Geotechnical Report.pdf

FINAL

Geological Engineering Report Rock Slope Investigation and Remediation Design George Washington Memorial Parkway Arlington County, Virginia

FHWA Project No. GWMP 4(1), 5(1)

Schnabel Project No. 05140065.02

May 11, 2007

"We are committed to serving our clients by exceeding their expectations."

Geotechnical • Construction Monitoring • Dam Engineering • Geoscience • Environmental

Mr. Andre Anderson Eastern Federal Lands, Highway Division Federal Highway Administration 21400 Ridgetop Circle Sterling, Virginia 20166

Subject: FINAL Geological Engineering Report

Rock Slope Investigation and Remediation George Washington Memorial Parkway Arlington County, Virginia Schnabel Project No. 05140065.02

EFLHD Reference No. HFTS-15

Dear Mr. Anderson:

Schnabel Engineering North, LLC (Schnabel) is pleased to submit this geological engineering report for the above referenced project. This report has been prepared in accordance with Federal

Highway Administration, Eastern Federal Lands Highway Division (EFLHD) Contract No. DTFH71-

06-D-00001 (dated November 15, 2005), and the Notice to Proceed for Task Order No. 07-001 (dated

January 4, 2007).

Scope of Services

Our services are provided in accordance with the EFLHD Notice to Proceed and Scope of

Work, dated November 6, 2006. Our understanding of the work is described herein. The scope of our services includes:

1. Review background information, topographic site plans and cross sections provided to us, as well as available geologic maps;

2. Field geologic structure mapping;

3. Kinematic and limit equilibrium rock slope stability, and rockfall hazard analyses;

656 Quince Orchard Road Ste 700 Gaithersburg, MD 20878

Phone (301) 417-2400 Fax (301) 417-2730 www.schnabel-eng.com

Mr. Anderson, EFLHD

Project No. 05140065.02 2 Schnabel Engineering North, LLC

4. Develop two slope stabilization design concept options (including rock gluing in conjunction with rock anchors as one of the options);

5. Develop comparative cost estimates for the two proposed design concepts; and

6. Prepare a draft geological engineering report dated February 21, 1007.

7. Schnabel personnel met with the EFL and NPS personnel to discuss our proposed slope remediation design concepts and comparative cost estimates on March 14, 2007.

Once an option is selected, we will develop a final geological engineering report and

PS&E for the chosen concept design option.

We appreciate the opportunity to be of service to you for this project. Please contact Mr.

Brian Banks at (301) 417-2400 or Mr. George Aristorenas at (703) 779-0773 if you have any questions regarding this report.

Very truly yours, SCHNABEL ENGINEERING NORTH, LLC

Prepared By:

Brian Banks, C.P.G.

Associate

Eric Rehwoldt, P.E.

Principal

Reviewed By:

Ken Derrenbacher, P.E.

Principal

BB/GA/PP/ER/KD/

G:\05140065.02\WP\DRAFT\05140065.02 Final Geologic Report 07_05-11.doc tkerrick Brian Banks tkerrick Eric Rehwoldt tkerrick Ken Darrenbacher tkerrick ER VA Stamp

Project No. 05140065.02/May 11, 2007 i Schnabel Engineering North, LLC

Table of Contents

1.0 Executive Summary

2.0 Introduction

2.1 Background

2.2 Site Description

3.0 Geologic Descriptions

3.1 Regional Geology

3.2 Field Measurements

3.3 Findings

3.3.1 Rock Description

3.3.2 Major Discontinuities

3.3.3 Groundwater Conditions

3.3.4 Typical Rockfall Block Sizes

4.0 Assessment of Geologic Hazards

4.1 Kinematic Analysis

4.2 Limit Equilibrium Analyses

4.2.1 Planar Sliding

4.2.2 Wedge Sliding

4.2.3 Toppling

4.3 Rockfall Hazard Analysis

4.4 Assessment of the 2002 Rockfall Event

4.5 Failure Modes for Stabilization Design

5.0 Recommendations

5.1 Rock Gluing

5.1.1 Scaling

5.1.2 Spot Bolting

5.1.3 Rock Gluing

5.1.4 Drains

5.2 Rock Anchoring

5.2.1 Scaling

5.2.2 Spot Bolting

5.2.3 Rock Anchoring

5.2.4 Drains

6.0 Cost Estimates

6.1 Rock Gluing

6.2 Rock Anchoring

7.0 Limitations

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List of Tables

Table 1: Summary of Limit Equilibrium Analysis Results Table 2: Cost Estimate for Rock Gluing Table 3: Cost Estimate for Rock Bolting Table B-1: Discontinuity Survey Data Summary Table Table C-1: Limit Equilibrium Analysis Results – Planar Sliding Table C-2: Limit Equilibrium Analysis Results – Wedge Sliding Table C-3: Limit Equilibrium Analysis Results – Toppling Table C-4: Rock Gluing Evaluation Summary Table – Planar Sliding Table C-5: Rock Gluing Evaluation Summary Table – Toppling

Appendices Appendix A: Figures

Figure 1: Site Vicinity Map Figure 2: Topographic Site Plan Figure 3: Typical Cross Section Figure 4: Geologic Map Figure 5: Discontinuity Contour Plots Figure 6: Kinematic Analysis Figure 7: Area of Rock Gluing Stabilization Figure 8: Rock Gluing Application – Poudre Canyon, Colorado Figure 9: Rock Gluing Typical Section Figure 10: Rock Anchoring Typical Section

Appendix B: Field Data Table B-1: Discontinuity Survey Data Summary Table (1 Sheet) Discontinuity Survey Data Sheet (5 Sheets) Descriptive Criteria for Rock Core Logging (2 Sheets) Appendix C: Calculations

Interface Friction Angle Calculations (6 Sheets) Limit Equilibrium Analysis Results – Planar Sliding (13 Sheets) Limit Equilibrium Analysis Results – Wedge Sliding (7 Sheets) Limit Equilibrium Analysis Results – Toppling (3 Sheets) CRSP Analysis Results (4 Sheets) Rock Gluing Evaluation (2 Sheets)

Appendix D: Photographs (9 Sheets) Appendix E: Selected References for Rock Gluing

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1.0 Executive Summary

Based on our evaluation of data from the field investigation and the project data furnished to us, we have developed the following conclusions and recommendations. This Executive

Summary has been prepared for the convenience of the users of this report and does not contain all of the information presented in this report. Therefore, the entire report should be read to assure transmission of all pertinent information. Detailed recommendations are presented in the body of this report.

1. Six major discontinuity sets (Planes 1 through 6) were identified during the geologic structure field mapping and considered in our analyses.

2. Based on our site observations analyses, and our understanding of the 2002 rockfall event, we believe there is a significant risk of a similar rockfall event in the future. Therefore, we recommend that rock slope stabilization measures be performed.

3. Based on our kinematic and limit equilibrium analyses, and consideration of the likely cause of the 2002 rockfall event, we conclude that the rock slope stabilization design must consider the following failure modes:

a. Global planar sliding along Plane 3 discontinuities.

b. Local planar sliding along Plane 3 discontinuities.

c. Local toppling failure along Plane 2 discontinuities.

4. Although the recommended remediation measures are expected to address large blocks in excess of 10 ft, the potential for localized rockfall involving small blocks will not be eliminated after the proposed slope remediation is completed.

Our rockfall hazard analysis, described in Section 4, indicates that, in at least one condition analyzed, some of these smaller rockfall blocks may reach the highway.

To mitigate this potential, we recommend a rockfall barrier (e.g., a PVC-coated

Project No. 05140065.02/May 11, 2007 2 Schnabel Engineering North, LLC wire mesh fence, permanent Jersey barrier, or fixed wall) be constructed at the edge of pavement to control small rockfall blocks. The rockfall barrier should be at least 4 ft high and capable of withstanding an impact up to 22 ft-tons. We suggest maintenance visits at least every two years and after major rain storm events to remove rockfall blocks from the road shoulder between the base of the slope and the rockfall barrier.

5. We developed recommendations for two design concepts to reduce the potential for global- and local-scale planar sliding and local-scale toppling failure mechanisms indicated in our limit equilibrium analyses, including rock gluing and permanent rock anchors. The total area to be reinforced is approximately 2,500 ft 2.

6. For the rock gluing option, we recommend injecting PUR grout to stabilize the rock cut slope. We recommend that the glue injection holes, typically 1 ½” in diameter and inclined 5 o into the rock slope, be spaced 10 ft apart on center. We also recommend that the injection holes are drilled 15 ft deep behind the rock face. Our analyses indicate this pattern can adequately stabilize both 5 ft and 10 ft deep planar sliding and toppling rock blocks.

7. For the permanent rock anchor option, we recommend tensioned rock anchors to stabilize the rock cut slope. We recommend 1-in diameter, grade 75 bars placed within a 4-in diameter bore hole and grouted. The bars should be protected against corrosion (e.g., epoxy coated and surrounded by cement grout). We recommend minimum 20 ft long anchors with minimum 10 ft long bond zone and spaced 10 ft apart on center (vertical and horizontal). We recommend that the anchors be tensioned to achieve a capacity of 50 kips. The anchors should be inclined a maximum of 15 o from horizontal. Our analyses indicate this rock anchor pattern and anchor capacity should adequately stabilize both planar sliding and toppling rock blocks.

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8. For both design concept options, we recommend that scaling is performed as the first step to prepare the rock face for further reinforcement, and as a safety precaution to protect workers by reducing the potential for localized rockfall during reinforcement construction activities.

9. In addition to scaling for both design concept options, some larger rock blocks may need to be individually bolted (“spot bolted”) to the slope face using tensioned rock anchors or passive rock dowels if it is concluded that further attempts at scaling may destabilize the slope.

10. For both design concept options, near-horizontal drains should be installed. We recommend the lowest row of drains to be installed at about 6 ft above the ground surface at the base of the slope, and spaced every 20 ft horizontally. Another row of drains should be installed about 10 ft above the lowest row of drains with a 20 ft horizontal spacing and staggered above the lower drains. The drains installed within the rock slope should penetrate at least 10 ft beyond the “Potential Planar

Sliding Surface” as shown on Figures 9 and 10. We expect the drains will vary from 25 to 50 ft long. The drains should be inclined at about 5 o upwards from horizontal at the excavation face. The drain holes should be no smaller than 3 in diameter and sleeved with slotted PVC casing. Maintenance visits should be made at least every 2 years to check for drain pipe clogging (e.g., with mineral or biological deposits). For the rock gluing option, drains should be installed after rock gluing. For the rock anchoring option, we recommend the drains should be drilled between every other row of rock anchors. For the rock anchoring option, the sequence of the drain and anchor installation (i.e., drains before or after anchors) should be determined by the contractor.

11. The estimated cost of the rock gluing and rock anchoring design concepts is

$233,250.00 and $186,050.00, respectively. These estimates do not include costs associated with a rockfall protection (fence, wall) at the slope, traffic control, construction administration, field observations by engineering personnel, other engineering consultation services during construction, aesthetic “touch-up” of

Project No. 05140065.02/May 11, 2007 4 Schnabel Engineering North, LLC rock anchor heads, and periodic maintenance to remove small rockfall blocks at the base of the slope.

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2.0 Introduction

2.1 Background

On December 13, 2002 a rockfall event occurred in a road cut slope along the George

Washington Memorial Parkway (GWMP). The rockfall event involved several large rock blocks that damaged curb and pavement sections of the adjacent road. Cleanup activities included limited scaling of the rock slope with a backhoe to remove the most critically unstable blocks.

However, there are concerns that potentially unstable rock blocks remain as a rockfall hazard.

Therefore, Schnabel Engineering North, LLC (Schnabel) has been asked to provide engineering services in support of FHWA-EFLD efforts to reduce the rockfall hazard potential of the slope.

The purpose of the project is to characterize the rock slope instability hazard and to develop technical plans, specifications and an engineering cost estimate (PS&E) for remediation of the unstable rock slope. The scope of the work for this project is defined in Task Order No.

07-001, dated January 4, 2007. The project includes four phases of work:

1. Phase I – Geologic Mapping and Field Investigation. The scope for this phase includes reviewing the topographic site plan and cross sections provided by

FHWA-EFLD, available geologic maps, and performing a field evaluation including geologic rock structure mapping to gather data for subsequent geo-engineering analysis.

2. Phase II – Draft Geologic Report and Preliminary Cost Estimates. The scope for this phase includes performing geological engineering analyses (i.e., kinematic and limit equilibrium rock slope stability, and rockfall hazard analyses) to evaluate and characterize the rock slope hazards, and to develop two remediation design concepts with preliminary cost estimates. This Draft Geological

Engineering Report presents the results of our geological engineering assessment, remediation design concepts, and preliminary cost estimates.

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3. Phase III – Draft Report Review Meeting. Schnabel personnel will attend a meeting to discuss our Draft Geological Engineering Report with the COTR, PM, and NPS personnel. We understand this meeting will occur after the interested parties have the opportunity to review and evaluate this report.

4. Phase IV – Final Geological Engineering Report. The scope for this phase includes performing additional geological engineering analyses as necessary to develop the Final Geological Engineering Report including the PS&E for the concept design option chosen.

It should be noted that core drilling, piezometer installation, and laboratory testing of rock strength characteristics was not included in our scope of services for this work. Rock mass properties were based on observations of the exposed rock cut slope face. Estimates of rock strength referenced herein are based on field index testing and our experience with similar materials in the Washington, D.C. area.

2.2 Site Description

The subject rock slope is an approximately 240 ft long road cut located along the right side of the northbound lane of the GWMP between Spout Run Parkway and the Key Bridge, Arlington County, Virginia (Appendix A - Figure 1). The northbound travel lane is lined with a raised concrete curb that marks the edge of pavement. The toe of the rock slope is about 12 ft from the curb. A concrete jersey barrier extends parallel to the road approximately 2 ft to 5 ft from the curb. The Jersey barrier was placed after the 2002 rockfall event to impede future rockfall blocks from reaching the travel lane. The area between the curb and the toe of the cut slope is open (except for the Jersey barrier) and is lightly vegetated with grasses, weeds, and vines.

The Topographic Site Plan (Appendix A - Figure 2) and Typical Cross-Section

(Appendix A - Figure 3) included in Appendix A illustrate the site topography. The elevation of the northbound travel lane at the project site is at about El 20, approximately 20 ft above the level of the Potomac River, located to the north of the site. The area between the curb and the toe of the cut slope generally slopes down about 3 ft from the curb to the toe of the slope at El 17. In

Project No. 05140065.02/May 11, 2007 7 Schnabel Engineering North, LLC some locations, however, the elevation at the toe is elevated where rock outcrop extends from the base of cut slope.

The cut slope face is generally a near vertical rock exposure up to about 35 ft high (El

52). Some rock ledges occur along the slope and are covered with soil, moss and weeds in some locations. The natural ground above the rock cut slope continues to slope up southward at about

32 degrees. The natural ground above the cut slope is vegetated with small trees, bushes, grasses, weeds, and vines, and contains several rock exposures, some of which may be loose.

The photographs included in Appendix D show site conditions after the 2002 rockfall event and during our field investigation on January 17, 2007.

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3.0 Geologic Descriptions

3.1 Regional Geology

The boundary between two geologic provinces - the central Appalachian Piedmont to the west and the Atlantic Coastal Plain to the east – occurs in the vicinity of the site. This boundary, commonly referred to as the “Fall Line” or “Fall Zone”, marks the western-most extent of where relatively young (lower Cretaceous- to Quaternary-aged) and poorly consolidated sedimentary formations overlie older metamorphic and intrusive Piedmont rocks. The depth of Coastal Plain sediments generally increases eastward away from the Fall Zone. In the Fall Zone, Piedmont rocks are exposed in the topographic lows (e.g., stream valleys), whereas Coastal Plain sediments are exposed in many topographic peaks and plateaus.

Based on the Geologic Map of the Washington West Quadrangle (Fleming et al., 1994;

Figure 4), the project site is underlain by Piedmont rocks of the lower Cambrian-aged Sykesville

Formation. The description on the map characterizes the Sykesville Formation as “light- to medium- gray sedimentary mélange with a quartzo-feldspathic matrix containing quartz augens and fragments of phyllonite; metagraywacke; migmatite; serpentinite; amphibolite; and actinolite schist; and mappable olistoliths of amphibolite; metagabbro; metagraywacke; phyllonite;

actinolite schist; and soapstone.” The geologic structure of the Sykesville Formation is characterized on a local scale by foliation and joints. Foliation planes mapped in the vicinity of the site indicate foliation dips from about 45 to 90 degrees with dip directions (azimuth) ranging from about 255 to 277 degrees (Fleming et al., Figure 4).

3.2 Field Measurements

Mr. Brian Banks, C.P.G. and Ms. Pamela Patrick, E.I.T. performed geologic rock structure mapping at the site on January 17, 2007 to collect field data in support of our geological engineering analyses and remediation design development. Measurements were made of typical rockfall block size and shape. In addition, other data was collected to characterize the rock mass including:

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1. Rock description information including rock type, strength, weathering, fracture spacing, color, and foliation character. The terms used to describe the rock mass are defined in the Descriptive Criteria for Rock Core Logging included in

Appendix B.

2. Data characterizing individual discontinuities, including:

a. Location (i.e., station),

b. Discontinuity type,

c. Orientation (i.e., dip and dip direction),

d. Estimate of the Joint Wall Compressive Strength (JCS) based on hardness index testing,

e. Joint Roughness Coefficient (JRC, following Barton and Choubey, 1977),

f. Infilling type and coverage,

g. Water condition, and

h. Persistence and termination.

3.3 Findings

3.3.1 Rock Description

The rock exposed in the cut slope is described as quartzo-feldspathic schistose gneiss:

strong, moderately weathered, moderately to very slightly fractured, brownish gray, and moderately to strongly foliated. The terms used to describe the rock mass are defined in the

Descriptive Criteria for Rock Core Logging included in Appendix B. This rock description is consistent with the description of the Sykesville Formation from the referenced geologic map.

3.3.2 Major Discontinuities

A total of 66 individual discontinuities were inspected during our rock structure survey.

Appendix B includes a table summarizing the discontinuity data gathered for use in our kinematic and limit equilibrium analyses and the Discontinuity Survey Data Sheet used to collect rock structure data in the field.

The orientations of these discontinuities were plotted and contoured using the computer program DIPS (produced by Rocscience Inc.). The DIPS stereonet contour plot attached as

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Figure 5 shows the discontinuity orientation measurements. The contours represent statistical concentrations of discontinuity orientation vectors, and help to visualize the clustering of orientation data. Each data cluster represents a discontinuity set, such as foliation and joint sets.

The orientations of the planes associated with the major contour peaks are considered to represent the average orientation of each discontinuity set. The contours also indicate the variability in the orientation of each set. We identified the following major discontinuity sets to consider in our analyses:

Plane 1: Joint set with an average orientation of 80°/200° (dip/dip direction). Plane

1 joints tend to be very persistent. Some Plane 1 joints were traceable for over 100 ft along strike. These joints form the majority of the rock slope face. A lineation feature observed on Plane 1 joint surfaces is defined by the intersection of Plane 1 joints with Plane 3 joints. The JCS is estimated to be 4,000 psi. The average JRC is about 7. Iron staining was observed on the majority of the Plane 1 joint surfaces.

Plane 2: Joint set with an average orientation of 62°/174° (dip/dip direction). Plane

2 joints do not tend to be very persistent. Plane 2 joints are typically traceable for less than 10 ft. The JCS of these joints is estimated to be

4,000 psi. The average JRC is about 7. Iron staining was observed on many of these joint surfaces.

Plane 3: Joint set with an average orientation of 28°/003° (dip/dip direction). Plane

3 joints are relatively persistent as they are traceable for up to about 30 ft along strike. These joints were identified as possibly a secondary foliation feature. The intersection of Plane 3 joints with the Plane 1 joints are identified by a lineation observed on Plane 1 joint surfaces. The JCS is estimated to be 4,000 psi. The average JRC is about 11. Plane 3 joints are particularly rough along strike. Some iron staining was observed on a few of the Plane 3 joint surfaces. Plane 3 joint surfaces are adversely oriented with respect to slope stability. Planar sliding along Plane 3 joint surfaces appear to have resulted in several localized rockfall events. It is believed

Project No. 05140065.02/May 11, 2007 11 Schnabel Engineering North, LLC that sliding along a Plane 3 joint surface contributed to the December 13, 2002 rockfall since it appears that a Plane 3 joint surface defines the base of the rockfall source area.

Plane 4: Joint set with an average orientation of 70°/063° (dip/dip direction). Plane

4 joints do not tend to be very persistent as they are typically traceable for less than 10 ft. Plane 4 joints intersect with Plane 6 foliation to form an adversely oriented wedge sliding surface. Wedge sliding along Planes 4 and 6 appear to have resulted in several localized rockfall events. The JCS is estimated to be 4,000 psi. The average JRC is about 11. Iron staining was observed on some of the Plane 3 joint surfaces.

Plane 5: Joint set with an average orientation of 33°/153° (dip/dip direction). Plane

5 joints tend to be very persistent and traceable for up to 60 ft. They have an average spacing of between about 15 ft and 20 ft. The JCS is estimated to be 1,000 psi, on average, with an average JRC of about 6. Plane 5 joints were observed to exhibit iron staining.

Plane 6: Foliation with an average orientation of 74°/280° (dip/dip direction). Plane

6 discontinuities are separations along foliation planes, and tend to be very persistent and traceable for up to 50 feet. Plane 6 foliation separations intersect with Plane 4 foliation to form adversely oriented wedge sliding surfaces. Wedge sliding along Planes 4 and 6 appear to have resulted in several localized rockfall events. The JCS is estimated to be 1,000 psi, on average, with an average JRC of about 10. Plane 6 foliation separations were observed to exhibit iron staining.

3.3.3 Groundwater Conditions

Water seepage was observed in localized areas along the slope. Seepage was observed from discontinuities within about 5 ft to 10 ft of the base of the slope predominately between approximate stations 20+50 and 21+00. Seepage from discontinuities located higher in the cut slope was observed between approximate stations 20+70 and 21+30. The photographs in

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Appendix D show seepage locations as vertical dark streaks extending downward from discontinuities.

Without reliable groundwater level data from piezometers installed behind the slope face, it is unclear whether the seepage represents stabilized groundwater conditions within the rock mass or if it reflects localized surface drainage pathways. We have based our analyses on the possibility of high groundwater levels behind the slope face as a worst-case condition.

3.3.4 Typical Rockfall Block Sizes

Typical rockfall block sizes were measured from rock blocks lying on the ground at the base of the existing rock slope at approximate stations 19+00 (Photo No. 17) and 21+40. These rocks are generally rectangular to irregularly shaped, and are typically up to approximately 3 ft in diameter. However, we acknowledge that rock blocks up to about 10 ft are possible based on the rock blocks that were shed from the slope during the 2002 rockfall event.

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4.0 Assessment of Geologic Hazards

4.1 Kinematic Analysis

Kinematic stability analysis was performed to evaluate potential rock slope failure modes controlled by planar rock mass discontinuities based exclusively on the geometric relationships of the 66 discontinuities observed and measured at the site (Figure 5). Potential rock slope failure modes include sliding of rock blocks along a single planar discontinuity, sliding of wedges formed in the slope by intersecting discontinuity planes, and toppling of rock blocks.

The computer program RockPack III for Windows (produced by RockWare Inc.) was used for the kinematic stability analyses. Inputs to the program for the analysis include:

1. Average orientation (dip magnitude and dip direction) of measured discontinuity sets within the rock mass, including bedding and joint discontinuities;

2. Cut slope orientation (slope magnitude and direction); and

3. Estimated rock mass discontinuity interface friction angle. Interface friction angle calculations are included in Appendix C for each of the six major discontinuity sets. The interface friction angle calculations are based on the

Barton (1976) method. For the kinematic analysis, an interface friction angle of

33 degrees was selected to check for sliding potential since it was the lowest friction angle calculated for each of the six discontinuity sets. Note that higher interface friction angles are used for specific discontinuities in our limit equilibrium analysis based on the interface friction angle calculations provided in

Appendix C.

A kinematic stability analysis was performed for the typical rock cut slope configuration presented in Figure 3. A RockPack stereonet plot was constructed for the typical rock cut slope configuration to illustrate the results of our kinematic analysis (Figure 6).

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The results of the kinematic stability analysis are summarized as follows:

1. Potential wedge sliding failures are indicated based on the average orientations of

Plane 4 and Plane 6 discontinuities.

2. Although the dip direction of the Plane 3 discontinuity set is adversely oriented with respect to potential planar sliding (i.e., the plane dips towards the slope face), potential planar sliding is not indicated based on the average dip of the Plane 3 discontinuity set since the dip of the plane is less than the interface friction angle assumed in our analysis. However, we recognize that groundwater pressures acting on Plane 3 discontinuities may trigger planar sliding.

3. A potential for toppling failure is indicated. Based on the kinematic analysis, toppling may occur along Planes 1 and 2. Toppling is considered to be a local, and not a global stability hazard.

4.2 Limit Equilibrium Analyses

Where potential failure modes were identified in the kinematic analysis, more detailed limit equilibrium analyses were performed to calculate a factor of safety (FS) against sliding based on slope geometry, measured discontinuity orientations, estimated groundwater conditions, calculated discontinuity interface friction angles, and estimated or measured material properties of the rock mass. Limit equilibrium analyses were performed for potential planar sliding identified along Plane 3, and for potential wedge failure formed by Plane 4 and Plane 6.

The limit equilibrium analyses were performed using RockPack III. The RockPack III program calculates the FS for a potential planar or wedge slide geometry as a function of driving and resisting forces. The RockPack III program also calculates the sum of moments for a given toppling scenario. The output for the toppling evaluation provides the normal force required of a rock anchor to hold the toppling block in place.

The program was used to evaluate existing conditions (i.e., without reinforcement), as well as the force required to achieve a FS of at least 1.5. A FS of 1.5 is typically considered acceptable for permanent slopes (FHWA Publication No. TS-89-045, 1989). Therefore, for cases where the FS was found to be less than 1.5, we also calculated the horizontal force required to

Project No. 05140065.02/May 11, 2007 15 Schnabel Engineering North, LLC obtain a FS of 1.5. In these cases, the horizontal force required to obtain a FS of 1.5 was used to evaluate potential reinforcement mechanisms including rock gluing and rock anchoring.

Our limit equilibrium analyses were based on the following considerations and input parameters:

1. Slope orientation. We considered a vertical slope face striking at 010 degrees

(azimuth).

2. Slope height. We considered slope heights of 35 ft and 10 ft for the global and local slope failure conditions, respectively. For the global case, we assumed a failure surface (i.e., plane or wedge intersection) extending through the toe of the slope as a worse-case scenerio (Appendix A - Figure 3). We assumed a 10-ft high rock block for the local case based on the size of the 2002 rockfall block.

3. Discontinuity orientations indicated by the kinematic analysis to be potential failure surfaces. Plane 3, oriented at 28o/003o (dip/dip direction) was considered for the planar sliding condition. Planes 4 and 6, oriented at 70o/063o and 76o/280o, respectively, were considered for the wedge sliding condition. Plane 2, oriented at

62o/174o, was considered as the worst-case toppling failure condition.

4. Discontinuity interface friction angles of 47o, 63o, and 49o were used for Planes 3, 4 and 6, respectively, based on the calculations provided in Appendix C. We assumed zero cohesion in our FS analyses.

5. A tension crack was assumed to occur 50 ft from the slope crest for analysis of global planar failure (based on geometrical considerations), and 5 ft and 10 ft from the slope crest during local planar failure. For the local wedge failure analyses, tension cracks were assumed to occur 5 ft and 9 ft from the slope crest.

A tension crack was not assumed for the global wedge failure condition.

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6. Both “dry” and “wet” conditions were considered to account for potential groundwater uplift pressures acting on sliding and toppling blocks. For the global failure analyses, we considered that the sliding surfaces and tension crack could be 0% and 100% filled with water. For the local failure analyses, we considered tension cracks could be 0% and 100% filled with water.

7. Rock Mass unit weight of 160 pcf.

8. Earthquake loading was not included in the rock slope stability analysis.

9. No other internal or external surcharges are considered in the analysis.

The limit equilibrium analysis summary tables (Tables C-1, C-2, and C-3) and the

RockPack III output summary sheets included in Appendix C show the input parameters and results of our FS calculations. Table 1 provides a summary of the limit equilibrium analysis results.

Table 1: Summary of Limit Equilibrium Analysis Results

Failure Mechanism

Controlling Discontinuity

Plane(s) Block

Height (ft)

Tension Crack Location -

Distance from Slope Crest

(ft)

Percent of Tension Crack

Filled with Water

Factor of Safety

Horizontal Force (lb)

Required to Achieve

FS ≥ 1.5

0 2.02 0 57 1.50 0 Global Planar

Sliding Plane 3 35 50 100 1.00 51,000

0 2.02 0 43 1.50 0 Plane 3 10 5

100 0.59 3,500 0 2.02 0 53 1.50 0

Local Planar Sliding

Plane 3 10 10 100 0.88 4,000

35 None 0 3.72 0 Global Wedge Sliding

Planes 4 and 6 35 None 100 2.58 0

10 5 0 3.72 0 Planes 4 and 6 10 5 100 2.09 0

10 9 0 3.72 0 Local Wedge

Sliding Planes 4 and 6 10 9 100 3.54 0

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4.2.1 Planar Sliding

Limit equilibrium analysis indicates a FS greater than 1.5 against potential global (i.e., 35 ft high failure blocks) and local (i.e., 10 ft high failure blocks) planar sliding along Plane 3 discontinuities assuming a dry condition for the failure surface. However, the FS against planar sliding is less than 1.5 for both the global and local cases if the tension crack is considered to be partially filled with water. This can occur when surface run-off during a heavy rainstorm fills up the crack before it drains, or if the tension crack is filled due to groundwater conditions.

Our analysis indicates the global-scale planar sliding mechanism becomes critical when a tension crack located 50 ft behind the face is 57% filled with water. Similarly, the local-scale planar sliding mechanism becomes critical when groundwater uplift pressures developed along the base of the sliding block contribute to the driving forces. Therefore, our remediation design concepts consider reduction in groundwater uplift pressure for the global- and local-scale planar sliding condition.

It is noted that we do not have reliable groundwater data upon which to assess the static groundwater conditions within the rock mass, or the groundwater conditions within the rock mass at the time of the previous rock slope failures. We interpret that the seepage observed from discontinuities near the top of the slope reflects local conditions; however, the amount of surface water inflow into the rock mass discontinuities during and shortly after an intense rainstorm, or sustained melt water runoff event, could contribute adverse groundwater uplift pressures within the rock mass that could lead to localized or global rock slope instability.

4.2.2 Wedge Sliding

Our analysis indicates a FS greater than 1.5 against potential wedge sliding along Planes

4 and 6 discontinuities for global- and local-scale failures assuming both dry and wet conditions.

4.2.3 Toppling

Limit equilibrium analysis indicates that rock blocks bounded by Plane 2 discontinuities have the potential for toppling failure. Toppling failure is considered a local, and not a global slope stability hazard. The tension force on a rock anchor installed in the middle of the block required to stabilize (i.e., FS = 1.5) a 10 ft tall block is 336 lbs and 5,018 lbs for dry and wet conditions, respectively.

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4.3 Rockfall Hazard Analysis

Even after the rock cut slope is remediated to increase the global stability and reduce the local-scale rockfall hazard, small rock blocks can still be shed from the slope over time. These small rocks may become loose due to weathering and freeze-thaw effects. Rock slope stabilization schemes cannot completely eliminate local rockfall occurrences. Therefore, we used the Colorado Rockfall Simulation Program (CRSP), Version 4.0 (2000) to evaluate the rockfall hazard of the rock cut slope. Our rockfall hazard analysis was based on the following considerations and input parameters:

1. Existing slope geometry and shoulder configuration. We considered two configurations, including:

a. 35 ft high vertical rock face with a 12-ft wide shoulder. This configuration represents the typical cut slope section.

b. 35 ft high rock face (including a 30 ft high vertical upper portion and a 5 ft high lower portion sloping at a 45 degree angle toward the highway) with a 7-ft wide shoulder. This configuration represents the western portion of the cut slope face where rock outcrop extends away from the base of the slope.

2. Estimated dimensions of rock debris likely to be shed from the rock cut slope due to the effects of weathering over time. We considered spherical rocks up to 5 ft in diameter as a typical shape and size of potential rockfall blocks after the slope is remediated as per the proposed slope stabilization concepts introduced herein.

This size is based upon our observations of rockfall blocks at the base of the rock cut slope, and based on our belief that the proposed stabilization concepts will significantly reduce the likelihood of rockfall blocks greater than 5 ft.

3. Rock unit weight. We estimated the unit weight of the falling rock to be 170 pcf, equal to that of the intact rock unit weight.

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4. Surface roughness along the slope profile. Slope roughness is estimated based on the assumed size and shape of intact rock material protruding from the slope face.

We considered a surface roughness of 1.0 for our slope, which indicates a relatively smooth surface.

5. Tangential coefficient, which is a measure of the resistance the rock will encounter parallel to the slope due to vegetation, soil type, or the presence of a hard rock surface. We used 0.85 as the tangential coefficient for the rock slope, which is in the middle of the range of values considered to be acceptable for hard rock. We used 0.65 for the tangential coefficient for the shoulder, which is considered acceptable for soft soil slopes.

6. Normal coefficient, which is a measure of the elasticity or hardness of the slope, perpendicular to the slope. We used 0.225 as the normal coefficient for the rock slope, which is in the middle of the range of values considered to be acceptable for hard rock. We used 0.015 as the normal coefficient for the shoulder, which is considered acceptable for soft soil slopes.

7. Location on the slope from which to begin rolling rocks. We started rolling rocks at the top of the rock cut slope. We did not evaluate rocks that may be shed from the natural slope above the rock cut slope.

8. Location of analysis point on the slope for which to analyze the rock’s trajectory and energy. We used the edge of pavement as the analysis point.

The CRSP analysis consisted of “rolling” 100 rocks from the top of each rock cut slope profile considered to see how close to the GWMP travel lane the rocks landed. The CRSP results are included in Appendix C. For the first rock cut slope profile (35 ft high vertical rock face), no rocks reached the analysis point at the edge of the travel lane because the energy of a rock falling from a vertical face is directed mostly downward such that there is not enough of a lateral component to its trajectory to carry it into the road. For the second rock cut slope profile

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(30 ft high vertical face above a 45 degree lower slope), 66 out of 100 rocks reached the travel lane because the rock falling from the vertical face was directed toward the highway by bouncing off of the 45 degree lower slope.

The CRSP model calculates the bounce height and the kinetic energy of the rocks as they cross the analysis point. The maximum bounce and maximum kinetic energy of the rocks that crossed the analysis point are 0.38 ft and 21.9 ft-lbs, respectively. We recommend a rockfall barrier (e.g., a PVC-coated wire mesh fence, permanent Jersey barrier, or fixed wall) at the edge of pavement to control small rockfall blocks. The rockfall barrier should be at least 4 ft high and capable of withstanding an impact up to 22 foot-tons. We suggest maintenance visits at least every two years and after major rain storm events to remove rockfall blocks from the road shoulder between the base of the slope and the rockfall barrier.

It is also possible to control small rockfall blocks by constructing a rockfall containment ditch at the base of the rock cut slope. Construction of a rockfall containment ditch would involve reconfiguration of the base of the cut slope to make the ground surface between the cut slope and the highway flat or sloped gently inward toward the cut slope. Rock outcrop extends from the base of the cut slope in some locations such that blasting would be required to reconfigure the slope for construction of a rockfall containment ditch. Even with the containment ditch, rocks still may shatter and spray rock pieces onto the highway. We recommend a rockfall barrier is used to control small rockfall blocks in lieu of a rockfall containment ditch because of the difficulty involved in removing the rock outcrop at the base of the cut slope, and because the barrier would reduce the potential for shattered rock to reach the highway. An additional consideration is that removing material from the toe of the cut slope may further destabilize the slope.

Our rockfall hazard analysis does not consider larger mass movements related to global-scale planar and wedge sliding, nor does it consider a buildup of rockfall debris at the base of the rock cut slope. Periodic maintenance to remove rockfall blocks from the road shoulder at the base of the slope will likely be required, particularly after significant rainfall events.

4.4 Assessment of the 2002 Rockfall Event

Site conditions soon after the December 13, 2002 rockfall event are illustrated in

Photograph Nos. 2 through 4 in Appendix D. The rockfall zone is approximately 20 ft tall by 10

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Based on the geometry of the release surfaces, we interpret that this rockfall event was controlled by planar sliding along the Plane 3 joint surface at the bottom of the rockfall block.

We believe the major contributing factor in this rockfall event was the relatively small surface area of the Plane 3 boundary surface compared with the relatively large size of the rock block.

The small surface area would provide limited shear resistance to stabilize the large rock block it supported.

The results of our FS analyses indicate that the potential for planar sliding along Plane 3 joints is greater with ground water uplift pressures along the release surfaces acting to increase the driving force. Therefore, water pressures may have also influenced the rockfall event based on the seepage markings that appear from many discontinuities in the rockfall area. A third contributing factor may have been ice wedging that acted along the discontinuity planes bounding the failed rock block over time.

4.5 Failure Modes for Stabilization Design

Based on our kinematic and limit equilibrium analyses, and consideration of the likely cause of the 2002 rockfall event, we conclude that there is a risk of future rock sliding events within the rock slope area. Our analyses show various conditions that yield factors of safety of less than 1.0, which is failure. We believe the following failure modes are possible if no stabilization measures are performed:

1. Global planar sliding along Plane 3 discontinuities.

2. Local planar sliding along Plane 3 discontinuities.

3. Local toppling failure along Plane 2 discontinuities.

The rock slope stabilization measures must account for these failure modes. Rock slope stabilization design options addressing the three failure modes identified above are described in the following report sections.

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5.0 Recommendations

We have developed recommendations for two design concepts to reduce the rockfall hazard potential for global- and local-scale planar sliding and local-scale toppling failure mechanisms indicated in our limit equilibrium analyses. These design concepts were developed to increase the slope stability FS to at least 1.5. Our recommendations consist of the following concepts:

1. Rock Gluing

2. Rock Anchoring

We recommend treatment of the reinforcement area presented in Figure 7. The total area to be reinforced is approximately 2,500 ft2. Both design concepts include elements of scaling, spot bolting, and installation of drains described in more detail below.

5.1 Rock Gluing

This rock slope stabilization design concept involves scaling to remove loose blocks from the rock face and rock gluing to reduce the local-scale rockfall hazard. Spot bolting may also be used, if necessary, in conjunction with scaling to secure potentially unstable local-scale rock blocks that can not be removed safely by scaling. Rock mass drains are proposed to address the potential global planar instability associated with groundwater pressure buildup on Plane 3 discontinuities behind the slope face.

5.1.1 Scaling

Scaling, also referred to as slushing, involves removal of loose or potentially dangerous rock blocks from the slope face by manual scaling (with pry bars), drag scaling (with something heavy such as a dozer track), or mechanical scaling (with hydraulic or pneumatic splitters).

Scaling does not involve blasting.

Scaling, by itself, is not a feasible long term stabilization solution since natural weathering processes and freeze-thaw cycles will act to open discontinuities and destabilize rock blocks on a local scale over time. However, we recommend that scaling is performed as the first step to prepare the rock face for further reinforcement, and as a safety precaution to protect

Project No. 05140065.02/May 11, 2007 23 Schnabel Engineering North, LLC workers by reducing the potential for localized rockfall during reinforcement construction activities.

5.1.2 Spot Bolting

In addition to scaling, some larger rock blocks can be individually bolted (“spot bolted”) to the slope face using tensioned rock anchors or passive rock dowels if it is concluded that further attempts at scaling may destabilize the slope. Spot bolting is not intended to be a major work scope for this stabilization scheme, and may not be required at all. However, an allowance for spot bolting to mitigate localized rock fall potential is recommended in our cost estimate (see below). Design of spot bolting should be evaluated on a case-by-case basis during construction.

5.1.3 Rock Gluing

Rock gluing is recommended as one of the possible slope stabilization design concepts to seal off discontinuities near the slope face to control local-scale planar sliding and toppling instabilities, and to reduce the long-term impacts of weathering and freeze-thaw effects.

Rock gluing is the process of injecting grout into the rock mass for the purpose of bonding individual blocks into a bigger, more stable, continuous mass. Rock gluing has been used in the tunneling and mining industries primarily to control water seeping into underground spaces, but as such it also provides a structural function of improving the structural integrity and slope stability of the rock mass. Rock gluing for rock stabilization is relatively new but it has been used successfully. For example, Central Federal Lands Highway Division of the Federal

Highway Administration, used rock gluing to stabilize the rock mass at a tunnel entrance in

Poudre Canyon, Colorado (Figure 8). Some references that describe rock gluing are included in

Appendix E.

Grouts used in rock gluing can be classified as either cement (i.e., suspension) grouts or chemical (i.e., solution) grouts. We recommend chemical grout for rock gluing since it is almost a pure liquid and is able to permeate narrow rock discontinuities better than cement grout. In particular, the chemical grout we recommend is called a Polyurethane Resin (PUR). The main advantages of PUR grouts include:

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1. PUR grout can penetrate finer cracks than a cement grout because it is a solution grout, instead of a suspension grout, and because there is pressure generated by foaming (to force the grout into discontinuities) in addition to the grouting pressure.

2. PUR grout is available with a variety of characteristics, including viscosity, set time, permeability and strength.

3. Pore volume of the rock mass is reduced, reducing the potential for opening new discontinuities.

4. Although the strength of the polyurethane grout itself can be smaller than that of a cement grout, polyurethane grouts adhere, or bond, better to the surface of rock.

5. PUR grouts form a water barrier (hydrophobic grout will even repel water), thereby reducing freeze-thaw effects in discontinuities.

6. The set time of PUR grout is shorter than that of cement grout, and it can be controlled. This allows a faster grouting operation for PUR grout.

7. As a rock stabilization technique, PUR grouting possesses more aesthetic appeal as it is less conspicuous than cement grout or rock bolting/anchoring methods.

There are some disadvantages to PUR grout that should be considered.