Analysis of Rockfall and Blasting Backbreak Problems, US 550, Molas Pass, CO

9 Pages·1.085 MB·English

by  Gates W.C.B., Ortiz L.T., Florez R.M.

ARMA/USRMS 05-671

Analysis of Rockfall and Blasting Backbreak Problems, US 550, Molas
Pass, CO

William C. B. Gates
Kleinfelder, Bellevue, Washington, USA
L. Ty Ortiz
Colorado Department of Transportation, Denver, Colorado, USA
Robert M. Florez
Colorado Department of Transportation, Denver, Colorado, USA

Copyright 2005, ARMA, American Rock Mechanics Association

This paper was prepared for presentation at Alaska Rocks 2005, The 40th U.S. Symposium on Rock Mechanics (USRMS): Rock Mechanics for Energy, Mineral and Infrastructure
Development in the Northern Regions, held in Anchorage, Alaska, June 25-29, 2005.
This paper was selected for presentation by a USRMS Program Committee following review of information contained in an abstract submitted earlier by the author(s). Contents of the paper,
as presented, have not been reviewed by ARMA/USRMS and are subject to correction by the author(s). The material, as presented, does not necessarily reflect any position of USRMS,
ARMA, their officers, or members. Electronic reproduction, distribution, or storage of any part of this paper for commercial purposes without the written consent of ARMA is prohibited.
Permission to reproduce in print is restricted to an abstract of not more than 300 words; illustrations may not be copied. The abstract must contain conspicuous acknowledgement of where
and by whom the paper was presented.

ABSTRACT: Molas Pass is located in Southwestern Colorado on US 550. In August 2003, Colorado Department of
Transportation (CDOT) began a rockfall mitigation project to alleviate potential dangers caused by rockfall and/or rockslides and
to improve the sight distance and reduce the blind corner caused from the rockslope. The work involved rock excavation by
production blasting, rock reinforcement with dowels and rock bolts and rockfall mitigation using wire mesh drape. A contractor’s
access road was constructed across the upper rock face about 85 feet above the shoulder of the highway to facilitate excavation of
the rockslope. The excavation created an unstable highwall specifically a large rock block above the access road. Blasting of the
access road exacerbated the rockslope instability of the high wall by day lighting the failure plane and unweighting the toe of the
slope of the highwall. The stability of the highwall was aggravated by the excessive backbreak created from blasting. A
combination of factors in the blasting such as over-stemming of the shot holes and short timing delays in the firing sequence
between the second and back shot hole rows may have led to the severe backbreak. Moreover, the adverse geologic structure
appears to have exacerbated the excessive backbreak.
investigation of the rock slope after the Colorado
1. INTRODUCTION
Department of Transportation (CDOT) observed a
1.1. Background large kinematically unstable block on the highwall
Excavation and stabilization of unstable rockslopes and severe backbreak problems created from
can present many challenges to the engineer. To blasting.
reduce some of the uncertainty it is critical that the
1.2. Location
engineer have a clear understanding of the problem
Molas Pass is located in Southwestern Colorado on
prior to developing a design for the rock excavation,
US 550 between Durango and Silverton. In August
blasting, stabilization and costs.
2003, CDOT, Region 5 began a rockfall mitigation
The following paper describes a case history project at mile marker 68.5 about 1.5 miles south of
attendant to analysis of an unstable rockslope Silverton to mitigate potential dangers caused by
resulting from rock excavation and blasting rockfall and/or rockslides and to improve the sight
backbreak problems. In this case the contractor distance and reduce the blind corner caused from
developed a construction, excavation and blasting the rockslope. At this location, the highway formed
plan without a clear understanding of the rock- a sharp blind corner and pinch point. The
engineering problem. As the project progressed, overhanging rock not only provided a ramp to pitch
excavation and blasting led to a critically unstable rockfall onto the highway; the overhang had a
slope. The authors conducted this geotechnical history of catching large trucks (Refer to Figure 1).
installed to reinforce portions of the upper and
lower slopes. However, the large unstable block on
the highwall had largely been ignored and further
had not been reinforced. Moreover, the contractor
was preparing to drape the entire slope with steel
wire mesh to mitigate rockfall before stabilizing the
unstable block.
2. GENERAL GEOLOGY
Between Durango and Silverton, US-550 runs north
south through the San Juan Mountains. Glaciation
has formed the alpine morphology of the area. Just
south of Silverton at the site location, the highway
cuts through Tertiary intrusive rock consisting
mostly of a fine-grained granodiorite with talus
deposits mapped at the southern end of the site [1,
Fig. 1: Trucks at pinch point below rock overhang on 2]. Gold and silver mining was prevalent in the
US-550 at MP 68.5. area in the late 1800’s and was still occasionally
The project was scheduled through the CDOT active into the mid 1980’s. The site itself is located
Rockfall Mitigation Project Plan. The site is above the Champion mine.
classified as a Category A Rockfall Hazard
according the CDOT’s Rockfall Hazard Rating
System.
To complete the work, CDOT retained a contractor
to provide rock excavation by production blasting,
rock reinforcement with dowels and rock bolts and
Fault
rockfall mitigation using mesh posts and wire mesh
drape.
Highwall &
1.3. Site Description Unstable Block
At the request of CDOT, the authors visited the
rockslope in June 2004. At the time of the visit, the
rockslope was under construction. The rockslope is
approximately 40 meters (130 feet) high (at the
highest point) and 122 meters (400 feet) long.
Access Lower
Based on a vertical scan line from approximately
Road Slope
the highest point of the rockslope, we observed two
discrete sections in the rockslope (Figure 2). The
upper near vertical section, referred to, as the high
wall, was about 14 meters (45 feet) high with a
slope angle of about 83 degrees. According to
CDOT the high wall was created as a result of an
initial pioneer access road. The lower 25 meters (86
feet) of the rockslope formed a slope angle of about
63 degrees dipping directly (southeast) towards the
road. A distinct undulating ledge dipped out of
slope at approximately 37 to 42 degrees and formed
the boundary between the upper and lower sections
of the rockslope.
The contractor had previously scaled the rockslope. Fig. 2: Profile of rock cut showing the highwall,
As well, spot rock bolts and dowels had been unstable block, the lower slope and pioneer access
road.
3. ENGINEERING GEOLOGY the horizontal and vertical scanlines. RQD is a
directionally dependent parameter and its value may
Our first objective when we arrived onsite was to
change systematically depending upon borehole
develop an accurate picture of the engineering
orientation. Therefore, using the volumetric joint
geology and its relationship to the rockslope.
count by Planström [6] is useful in reducing this
Because many of the geological features on the
directional dependence. To evaluate J in the field
rockslope were inaccessible for standard horizontal v
one must select an open cleft in the rock which
scan line mapping; the authors employed a
displays x, y and z dimensions. One then sums the
combination of horizontal and vertical scan line
fractures along a 1-meter length in 3-dimensions to
mapping techniques. Mountaineering rappelling
obtain a volumetric joint count.
techniques were used to evaluate the steep face of
the rockslope. Critical geomechanical information 3.2. Vital Geomechanical Statistics
on the rock mass is collected from the vertical and The rockslope consists of slightly to moderately
horizontal scanlines. weathered to fresh, fine-grained gray granite. The
granite outcrop is part of a larger Tertiary intrusive
3.1. Geomechanical Rock Mass Classification
body [1, 2]. The raw RMR (RMR without
Two of the more widely accepted classifications
consideration of orientation of discontinuities)
systems for evaluating the rockmass are the Rock
estimated during our vertical scan line ranged from
Mass Rating System (RMR) by Bieniawski [3], and
fair rock on top to good rock at the base of the
the Geological Strength Index (GSI) by Hoek and
rockslope. However, the orientations of the
Brown [4]. The GSI system is more applicable than
discontinuities dip unfavorably out of the rockslope,
the RMR system when rating very weak rock
which decreases the overall RMR classification to
masses where the stability of the rock mass is
poor and fair rock.
controlled rockmass strength. It was apparent from
the massive nature of the rock that structure The geologic hammer is an excellent tool to
controlled the slope stability and not the overall estimate rock strength [7]. The hammer responded
rockmass strength so the authors used RMR system with a distinct bounce and loud ring after striking
during our field evaluation. the granitic rock suggesting the rock is very strong
with an average compressive strength for the intact
The RMR classification system considers the
rock of about 150 MPa (21,750 psi). The RQDs
following geomechanical characteristics of the
ranged from 50 to 90 (fair to good rock quality).
rockmass:
Spacing of the discontinuities average about 140mm
• Strength of the intact rock material (5.5 inches). 90 percent of the discontinuities
• Rock quality designation (RQD)* exhibit apertures ranging from 1mm to 5mm (0.04 –
• Spacing of discontinuities 0.2 inches). The surface of the fractures appeared
• Condition of discontinuities rough and undulating (Joint Roughness Coefficient
• Groundwater conditions (JRC) ~20). In some of the sub-vertical fractures
we observed dry clay. We did not observe clay on
• Orientation of the discontinuities relative to
the major discontinuity that separates the upper and
the rockslope
lower portions of the rockslope. Moreover, we did

not observe seeps or springs on the slope; the slope
*Note: Deere [5] developed the rock quality
was dry.
designation (RQD) technique, which is simply
3.3. Critical Geologic Structures
estimated from percent of rock core recovery ≥ 10
Simple examination of the rockslope from the
cm (4 inches) compared to the total run. However,
shoulder of the roadway indicated major
in the field one may not have access to drill core.
discontinuities dipping out of slope that could be
Therefore, Planström [6] developed a technique to
kinematically unstable. Both historical planar and
estimate RQD while evaluating the rock face,
wedge block failures had obviously occurred
where:
previous to construction of the highway. Moreover,
RQD % = 115 – 3.3J . (1)
v during production blasting the rock separated from
In this case, J is the total number of discontinuities the slope face and slid downslope naturally as
v
per cubic meter or volume of rock selected along planar and wedge failures.
One of the structures of additional concern was a angle (φ), the factor of safety is less than 1 and
fault clearly separating the upper unstable block kinematicly unstable.
from the main rock face (Figure 2).
The basic concept of kinematic analysis for plane
Figure 3 is an equal-area stereonet with a Markland failure is straightforward. Two conditions must be
test plot for the rockslope developed using the met for sliding to occur. First, the discontinuity
computer code Rockpack®. Watts [8] has developed must dip more steeply than its friction angle.
a computer code Rockpack® which applies the Secondly, for sliding to occur, the discontinuity
must daylight the slope face in a down-dip
direction. These two conditions may be represented
on a stereonet in the form of a crescent-shaped
critical zone (Refer to Figure 3). Dip vectors that
plot in this critical zone are considered
kinematically unstable. The most vulnerable area
within the critical zone occurs with in ± 20° of the
slope face dip direction. Figure 3 displays at strong
population of dip vectors for discontinuities which
daylight the critical zone suggesting that the
possibility for planar failures are very likely. Figure
4 is a stereonet displaying pole concentrations for
major joint sets observed in the rockslope. Figures 3
and 4 display three prominent joint sets (Joint sets
1,2 and 7) that plot in the critical zone and control
the kinematic stability of the rockslope.
Fig. 3: Dip vector plot equal – area stereonet with
Markland analysis of rockslope.
Markland analysis and dip vectors to facilitate
evaluation of the kinematic stability of a rockslope.
The Markland analysis indicates that the rockslope
is kinematicly unstable. Markland uses dip vectors
of the individual discontinuities to analyze the
kinematic stability of the slope. This plot may be
used to analyze potential translational failures
including planar and wedge failures. A Markland
analysis is also useful for identifying potential Fig. 4: Pole plot equal – area stereonet displaying major
joint sets. Arrow points in general direction (110º) of
toppling failures.
potential planar failure and plunge of potential wedge
According to Watts [8], the kinematic analysis is a failures.
“cohesion-equals-zero” analysis in which the effects
Stereonet analysis for potential wedge failures is
of cohesion are ignored. With cohesion equal to
similar to that for plane failures. For a wedge
zero, the fundamental limiting equilibrium equation
failure to occur, the line made by the intersection of
for computing the safety factor of a block with the
the planes creating the wedge must plunge more
potential to slide out of the slope face on one
steeply than the friction angle and less steeply than
discontinuity (joint) reduces to the following:
the dip of the slope face and in a direction that
Tanφ daylights the slope face. To analyze for potential
FS = (2)
wedge failure, two great circles are constructed
Tanθ
which bound the ranges of dip vector clusters. If
As a result, whenever the dip value of the potential
the wedge formed by the intersection of the great
failure plane (discontinuity) (θ) exceeds the friction circles plots within the critical zone, potential
wedge failures may occur. Figures 3 and 4 show
large steeply bound wedge intersections within the
critical zone suggesting the potential for wedge
failure. Joint set 5 intersects Joint sets 1, 2 and 7
forming steeply bounded unstable wedges.
Stereonet analysis for potential toppling failure
using Markland analysis is relatively
straightforward. The presence of discontinuities
that dip approximately 180 degrees from the dip
Fault
direction of the slope face (± 30°) and dip into the
slope at a steep angle suggest the potential for
toppling failure. On the Markland test plot, these
discontinuities plot within the shaded triangle.
Figures 3 and 4 display the kinematic potential for
toppling. Joints with orientations that plot within
the shaded triangle (Figure 3) strike nearly parallel
to the slope and therefore could also form tension
fractures bounding potentially unstable blocks
(including planar and wedge shaped blocks).
After evaluation of the upper unstable block, it was 37º
clearly apparent that Joint sets 1,2 and 7 were
critical to the stability of the upper slope. Moreover,
planar failure appeared to be the primary mode of
failure of the large block on the highwall.
3.4. Stability Analysis
Figure 5 displays the critical block on the south
Fig. 5: Unstable block displaying potential planar failure
aspect of the highwall of the upper slope. The
plane. Construction of access road day lighted the failure
block rests on a potential failure plane that dips out
plane. Fault is on the left inside of the block.
of slope (daylights) at approximately 37 to 41
To evaluate the stability of the block, we assumed
degrees. The dip of this discontinuity falls with the
the following:
range of friction angles expected along an unfilled
joint in granitic rock [4, 6, 8]. The toe of the block • The block is marginally stable (i.e. factor of
and failure plane was exposed during construction safety (FS) = 1).
of the access road to the face. The critical block is
• Cohesion (C) is near zero along the failure
about 14 meters (45 feet) high, and the face of the
plane.
block dips to the southeast to the highway at
approximately 83 degrees. The top of the block • The peak friction angle (φ) for the
slopes back at about 35 degrees. The back of the discontinuity is about 37º (same as the dip
block is bounded by a fault that strikes to the plane).
southeast (Refer to Figures 2 and 5).
• Currently, the block is free draining.
Because the rock block on the high wall was
The authors estimated 37º for peak friction by
kinematicly unstable, our next step was to
conducting a series of tilt tests using rock blocks
investigate the global stability. We assumed the
with joint faces similar to that of the rock mass. Of
block had moved because the toe of the block had
interest 37º is the same value as the dip of the
been un-weighted and the potential failure plane
failure plane. This suggests the factor of safety for
daylighted by excavation of the access road. During
the rock block is near 1 (Refer to equation 2).
our investigation we observed no open tension
Because of the roughness of the fractures, we
fractures behind the block suggesting that it has not
assumed 37º to be peak friction. Residual friction,
moved recently.
after smoothing out the rough asperities would be
closer to 34º.
To verify the strength characteristics of friction and
cohesion, we conducted a back analysis of the
stability of the block using the limit of equilibrium
method facilitated by computer code ROCPLANE®
by Rocscience®[9]. Figure 6 displays the results of
the back analysis on the unstable block. Based on
Fig. 6: Back analysis of unstable block using computer
code Rockplane® by Rocscience®: φ = 37º, C = 0, FS ~ 1. Fig 7: General patterned layout for rock dowels on the
face of the unstable rock block.
the field evaluation and back analysis the block was pressures behind the block and along the basal
marginally stable. However, any critical outside discontinuity.
forces such, as hydrostatic pressure from water
4. EVALUATION OF OVERBREAK
along the failure plane or seismic activity will
induce failure. 4.1. General
Wyllie and Mah [10] have shown that slope
3.5. Stabilization of Upper Unstable Block
instability is often related to blast damage of the
Based on the results of the kinematic and stability
rockslope behind the face. The damage is often
analysis we concluded that the unstable block must
related to overbreak. Backbreak can originate from
be reinforced to increase the factor of safety to at
a poor blasting design exacerbated by unfavorable
least 1.3. We recommended to the contractor to
structural geology.
install post-tensioned rock bolts or passive dowels
on a patterned basis on the high wall face. Overbreak is excessive rock breakage beyond the
excavation limits [11, 12]. Overbreak includes
Figure 7 is a photo of the face of the hanging block
backbreak, which is breakage behind the last row of
with a general patterned layout of the rock dowels.
shot holes and endbreak that occurs at the end of the
The rock face measures about 12m (40 ft) high by
shot line. One of the problems that lead to
11m (35 ft) wide. The contractor elected to install
destabilization of the upper rockslope was at least 6
25 rock dowels on a patterned layout and spaced at
meters (20 feet) backbreak beyond the planned
approximately 2x2m (7x7 ft). Locations for the
excavation line for the present rock cut (Figure 8).
dowels were adjusted depending on the asperities of
This pushed the access road into the toe of the upper
the rock face and previous. Each dowel was seated
slope day lighting the failure plane of the upper
to about 111 kn (25 kips) using a torque wrench.
block.
For our stability analysis, we assumed that the slope
4.2. Structural Geology
is free draining. However, since the rockslope is in
Geology plays a very important role on the outcome
an area that experiences extreme weather changes,
of a blasting job on any type of rock excavation
we recommended the contractor install PVC drains
[13]. Typically poor blasting results occur in rock
to intersect the fault trace behind the block. This
where the strike of the structures (faults and joints)
would help mitigate potential buildup of water
• Burden dimension (distance to free face
from the shot hole).
• Stiffness ratio (ratio of burden to bench
height).
• Stemming depth of shot hole.
• Subdrilling depth of shot hole.
• Loading density of shot hole.
• Powder column length of shot hole.
• Total weight of explosive per shot hole.
• Powder factor per cubic volume of rock.
Fig. 8: Severe backbreak caused from poor blasting • Vibration and scale distance formula.
design exacerbated by adverse structural geologic
• Timing sequence between shot holes.
conditions.
• Spacing between shot holes.
are parallel or perpendicular to the free face of the
• Potential for violence to establish if mats are
rock slope. Blasting results are typically good if the
strike of the structures is oblique to the free rock needed.
face. Strike parallel to the free face of the rockslope
The authors carefully evaluated the design for each
may cause severe backbreak and endbreak [13].
blast performance using these steps. As we went
On this rockslope, several joint sets dip out of slope
through each step, we would establish if the design
forming potential planar and wedge blocks (See
was within design limits. We asked the following
Figures 2 through 5). Where the blast occurred with
question. Did the design make sense and was it
the dip of the slope, CDOT observed severe
reasonable?
backbreak. Blasting with the dip of the rock
There are several design factors that may lead to
structure apparently forced the explosive energy up
backbreak. Konya and Walter [11, 12] describe
and along the preexisting joint sets. This loosened
some of the causes for backbreak:
and displaced the rock along the dip planes
backbreaking beyond the design excavation line • Excessive burden
resulting in planar and wedge rock failures. This is
• Excessively stiff benches
very evident on Figure 2, where the dip of the
structure runs uphill into the access road and • Long stemming depths on stiff benches
highwall.
• Improper timing delay
4.3. Blasting Design Evaluation
Burden (B) is the shortest distance of rock between
Backbreak may result from several problems in the
the shot line and the nearest free face or open face
blasting design performance. Konya and Walter
at the time of detonation. Backbreak can result from
[11, 12] have developed a logical set of steps to
excessive burden on the shot holes, thereby causing
follow when evaluating blasting designs. The
the explosive to break and crack radially further
following steps for evaluation of the blasting design
behind the last row of shot holes.
were modified after Konya and Walter [11, 12]:
The stiffness ratio (L/B) is the ratio of bench height
• Type of blasting (production or controlled)
(L) to burden (B). Excessively stiff benches (L/B <
• Drilling equipment capability and proposed 2) cause more uplift and backbreak near the collar
borehole diameter. of the shot hole. According to Konya and Walter
[11, 12] for better blast results, stiffness ratios
• Explosive selection for specific site
should be in the range of 3 to 4.
conditions.
• Specific gravity of rock and explosives.
Stemming (T) is inert material such as drill cuttings According to Wyllie and Mah [10] control of
packed around the collar of the blast hole to confine overbreak damage to the final walls can be limited
the gasses during detonation. Long stemming by implementing a combination of the following
depths on stiff benches promotes backbreak. measures. Production blasting should be designed to
limit backbreak behind the final wall. Controlled
Improper timing delay from row-to-row may cause
blasting techniques such as line drilling, pre-
backbreak if the timing is too short, thereby
shearing, and cushion blasting will limit the fracture
resulting in excessive confinement of the gases in
creep up the dip planes during detonation by
the last row of the shot.
defining a final rock face.
4.4. Results of blasting design evaluation
5.2. Test Shots
Typically for rockslopes, a knowledgeable
Before blasting, a series of test shots evaluating the
contractor will employ controlled blasting
performance of the controlled and production
techniques such as line drilling, pre-shearing or
blasting should be conducted to evaluate how the
cushion blasting for backbreak control of the back
rock and geologic structure responds to the blast
wall coupled with production blasting to break up
design. This is critical to a good design and
the rest of the rock. On this project, the contractor
blasting performance.
used production blasting to bring the rock down. It
was apparent that the structural geology was not 6. CONCLUSIONS
considered in the design. Moreover, test blasting to
From the investigation, the authors concluded the
evaluate the design was not done.
following:
After review of design issues that lead to backbreak
• The upper rockslope was kinematicly
such as excessive burden, stiff benches, excessive
unstable with major blocks dipping out of
stemming and short timing delays, the burden and
slope that required reinforcement with post-
stiffness ratio appeared within recommended design
tensioned rock bolts or dowels.
limits. However, two design issues appear to have
worsened the backbreak along the wall; over • Construction of the contractor’s access road
stemming of the shot holes and short timing delay. exacerbated the rockslope instability of the
high wall by day lighting the failure plane
Konya and Walter [11, 12] recommend stemming of
and unweighting the toe of the slope of the
the shot holes to be about 0.7 of the burden. In this
high wall rendering the slope kinematicly
case the stemming material was twice as long as
unstable.
recommended. This led to backbreak of the last row
of shot holes near the excavation limit. • Back analysis of the stability of the high
wall using the limit of equilibrium method
In addition, the firing timing delay appeared to short
indicated a factor of safety of 1 or the slope
between the last row and the second row of shot
is marginally stable.
holes. Timing delay was about 2 milliseconds.
Konya and Walter [11, 12] recommend a time delay • To achieve a factor of safety of 1.3 or
between the back holes of 4 to 6 milliseconds. This greater the high wall required reinforcement
will result in a scattered muck pile with minimum with post-tensioned rock bolts or dowels.
backbreak. An increase by 2 or 3 milliseconds
• A combination of factors in the blasting
between the second and last row of shot holes
design such as over-stemming of the shot
would have allowed the earlier shots to move the
holes and short timing delays in the firing
burden out of the way. This would have reduced the
sequence between the second and back shot
resistance on the last row of shot holes and reduced
hole rows led to the severe backbreak.
the pressure on the back wall. The end result would
have been cleaner rock breaks with fewer overbreak • The adverse geologic structure dipping out
problems. of slope exacerbated the excessive
backbreak.
5. SOLUTIONS TO MITIGATE POOR BLAST
PERFORMANCE • Controlled blasting techniques such as line
drilling, cushion or pre-shearing blasting
5.1. Controlled Blasting
may have reduced or prevented the
excessive backbreak.
• Blasting test shots should have been
conducted to finalize the blast design to
establish optimum shot hole layout and shot
hole charge compatible with the structural
geology.
• The contractor should have become familiar
with the geologic problems to establish
construction, blasting methods and costs
prior to excavation of the rock.
REFERENCES
1. Ludke, R.G. and W. S. Burbank, 2000, Geologic Map of
the Silverton and Howardsville Quadrangles,
Southwestern Colorado, United States Geological Survey.
2. Chronic H. and Williams F. 2002. Roadside Geology of
Colorado. Mountain Press Publishing Company, 269.
3. Bieniawski, Z.T. 1989. Engineering Rock Mass
Classifications, John Wiley & Son, 251.
4. Hoek, E. and E. T. Brown. 1997. Practical Estimates of
Rock Mass Strength, International Journal of Rock
Mechanics and Mineral Sciences, Vol. 34, No. 8, pp.
1165-1186.
5. Deere, D. U.1963. Technical description of rock cores for
engineering purposes, Felsmechanik und
Ingenieugeologie, 1(1). 16-22.
6. Planstrom, A. 1975. “Karakterisering av
Oppsprekningsgrad og Fjellmassers Kvalild” Internal
Report, Ing. A. B. Berdal A/S, Oslo, Norway. 1-26.
7. International Society of Rock Mechanics (ISRM). 1981.
Suggested Methods for the Quantitative Descriptions of
Discontinuities in Rock Masses. ed. E.T. Brown,
Pergamon Press, Oxford, UK. 211.
8. Watts, C.F. 2001. Rockpack III, Slope Stability
Computerized Analysis Package Reference Manual,
Radford University, 48pp. and Appendices.
9. ROCPLANE Version 2.0. 2003. Stability Analysis
Software for Translational Failures. RocSceince Inc.
Toronto, Canada.
10. Wyllie, D.C. and C. W. Mah. 2004. Rock Slope
Engineering, Civil and Mining, 4th ed., Spoon Press, 431.
11. Konya, C.J. and E. J. Walter. 1991. Rock Blasting and
Overbreak Control, 1st Edition, FHWA-HI-92-001,
National Highway Institute, 430.
12. Konya, C.J. 2003. Rock Blasting and Overbreak Control,
2nd Edition, FHWA-HI-92-001, National Highway
Institute, 399.
13. International Society of Explosive Engineers (ISEE).
1998. Blasters Handbook, 17th ed., ed R.B. Hopler,
International Society of Explosive Engineers, 742.