16

KEY
A Before installation
106 After installation
2 iN 1 Mining sequence
4 106 stope
! 110
ia al
K an “7
3
W Wal J
LFUL stope
B 106 cut 2 LFUL cut 2 LFUL cut 3
sl 1 i 1 f 1

wee

N
I

KEY
@ West closure extensometer
A East closure extensometer

CLOSURE, in

0 ! | | | | !
400 440 480 520 560 600 640 680 720 760
TIME, days since 12/31/84

Figure 18.—Excavation sequence in 106 and LFUL stopes (A); response of closure extensometers in 106 stope to mining in
LFUL stope (B).


10

Ore from the first cut was hauled to the 5100 level and
dumped into rail cars until an ore pass to the 5300 level
was completed in September 1985. The 5-ft-diam ore pass
was excavated with standard raise boring equipment.
However, the raise deteriorated rapidly and became a 10-
by 15-ft oblong before it was stabilized with shotcrete.
The 15-ft diam of the raise was perpendicular to the
principal stress direction (12).

Figures 10 and 11 show data from manual tape extenso-
meters taken from cut 1 of the LFUL stope during

completion of the cut and mining of cut 2 in the 5100-106
stope above. Jumps in the closure rate after initial sta-
bilization were attributed to mining of various 106 stope
cuts. Closure at location E2 was nearly double that at
other locations. This result may be related to the presence
of a minor fault that crosses the vein at E2, stress concen-
trations associated with the bend in the vein (fig. 114),
and the end-line of the 106 stope directly above.

As noted earlier, two MPBX’s were placed in the walls
of the first LFUL cut. Measurements showed that the

106 raise
‘Cut 2-106
4 Cut 1-106" Second backstope round 106 stope
2 ea First backstope round
11
1 drift i ie MPBX-R1 AY
3 Cx) O SI ALE, e e rT
WS W4 W3 Wi £12] E3 E4 ES
LFUL stope
MPBX SC1-E1
106 raise

Offset fault

KEY
Before second back round
Second back round

After second back round

After all closure points
Fault
Mining sequence

Figure 10.—Location of monitoring stations and MPBY’s In cut 1 of LFUL.

5100 level

110

North Control Fault



20

UFUL_eut1 curt cut’2 ur? cut

cu cul cu cu

0.08 -—— | tt mes FS

] I T TT I I
KEY
07 |— @ S-ft anchor +
|A 15-ft anchor|
06 (Ml 25-ft anchor, i

|W 40-ft anchor}

| | |

300 400 500

600 700 800 900

TIME, days since 12/31/84

Figure 22.—Response of MPBX-R1.

SECOND INSTRUMENTATION CYCLE

The second set of four pairs of closure extensometers
and pressure cells was installed in cut 5 of the LFUL stope
along with the fill reinforcement. A lean cemented backfill
was placed 6 to 7 ft high throughout the stope.

All four pressure cells (fig. 23) showed an increase in
pressure between 300 and 400 psi during mining of cut 6,
but failed to survive excavation of cut 7. Three of the four
corresponding stope closure extensometers survived back-
filling to record closure during mining of cut 6 (fig. 24).
These instruments showed wide variations in the amount
of stope closure between locations (1 to 3 in) and did not
vary in concert with fill pressure measurements. This lack
of correlation suggests that fill properties continued to vary
widely within the stope.

The closure extensometer at W1 survived to monitor
excavation of cuts 7 and 8, showing reduced, but still
significant, closure with each cut. Visual inspection of
backfill in the stope (possible through an unfilled portion
of the ramp crosscut) just prior to excavation of cut 8
showed cut 5 fill was buckling up into the gap between
stope fill levels (fig. 25). Two additional closure exten-
someters were installed at this time at locations W1A and
WIB, between W1 and the ramp crosscut at cut 5. These
closure extensometers survived mining of cut 8 through
part of cut 11 and indicated continued closure at signif-
icant, but declining, rates with mining (fig. 24). These
measurement readings reinforced the pattern of local var-
iation in closure registered by the extensometers placed in
cut 5 fill.


ROCK MECHANICS INVESTIGATIONS AT THE LUCKY FRIDAY MINE

(In Three Parts)
1. Instrumentation of an Experimental Underhand Longwall Stope

By T. J. Williams,’ J. K. Whyatt,’ and M. E. Poad?

ABSTRACT

Researchers at the U.S. Bureau of Mines monitored rock mass response to mining of an experimental
underhand longwall stope in Hecla Mining Co.’s Lucky Friday Mine, Mullan, ID. This stope design, the
Lucky Friday underhand longwall (LFUL), was proposed as a means of controlling rock bursting while
also allowing increased mechanization of mining operations. Instruments were specifically manufactured
to monitor three geomechanical factors (backfill performance, ramp stability, and rock mass response)
that directly affect the success of stope design. Despite considerable difficulties with instrument
installation and survival in the harsh mine environment, especially within the stope fill, sufficient
information was collected to suggest that the LFUL method was successful in achieving project goals.
This conclusion, in addition to favorable economic and operational evidence, led the mine to adopt the
method throughout the mine. Indeed, this study and mining experience suggest that underhand longwall
mining could be considered a feasible mining method for rock-burst-prone ground throughout the Coeur
@Alene District. Further Reports of Investigations (RI) in this series present in-depth analyses of the
effects of backfill on rock bursting and calibration of a stope numerical model.

‘Mining engineer.
?Supervisory mining engineer.
Spokane Research Center, U.S. Bureau of Mines, Spokane, WA.


Closure measurements were continued in the ramp
after initial stabilization to check the response of the ramp
to mining in the LFUL stope. In general, only the por-
tions of the ramp nearest the stope (R2, R3, R6, R7, and
R8) showed any change. This was especially evident for
closure measurements in the second ramp spiral (R6, R7,
and R8), which were quite small until mining reached cut
4, the first cut to use the second ramp spiral.

Overall, measured closure throughout the monitoring
period was modest (<2 in), suggesting that the ramp was
stable.

Three reference points at RX1, RX2, and RX3 were
established in the crosscut between the spiral ramp and
the vein to measure closure during mining of the first cut
of the LFUL stope (fig. 21). RX2 was lost to a rock burst
just after this cut was completed, but the remaining refer-
ence points were available for measuring crosscut closure
associated with mining in the 106 stope above. These
points were lost when the crosscut ramp was decpened to
access the second cut of the LFUL stope.

MPBX-R1 was installed in the wall of the ramp near
the LFUL stope just after excavation of the first cut of the
LFUL stope. The purpose was to monitor the stability of
the pillar between ramp and stope (fig. 22). Measure-
ments were taken during mining in the 106 stope as well
as the second and third cuts in the LFUL stope. The
relative movement measured between anchors suggested
minor inelastic deformation of rock between the 15-, 25-,
and 40-ft anchors, which were nearest the stope, and
between the collar (at the ramp wall) and the 5-ft anchor
next to the ramp. The remaining portion of the pillar
between the 5- and 15-ft anchors was very stable. The
small magnitude of these displacements, less than 0.1 in,
indicated a stable ramp pillar.

7

LEGEND
Ri Reference point
RX1_ Crosscut reference point

3
3
2
3
b

Figure 19.—Location of ramp (R1 to R4) and ramp crosscut
(RXS to RX9) monitoring stations.


Figure 25.—Gap above sandfill in cut 5.

THIRD INSTRUMENTATION CYCLE

The final set of four closure extensometers and six
pressure cells were installed in cut 10. The fill buckling
and continued high rate of closure in cut 5 led researchers
and mine staff to discuss ways to improve the speed at
which support pressure could be generated in the fill. The
easiest way appeared to be filling the stope as tight to the
back as possible, hastening the attainment of fully confined
conditions. With significant extra effort, miners were able
to fill to within an average of 6 in of the back.

All the closure extensometers and five of the pressure
cells survived backfilling operations to monitor mining of
cut 11. The closure extensometers showed relatively

modest stope closures averaging less than 0.8 in. in
response to mining of cut 11 (fig. 264). The pressure cells
(fig. 26B) showed generally impressive support pressures
after excavation of cut 11. Four of the five pressure cells
measured 600 to 1,000 psi. The lowest pressure, measured
at W2A, registered nearly 300 psi, which was in line with
the average pressure generated in cut 5 and may have
been the result of a localized void. Thus, support pressure
measured by all five cells averaged 800 psi. These instru-
ments failed either during filling of cut 11 or early in the
excavation of cut 12; pressure cell W1B measured a peak
pressure of over 1,400 psi before it failed during excavation
of cut 12.


106

RX1 LFUL cut 1

LFUL cut 2

Ramp bottom for cut 2 access

8 106 LFUL
LFUL cut 1 {98 cut / cut 2 cut 2

CLOSURE, in
&
]

RX2 ruined by burst

0 150 250 350 450 550
TIME, days since 12/31/84

Figure 21.—Location of monitoring stations in ramp crosscut (A) and measured closure (B).

19


14

Figure 14.—Preparing Dywidag bolts for sandifill. Figure 15.—Hanging chain link fencing for
tection.

psi
uw

FILL PRESSURE, 102
N

~200 -100 0 100 200
DISTANCE FROM MINING FACE, ft

Figure 16.—Response of pressure cells in cut 1 to mining in cut 2.



fill (especially cemented fill) is failing and is being held in
place only by reinforcement members. Because mining
made maintaining a reference point on the floor difficult,
a sagmeter (fig. 5) was designed to measure displacement
of the fill back between the bottom of the fill and a fixed
10- by 10-in timber cap installed across the middle of the
stope. Measurements are taken with a string potentio-
meter protected by a telescoping tube. The bottom end of
the tube is attached to a plate in the sandfill, and the top
end is attached to the timber cap. The potentiometer and
cables are mounted in the top of the instrument, away
from blasting.

These instruments were not found to be useful. The
cemented fill fractured more quickly and discretely than
expected. Subsequent lateral shifts along these fractures
damaged the instruments and their readout cables. Such
conditions were not properly anticipated when the sag-
meters were designed. In addition, when the fill rein-
forcement system was redesigned, the timber caps were
removed, eliminating the sagmeter’s stable reference point.
Thus, measurement of sag was abandoned and back condi-
tions were judged entirely by on-the-spot observations.

12-in string
potentiometer

Broken floor

Raa mor

Figure 5.—Schematic showing ‘sagmeter design and instal-
lation between mine floor and timber cap.

Support Pressure and Stope Closure

The performance of backfill as a ground support ele-
ment lies in its ability to build up support pressure in
opposition to stope closure. From this standpoint, the
quality of a backfill system can be judged by how quickly
pressure develops as a function of stope closure. Thus, the
LFUL instrumentation package included both fill pressure
and stope closure measurements. A previous Bureau in-
vestigation (1) showed that the lateral (cross stope) fill
pressure was the only significant component and provided
some important lessons on how to protect pressure cells.

Stope wall closure was monitored between pairs of ref-
erence points installed at stations in the first cut be-
fore and after fill placement to provide a complete history
of closure to complement fill pressure measurements
(table 1), Monitoring with a tape extensometer began with
installation of grouted reference points 6 ft behind the
face. Points mounted closer to the face were frequently
blasted out. After a cut was completely mined out, elec-
tronic cross-stope closure extensometers (fig. 6) were
installed at these same stations to monitor closure after
backfill placement. They were built around a 24-in dis-
placement potentiometer, which was protected with a 2-in-
diam sealed plastic pipe. A plunger placed in one end of
the pipe maintained contact with the opposite wall.

Table 1,—instrument installation Information

Instrument and location

Tape extensometer:
LFUL, cut 1 .
Ramp .

Monitored excavation

LFUL, cut 1; 106, cuts 1 and 2.
LFUL, cuts 1 to 10; 106, cuts 1
and 2.

Ramp crosscut . . 106, cuts 1 and 2.

MPBX-1 and MPBX-2:
LFUL, cut 1. LFUL, cut 2; 106, cuts 1 and 2.
Ramp pillar . LFUL, cuts 1 to 3; 106, cuts 1

and 2.
Closure meters and pressure
calls:

Cutt. LFUL, cuts 2 and 3,
Cuts. LFUL, cuts 6 to 11.
Cut10 .. LFUL, cuts 11 and 12,

Fracturing and movement of the cemented backfill
apparently broke either the housing or cables of most of
the closure extensometers, although some survived excava-
tion of the following cut. During the project, the housing
was upgraded from plastic to aluminum, and extra coils of
cable were left near the closure extensometer to allow
for cable movement. However, only marginal improve-
ments in instrument survivability were realized.

Of the 13 closure extensometers installed in the fill,
10 survived through excavation of one cut, two survived


REFERENCES

1. Bush, D. D., W. Blake, and M. P. Board. Evaluation and Demon-
stration of Underhand Stoping To Control Rock Bursts (contract
10292013). BuMines OFR 62-85, 1982, 191 pp.

2. Cuvelier, D. (Hecla Mining Co.). Private communication, 1986.

3. Hedley, D. G. F, and R. J. Wetmiller. Rockbursts in Ontario
Mines During 1984. Can. Cent. Miner. and Energy Technol., Elliot
Lake, ON, CANMET Rep. MRL SP85-5, July 1985, 36 pp.

4. 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.

5. Jenkins, F. M,, T, J. Williams, and C. J. Wideman, Analysis of
Four Rockbursts 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
(Johannesburg, S. Afr., Sept. 17-21, 1990). S. Afr. Inst. of Min. and
Metall., Johannesburg, S. Afr., 1990, v. 2, pp. 1201-1212,

6. Leighton, F. A Case History of a Major Rock Burst. BuMines
RI-8701, 1982, 14 pp.

7. McMahon, T. Rock Burst Research and the Coeur d’Alene Dis-
trict BuMines IC 9186, 1988, 49 pp.

8. Noyes, RR, G. R. Johnson, and S, D, Lautenschlager. Under-
hand Stoping at the Lucky Friday Mine. Pres. at 4th annu. meet., NW
Min. Assoc. (Spokane, WA, Dec. 2, 1988). Paper No. 12, 8 pp; available
from J. K. Whyatt, Spokane Research Center, BuMines, Spokane, WA.

9. 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.

10. Scott, D. F. Relationship of Geologic Features to Seismic Events,
Lucky Friday Mine, Mullan, Idaho. Paper in Rockbursts and Seismicity
in Mines. Proceedings of the 2nd International Symposium on Rock-
bursts and Seismicity in Mines (Minneapolis, MN, June 8-10, 1988), ed.
by C. Fairhurst. A. A. Balkema, 1990, pp. 401-406.

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

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

13. 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.

14, Whyatt, J. K, and E. L. Hardin. Microcomputer-Based Instru-
mentation System for Monitoring Ground Support in a Deep Mine
Shaft. BuMines IC 9062, 1986, 8 pp.

15. Whyatt, J. K,, T. J. Williams, and M. P. Board, Rock Mechanics
Investigations at the Lucky Friday Mine (In Three Parts): 2. Evalua-
tion of Underhand Backfill Practice for Rock Burst Control, BuMines
RI 9433, 1992, 10 pp.

16. Williams, T. J., and D. J. Cuvelier. Report on a Field Trial of an
Underhand Longwall Mining Method to Alleviate Rock Burst Hazards,
Paper in Rockbursts and Seismicity in Mines. Proceedings of the 2nd
International Symposium on Rockbursts and Seismicity in Mines
(Minneapolis, MN, June 8-10, 1988), ed. by C. Fairhurst. A. A. Balkema,
1990, pp. 349-354,



‘Osburn Fault

Montana

Ong / Idaho
20 Boise

Figure 1.—Location of Lucky Friday Mine.

Haulage drift

Figure 3.—Schematic of cut-and-

North Control Fault

Fold axis aoe
. 4
N i
= SS j
oe
oN
Lucky Friday
Vein

South Control Fault

Figure 2.LFUL geologic map.

Haulage drift

mining methods. A, Overhand; B, underhand.


INTRODUCTION

The U.S. Bureau of Mines has sought to improve miner
safety and productivity through the development of im-
proved mining methods. One of these methods, under-
hand longwall cut-and-fil, was tested at the Lucky Friday
Mine, Mullan, ID (fig. 1), under a cooperative agreement
between the Bureau, Hecla Mining Co. of Coeur d’Alene,
ID, and the University of Idaho, Moscow, ID.

This first RI in a three-part series covers instrumenta-
tion and evaluation of ground control performance during
mining of the first 10 cuts of the experimental Lucky Fri-
day underhand longwall (LFUL) stope. Other RI’s in the
series present detailed analyses of the effects of backfill
practice on rock bursting (15),? and calibration of a finite-
element model to measured stope displacements (9).

ACKNOWLEDGMENTS

This investigation could not have been conducted with-
out the close cooperation of the staff of Hecla Mining Co.
and the College of Mines and Earth Resources, University
of Idaho, Fred Brackebush, manager of mining research
for Hecla (now president of Mine Systems Designs,
Kellogg, ID), was instrumental in organizing research

efforts. Dave Cuvelicr and Mike Werner, mining engi-
neers, Hecla Mining Co., coordinated mine access. In
addition, the contributions of Bureau of Mines staff,
including Doug Scott, geologist, and Mark Board, mining
engineer (now with Itasca Consulting Group), were inval-
uable to planning and monitoring the instrumented stope.

MINE GEOLOGY AND STOPE DESIGN

The LFUL test stope is located at the east end of the
Lucky Friday vein between the 5100 and 5300 levels in the
5300-107 block. This block lies in the upper submember
of the lower member of the Revett Formation, approxi-
mately 5,100 ft below the surface and 1,700 ft below sea
level. The vein ranges from several inches to 14 ft thick
and is nearly conformal to an anticline plunging 75° to the
southeast, Because the vein itself dips steeply (70° to 90°)
to the south and east, it comes in contact with progres-
sively older rocks with depth (fig. 2). Some of the many
faults and secondary folds in the anticline also intersect the
vein,

The rock mass surrounding the vein is made up of beds
of vitreous quartzite and sericitic quartzite from 12 to
36 in thick. Thin argillite beds are often found between
quartzite beds, The vein itself contains massive galena,
sphalerite, and tetrahedrite. Scott (10) provides an in-
depth description of stope geology.

CONVENTIONAL STOPING TECHNOLOGY

The Coeur d’Alene District in northern Idaho has a
long history and strongly established traditional methods
of mining. The ore, which is typically found in steep,
narrow veins throughout the district, is usually accessed
through rectangular timbered shafts and extracted using
the labor-intensive, overhand cut-and-fill method. The
name is derived from the fact that a vein is divided into
levels, usually 200 ft apart vertically, and each level is
mined overhand from bottom to top. Access to the veia

on each level is through crosscuts from a lateral haulage
drift paralleling the vein. Mining proceeds from the cross-
cuts by “raising up" into the vein, and a sequence of 200-ft-
long horizontal cuts centered on the raise is then exca-
vated. After a cut has been mined out, it is backfilled with
sand or waste material, providing support for the wall rock
and a floor for mining the next cut. Figure 34 illustrates
a typical overhand stope with its raise, ore chute, backfill,
and active stoping area.

One characteristic of the overhand stoping method is
that it produces a pillar of rock as mining proceeds toward
previously mined areas above. Eventually, the entire pillar
is removed. As the pillar becomes smaller, stress levels
increase until the pillar fails or is completely mined out.
Failure may be violent or nonviolent, depending upon the
brittleness of the pillar and the relative stiffnesses of the
pillar and wall rock. A nonviolent failure is characterized
by gradual fracturing and yielding, while violent failure is
characterized by the release of seismic energy. When suf-
ficient seismic energy is produced to damage mine open-
ings, the failure is called a rock burst.

Rock bursts have become increasingly troublesome with
depth (7) and pose major operational and safety problems
for mining at current depths. They are especially common.
during excavation of the last portion of a pillar. Leighton
(© has provided a detailed description of ground condi-
tions before and after a typical pillar burst at the nearby
Galena Mine.

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


Closure in the fill above mining clearly depended on the
progress of mining in cut 2 (fig. 17). The closure exten-
someters in the 106 raise, which were farther removed
from mining, showed a more gradual response (fig. 18).

Ramp

Instruments were installed in the ramp system to
monitor ground stability yet not impede access by LHD’s.
Reference points for tape extensometers (fig. 19) were
placed after every 40 to 50 ft of advance in the ramp spiral
and at 20-ft intervals in the crosscuts connecting the ramp
to the stope. In addition, an MPBX (MPBX-R1) was used
to monitor the pillar between the ramp and the stope.

The reference points were monitored to provide infor-
mation on initial stability, the effects of mining, and the
influence of extending the ramp spiral below the existing
ramp. The ore haulage section, which connected the stope
to the ore pass, was the most sensitive part of the offset
spiral ramp. Not only did the most traffic pass through it
as ore was hauled from three successive cuts of the stope
(30 ft of mining) to the ore pass, it also had a relatively
small, 20-ft-thick floor pillar. Although ore haulage ceased

15

before a second haulage ramp was excavated in the next
ramp spiral below, this section continues to be the sole
means of stope access for the life of the stope. The offset
spiral ramps connecting these haulage ramps alternated to
the north and to the south, providing one ramp for every
60 ft of mining and, therefore, a 50-ft-thick floor pillar.

Reference points were installed at R1 to R9 in the main
ramp spiral as it was driven for manual measurements of
ramp closure. Measurements at R1 and RS5 to R9 showed
quick stabilization (fig, 20), but measurements at R4, and
to a lesser extent, R2 and R3, showed significant amounts
of closure prior to stabilization. Apparently, points at R2,
R3, and R4 were located in a weaker portion of the rock
mass, probably near a fault that runs through the ramp
haulage section and the ore pass. This interpretation is
supported by raveling of the ore pass adjacent to R4
encountered during stope development.

Since ore pass raveling was stabilized by applying a
layer of shotcrete, the same approach was tried in the
second ramp haulage section. Reference points at R8 and
R9 were installed in this section along with a layer of
shotcrete. Measurements showed reduced ramp closure in
the shotcreted section, especially at R9.

LFUL cut 2

2.0

1.8
1.6
1.4
2)

eo

CLOSURE, in
a

590 595

600 605 610 615 620

TIME, days since 12/31/84

Figure 17—Response of closure extensometers in cut 1 to mining in cut 2.


(\ememmmmt 110 crosscut

No. 2 shaft

Access ramp

Ore pass

106 crosscut

,100-tt level

5,300-ft level

Figure 4,—Schematic of LFUL.

FILL BEHAVIOR

Information on the ability of the fill to provide a safe
back and some regional support was sought from a pack-
age of instruments designed to monitor stope closure and
fill behavior, As the safety of the back was a prime opera-
tional concern, and the back was regularly checked by
mine staff, the bulk of the instrumentation effort was
aimed at monitoring development of support pressure in
the backfill, which received little or no attention from
mine staff.

Back Safety

Safety practices used with overhead fill methods do not
differ significantly from the routine used in overhand
stopes. The back is inspected and loose material is re-
moved (barred down) at the beginning of each shift, and
reinforcement, including timber sets, is installed in
hazardous areas.

One indication of a potentially dangerous back is the
degree of fill sag. A large amount of sag suggests that the


LFUL
cut 5 LFUL cut 6 LFUL cut 7 LFUL cut 8 LFUL cut 9
B00 T T T T
eat
700 |— —
600 }— ap bo 4
._ 500 4
7
a
uj 400 KEY +
a O&Ft
a Awi
wm 300 O wa) aa
v @ ws
= 200 +
100 4
0 =
~100 | | |
1,140 1,180 1,220 1,260 1,300 1,340 1,380

TIME, days since 12/31/84

Figure 23.—Response of pressure cells in cut 5 to mining in cuts 6 through 9.

21


| REPORT OF INVESTIGATIONS/1992

Rock Mechanics Investigations
at the Lucky Friday Mine

(In Three Parts)
1. Instrumentation of an Experimental

Underhand Longwall Stope

By T. J. Williams, J. K. Whyatt,
and M. E. Poad

UNITED STATES DEPARTMENT OF THE INTERIOR

"BUREAU OF MINES



Extent of Fractured Rock

The extent of the fractured rock zone is not casily
predicted. An attempt to determine the dimensions of the
zone was made using a time-domain reflectometry (TDR)
method. In this method, a cable grouted into a hole

drilled ahead of mining is monitored, and the location of
breaks in the cable, presumably caused by rock fracturing,
are pinpointed. Unfortunately, the grouting operation
failed after the borehole was only one-third full. Whether
this was the fault of the grouting equipment, the installa-
tion procedure, or raveling of the borehole is not known.

PROJECT CHRONOLOGY

A full interpretation of the many measurements obtained
by the variety of instruments installed around the LFUL
stope required careful consideration of ever-changing
LFUL mine geometry. The following section presents a
chronology of data collection, mining progress, and other
pertinent information. Instruments were concentrated in
the first, fifth, and tenth cuts of the LFUL stope. A
summary timeline of mining activity and an outline of
mining geometry are presented in figures 7 and 8 to
complement the summary of instrument locations and
excavations monitored given in table 1.

FIRST INSTRUMENTRATION CYCLE

Development, 106 I-Drift, and First
106 Backstope Round

Standard development procedures were followed on the
5100 level, which serves several overhand stopes as well as
the LFUL stope. An I-drift was mined from the crosscut
for the 106 overhand stope, under which half of the LFUL
stope was later developed. Excavation of the LFUL ramp
was initiated in January 1985. After taking backstope

ia
vadalpe baisope
‘estore sostaik Fil round round Rais prop cut | Fit out Fily
* * >| | Hatt 108 mining
vw Gu rre (al bailey and watrsandpat Fl Gaming Fu GaSe
uu [oars Exend ap ta pase
ee —
1985 L 1986 |
T T T T T T T T
H a ao BS wo Ba oto 0
TIME, days since 12/31/84
ene, Fit Cuts Fill | Cutandfil§ | Cutandfil Cutandfil7 Cutandfil8 _ Cut and {fill Cut and fil10_ Cut and fil 14. Cut and fil 12
Ramp to out 4
pit
NN = = =
T T T T T T T
0 ‘to ‘ho 1x0 0 1100 10 1600

TIME, days since 12/31/84

Figure 7Project timeline.


aa

— t

ww,

a
wre

2. 1‘ Telescoping
Arrow rod
Spring loaded 24-in
linear potentiometer

Figure 6.—Closure extensometer.

through two cuts, and one through three cuts. Thus, the
instruments proved to be suitable for monitoring one or
two cuts, but failed to monitor long-term stope closure.

Pressure measurements were made with Geokon‘ hy-
draulic pressure cells installed with the closure extensom-
eters at stations in the first cut. The cells were modified
with additional moisture protection for the electronic
transducers and cables. Proper orientation of the cells
with the stope walls was ensured by attaching the cells to
6-in-square hog wire mounted in a frame of rock bolts
driven into the loose muck on the floor. The stope was
then filled with hydraulically placed cemented sandfill.

The pressure cells also proved to be suitable only for
short-term monitoring. Of the 15 cells installed during the
project, two failed almost immediately, and six more failed
after excavation of the following cut. The seven remaining
cells continued to function through mining two cuts, but
none survived excavation of three cuts.

RAMP STABILITY

The ramp system provides the only access to the stope.
Since a fully implemented longwall system concentrates
mining on a single level, closing down a single ramp for
repair disrupts as much as 20 pct of the mine’s production.
Thus, ramp stability is a key factor in preserving the
economic viability of the underhand longwall system.
Two types of measurements indicate ramp instability:
high rates of closure and an acceleration of those rates.
Closure itself can result in narrowing of the ramp, thereby

‘Reference to specific products does not imply endorsement by the
U.S. Bureau of Mines.

blocking LHD access, while acceleration can indicate
impending collapse of the pillar. Manual tape extenso-
meters were used to measure closure rates, and an elec-
tronically read multipoint borehole extensometer (MPBX)
was used to monitor rock deformation between the ramp
and the stope to check for fracturing and impending
failure of the pillar (table 1). These instruments were
generally successful in providing information on ramp
stability.

ROCK MASS RESPONSE

The rock mass responds to mining by fracturing at the
skin of underground openings and by the global redis-
tribution of stress. Stress redistribution often results in
overstressing intact rock and/or creating discontinuities
that may fail violently as a rock burst. It is this global
behavior that is analyzed by stope models. Indices of rock
burst hazard, such as the energy release rate, are drawn
from the results of these models. Evidence of rock mass
response to corroborate a stope model was sought by
measuring rock mass deformation and stress changes in
the rock mass with continued mining, and by observing the
progression of rock fracturing ahead of mining,

Rock Mass Deformation

Rock mass deformation was monitored with Geokon
two- and four-anchor MPBX’s placed in several locations
in the stope and ramp system. The first anchor of the
two-anchor MPBX’s and the second anchor of the four-
anchor MPBX’s were set at shallow depths (5 to 15 ft) to
capture fracture zone deformations. The other anchors
were set 25 to 50 ft in the rock and were expected to show
small, elastic deformations if the rock mass remained
intact.

Two MPBX’s (MPBX-1 and MPBX-2) were installed
in the stope walls prior to backfilling the first LFUL cut
(table 1). MPBX-1 failed during placement of the backfill,
probably from failure of instrument waterproofing.
MPBX-? failed after the second cut was completed; it also
showed signs of water interference with electronic read-
ings. In-stope MPBX installations were discontinued after
this experience.

Change in Stress

Vibrating wire stressmeters were installed in the wall
tock below the stope in an effort to detect mining-induced
stress concentrations ahead of mining. These stresses lic
beyond a zone of fractured rock that pushes stress into
intact rock farther below the mining floor. Unfortunately,
an unexpected reversal in the dip direction of the vein
during the second cut brought mining through the stress-
meter instrument wires. Attempts to repair the wires
failed to provide valid instrument readings.


Rock bursts have also been traced to sources other than
pillars, including slip along faults and shear failure of
highly stressed wall rock (5). Unfortunately, these mech-
anisms remain poorly understood (13).

LFUL EXPERIMENTAL STOPE PLAN

The underhand longwall method uses the principle of
a single advancing face in conjunction with the underhand
cut-and-fill method to reduce rock burst hazards. The
South African High-Level Committee on Rock Bursts and
Rock Falls recommended longwall mining as a means of
reducing rock burst hazards associated with mining
remnants (or sill pillars) as early as 1924 (4); this method
is now standard practice in South Africa. Hecla Mining
Co. has experimented with an underhand cut-and-fill
method at its Star Mine in northern Idaho (1).

Underhand cut-and-fill stoping is a method in which a
block of ore is mined by cutting and filling in sequence
from the top of the block to the bottom, rather than from
the bottom to the top as in the overhand cut-and-fill
method. Therefore, the intact vein forms the stope floor
instead of the stope back. As in the overhand cut-and-fill
method, the vein is accessed through crosscuts from
laterals, and mining is conducted with a conventional drill,
blast, and muck cycle. Figure 3B shows a typical under-
hand cut-and-fill stope with ore chutes and a filled area
above the stope.

In the underhand longwall method, the open stope is
backfilled following each cut with a reinforced, cement-
stabilized fill material that provides a safe stope back or
roof. Reinforcing members typically include wire mesh,
timber, and rock bolts. Maintenance of a safe back is a
critical issue, a point underscored by a tragic rock-burst-
induced backfill collapse that killed four miners at the

Falconbridge Mine, Ontario, on June 20, 1984 (3).
However, a well-designed fill back should have safety
advantages over the burst-prone vein rock that it replaces.
For instance, vein destressing (in which long blast holes
drilled ahead of mining are fired to relieve stress
concentrations) is often used to prevent bursting in the
back, and the fractured rock that results can pose slabbing
and caving hazards. Destressing the floor of an underhand
stope avoids this problem entirely.

The LFUL stope uses conventional underhand cut-and-
fill methods and a ramp system to provide access for
mechanized mining equipment (17). Slushers are replaced
with load-haul-dump equipment (LHD’s), which in turn
increase mucking efficiency sufficiently to allow stope
lengths to be increased from 200 to 500 ft, thereby
reducing the number of ramps required. Both 1/2- (in
areas under 5 ft wide) and 1-yd? LHD’s can be used,
depending on local vein width. In addition, all mining is
organized on a single longwall to eliminate the formation
of rock-burst-prone pillars. Figure 4 shows the resulting
mining plan, as implemented in the experimental LFUL
stope.

Each mining cut is about 10 ft high and extends
approximately 250 ft along the vein to each side of the
ramp. The declination of the ramp provides one turn
adjacent to the ore body every 30 vertical feet. Crosscuts
are driven from the turning point of the ramp to provide
access for services and LHD’s. Once a cut is complete,
the stope and crosscut are filled with cemented fill. After
the fill sets up, a new crosscut is driven from the ramp to
the next stope level. Noyes, Johnson, and Lautenschlager
(8) provide a detailed description of the method as it
emerged during the development of the experimental
LFUL stope.

INSTRUMENTATION PLAN

Three conditions were identified as crucial to successful
implementation of the underhand longwall mining method.
These were that:

1, The fill must provide a safe back and some regional
support,

2, The ramp system must remain stable and opera-
tional throughout the life of the stope, and

3. The rock mass response to changing the mining
method should not exacerbate, and should hopefully
reduce, rock burst hazards, especially for rock bursts
originating at the mining face.

An instrumentation plan was developed for.the LFUL
stope with the goal of monitoring conditions during mining

the first half of the stope block (about 100 vertical feet).
Experience with prior instrumentation projects in the
district suggested that electronic instruments would be
vulnerable to the harsh conditions underground. Nonethe-
less, the need for remote monitoring of many instruments
and the desirability of being able to read instruments
automatically several times daily necessitated a largely
electronic system. Therefore, the electronic instruments
were treated with additional waterproofing and other pro-
tective measures and monitored with a computerized data
acquisition system (14). Manually read instruments were
installed in the ramp where they could be read on a regu-
lar basis by research staff. Stope seismicity was monitored
by the mine’s microseismic system (10).


18

CLOSURE, in

LFUL LFUL © LFUL_ LFUL
cut cut cut cut
6 8 (Cana
LFUL = LFUL) ss LFUL FUL
LFUL 106 106 LFUL LFUL Mine LFUL cut cut. — cut cut
cut 1 cut 1cut 2 cut 2 cut 3 shutdown cut 4_ 5 7 9 4
i Alou feed <2 local. ‘ioe mnt ne re
alba | T T
Late +
13 | |
vy |_|
+

1} |_|
A
ab a | |
ee 4 1
TH a |
.
6 v la |
5 ° 4
Ab 7
3 - { |
2e | |
aR 1 |
or
“1 l | | |
200 400 600 800 1,000 1,200 1,400

TIME, days since 12/31/84

Figure 20.—Closure measured at ramp monitoring stations during mining. A, RX1 to RX4; B, RX5 to RX9.

1,600


24

PRESSURE,

CLOSURE, i

LFUL
cut 10 LFUL cut 11 LFUL cut 12
T T

1.4
I J I I

oN &

wra
rPeDO
<
NS

4 f ] l l | |

1,500 1,520 1,540 1,560 1,580 1,600 1,620
TIME, days since 12/31/84

cut 10 LFUL cut 11 LFUL cut 12
T T

1,640

a T

° e l | |

<4
1,500 1,520 1,540 1,560 1,580 1,600 1,620
TIME, days since 12/31/84

1,640

Figure 26.—Response of closure extensometers (A) and pressure cells (B) in cut 10 to mining.


BRRBRRRS

ILLUSTRATIONS—Continued

Location of ramp and ramp crosscut monitoring stations ......... bees
Closure measured at ramp monitoring stations during mining .........
Location of monitoring stations in ramp crosscut and measured closure . .
Response of MPBX-R1
Response of pressure cells in cut 5 to mining in cuts 6 through 9...
Location of stope monitoring stations and measured closure in cut 5 to mining in cuts 6 through 12
Gap above sandfill in cutS ...... .

Response of closure extensometers and pressure cells in cut 10 to mining . .

TABLE

Instrument installation information

UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT

ft foot psi pound per square inch
in inch yd yard
Ib pound yd cubic yard

pet percent



LFUL cut 1

106 cut 1 106 cut 2 LFUL cut 2
1 t 1 F

1 r

KEY
wi
w3
w4
WS
Et
£2
ES
E4

T
<mpeosn0

Second 106 backstope
4r round shot

i LN

:
5

T T

ne
ia

CLOSURE, in
o

5h

Second 106 backstope
round shot

Sar

T T

IT)
an

Sa aaa
!

100 200

300
TIME, days

400 500
since 12/31/84

600

Figure 11.—Horizontal stope closure at monitoring stations in cut 1 prior to backfill placement. A, Stations west (W)

of ramp; B, stations east (E) of ramp.


Lal hdgulijds oY
6-in broken rock

Figure 13.—Schematic of fill support system.

as was done in the pull test, actual load-carrying capacity
would be higher.

Five pressure cells and five closure extensometers were
installed along with fill reinforcement. Two additional
closure extensometers were installed in the 106 raise prep,
where they would not be subject to damage during back-
filling.

The LFUL stope was eventually filled with 3 to 4 ft of
cemented total mill tailings (Garpenburg fill) on the bot-
tom, while the rest of the stope was filled with uncemented
classified tailings. Despite the precautions taken to protect
and waterproof instruments installed in this cut, MPBX-1,
two of the five closure extensometers, and one of the five
pressure cells were damaged during backfilling.

The great length of this cut required that filling be done
in several stages behind several sandwalls. Variations in
fill properties and height were observed. The irregular
nature of the stope back and the 1-pct grade maintained
for drainage made filling the stope entirely to the back
impossible. Although no exact measurements were taken
of the void, it was readily observed to be at least 1 ft high
and could have been as high as 2 to 3 ft in places. This
void would not appear in an overhand mining sequence.
The consequences of such a space on rock bursting are
examined in the second report of this series (15).

Second Through Fourth LFUL Cuts

Cuts 2 and 3 of the LFUL were mined between July
1986 and January 1987, and cut 2 was filled during this

B

period. The fill back in these cuts performed better than
that in the I-drift, although some localized fill failures
occurred.

A mine shutdown in January 1987 provided mine engi-
neers with an opportunity to modify the fill reinforcement
system by replacing cable slings with 6-ft-long Dywidag
bolts. These bolts were installed vertically on 2- to 3-ft
centers and behaved in a manner similar to rebar in con-
crete (fig. 14). The bottom end of each bolt was driven
through wire mesh into loose muck on the floor. (The
wire mesh was eventually eliminated in later cuts without
any problems.) A nut and a plate were then placed on the
top of the bolt to serve as an anchor in the fill. As the
next cut was mined, the bottom ends of the bolts were
exposed. At this time, a second set of plates and bolts
were installed to tension the rebar and provide anchors for
fencing, Fencing extended along the entire back and down
the stope walls to within 1.5 ft of the bottom. Split-Set
bolts were used to fasten the fencing to stope walls. The
fencing protected miners from fly rock and localized rock
bursts (fig. 15).

During the mine shutdown, the decision was made that
all further mining on and below the 5300 level would use
the LFUL method. Mining resumed in August 1987 with
backfilling of cut 3, followed by excavation of cut 4. Cuts
2 through 4 were monitored by instruments installed in cut
1 and by two closure extensometers in the 106 raise.

Pressure cells monitored support pressure development
as mining proceeded past each cell. Three of the four
cells recorded substantial pressures when the mining face
was 10 ft away horizontally; these pressures peaked when
the face was 15 ft past and dropped off sharply with
further mining, indicating fracture of the cemented fill
(fig. 16). Fracturing was then confirmed by visual in-
spection. The fourth pressure cell (W3) was located in the
upper uncemented portion of the fill. This cell showed a
much smaller increase in pressure and no degradation of
Pressure as the underlying cemented fill fractured. The
Pressures measured in the cemented fill after fracturing
were surprisingly similar to those measured in the
uncemented fill.

Three of the pressure cells survived mining of the third
cut. The two installed near the center of the stope showed
increases of 40 to 70 psi, while the cell near the end of the
stope showed increases of only 10 psi. This difference may
have indicated a void at the top of the fill above the third
cell, or greater closure towards the middle of the stope, as
would be expected in an elastic rock mass.

MPBX-2 (fig. 12) showed that fractures between the
stope wall and the 5-ft anchor were compressed while
fractures between the 5-ft anchor and the 25-ft anchor
opened. Apparently, the loose rib rock was subjected to
confining pressures as fracturing rock expanded deeper in
the wall.


CONTENTS

Page

Abstract
Introduction ....
Acknowledgments
Mine geology and stope design
Conventional stoping technology
LFUL experimental stope plan ....
Instrumentation plan . .
Fill behavior ......
Back safety
Support pressure and stope closure .
Ramp stability
Rock mass response .
Rock mass deformation :
Change in stress . .
Extent of fractured rock .
Project chronology
First instrumentation cycle
Development, 106 I-drift, and first 106 backstope round

First LFUL cut mining and instrumentation
Second through fourth LFUL cuts .
Ramp
Second instrumentation cycle .
Third instrumentation cycle
Discussion and conclusions .
Fill performance .
Ramp stability . .
Rock mass response .

RRRRBSEBcmmmmiwauawaunneannnne

Mining method 25
References seeeeee 26
ILLUSTRATIONS
1. Location of Lucky Friday Mine .. 3
2. LFUL geologic map .. _ 3
3. Schematic of cut-and-fill mi 3
4, Schematic of LFUL . beens 5
5. Schematic showing sagmeter design and installation between mine floor and timber cap . 6
6. Closure extensometer . : . . 7
7, Project timeline ...... 8
8. Location of monitoring stations ‘and excavation sequence for 106 and LFUL stopes . 9
9, 106 I-drift fill reinforcement system ............c000ceeeeee beens 9
10. Location of monitoring stations and MPBX’s in cut 1 of LFUL ... sees 10
11. Horizontal stope closure at monitoring stations in cut 1 prior to backfill placement ... . cre
12. Response of MPBX-2 and MPBX-1 to cut excavation and location of MPBX’s 12
13. Schematic of fill support system .... : ee 13
14. Preparing Dywidag bolts for sandfill . 14
15. Hanging chain link fencing for overhead protection . 4
16. Response of pressure cells in cut 1 to mining in cut 2 . 4
17. Response of closure extensometers in cut 1 to mining in cut 2. 15

18. Excavation sequence in 106 and LFUL stopes and response of closure extensometers in 106 stope to
mining in LFUL stope ... 0.0.0.0. ccceeceeeeeeneeeaeenetenneenaee feet eeneeeeeees we 16


22

KEY
Before backfill gap
‘| After backfill gap
wife Instrumentation
110
fal
106 10
20
52 win
WS WI E1
oO wipe @ Of &
6
fat
8
9
10:
14
B LFUL LFUL. LFUL LFUL LFUL LFUL LFULLFUL
cut 5 cut 6 cut 7 cut 8 cut 9 cut 10 cut 11 cut 12
T T T T T T T
6
I I I I I I I T
sit pare | |
ate
£
3H eo +
ld
a
2 ge B
8
3] nas | |
oO
KEY
1 owl 4
@wia
@ wis
: Ows
oO #? Installed at cut 5 gap during @r a
cut 8 excavation
| | ! | | | | |

-1
1,150 1,200 1,250 1,300 1,350 1,400 1,450 1,500 1,550 1,600
TIME, days since 12/31/84

Figure 24.Location of stope monitoring stations (A) and measured closure (B) in cut 5 to mining in cuts 6 through 12.


Library of Congress Cataloging in Publication Data:

Rock mechanics investigations at the Lucky Friday Mine.
p. em, — (Report of investigations; 9432)
Includes bibliographical references (p. 26).

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 JK. 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 TJ. 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- . Il. Whyatt, J. K. (Jeff K) Ill. Poad, M. B. IV. Series:
Report of investigations (United States. Bureau of Mines); 9432,

TN23.U43 [TN317] 622 s—dc20—-[622’.28]__ 90-15069 CIP



DISCUSSION AND CONCLUSIONS

FILL PERFORMANCE

Although the survival rate for instruments installed in
the fill was poor, enough instruments did survive to give
g00d indications of fill performance. For instance, despite
significant localized variations, closure after one additional
cut averaged about 1.9 in for cut 1, 1.8 in for cut 5, and
less than 0.8 in for cut 10. This decrease in closure
appears to be in opposition to the tendency for greater
closure as mining opens a larger area. The corresponding
fill pressure and total load supported per foot of stope
length after one additional cut improved also. The instru-
ments showed that fill pressure increased from 40 psi in
cut 1, to 350 psi in cut 5, and finally to 800 psi in cut 10.
Adjusting for the portion of each cut actually filled, cut 1
fill (90 pet filled) produced about 52,000 Ib of support
force per foot of cut, cut 5 (60 pet filled) about 300,000 Ib,
and cut 10 (95 pet filled) more than 1,000,000 Ib.

Thus, fill performance steadily improved through the
life of this project. Continual improvement of the fill sys-
tem reduced the number and severity of local fill failures,
slowed stope closure, and increased the regional support
pressure generated in the fill. Whyatt, Williams, and
Board (15) present further analysis of LFUL backfill per-
formance and the implications for rock bursting in part 2
of this series.

RAMP STABILITY

Instrument readings generally showed good ramp
stability. The fracture zone ahead of mining appeared to
stabilize quickly after excavation and did not propagate far
into the ramp-stope pillar. However, some localized zones
of weaker rock were encountered that significantly in-
creased ramp closure and created ground control problems
in the ore pass. The ground support measures undertaken
were sufficient to stabilize these sections. The ramp
should not be in danger unless support is neglected or
zones of rock significantly weaker than those already
encountered are intersected. Rock bursts still pose a
threat in the ramp, as they do throughout the mine.

ROCK MASS RESPONSE
Monitoring the response of the rock mass was frustrat-

ed by instrument failures. Attempts to measure changes
in stress and fracture zone development below the mining

front failed completely. The only direct measurements of
rock mass response were obtained by MPBX-R1 installed
in the ramp pillar and the two MPBX’s installed in the
stope. However, these instruments were located above the
mining front. Thus, although they do provide some infor-
mation on the depth of the fracture zone in stope and
ramp walls, they do not describe the fracture and stress
concentration zones ahead of mining.

However, some indirect information on the relative rock
burst hazard associated with different backfill reinforce-
ment designs can be gleaned from the stope closure data.
Since the energy release rate is a known indicator of rock
bursting and is a direct function of closure, improvements
in backfill design that reduce closure do have a positive
influence on the probability of rock burst occurrence. An
analysis of this influence, emphasizing the role of gaps left
in the fill between levels, is presented in part 2 of this

series.
MINING METHOD

Geomechanical evaluation of a new mining method,
especially in a geologically complex environment such
as the Cocur d’Alene Mining District, is anything but
straightforward. The effort to monitor the geomechanical
performance of the experimental LFUL stope encountered
serious problems because of difficulties during instrument
installation and because most of the instruments failed,
Nonetheless, some insights were gained regarding three
geomechanical factors (backfill performance, ramp stabil-
ity, and rock mass response) that directly affected stope
performance. Surviving instruments, the analyses pursued
in later parts of this series, and mining experience showed
that the method did not suffer from serious problems in
any of these areas.

This conclusion, and positive operational and economic
considerations, have led to mine-wide adoption of the
mining method. Indeed, this study and mining experience
suggest that the underhand longwall method of mining
could be considered a practical and economical mining
method for rock-burst-prone ground throughout the Coeur
d’Alene District. Mining this experimental stope is the
culmination of years of research by Bureau investigators,
cooperating companies in the mining industry, and aca-
demia. This effort has successfully adapted the under-
hand longwall cut-and-fill mining method to conditions in
the Coeur d’Alene Mining District.


106 raise

[
[106 overhand stope

Gute
Cutt
PELE: -
o Fecers mo e@ oni]
{ cuz]
I MPBX Ri MPBX SC1-Et cuts T
wip y Guta ]
ews ow eof ef cus
yp 1a Cuts |
ia Cut7
ial ‘Cuts ]
{ it cuts
I ews ews OI @re Cutt}
[ ial cut]
t ial Cutt2 TT
L al Cut 13
[ ial Cuti4 ]
z z z
g & g

rounds in 106, the I-drift was prepared to be backfilled as
an underhand stope. Ten- by ten-inch caps were placed
from wall to wall 3 ft off the floor, and wire mesh was
hung at floor level from J-bolts and straps (fig. 9).

The I-drift was filled with 5 ft of cemented total mill
tailings (called Garpenburg fill at the mine) with a cement-
to-sand ratio of about 1:5 and a slurry density of about
70-pet solids, Laboratory tests were run to define the
uniaxial strength and friction angle of the fill material.
Samples of cemented fill tested after curing for 30 to 45
days showed uniaxial strengths from 300 to 800 psi,
depending primarily on sampling location. Segregation
and other problems hampered placement of a uniform fill
mix.

First LFUL Cut Mining and instrumentation

The LFUL ramp system reached the vein for the first
LFUL cut on April 9, 1985. This cut was driven from the
ramp as an I-drift east through the 110 block and as an
underhand stope west under the backfilled 106 I-drift. The
LFUL and the 106 stopes were mined simultancously with
only 5 ft of intervening fill. The first LFUL cut was
completed in October 1985, Mining in the 5100-106 stope
continued until April 11, 1986, when it was halted prior to
mining the second LFUL cut. This left the 106 mining
floor 40 ft above the back of the first LFUL cut.

Figure 8.—Location of monitoring stations and excavation sequence for 106 and LFUL stopes.

Nos}

Broken muck

"J" bolt hangers

Figure 9.—106 I-drift fill reinforcement system.


n

south wall was converging and fracturing more rapidly
than the north wall (fig. 12) as cut 1 was finished. This
phenomenon was visually evident and agreed with data col-
lected by the mine-wide microseismic system (16), which
indicated that most seismic activity and fracturing were
occurring in the south wall. The seismicity pattern in the
south wall did not appear during subsequent cuts, and an
explanation for this pattern was never established.

The LFUL stope remained unfilled from October 1985
to June 1986 while mining continued in the 106 stope

above. The LFUL stope was prepared for sandfill place-
ment by drilling holes in the wall at floor height, from
which 1/2-in wire cable slings were installed from wall to
wall (fig. 13). A wire mesh mat was laid over the cables.
Engineers at Hecla Mining Co. (2) conducted pull tests of
cable wedged into Split-Set rock bolts to evaluate the load-
carrying capacity of the anchor. It was found that a typical
cable held 8,000 Ib of load before sliding out of the Split-
Set anchor. Since actual loading direction in the LFUL
stope is at an angle to, instead of parallel to, the Split-Set,

LFUL cut 1 106 cut 1 106 cut 2 LFUL cut 2 LFUL cut 3
1 of ot 1 cr
24
A T T T T li
= KEY 4
18 © 5-ff anchor =|
4 15-ft anchor
@ 25-ft anchor
1B y 40-ff anchor 4
12 4
2 Val 4
-2-o—v
£
“6 +
ur a
[4
2
6 3- 4
ot
ra)
ol wef | 1 l
250 350 450 550 650 750
0.6
Cc N
B T I
5 y- waanad q No. 2 shaft
4h 4
Be 4
2h 4
ah 4
° | |
250

350 450
TIME, days since 12/31/84

Figure 12._Response of MPBX-2 (A) and MPBX-1 (B) to cut excavation. Location of MPBY’s (C).