-and-fill method (fig. 1B), was
chosen for testing at the Lucky Friday Mine, Mullan, ID.
The experimental stope, dubbed the Lucky Friday under-
hand longwall (LFUL) stope, was tested under a cooper-
ative agreement among the Bureau, Hecla Mining Co.,
Coeur d'Alene, ID, and the University of Idaho,
Moscow, ID.

The principle of a single advancing face, which is
central to the underhand longwall method, is not new. In
fact, the South African High-Level Committee on Rock
Bursts and Rock Falls recommended the use of longwalls

3Italic numbers in parentheses refer to items in the list of references
at the end of this report.

Haulage drift

as a means of reducing rock burst hazards associated with
mining remnants (or sill pillars) as early as 1924 (9).
Longwall mining is now standard practice in the deep gold
mines of South Africa. However, other mechanisms of
rock bursting, such as shear failure and slip, are not
addressed by the energy release rate (ERR) index. Such
mechanisms have been identified in the Coeur d’Alene
Mining District where the greatest premining stress is
horizontal compression (10). Recent research (19) on
explicit modeling of shear fractures has also pointed
toward adopting the underhand longwall method as a
means of reducing rock burst hazards.

However, these studies were based on the assumption
that backfill behavior does not differ appreciably between
the overhand and underhand methods. Experience with
backfill in the LFUL stope showed that while overhand
stopes are easy to fill completely, it is difficult to fill
underhand stopes completely. Several researchers, includ-
ing Salamon (16), have expressed concern that the result-
ing gap may seriously increase the rock burst hazard. A

Haulage drift

Figure 1.—Schematic of cut-and-fill mining methods. A, Overhand; B, underhand.


NUMERICAL ANALYSIS OF EFFECTS OF PLACEMENT GAPS IN FILL

A numerical investigation of cemented fill behavior in
an underhand stope was undertaken using an experimental
strain-softening constitutive law implemented in the finite-
difference program FLAC (fast Lagrangian analysis of
continua) (7). Since FLAC does not model consolidation,
strain-to-consolidation must be modeled as an additional
gap in the backfill. Standard elastic parameters and a
Mohr-Coulomb strength criterion with cohesion-softening
plasticity (table 1) were defined to match typical laboratory
tests (fig, 5). Recent numerical modeling efforts (79) have
shown that such a procedure will produce sample shape
effects similar to those observed in laboratory tests.
However, this procedure must be considered experimental
because of mesh effects on softening.

‘Table 1.—Backfill properties used
In FLAC numerical simulations
of stope and gap closure

Property Value
Shear modulus... . 3,200
Bulk modulus 9,300
Cohesion»? 30
Friction angle’ 37

Density ... - 125
Cohesion varies linearly from 30
psi at 0 pet plastic strain to 0 psi at 4
pot plastic strain,
Estimated values taken from back-
{ill of similar composition (2).

A FLAC grid was defined with several fill levels and
intervening gaps (fig. 64) to examine variations in the
stope closure-fill pressure relationship when the unfilled
volume was changed. Support hardware (including rock
bolts, timber, and wire mesh) in the stope and fill were
ignored in this model. These clements play a structural
role but only initially; their primary effect is to maintain a
safe working back during mining of the following cut.

Program logic automatically accounted for large strains
and detected contact between fill pillars as gaps closed.

Once contact was detected, normal and shear forces were
transmitted across the interface. Thus, the model followed
the transition from unconfined to fully confined compres-
sion. A series of grid plots for the 10-pct-gap case shows
this progression (fig. 6B).

Complete stress-strain curves (fig. 7) were developed
for this progression using backfill gaps of 0, 10, and 20 pet.
In the numerical model, the plotted stresses were calcula-
ted for the center of the fill pillars so that plotted
measurements could be compared with field measure-
ments. Thus, these stresses were slightly larger than
average support pressures.

The 10- and 20-pet-gap curves peaked initially at ap-
proximately 350 psi. The model pillar, with a 1:1-aspect
ratio, was appropriately stronger than the reference
2:1-aspect ratio of the simulated laboratory test sample
and compared well with field measurements of 350 psi
peak pressure. The 3-pet strain shown by the model at
peak load was somewhat greater than the 0.7 to 1.2 pet
measured in the stope (0.8 to 1.5 in for a 10-ft-wide stope)

2.0

wo

STRESS, 102 psi
roy

a

0.02

0.03 0.04

STRAIN

Figure 5.—Numerical model simulation of unconfined compres-
sion test with backfill properties.


RI

9433

REPORT OF INVESTIGATIONS /1992

Rock Mechanics Investigations
at the Lucky Friday Mine

(In Three Parts)
2. Evaluation of Underhand Backfill
Practice for Rock Burst Control

By J. K. Whyatt, T. J. Williams,
and M. P. Board

UNITED STATES DEPARTMENT OF THE INTERIOR _

BUREAU OF MINES __



Model properties were chosen in an attempt to cover
the range of probable rock mass property values for
formations in the Coeur d’Alene Mining District (table 2).
A Mohr-Coulomb failure criterion was used for those runs
allowing rock mass failure. A further simplification was
made by considering only the extreme cases of perfect-
backfill models (fully consolidated fill with no gap), as
approximated by a linear spring, and no-backfill models.
This simplification was forced by the inability of the
program to use the nonlinear backfill response curves
developed for backfill with gaps.

Table 2.—Rock mass properties used to model
cross section of underhand stope

Property Value

Shear modulus . . psi. 1,250,000
Bulk modulus 16,700,000
Cohesion .. 1,500-3,000
Friction angle .. 30-40

The following assumptions were made for the six model
runs:

1, Elastoplastic rock mass with a cohesion of 1,500 psi,
friction angle of 40°, and no backfill (these are considered
to be realistic rock mass properties).

2. Elastoplastic rock mass with a cohesion of 1,500 psi,
friction angle of 30°, and backfill (this run checks fill
performance when a weak rock mass is assumed).

3. Elastic rock mass, backfill.

4. Elastic rock mass, no backfill.

5. Elastoplastic rock mass with a cohesion of 1,500 psi,
friction angle of 40°, and backfill (model 1 with backfill).

6. Elastoplastic rock mass with a cohesion of 3,000 psi,
friction angle of 40°, and backfill (this model checks back-
fill performance when the strongest rock mass is assumed).

Figures 9 and 10 show the half-closure (i.¢., closure
divided by 2) of the stope as a function of distance from
stope midspan and horizontal stress from the stope face

0

} oj | ot

HALF—CLOSURE, ft

0 1 2 3 4 5

DISTANCE FROM CENTERLINE, 102 ft

6 7 8

Figure 9.—Range of stope half-closures for various rock mass

property and fill assumptions.

HORIZONTAL STRESS, 103 psi
°

2 4
DISTANCE FROM FAC!

6 8
E, 102 ft

Figure 10.—Horizontal stress at stope face for various modi



STOPE MODEL PARAMETRIC STUDY

Backfilled stopes in deep mines provide an abutment
behind mining that will take on stress, thereby reducing
stress concentrations at the mining face. The problem of
locating the backfill abutment, and the effects of rock mass
and backfill properties on this location, can be addressed
with numerical models. However, recent work by Barrett,
Kirsten, and Stacey (1) has shown that traditional models
based on an elastic or elastoplastic constitutive law under-
estimate closure and hence the development of backfill
pressure.

In the third report of this series, this shortcoming is
addressed by reducing rock mass stiffness and strength to
replicate field measurements in a three-dimensional, finite-
element model (13), a procedure referred to as model
calibration. This model follows mining of the first 10 cuts
of the LFUL stope in considerable detail, but does not
include a backfill model. Backfill was not considered
important because 10 cuts was a clearly insufficient span
over which to develop a fully confined backfill abutment,
especially when the backfill gap was considered.

Experimental numerical models that simulate fracture
zone behavior (¢.g., 1) are another alternative. They have
shown that backfill does reduce rock bursting, even for
short roof spans. For this investigation, however, the time
and effort involved in applying an experimental fracture
zone modeling method were excessive.

Rather than adding a backfill model and expanding the
scope of an already large calibrated model, or applying an

experimental numerical model, simple two-dimensional
models were used in a comparative analysis of a highly
idealized geometry. This approach sacrifices precision in
modeling the LFUL stope for a very considerable
reduction in analytical effort. Besides, as the underhand
longwall method is adopted mine-wide, the series of
adjacent stopes will extend about 1,500 ft on strike, which
is moving toward the long-slit, two-dimensional case
assumption in a vertical cross section.

The first two-dimensional model explored the sensitivity
of stope closure to rock mass properties. The model was
formulated with the FLAC program used previously with
the mesh shown in figure 8 (which was halved along its
line of symmetry), The perfect backfill force-displacement
curve from figure 7 was approximated by a piecewise
linear spring across the vein.

Initially, a 1,300-ft-tall by 10-ft-wide portion of the mesh
was mined and filled to represent prior mining. The
model was then used to follow the development of stress
concentrations at the face and loading of the fill. Since
shear components of the in situ stress field have little
influence on these factors, the stope section was assumed
to be oriented with the principal stress directions.
Principal stresses of 4,500 psi vertical and 9,000 psi
horizontal were assumed on the basis of a survey of
overcoring measurements (18).

Vein

Roller

Fixed in x,y

Fixed in x,y

Fixed in x,y

Figure 8.—Large-scale finite-difference mesh for vertical cross section of LFUL. Vein Is 10 ft wide In a mesh 4,000 ft wide by 1,300 ft

high. Mining progresses 650 ft down vein.


There was a lag between fracturing of the cemented fill
and generation of significant support loads as the fill
consolidated. That is, as long as the backfill particles
could consolidate, or repack, to fill all void spaces, large
loads could not be generated in the fill. Estimating the so-
called closable void ratio of LFUL fill at placement was

difficult. Consolidation of hydraulically placed sandfill
4
A T T T
3 +
2b 4
PA KEY
a ——= Cemented fill
SIT = — Uncemented fill om
uf
c
2 1.0 15 2.0
n
n
uw
uw 4
oa
. 4
a
77 4
|
0 0.5 1.0

WALL CLOSURE, in_

Figure 3.Measured stope closure-fill pressure curves re-
sponding to excavation of following cut in LFUL. A, Cemented
and uncemented fill placed with no gap, cut 1; B, cemented fill
placed with no gap, cut 10.

(total tailings with slimes removed) was monitored in situ
by Corson and Wayment (5) in an overhand cut-and-fill
stope where each cut was totally filled. In this study, the
sandfill became totally consolidated at 15 pct strain
(closure), which was judged to be a good, conservative
estimate for the present work because the inclusion of
slimes would result in a denser fill requiring less closure
for consolidation.

The reduction in backfill viscosity required for pumping
fill into the LFUL stope through the existing system also
reduced the fill’s angle of repose (i.e., increased slump).
Since the stope cuts are slightly inclined to provide
drainage and the back is irregular, it was difficult to fill
tight to the back throughout the stope. Early stope cuts
were filled to within 1.5 ft of the back on the average,
leaving about 15 pct of the cut as gap (fig. 4). Since the
fill must close this gap as well as consolidate, the gap
dramatically delayed the development of appreciable
support loads. A 15-pct consolidation combined with a 15-
pet gap between fill lifts combine to require roughly 30-pet
closure before confined compression conditions are
attained.

As part of this investigation, mine personnel agreed to
take extra efforts to fill cut 10 of the LFUL stope
completely and so reduced the gap to an average of less
than 5 pet. Stope closure and fill pressure instrument
pairs placed in this cut measured significant improvements
in support pressure and a reduction in stope closure
during mining of the following cut (fig. 3B). It would
appear that the reduction in backfill gap was responsible
for the improvement. Although variations in lithology and
geologic structure between cuts 1 and 10 might have
affected these measurements, the simultaneous increase in
support pressure and reduction of closure demonstrated
that cut 10 backfill was performing significantly better than
cut 1 backfill.

‘Methods for placing fill at close-to-maximum density have also been
investigated. For instance, Corson (4) reported an increase from the
normally observed 85 pet of maximum density to about 91 pet with the
use of vibrators during hydraulic placement of sandfill.

Active stope
level

Sandwall_

Figure 4.—Backfill placement in typical underhand stope cross and longitudinal sections. Stope
gradation for drainage and uneven backfill complicate filling operations.


deg

UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT

degree Ib/f® —_ pound of mass per cubic
foot
foot
pet percent
square foot
psi pound of force per square

inch inch



CONTENTS
Page

Abstract ....
Introduction .
Acknowledgments
Backfill behavior... 0.0... 0.6 cece cece eee eee
Numerical analysis of effects of Placement gaps in fill .
Stope model parametric study .
Rock burst hazard study .
Discussion and conclusions on
References .......... teens

BScxuwenn

e

Schematic of cut-and-fill mining methods .........
Unconfined compression stress-strain curves for 6-pet cement fill using fine and coarse total mill tailings
Measured stope closure-fill pressure curves responding to excavation of following cut in LFUL .......
Backfill placement in typical underhand stope ee
Numerical model simulation of unconfined compression test with backfill properties . .
Original problem mesh and deforming paste fill pillars during numerical model simulation of stope
closure 2... eee ee eee eee eee se edenee
Stress-strain response at center of model backfill pillar ee
Large-scale finite-difference mesh for vertical cross section of LFUL ..... .
Range of stope half-closures for various rock mass property and fill assumptions
Horizontal stress at stope face for various models . an
Effective fill support pressure

URROWN

AWS YNE

FSpan

wwwmIidaan

TABLES

1. Backfill properties used in FLAC numerical simulations of stope and gap closure
2. Rock mass properties used to model cross section of underhand stope .........

Coney



Library of Congress Cataloging in Publication Data:

Rock mechanics investigations at the Lucky Friday Mine.
p. em. — (Report of investigations; 9433)
Includes bibliographical references (p. 10).

Contents: 1. Instrumentation of an experimental underhand longwall stope / by
TJ. Williams, J.K. Whyatt, and M.E. Poad—2. Evaluation of underhand backfill
practice for rock burst control / by J.K. Whyatt, TJ. Williams, and M.P. Board—
3. Calibration and validation of a stope-scale, finite-element model / by W.G.
Pariseau, J.K. Whyatt, and T.J. McMahon.

1. Rock bursts—Idaho—Lucky Friday Mine. 2. Longwall mining—Idaho—Lucky
Friday Mine. 3. Rock mechanics—Idaho—Lucky Friday Mine, I. Williams, Theo-
dore Joseph, 1923- . I. Whyatt, J. K (Jeff K). III. Poad,M.E. IV. Series:
Report of investigations (United States. Bureau of Mines); 9433.

TN23.U43 [TN317] 622 s—de20 [622.28] 90-15069 CIP



for all six models. The effective ground support pressure
generated by the fill as a function of distance from the
face is shown in figure 11. The effective ground support
pressure was calculated by dividing spring force by load-
bearing area (approximately 10 ft2).

The following conclusions were drawn from this model
study.

1. There was little difference in stope closure between
the elastic and the elastoplastic models. The yield zone
surrounding the stope was limited and resulted in only
slightly greater closures.

2. The closure of the stope was, however, affected by
backfill. Closure at midspan was reduced by roughly 20
pet. Fill strain was approximately 35 pct at midspan,
decaying only near the mining face.

3. Backfilling reduced rock mass stress by roughly 10
pet immediately ahead of the face, but there was virtually
no difference 60 ft from the face.

4, The various rock mass property assumptions caused
only minor changes in backfill performance.

KEY
—— Case 2
=== Case 5 >
——"= Cases 3,6

SUPPORT PRESSURE, 103 psi
e

| ee ee |
o 141 2 3 4 5 6 7

DISTANCE FROM FACE, 102 ft

Figure 11.—Effective fill support pressure.

ROCK BURST HAZARD STUDY

Assessing rock burst hazards for a mining system is far
from an exact science. Factors contributing to this uncer-
tainty include the variety of rock burst mechanisms that
may be active, the lack of knowledge about these mecha-
nisms, and the tendency of rock bursts to be associated
with unusual localized conditions rather than average rock
mass conditions.

One existing technique for comparing the relative rock
burst hazard is the empirical ERR index (15). ERR is a
measure of the energy change accompanying an increment
of mining, and is calculated as one-half the product of the
preexisting tractions on the block mined and the change in
closure divided by the area of the block mined.

The ERR index was developed and implemented in
South African mines that have fairly simple geology and
has been linked directly to the overall level of rock burst
activity in these mines, primarily in pillars or ribs. As
such, it should provide some relative indication of the
effect of various backfill gaps on the level of rock burst
activity. Using the ERR index in the Coeur d’Alene
Mining District, which has significantly different conditions
from those found in South Africa and possibly different
rock burst mechanisms (16), introduces some uncertainty,
especially in regard to the absolute level of rock bursts
activity. However, the ERR index is useful as a relative
measure of activity. That is, if two mining schemes are
compared, the option with a significantly lower ERR index
should have fewer rock bursts.

The displacement-discontinuity, boundary-element pro-
gram MINAP (6) was chosen to calculate ERR for the
various gap proportions. This program assumes the rock
mass to be an elastic continuum, which the second FLAC
study showed to be appropriate. The nonlinear backfill
force displacement curves in figure 7 developed with the
first FLAC model were entered into MINAP as best-fit,
second-order polynomials. MINAP underhand mining
models were run with backfill load-deformation responses
representing no fill, no gap, 10 pet gap, 20 pet gap, and an
extrapolated 30 pct gap. The ERR results are presented
as the relative percentage change from the present practice
case (30 pet gap including consolidation), since a corre-
lation between magnitude of ERR and rock burst hazard
has yet to be established in the Coeur d’Alene Mining
District. The results were that:

1. Underhand mining without backfilling increased the
ERR by 42 pct over the ERR with present practice. Thus,
present backfill practice provides an important contri-
bution to the reduction of rock burst hazards.

2. Reducing the stope closure required for the backfill
to attain confined compression conditions (either by filling
the gap or densifying the backfill) by 1 pet reduced the
ERR about 1 pet. Implementing an ideal backfill practice
(no gaps and full consolidation) would reduce the ERR by
28 pet over present practice.