ARMA/USRMS 05-671 Analysis of Rockfall and Blasting Backbreak ProblemPass, 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 MechDevelopment 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 eas presented, have not been reviewed by ARMA/USRMS and are subject to correction by the author(s). The material, as presented, does noARMA, their officers, or members. Electronic reproduction, distribution, or storage of any part of this paper for commercial purposes withouPermission to reproduce in print is restricted to an abstract of not more than 300 words; illustrations may not be copied. The abstract must conand by whom the paper was presented.

1. INTRODUCTION

1.1. Background Excavation and stabilization of unstable rockslopes can present many challenges to the engineer. To reduce some of the uncertainty it is critical that the engineer have a clear understanding of the problem prior to developing a design for the rock excavation, blasting, stabilization and costs.

The following paper describes a case history attendant to analysis of an unstable rockslope resulting from rock excavation and blasting backbreak problems. In this case the contractor developed a construction, excavation and blasting plan without a clear understanding of the rock-engineering problem. As the project progressed, excavation and blasting led to a critically unstable slope. The authors conducted this geotechnical

investigation of the rockDepartment of Transporlarge kinematically unstand severe backbreak prblasting.

1.2. Location Molas Pass is located inUS 550 between Durang2003, CDOT, Region 5 project at mile marker 6Silverton to mitigate porockfall and/or rockslidedistance and reduce the the rockslope. At this loa sharp blind corner andoverhanging rock not onrockfall onto the highwahistory of catching large

ABSTRACT: Molas Pass is located in Southwestern Colorado on US 550. In August 2003, ColoTransportation (CDOT) began a rockfall mitigation project to alleviate potential dangers caused bto improve the sight distance and reduce the blind corner caused from the rockslope. The work inproduction blasting, rock reinforcement with dowels and rock bolts and rockfall mitigation usingaccess road was constructed across the upper rock face about 85 feet above the shoulder of the hithe rockslope. The excavation created an unstable highwall specifically a large rock block aboveaccess road exacerbated the rockslope instability of the high wall by day lighting the failure planeslope of the highwall. The stability of the highwall was aggravated by the excessive backbreak crcombination of factors in the blasting such as over-stemming of the shot holes and short timing dbetween the second and back shot hole rows may have led to the severe backbreak. Moreover, thappears to have exacerbated the excessive backbreak.

s, US 550, Molas

anics for Energy, Mineral and Infrastructure

arlier by the author(s). Contents of the paper, t necessarily reflect any position of USRMS, t the written consent of ARMA is prohibited. tain conspicuous acknowledgement of where

slope after the Colorado tation (CDOT) observed a able block on the highwall oblems created from

Southwestern Colorado on o and Silverton. In August began a rockfall mitigation 8.5 about 1.5 miles south of tential dangers caused by s and to improve the sight blind corner caused from cation, the highway formed pinch point. The ly provided a ramp to pitch y; the overhang had a trucks (Refer to Figure 1).

rado Department of y rockfall and/or rockslides and volved rock excavation by

wire mesh drape. A contractor’s ghway to facilitate excavation of the access road. Blasting of the and unweighting the toe of the eated from blasting. A elays in the firing sequence e adverse geologic structure

The project was scheduled through the CDOT Rockfall Mitigation Project Plan. The site is 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 rockfall mitigation using mesh posts and wire mesh drape.

1.3. Site Description 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. Based on a vertical scan line from approximately 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. As well, spot rock bolts and dowels had been

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, 2]. Gold and silver mining was prevalent in the area in the late 1800’s and was still occasionally active into the mid 1980’s. The site itself is located above the Champion mine.

Fig. 1: Trucks at pinch point below rock overhang on US-550 at MP 68.5.

Fault

Highwall & Unstable Block

Access Road

Lower Slope

Fig. 2: Profile of rock cut showing the highwall, unstable block, the lower slope and pioneer access road.

3. ENGINEERING GEOLOGY

Our first objective when we arrived onsite was to develop an accurate picture of the engineering geology and its relationship to the rockslope. Because many of the geological features on the rockslope were inaccessible for standard horizontal scan line mapping; the authors employed a combination of horizontal and vertical scan line mapping techniques. Mountaineering rappelling techniques were used to evaluate the steep face of the rockslope. Critical geomechanical information on the rock mass is collected from the vertical and horizontal scanlines.

3.1. Geomechanical Rock Mass Classification Two of the more widely accepted classifications systems for evaluating the rockmass are the Rock Mass Rating System (RMR) by Bieniawski [3], and the Geological Strength Index (GSI) by Hoek and Brown [4]. The GSI system is more applicable than the RMR system when rating very weak rock masses where the stability of the rock mass is controlled rockmass strength. It was apparent from the massive nature of the rock that structure controlled the slope stability and not the overall rockmass strength so the authors used RMR system during our field evaluation.

The RMR classification system considers the following geomechanical characteristics of the rockmass:

• Strength of the intact rock material • Rock quality designation (RQD)* • Spacing of discontinuities • Condition of discontinuities • Groundwater conditions • Orientation of the discontinuities relative to

the rockslope

*Note: Deere [5] developed the rock quality designation (RQD) technique, which is simply estimated from percent of rock core recovery ≥ 10 cm (4 inches) compared to the total run. However, in the field one may not have access to drill core. Therefore, Planström [6] developed a technique to estimate RQD while evaluating the rock face, where:

RQD % = 115 – 3.3Jv. (1)

In this case, Jv is the total number of discontinuities per cubic meter or volume of rock selected along

the horizontal and vertical scanlines. RQD is a directionally dependent parameter and its value may change systematically depending upon borehole orientation. Therefore, using the volumetric joint count by Planström [6] is useful in reducing this directional dependence. To evaluate Jv in the field one must select an open cleft in the rock which displays x, y and z dimensions. One then sums the fractures along a 1-meter length in 3-dimensions to obtain a volumetric joint count.

3.2. Vital Geomechanical Statistics The rockslope consists of slightly to moderately weathered to fresh, fine-grained gray granite. The granite outcrop is part of a larger Tertiary intrusive body [1, 2]. The raw RMR (RMR without consideration of orientation of discontinuities) estimated during our vertical scan line ranged from fair rock on top to good rock at the base of the rockslope. However, the orientations of the discontinuities dip unfavorably out of the rockslope, which decreases the overall RMR classification to poor and fair rock.

The geologic hammer is an excellent tool to estimate rock strength [7]. The hammer responded with a distinct bounce and loud ring after striking the granitic rock suggesting the rock is very strong with an average compressive strength for the intact rock of about 150 MPa (21,750 psi). The RQDs ranged from 50 to 90 (fair to good rock quality). Spacing of the discontinuities average about 140mm (5.5 inches). 90 percent of the discontinuities exhibit apertures ranging from 1mm to 5mm (0.04 – 0.2 inches). The surface of the fractures appeared rough and undulating (Joint Roughness Coefficient (JRC) ~20). In some of the sub-vertical fractures we observed dry clay. We did not observe clay on the major discontinuity that separates the upper and lower portions of the rockslope. Moreover, we did not observe seeps or springs on the slope; the slope was dry.

3.3. Critical Geologic Structures Simple examination of the rockslope from the shoulder of the roadway indicated major discontinuities dipping out of slope that could be kinematically unstable. Both historical planar and wedge block failures had obviously occurred previous to construction of the highway. Moreover, during production blasting the rock separated from the slope face and slid downslope naturally as planar and wedge failures.

One of the structures of additional concern was a fault clearly separating the upper unstable block from the main rock face (Figure 2).

Figure 3 is an equal-area stereonet with a Markland test plot for the rockslope developed using the computer code Rockpack®. Watts [8] has developed a computer code Rockpack® which applies the

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 toppling failures.

According to Watts [8], the kinematic analysis is a “cohesion-equals-zero” analysis in which the effects of cohesion are ignored. With cohesion equal to zero, the fundamental limiting equilibrium equation for computing the safety factor of a block with the potential to slide out of the slope face on one discontinuity (joint) reduces to the following:

θφ

TanTanFS = (2)

As a result, whenever the dip value of the potential failure plane (discontinuity) (θ) exceeds the friction

angle (φ), the factor of safety is less than 1 and kinematicly unstable.

The basic concept of kinematic analysis for plane failure is straightforward. Two conditions must be met for sliding to occur. First, the discontinuity must dip more steeply than its friction angle. 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.

Stereonet analysis for potential wedge failures is similar to that for plane failures. For a wedge failure to occur, the line made by the intersection of the planes creating the wedge must plunge more steeply than the friction angle and less steeply than the dip of the slope face and in a direction that daylights the slope face. To analyze for potential wedge failure, two great circles are constructed which bound the ranges of dip vector clusters. If the wedge formed by the intersection of the great circles plots within the critical zone, potential

Fig. 3: Dip vector plot equal – area stereonet with Markland analysis of rockslope.

Fig. 4: Pole plot equal – area stereonet displaying majorjoint sets. Arrow points in general direction (110º) of potential planar failure and plunge of potential wedge failures.

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 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).

Fault

After evaluation of the upper unstable block, it was 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 aspect of the highwall of the upper slope. The block rests on a potential failure plane that dips out of slope (daylights) at approximately 37 to 41 degrees. The dip of this discontinuity falls with the range of friction angles expected along an unfilled joint in granitic rock [4, 6, 8]. The toe of the block and failure plane was exposed during construction of the access road to the face. The critical block is about 14 meters (45 feet) high, and the face of the block dips to the southeast to the highway at approximately 83 degrees. The top of the block slopes back at about 35 degrees. The back of the block is bounded by a fault that strikes to the southeast (Refer to Figures 2 and 5).

Because the rock block on the high wall was kinematicly unstable, our next step was to investigate the global stability. We assumed the block had moved because the toe of the block had been un-weighted and the potential failure plane daylighted by excavation of the access road. During our investigation we observed no open tension fractures behind the block suggesting that it has not moved recently.

To evaluate the stability of the block, we assumed the following:

• The block is marginally stable (i.e. factor of safety (FS) = 1).

• Cohesion (C) is near zero along the failure plane.

• The peak friction angle (φ) for the discontinuity is about 37º (same as the dip plane).

• Currently, the block is free draining.

The authors estimated 37º for peak friction by conducting a series of tilt tests using rock blocks with joint faces similar to that of the rock mass. Of interest 37º is the same value as the dip of the failure plane. This suggests the factor of safety for the rock block is near 1 (Refer to equation 2). Because of the roughness of the fractures, we assumed 37º to be peak friction. Residual friction,

Fig. 5: Unstable block displaying potential planar failure plane. Construction of access road day lighted the fplane. Fault is on the left inside of the block.

ailure

37º

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

the field evaluation and back analysis the block was marginally stable. However, any critical outside forces such, as hydrostatic pressure from water along the failure plane or seismic activity will induce failure.

3.5. Stabilization of Upper Unstable Block Based on the results of the kinematic and stability analysis we concluded that the unstable block must be reinforced to increase the factor of safety to at least 1.3. We recommended to the contractor to install post-tensioned rock bolts or passive dowels on a patterned basis on the high wall face.

Figure 7 is a photo of the face of the hanging block with a general patterned layout of the rock dowels. The rock face measures about 12m (40 ft) high by 11m (35 ft) wide. The contractor elected to install 25 rock dowels on a patterned layout and spaced at approximately 2x2m (7x7 ft). Locations for the dowels were adjusted depending on the asperities of the rock face and previous. Each dowel was seated to about 111 kn (25 kips) using a torque wrench.

For our stability analysis, we assumed that the slope is free draining. However, since the rockslope is in an area that experiences extreme weather changes, we recommended the contractor install PVC drains to intersect the fault trace behind the block. This would help mitigate potential buildup of water

pressures behind the block and along the basal discontinuity.

4. EVALUATION OF OVERBREAK

4.1. General Wyllie and Mah [10] have shown that slope instability is often related to blast damage of the rockslope behind the face. The damage is often related to overbreak. Backbreak can originate from a poor blasting design exacerbated by unfavorable structural geology.

Overbreak is excessive rock breakage beyond the excavation limits [11, 12]. Overbreak includes backbreak, which is breakage behind the last row of shot holes and endbreak that occurs at the end of the shot line. One of the problems that lead to destabilization of the upper rockslope was at least 6 meters (20 feet) backbreak beyond the planned excavation line for the present rock cut (Figure 8). This pushed the access road into the toe of the upper slope day lighting the failure plane of the upper block.

4.2. Structural Geology Geology plays a very important role on the outcome of a blasting job on any type of rock excavation [13]. Typically poor blasting results occur in rock where the strike of the structures (faults and joints)

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.

are parallel or perpendicular to the free face of the rock slope. Blasting results are typically good if the strike of the structures is oblique to the free rock face. Strike parallel to the free face of the rockslope may cause severe backbreak and endbreak [13]. On this rockslope, several joint sets dip out of slope forming potential planar and wedge blocks (See Figures 2 through 5). Where the blast occurred with the dip of the slope, CDOT observed severe backbreak. Blasting with the dip of the rock structure apparently forced the explosive energy up and along the preexisting joint sets. This loosened and displaced the rock along the dip planes backbreaking beyond the design excavation line resulting in planar and wedge rock failures. This is very evident on Figure 2, where the dip of the structure runs uphill into the access road and highwall.

4.3. Blasting Design Evaluation Backbreak may result from several problems in the blasting design performance. Konya and Walter [11, 12] have developed a logical set of steps to follow when evaluating blasting designs. The following steps for evaluation of the blasting design were modified after Konya and Walter [11, 12]:

• Type of blasting (production or controlled)

• Drilling equipment capability and proposed borehole diameter.

• Explosive selection for specific site conditions.

• Specific gravity of rock and explosives.

• 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.

• Vibration and scale distance formula. Fig. 8: Severe backbreak caused from poor blasting design exacerbated by adverse structural geologic conditions. • Timing sequence between shot holes.

• Spacing between shot holes.

• Potential for violence to establish if mats are needed.

The authors carefully evaluated the design for each blast performance using these steps. As we went through each step, we would establish if the design was within design limits. We asked the following question. Did the design make sense and was it reasonable?

There are several design factors that may lead to backbreak. Konya and Walter [11, 12] describe some of the causes for backbreak:

• Excessive burden

• Excessively stiff benches

• Long stemming depths on stiff benches

• Improper timing delay

Burden (B) is the shortest distance of rock between the shot line and the nearest free face or open face at the time of detonation. Backbreak can result from excessive burden on the shot holes, thereby causing the explosive to break and crack radially further behind the last row of shot holes.

The stiffness ratio (L/B) is the ratio of bench height (L) to burden (B). Excessively stiff benches (L/B < 2) cause more uplift and backbreak near the collar of the shot hole. According to Konya and Walter [11, 12] for better blast results, stiffness ratios should be in the range of 3 to 4.

Stemming (T) is inert material such as drill cuttings packed around the collar of the blast hole to confine the gasses during detonation. Long stemming depths on stiff benches promotes backbreak.

Improper timing delay from row-to-row may cause backbreak if the timing is too short, thereby resulting in excessive confinement of the gases in the last row of the shot.

4.4. Results of blasting design evaluation Typically for rockslopes, a knowledgeable contractor will employ controlled blasting techniques such as line drilling, pre-shearing or cushion blasting for backbreak control of the back wall coupled with production blasting to break up the rest of the rock. On this project, the contractor used production blasting to bring the rock down. It was apparent that the structural geology was not considered in the design. Moreover, test blasting to evaluate the design was not done.

After review of design issues that lead to backbreak such as excessive burden, stiff benches, excessive stemming and short timing delays, the burden and stiffness ratio appeared within recommended design limits. However, two design issues appear to have worsened the backbreak along the wall; over stemming of the shot holes and short timing delay.

Konya and Walter [11, 12] recommend stemming of the shot holes to be about 0.7 of the burden. In this case the stemming material was twice as long as recommended. This led to backbreak of the last row of shot holes near the excavation limit.

In addition, the firing timing delay appeared to short between the last row and the second row of shot holes. Timing delay was about 2 milliseconds. Konya and Walter [11, 12] recommend a time delay between the back holes of 4 to 6 milliseconds. This will result in a scattered muck pile with minimum backbreak. An increase by 2 or 3 milliseconds between the second and last row of shot holes would have allowed the earlier shots to move the burden out of the way. This would have reduced the resistance on the last row of shot holes and reduced the pressure on the back wall. The end result would have been cleaner rock breaks with fewer overbreak problems.

5. SOLUTIONS TO MITIGATE POOR BLAST PERFORMANCE

5.1. Controlled Blasting

According to Wyllie and Mah [10] control of overbreak damage to the final walls can be limited by implementing a combination of the following measures. Production blasting should be designed to limit backbreak behind the final wall. Controlled blasting techniques such as line drilling, pre-shearing, and cushion blasting will limit the fracture creep up the dip planes during detonation by defining a final rock face.

5.2. Test Shots Before blasting, a series of test shots evaluating the performance of the controlled and production blasting should be conducted to evaluate how the rock and geologic structure responds to the blast design. This is critical to a good design and blasting performance.

6. CONCLUSIONS

From the investigation, the authors concluded the following:

• The upper rockslope was kinematicly unstable with major blocks dipping out of slope that required reinforcement with post-tensioned rock bolts or dowels.

• Construction of the contractor’s 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 high wall rendering the slope kinematicly unstable.

• Back analysis of the stability of the high wall using the limit of equilibrium method indicated a factor of safety of 1 or the slope is marginally stable.

• To achieve a factor of safety of 1.3 or greater the high wall required reinforcement with post-tensioned rock bolts or dowels.

• A combination of factors in the blasting design such as over-stemming of the shot holes and short timing delays in the firing sequence between the second and back shot hole rows led to the severe backbreak.

• The adverse geologic structure dipping out of slope exacerbated the excessive backbreak.

• Controlled blasting techniques such as line drilling, cushion or pre-shearing 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.

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the Silverton and Howardsville Quadrangles, Southwestern Colorado, United States Geological Survey.

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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.

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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.

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13. International Society of Explosive Engineers (ISEE). 1998. Blasters Handbook, 17th ed., ed R.B. Hopler, International Society of Explosive Engineers, 742.