Governs the Making of Photocopies Or Other Reproductions of Copyrighted Materials

Home , Bingham Canyon Mine, Copper sulfide, Kennecott Utah Copper, Quartz latite

Warning Concerning Copyright Restrictions

The Copyright Law of the United States (Title 17, United States Code) governs the making of photocopies or other reproductions of copyrighted materials.

Under certain conditions specified in the law, libraries and archives are authorized to furnish a photocopy or other reproduction. One of these specified conditions is that the photocopy or reproduction is not to be used for any purpose other than private study, scholarship, or research. If electronic transmission of reserve material is used for purposes in excess of what constitutes "fair use," that user may be liable for copyright infringement.

(Photo: Kennecott)

Bingham Canyon Landslide: Analysis and Mitigation

GE 487: Geological Engineering Design Spring 2015 Jake Ward 1

Honors Undergraduate Thesis Signatures:

Abstract

On April 10, 2013, a major landslide happened at Bingham Canyon

Mine near Salt Lake City, Utah. The Manefay Slide has been called the largest non-volcanic landslide in modern North American history, as it is estimated it displaced more than 145 million tons of material. No injuries or loss of life were recorded during the incident; however, the loss of valuable operating time has a number of slope stability experts wondering how to prevent future large-scale slope failure in open pit mines. This comprehensive study concerns the analysis of the landslide at Bingham Canyon Mine and the mitigation of future, large- scale slope failures. The Manefay Slide was modeled into a two- dimensional, limit equilibrium analysis program to find the controlling factors behind the slope failure. It was determined the

Manefay Slide was a result of movement along a saturated, bedding plane with centralized argillic alteration. Recommendations for mitigating future slope failure are provided based on the results of the limit equilibrium analyses.

Table of Contents

Abstract ...... 2

List of Figures ...... 5

List of Tables ...... 8

Introduction ...... 10

History of Bingham Canyon Mine ...... 12

i. Utah and Mining ...... 12

ii. Bingham Canyon Mine: Discovery and Development ...... 13

iii. “The Richest Hole on Earth” ...... 15

iv. World Wars and the Great Depression ...... 18

v. The Latter 20th Century and Bingham Canyon Mine Today . . . 20

Environmental Legacy of Bingham Canyon Mine ...... 22

Social and Economic Impact of the Manefay Slide ...... 26

Geology of the Bingham Canyon Mine ...... 28

i. Regional Geology: The Oquirrh Mountains ...... 28

ii. Geologic Rock Types ...... 29

iii. Structural Geology ...... 41

iv. Economic Geology ...... 47

v. Groundwater Hydrology ...... 58

vi. Seismicity ...... 60

Geotechnical Monitoring ...... 61

Theory: Slope Stability ...... 69

i. Contributory Processes ...... 69

ii. Challenges of Slope Stability ...... 74

iii. Typical Failure Mechanisms ...... 78

iv. Mohr-Coulomb Failure Relationship ...... 84

v. Slope Stability Analysis Methods ...... 86

Open Pit Mine Design & Slope Stability ...... 96

i. Economics of Open Pit Mining ...... 97

ii. Architecture of Open Pit Mining ...... 99

Methods ...... 101

i. Slide by Rocscience, Inc...... 101

ii. Determining Rock Mass Properties ...... 104

iii. Slide Testing ...... 105

Analysis ...... 107

Discussion ...... 116

i. Discussion ...... 116

ii. Recommendations for Mitigating Future Slope Failure . . . 119

iii. Run Out Prediction: Volume-Fahrböschung Relationship . . 121

iv. Future Work ...... 123

References ...... 124

Appendix A: Tabled Data from Direct Shear Tests ...... 129

Appendix B: Tabled Data from Slide ...... 133

Appendix C: Slope Geometry Input for Slide ...... 145

List of Figures

Figure 1. Manefay Slide at Bingham Canyon Mine.

Figure 2. Utah Territory map.

Figure 3. Artist’s view of Bingham Canyon Mine.

Figure 4. Bingham Canyon Mine today.

Figure 5. Boundary of Bingham NPL Zone.

Figure 6. Uinta Axis and Wasatch Fault map.

Figure 7. Stratigraphic column.

Figure 8. Dikes and sills in Bingham Canyon Mine open pit

Figure 9. Distribution of igneous rock map.

Figure 10. Bingham Syncline.

Figure 11. Structural map of Bingham Canyon Mine.

Figure 12. Alteration map.

Figure 13. Ore body cross section.

Figure 14. Economic mineral distribution map.

Figure 15. Aquifer map, Salt Lake County.

Figure 16. USGS 2014 Earthquakes Hazard Map.

Figure 17. Rio Tinto’s TRACK program.

Figure 18. A Robotic Total Station (RTS).

Figure 19. GroundProbe Slope Stability Radar (SSR).

Figure 20. IBIS System.

Figure 21. Uniaxial strength and alteration.

Figure 22. Effect of water pressure on Mohr’s Circle.

Figure 23. Drainage methods.

Figure 24. Spatial variability of rock mass. 6

Figure 25. Discontinuities and scale.

Figure 26. Plane failure diagram.

Figure 27. Wedge failure diagram.

Figure 28. Circular failure diagram.

Figure 29. Toppling failure diagram.

Figure 30. Buckling failure diagram.

Figure 31. Mohr-Coulomb failure envelope.

Figure 32. Typical kinematic analysis plots.

Figure 33. Hoek and Bray (1981) slope height vs slope angle study.

Figure 34. Inverse velocity plot.

Figure 35. Inclined block.

Figure 36. Inclined block with water pressure.

Figure 37. Inclined block with rock bolt.

Figure 38. Circular failure/ method of slices.

Figure 39. Bingham Canyon Mine.

Figure 40. 25 year copper price

Figure 41. Slope geometry of open pit mine.

Figure 42. Typical Slide plot.

Figure 43. Slopes of failure surface in Slide.

Figure 44. Water table in Slide.

Figure 45. Quartzite Mohr-Coulomb plot.

Figure 46. Argillic Clay Mohr-Coulomb plot.

Figure 47. Preexisting discontinuity in quartzite, Slide.

Figure 48. Water table in Slide.

Figure 49. Factor of safety rankings.

Figure 50. FoS plot, preexisting discontinuity in quartzite. 7

Figure 51. Average FoS by failure plane angle.

Figure 52. FoS plot, failure plane in argillic clay.

Figure 53. FoS plot, quartzite sliding on argillic clay.

Figure 54. NOAA precipitation map, April 9, 2013.

Figure 55. Side scarp diagram.

Figure 56. Predicted vs actual run out. 8

List of Tables

Table 1. Advanced limit equilibrium methods.

Table 2. Quartzite strength properties, Styles et al. (2011)

Table 3. Quartzite: 500 psi Normal Load.

Table 4. Quartzite: 1000 psi Normal Load.

Table 5. Quartzite: 1500 psi Normal Load.

Table 6. Quartzite: 2000 psi Normal Load.

Table 7. Quartzite: 2500 psi Normal Load.

Table 8. Peak and residual values for quartzite.

Table 9. Argillic Clay-Quartzite: 500 psi Normal Load.

Table 10. Argillic Clay-Quartzite: 1000 psi Normal Load.

Table 11. Argillic Clay-Quartzite: 1500 psi Normal Load.

Table 12. Argillic Clay-Quartzite: 2000 psi Normal Load.

Table 13. Argillic Clay-Quartzite: 2500 psi Normal Load.

Table 14. Argillic Clay-Quartzite: 3000 psi Normal Load.

Table 15. Failure Plane in Quartzite; Failure Plane: 15˚. DRY.

Table 16. Argillic Clay Sliding on Quartzite; Failure Plane: 15˚. DRY.

Table 17. Failure Plane in Argillic Clay; Failure Plane: 15˚. DRY.

Table 18. Altered Limestone Sliding on Quartzite; Failure Plane: 15˚.

Table 19. Failure Plane in Altered Limestone; Failure Plane: 15˚. DRY.

Table 20. Quartzite Sliding on Argillic Clay; Failure Plane: 15˚. DRY.

Table 21. Failure Plane in Quartzite; Failure Plane: 15˚. SAT.

Table 22. Argillic Clay Sliding on Quartzite; Failure Plane: 15˚. SAT.

Table 23. Failure Plane in Argillic Clay; Failure Plane: 15˚. SAT.

Table 24. Altered Limestone Sliding on Quartzite; Failure Plane: 15˚. 9

Table 25. Failure Plane in Altered Limestone; Failure Plane: 15˚. SAT.

Table 26. Quartzite Sliding on Argillic Clay; Failure Plane: 15˚. SAT.

Table 27. Failure Plane in Quartzite; Failure Plane: 20˚. DRY.

Table 28. Argillic Clay Sliding on Quartzite; Failure Plane: 20˚. DRY.

Table 29. Failure Plane in Argillic Clay; Failure Plane: 20˚. DRY.

Table 30. Altered Limestone Sliding on Quartzite; Failure Plane: 20˚.

Table 31. Failure Plane in Altered Limestone; Failure Plane: 20˚. DRY.

Table 32. Quartzite Sliding on Argillic Clay; Failure Plane: 20˚. DRY.

Table 33. Failure Plane in Quartzite; Failure Plane: 20˚. SAT.

Table 34. Argillic Clay Sliding on Quartzite; Failure Plane: 20˚. SAT.

Table 35. Failure Plane in Argillic Clay; Failure Plane: 20˚. SAT.

Table 36. Altered Limestone Sliding on Quartzite; Failure Plane: 20˚.

Table 37. Failure Plane in Altered Limestone; Failure Plane: 20˚. SAT.

Table 38. Quartzite Sliding on Argillic Clay; Failure Plane: 20˚. SAT.

Table 39. Failure Plane in Quartzite; Failure Plane: 25˚. DRY.

Table 40. Argillic Clay Sliding on Quartzite; Failure Plane: 25˚. DRY.

Table 41. Failure Plane in Argillic Clay; Failure Plane: 25˚. DRY.

Table 42. Altered Limestone Sliding on Quartzite; Failure Plane: 25˚.

Table 43. Failure Plane in Altered Limestone; Failure Plane: 25˚. DRY.

Table 44. Quartzite Sliding on Argillic Clay; Failure Plane: 25˚. DRY.

Table 45. Failure Plane in Quartzite; Failure Plane: 25˚. SAT.

Table 46. Argillic Clay Sliding on Quartzite; Failure Plane: 25˚. SAT.

Table 47. Failure Plane in Argillic Clay; Failure Plane: 25˚. SAT.

Table 48. Altered Limestone Sliding on Quartzite; Failure Plane: 25˚.

Table 49. Failure Plane in Altered Limestone; Failure Plane: 25˚. SAT.

Table 50. Quartzite Sliding on Argillic Clay; Failure Plane: 25˚. SAT. 10

Introduction

Bingham Canyon Mine is located 18 miles southwest of Salt Lake

City, Utah, and is one of the largest open pit mines in the world. The pit is approximately 3,900 feet deep and is more than 2.75 miles across. Bingham Canyon Mine has produced more copper—about 19 million tons—than any other mine in the world, and also produces gold, silver, and molybdenum (Rio Tinto Kennecott, 2013a).

On April 10, 2013, a major landslide occurred in the northeastern pit wall, moving material at an average speed of 70 mph and burying the pit floor in 300 feet of dirt and rock (Carter, 2014; Pankow and

Moore, 2014). The Manefay Slide has been called the largest non- volcanic landslide in modern North American history, displacing more than 145 million tons of material. The Manefay Slide even caused two seismic events registering 2.5 and 2.4 in magnitude, as well as 14 smaller earthquakes (Pankow and Moore, 2014; Carter, 2014).

Fortunately, no one was injured or died because of the landslide.

The mine’s geotechnical operations team used the data collected from several slope monitoring instruments to administer a preemptive response, suspending mining activities and clearing all personnel from the open pit. Bingham Canyon Mine’s “9 Layers of Protection” includes advanced employee training, strict adherence to MSHA standards, and state-of-the-art technologies used to monitor and anticipate the occurrence of the Manefay slide (Rio Tinto Kennecott, 2013).

While the timing of the Manefay Slide was accurately predicted, the run out distance of the material in the landslide was 11 underestimated. The landslide traveled 10,000 feet across the pit floor, and buried 13 haul trucks, three shovels, and 60,000 lb. of diesel fuel underneath the material from the pit slope. Early estimates forecast Rio Tinto would lose $5 million a day that new ore was not delivered to the concentrator (Carter, 2014).

The economic consequence and potential loss of life created by the Manefay Slide has a number of experts in the field of slope stability speculating how to prevent future, large-scale slope failures in open pit mines. This report is concerned with the analysis of the Manefay Slide at Bingham Canyon Mine, with the ultimate goal of mitigating future, large-scale slope failures in open pit mines.

Figure 1. The Manefay Slide at Bingham Canyon Mine displaced 145 million tons of material. 12

History of Bingham Canyon Mine

Utah and Mining

Long after the regional territories that make up the state of

Utah were settled by American Indian tribes, such as the Shoshones,

Utes, and Southern Paiutes, white explorers and trappers came to Utah in search of tradable goods and the infamous, although mythical,

Buenaventura River (Layton, 2014) . In 1847, Brigham Young and the

Mormons settled the Salt Lake City valley in hopes of establishing a religious sanctuary. Not long after, the Treaty of Guadalupe Hidalgo was signed in 1848, ending the Mexican-American War. The terms of the treaty included the transference of much of the Southwestern United

States from the Mexican Government to the American Government, and, ultimately, encouraged the formation of the Utah Territory under the

Figure 2. The Utah Territory was formed under the National Compromise in 1850. 13

National Compromise of 1850. After several delays, including the prevailing anti-Mormon sentiment of the United States Congress and the transition from an autonomous government under the direction of the

Church to one organized by a newly-appointed territorial government,

Utah was granted admittance to the Union in 1896 by President Grover

Cleveland (Powell, 1992).

In the 1860s, mostly non-Mormon settlers came to Utah in search of precious metals, despite Brigham Young’s discouragement against the exploitation of mineral resources. Colonel Patrick E. Connor, a Civil

War commander assigned to protecting the communication lines across the state of Utah, encouraged his men to prospect for precious metals, starting Utah’s mining era. By the 1870s and 1880s, several mines and ore processing facilities were established, as well as newly developed infrastructure to support the increasing number of non-Mormon settlers entering Utah to prospect for precious minerals. Mining and its ability to generate immense wealth played an important role in the development of the social, economic, and political character of the state of Utah (Rood and Thatcher, 2014).

Bingham Canyon Mine: Discovery and Development

The earliest Mormon settlers to ranch and prospect at Bingham

Canyon were brothers Thomas and Sanford Bingham during the 1840s. The brothers were the first to discover outcroppings of ore. The first mining claim made in the Utah territory was at Bingham Canyon

(Whitehead and Rampton, 2006). The claim was made by Colonel Patrick 14

E. Connor after several of his occupying troops had collected what they thought to be ore bearing rock. After assaying the samples and confirming his troops’ theory, the Colonel formed the “Western

Mountain Mining District” in the central Oquirrh Mountains and encouraged his men, many of whom had prospected throughout the gold fields of California and Nevada, to prospect (Arrington and Hansen,

1963). By 1868, the Walker brothers had successfully shipped lead- silver ore by the first wagonloads to Weber County, Utah, so that it could be shipped by train to Baltimore for smelting (Arrington and

Hansen, 1963; Notarianni, 2014). When the transcontinental railroad was completed in 1869 and a branch line to Bingham Canyon was completed in 1873, several million dollars-worth of lead-silver ore was shipped to smelters across the United States (Arrington and

Hansen, 1963). Soon, however, the Panic of 1893 prevented further lead-silver mining at Bingham Canyon (Goin and Raymond, 2004).

As the demand for copper increased with the development of electric power during the mid-1890s, Thomas Weir and Samuel Newhouse started mining and smelting low-grade copper from Bingham Canyon. The new industry attracted several important millionaire-investors, including William Rockefeller and the Guggenheim family (Notarianni,

2014). In 1903, a metallurgical engineer from Missouri and the “father of Utah copper mining”, Daniel C. Jackling, organized the Utah Copper

Company to start copper mining at Bingham Canyon using his newly developed “open-pit” mining method (Notarianni, 2014; Goin and

Raymond, 2004). Workings at the site of Bingham Canyon Mine officially commenced as steam shovels co-financed by the Utah Copper Company and 15 the Boston Consolidated Mining Company broke ground in 1906 (Goin and

Raymond, 2004). In 1910, the Utah Copper Company, the Boston

Consolidated Mining Company, and smaller claim holders were consolidated under financing by the Guggenheims’ to form one large corporation, symbolizing the beginning of formally organized ventures at Bingham Canyon Mine (Notarianni, 2014).

Utah Copper was completely absorbed by the Kennecott Copper

Company, another Guggenheim family interest, in 1935. Bingham Canyon

Mine and the Kennecott Utah Copper Company were later acquired by the minerals division of British Petroleum (BP) in 1981, and were again purchased by Rio Tinto Minerals (then RTZ Minerals) in 1988 (Goin and

Raymond, 2004).

“The Richest Hole on Earth”

“When the Utah Copper Mine was getting started with its novel idea

of open-pit mining, the individual miners were offered stock in

the company as part of their wages. Not many took advantage of the

offer. Most of the miners were underground men and the open-pit

method of mining was not acceptable.” (Dunn, 1973)

By 1911, activity at the Bingham Canyon Mine was in full swing under the direction of Daniel C. Jackling and the Utah Copper Company

(Smith, 1973). With financial backing from the Guggenheims, Jackling had the means to apply his innovative open-pit mining method. Jackling learned that steam shovels significantly reduced the per-ton cost 16 required to move overburden while visiting the Minnesota Mesabi Iron

Range. As the steam shovels started digging at Bingham, an unprecedented 100,000 tons of the overburden “cap” was removed per month. In eighteen months, the shovels had removed 3.2 million cubic yards of overburden to reveal more than six acres of minable ore

(Whitehead and Rampton, 2006).

Open-pit mining proved to be less technical than underground mining (Goin and Raymond, 2004). Because of this, unskilled and immigrant laborers were attracted to the Bingham Canyon Mine. As the physical landscape of Bingham Canyon changed with mining activity, so did the social landscape as the mine needed more laborers to continue growth. For the next two decades, Finns, Swedes, Italians, Croatians,

Figure 2.An artist’s representation of early Bingham Canyon Mine. 17

Serbs, Greeks, Armenians, Japanese, Koreans, and Chinese populated and established mining camps and “villages” around the Bingham Canyon

Mine. The settlements reflected the origin of those living there and included Copperfield, Frogtown, Dinkeyville, and Japtown. Although

Bingham Canyon’s largest settlement, Bingham was mostly Anglo settlers, an estimated 60 percent of Bingham Canyon’s residents were foreign born. Before World War I, immigrant populations at Bingham even joined labor unions like the International Workers of the World spread the doctrine of workers’ rights (Whitehead and Rampton, 2006).

“Upper Bingham with its dark residents and equally dark

businesses with puppet shows, belly dancing, smells of

exotic tobacco, wines and food, never failed to make a

lasting impression on Anglo-Americans.” –Lynn Bailey,

Bingham Historian (Goin and Raymond, 2004).

As the population of the settlements surrounding the Bingham

Canyon Mine continued to expand, so did their general infrastructure.

By 1900, grocery stores, clothiers, hotels, and salons lined Bingham’s main street (Whitehead and Rampton, 2006). Bingham’s streets had an open sewer called the “creek” that was a solution of Copper Sulfide.

The residents considered the solution to have antiseptic properties as it came from the mine that produced their wealth. At times, Bingham’s population grew over 15,000 and had a local government (Goin and

Raymond, 2004). 18

In 1926, the Utah Copper Company established the Copperton community in lower Bingham for its affluent employees and executives.

Copperton featured a collection of “modern” style homes with copper roofs, tiles, nails, and wiring. Unlike the immigrant settlements,

Copperton homes had central heating, yards, underground sewers, and parks. By 1931, there was even a high school in Copperton. However,

Home ownership was based on an employee’s hierarchy in the company, seniority at his position, and his work record. Everything in

Copperton was run by the company, meaning its residents often paid additional prices on store credit that was subtracted from their paychecks. Copperton was closed in 1955 as the mine needed new space to dump waste rock (Goin and Raymond, 2004).

World Wars and the Great Depression

The onset of World War I created a slump in market copper prices as Great Britain placed copper on a conditional contraband list, to reduce the flow of strategic metals to Germany. Bingham Canyon Mine and the Utah Copper Company were not able to escape the effects of the war and were forced to curtail their production by 50 percent. As the

United States entered the war, copper prices quickly rebounded and the

Bingham Canyon Mine was producing at 33 percent above average production. After the war, copper demand fell for a brief period.

In the 1930s as the Great Depression weighed on the already volatile copper market, the Utah Copper Company’s operations were again curtailed. At the lowest point of the Great Depression in 1933, 19

Bingham Canyon Mine was at one fifth of its normal production. Despite the economic downturn, the Utah Copper Company continued to invest in

Bingham Canyon Mine, upgrading mills, laying more than 148 miles of new rail, and adding to the fleet of electric shovels and haul trains.

In 1936, the Bingham Canyon Mine was renewed as Kennecott acquired full retention over the Utah Copper Company as the Guggenheim family consolidated their holdings into a single, publicly traded company

(Whitehead and Rampton, 2006).

Until World War II, unionizing efforts at Bingham Canyon Mine had faltered (Goin and Raymond, 2004). However, in 1944, Kennecott agreed to the first collective bargaining agreement covering wages and safety conditions. Despite these efforts, wartime preparations took its toll on Bingham Canyon Mine as workers left the mining industry for jobs in the defense industry. The shortage of skilled laborers led to an agreement that allowed workers to defer the draft so Kennecott could continue production. Also, a six-month contract enabled workers from

Puerto Rico to come to Bingham Canyon. While most of Bingham Canyon’s men left to take defense industry jobs or were drafted, women started to fill in on maintenance gangs, railroad crews, and supervising mill operations at Bingham Canyon Mine. Kennecott even branded its own version of “Rosie the Riveter”, called “Millie the Miner”, to encourage women to seek employment at Bingham Canyon (Whitehead and

Rampton, 2006).

Bingham Canyon Mine’s contribution to the war effort was tremendous. The government subsidized copper production during the war, allowing Kennecott and the Bingham Canyon Mine to contribute more 20 than half of the three billion pounds of copper produced in the

Bingham mining district from 1941 to 1944 (Whitehead and Rampton,

2006).

The Latter 20th Century and Bingham Canyon Mine Today

In 1977, Kennecott Utah Copper Company spent $300 million on expanding its smelting facility to comply with Clean Air Act (1970) standards. The Kennecott Copper Corporation acquired by British

Petroleum following a down turn in the global copper market in 1980.

Under new leadership and a newly formulated strategic plan, the

Figure 3. Bingham Canyon Mine today. 21

Kennecott Utah Copper Company announced a $400 million revitalization effort to modernize peripheral tailing facilities and other facilities. By 1993, Kennecott Utah Copper employed over 2,400 individuals and produced 300,000 tons of copper annually, as well as significant amounts of gold, silver, and molybdenum. Bingham Canyon

Mine’s smelter and refinery were modernized in 1993 for $880 million, making it the most expensive private investment in Utah history.

Bingham Canyon Mine was recognized as a National Historic

Monument in 1972 (Goin and Raymond, 2004). Bingham Canyon Mine is still in operation today and is expected to be able to mine profitably until 2028 (Whitehead and Rampton, 2006).

Environmental Legacy of Bingham Canyon Mine

After more than a century of open pit mining, Bingham Canyon Mine has been a source of numerous environmental offenses, as environmental stewardship was not a concern of government or industry leaders during its early years of operation. At the “South Zone”, near the open pit, unlined waste rock piles contaminated groundwater as they leached dissolved solids, sulfates, and heavy metals. The “North Zone”, at the end of the Oquirrh Mountains and at the edge of the Great Salt Lake, was where the smelting facility contaminated groundwater and nearby wetlands with a plume of lead, arsenic, and selenium. Passing of the

Comprehensive Environmental Response, Compensation, and Liability Act, or “Superfund Act”, in 1980, allowed the U.S. Environmental Protection

Agency (EPA) to launch an investigation into the contamination at

Bingham Canyon Mine in 1983 (U.S. Environmental Protection Agency,

2006).

In recent years, Rio Tinto Kennecott has been held responsible for reclaiming the several mining-related, contaminated sites. By cooperating with the EPA and the State of Utah Department of

Environmental Quality (UDEQ), Rio Tinto Kennecott avoided Bingham

Canyon Mine making the Superfund National Priorities List (NPL). A

Memorandum of Understanding (MOU) was signed in September of 1995 by

Rio Tinto Kennecott, the EPA, and UDEQ, outlining the scope of cleanup work and environmental studies that Rio Tinto Kennecott would perform at and around Bingham Canyon Mine (U.S. Environmental Protection

Agency, 2006; Rio Tinto Kennecott, 2008). The reclamation projects 23 include mine waste cleanup, groundwater contamination removal, soil cleanup, and facilities demolition, as well as a number of other projects (Rio Tinto Kennecott, 2008).

Figure 5. A five-year review of the Kennecott South Zone shows the NPL listing boundary (U.S. Environmental Protection Agency, 2015). 24

Groundwater Contamination – The Utah Water Quality Board, in conjunction with the United States Geological Survey (USGS) and the

EPA, has established limits on the amount of total dissolved solids

(TDS) allowed in public-supply wells. The Utah Water Quality Board defines “Drinking Water” as having a TDS concentration of 500 to 3,000 mg/L. “Limited Use” and “Saline” water is defined as having a TDS concentration above 3,000 mg/L (Wallace and Lowe, 2009). Findings by

Wallace and Lowe (2009) show that groundwater at the mouth of Bingham

Canyon has a TDS concentration as high as 75,000 mg/L. Contaminants found in groundwater at the mouth of Bingham Canyon include arsenic, cadmium, chromium, copper, lead, nickel, selenium, silver, acids, sulfate, and zinc (U.S. Environmental Protection Agency, 2006).

A fifty-square-mile plume of contaminated groundwater is slowly migrating through the aquifer and is threatening the drinking water of the greater Salt Lake City area. Fortunately, the naturally occurring limestone aquifer is able to neutralize some of the acidic content from the contaminated plume, causing heavy metals to precipitate from solution. Still the sulfates are able to stay in solution in levels that exceed both state and federal concentration limits. The EPA has estimated Kennecott will spend as much as $2.2 billion to pump and treat the contaminated groundwater underneath a growing number of homes in Salt Lake City suburbs (Christensen, 1994).

Groundwater remediation efforts by Rio Tinto Kennecott started in the early 1990s and have been extensive in scope. The application of leachate water on waste rock was discontinued, stopping 95 percent of the contaminated flows from the base of waste rock disposal areas. 25

Collection systems have been upgraded to capture the remaining contaminated runoff from waste rock areas. A triple liner system with electronic leak detection and groundwater monitoring systems was installed under the reservoirs containing leachate solution. Acidic groundwater has been collected by large-capacity production wells and recycled for industrial use at the mine. Also, Groundwater with high sulfate concentrations has been treated to drinking water quality by two reverse osmosis plants for the towns of South Jordan, West Jordan,

Herriman, and Riverton (Rio Tinto Kennecott, 2008).

Social and Economic Impact of the Manefay Slide

Following the Manefay Slide in 2013, the immediate economic impacts for Rio Tinto Kennecott were questioned. The Wall Street

Journal reported that Commonwealth Bank estimated Rio Tinto would lose six percent of its total earnings, or $780 million, assuming Bingham

Canyon Mine would be inoperable for the rest of that year. Other cost expense estimates for cleaning up after the Manefay Slide were less optimistic. A global natural resources fund portfolio manager with

U.S. Global Investors assessed the cleanup costs of the Manefay Slide at more than $1 billion (Petley, 2013; Romboy, 2013).

Rio Tinto Kennecott responded by confirming they would continue production from previously mined ore reserves. However, initial reports suggested Rio Tinto Kennecott had only twenty days of reserves to process (Petley, 2013). Within sixteen days, new ore mining resumed at reduced levels and the year’s production outlook was estimated to be 50 percent of the previous year’s (Kosich, 2013).

Rio Tinto Kennecott initially encouraged 800 employees to take voluntary, unpaid vacation time, before extending the offer to all

2,100 of its direct employees. Kennecott offered 270 employees over 50 years old a one-time retirement incentive of $20,000 per person. Soon, however, it was rumored that 35 percent of Bingham Canyon Mine’s salaried employees would be laid off, and by May of 2013, 100 employees had permanently lost their jobs with more staff reductions occurring layer in the year. Several of mining contractors announced layoffs during the weeks following the slide. Cementation USA, a 27 mining construction contractor, was forced to lay off 45 employees in the wake of the Manefay Slide (Romboy, 2013).

Utah’s economy also felt the effects of the Manefay Slide. In

2011, Bingham Canyon Mine contributed $1.2 billion to the state’s economy, including $270 million in salaries and benefits, $765 million on in-state companies, and $140 million in taxes. The Manefay Slide was expected to have a significant impact on these figures (Romboy,

2013). By the end of 2013, the Manefay Slide had an effect on the overall sales-tax revenue collected in Salt Lake County. The sales-tax revenues were reduced by 3.4 percent, or $444,000, and was expected to be a direct outcome of Kennecott spending less on goods and services.

Revenue from a local option tax imposed on sales in unincorporated areas was reduced by 7.2 percent, as Bingham Canyon Mine is part of an unincorporated part of the greater Salt Lake City area (Gorrell and

Fahys, 2013).

Geology of Bingham Canyon Mine

Regional Geology: The Oquirrh Mountains

The Bingham Canyon Mine is centered on a polymetallic, Tertiary porphyry deposit located 30-km southwest of Salt Lake City, UT, and in the north-central region of the Oquirrh Mountain Range (Babcock et al., 1997). The Oquirrh Mountains are part of three distinct, allochthonous sheets of similarly aged sedimentary rocks, overlying

Precambrian cratonal terrane. The thrust nappes that compose the

Oquirrh Mountains are the Pass Canyon, Bingham, and Rogers Canyon. The

Bingham Canyon Mine is located over an intrusive complex formed along the leading most edge of the Bingham thrust nappe (Tooker and Roberts,

1988). Tooker and Roberts (1988) hypothesize the three thrust systems were generated during the Mesozoic Sevier orogeny and were moved into the region to form an imbricate thrust complex.

The Bingham nappe traveled several miles from the West-Southwest along an underlying basal thrust--the Midas Thrust system--before overcoming the Pass Canyon thrust nappe. The Bingham nappe terminates northeast of the Bingham Canyon Mine as it was truncated by the steeply dipping AJ fault (Tooker and Roberts, 1988). The Bingham nappe is 7,989-m thick with a stratigraphic section that ranges from

Cambrian Quartzite to the Pennsylvanian Bingham Canyon Mine Formation of the Orquirrh Group (Tooker, 1999).

The regional setting of the Oquirrh Mountains contributed significantly to the formation of the Bingham porphyry deposit. The deposition of carbonaceous shelf sediments that make up the Bingham 29 nappe occurred before the Midas thrust system moved the plate. These carbonaceous sediments allowed the North Ore Shoot and Carr Fork skarn mineralization zones proximal to the Bingham porphyry deposit (Babcock et al., 1997). An east-west running cratonal margin called the Uinta

Axis and the north-south trending Wasatch fault line (see Figure 1) have been active since the Archean period and have furthered the deformation of the Oquirrh Mountains through periodic lifting and erosion (Babcock et al., 1997), exposing the plutons which host the porphyry deposit. The Uinta Axis is a source of localized igneous stocks, sills, and dikes that formed from sub-crustal material that has assimilated with autochthonous and allochthonous rocks above the long-lived crustal flaw (Tooker, 1999).

The Oquirrh Mountains are the eastern most, north-trending mountain range in the Basin and Range system and are subject to tensional stress regimes that create an alignment of Horst and Graben fault blocks across the Salt Lake basin. Also, the Oquirrh Mountains are within the Cordilleran fold and thrust belt and are in line with compressional stress regimes (Babcock et al., 1997). These regional stresses created large scale structural features that influenced the formation of local folds, faults, and fractures for hydrothermal solutions to infiltrate.

Geologic Rock Types

The Bingham Canyon Mine is at the leading most edge of the

Bingham nappe. The Bingham nappe is a thick stratigraphy of 30 interbedded Paleozoic quartzites, shales, limestones, and siliceous clastics. The Bingham porphyry deposit is hosted by a complex of

Tertiary intrusions of equigranular and porphyrytic stocks, dikes, and sills (Tooker, 1999).

Sedimentary Rocks –The Pennsylvanian Butterfield Peaks and

Bingham Mine Formation are part of the Oquirrh Group and are the primary sedimentary rock units found at the Bingham Canyon Mine.

Quartzite is the predominant sedimentary rock type within these units

Figure 6. The Uinta Axis and the Wasatch Fault Line are shown in relation to the Bingham Canyon Mine. (Modified from Babcock et al., 1997). 31 and is the most observed throughout the open pit surface. The

Sedimentary rocks units dip moderately to the north-northwest (Bray et al., 1975).

Butterfield Peaks Formation (oldest to youngest) – The middle and upper sections of the Butterfield peaks formation are exposed at the southern part of the Bingham Canyon Mine and are the oldest geologic unit in the mining area (Lanier et al., 1978). The Butterfield Peaks formation has a thickness of 2,765-m and overlies the West Canyon

Limestone to the south of the Bingham Canyon Mine (Swensen, 1975). The lower section of the Butterfield Peaks Formation (exposed just south of the Bingham Canyon Mine) is an alternating series of thinly bedded limestone, quartzite, sandstone, siltstone, and shale.

The middle section of the Butterfield Peaks formation is a series of alternating beds of bluish-gray cherty, argillaceous, fossiliferous, or arenaceous limestones, and light gray calcareous quartzites or feldspathic orthoquartzites (Swensen, 1975). This characteristic series of limestones and quartzites has been identified as the Alphabet Series and host Pb-Zn-Ag mineralization in a mine located near the Bingham Canyon Mine, but are not ore-bearing at the

Bingham Canyon Mine (Babcock et al., 1997).

The upper section of the Butterfield Peaks formation is mainly light-gray to tan orthoquartzite and calcareous quartzite with interbedded and cross bedded medium-gray, arenaceous or cherty limestone and calcareous sandstones.

Bingham Mine Formation – The Bingham Mine Formation is the primary stratigraphic unit that the Bingham Canyon Mine operation intersects.

The Bingham Mine Formation is 2,680-m thick and is terminated to the north of the open-pit mine by the southwest trending Bear Fault and the underlying Midas Fault (Swensen, 1975). The Bingham Mine

Formation is further subdivided into the lower Clipper Ridge Formation and the upper Markham Peak Formation (Tooker and Roberts, 1988).

The Clipper Ridge formation is 944-m thick. At the base of the

Clipper Ridge formation is the Jordan limestone--a dark gray, argillaceous, cherty, and arenaceous limestone (Swenson, 1975).

Overlying the Jordan limestone is a layer of light gray, calcareous quartzite, orthoquartzite, and calcareous sandstones. The Commercial limestone overlies the quartzite-sandstone layer and is a thin bedded, dark gray, argillaceous, silty, and cherty limestone. The Jordan and

Commercial limestones host Cu-Au skarn mineralization in the North Ore

Shoot and Carr Fork deposits (Babcock et al., 1997). Above the

Commercial limestone is a 720-m section of light gray, orthoquartzites, and calcareous quartzites with interbedded calcareous, fine-grained limestones, sandstones, siltstones, and shales (Swenson, 1975). Ore-grade copper mineralization occurs to 800- m horizontally from the porphyry intrusive complex, into these calcareous layers (Lanier et al., 1978). The upper most layer of the

Clipper Ridge formation are the Manefay beds, three thin beds of arenaceous limestone and calcareous sandstone (Lanier et al., 1978).

The Markham Peak formation is approximately 670-m thick and is a light-gray, thin-bedded, and locally cross-bedded section of 33

Figure 7. The Paleozoic and Mesozoic sequences of sedimentary rocks at the Bingham Canyon Mine. (Modified from Swensen, 1975). orthoquartzites and calcareous quartzites. The orthoquartzites and calcareous quartzites are interbedded with light gray, thin, 34 calcareous, fine-grained sandstones, limestones, and siltstones

(Swenson, 1975).

Curry Peak Formation and Freeman Peak Formation – The Curry Peak and

Freeman Peak formations are overlying the Bingham Mine Formation, but due to faulting, rotation, and erosion, are only exposed to the north of Bingham Canyon Mine. The formations may be within zones of certain alterations.

Igneous Rocks – Several episodes of igneous intrusion occurred during the late Eocene (39.8±0.4 Ma) and early Oligocene (38.8±0.4 Ma) to form the Bingham Stock (Lanier et al., 1978). The Bingham Canyon

Mine is located on the northeastern most edge of the Bingham thrust nappe, which has deformed through tear faults associated with local folding and thrust faults developed during the Mesozoic Sevier orogeny. These openings have provided sufficient pathways for the introduction of magmatic bodies and hydrothermal solutions (Tooker,

1999).

The Bingham Stock is interconnected to the Last Chance stock and surrounding plutons of similar texture and composition by dikes, suggesting a possible convergence of the stocks at depth (Lanier et al., 1978). The hypothesized emplacement of intrusive bodies may not be exact due to an absence of cross-cutting relationships and available material for radiometric dating (Bray et al., 1975). Phases of the Bingham Stock have occurred as extrusive volcanics that have 35 been observed to have similar mineralogical compositions and radiometric age dates as the intrusive stocks.

Equigranular Monzonite (Eocene) – Monzonite makes up a third of the pit exposure and is the most common igneous rock type at the Bingham

Canyon Mine (Bray et al., 1975; Lanier et al., 1978). The Bingham

Canyon Mine is centered over a Monzonite pluton which extends south and west from the center of the pit.

Unaltered Monzonite is dark-gray, fine-grained, and equigranular.

The unaltered Monzonite contains near equal percentages of Orthoclase

(30 percent) and Plagioclase (33 percent), interstitial quartz (7 percent) and feldspar phenocrysts (1-2 percent), with sub- to anhedral augite (11 percent), uralitic amphibole (7 percent), biotite (8 percent), and magnetite (4 percent) (Lanier et al., 1978; Tooker and

Roberts, 1988). In the highly altered rock found throughout the open- pit surface, ferromagnesian minerals have been replaced and make up approximately 25 percent of the rock. Quartz veinlets, sulfide minerals (ore grade chalcopyrite and associated pyrite, disseminated and in veins), apatite, zircon, and rutile are accessory minerals and inclusions (Bray et al., 1975). The sulfides are an important host for the porphyry copper deposit.

Crosscutting relationships show the Equigranular Monzonite to be cut by the later porphyry intrusions, suggesting it was the first intrusive body to rise while assimilating with surrounding crustal material (Lanier et al., 1978).

Porphyrytic-Equigranular Quartz Monzonite (Ohio Copper Dike) (Eocene)

– The Porphyritic-Equigranular Quartz Monzonite, or Ohio Copper Dike, makes up approximately two to five percent of the Bingham Mine’s pit exposure (Swensen, 1975). The Ohio Copper Dike is a northwest- trending, rectangular intrusion that outcrops in the southeast quadrant of the Bingham Canyon Mine. The Ohio Copper Dike is approximately 150- by 300-m and plunges 55⁰ to the northwest (Lanier et al., 1978). The Ohio Copper Dike is gray to greenish-gray and is porphyritic with an equigranular phaneritic groundmass. The porphyritic grains (15 percent) are feldspar phenocrysts with an

Figure 8. Several Dikes and sills, some ore bearing, cross the Bingham Canyon Mine. 37 average diameter of 5-mm, specifically plagioclase phenocrysts with an average diameter less than 5-mm and orthoclase phenocrysts with an average diameter greater than 5-mm (Swenson, 1975). The groundmass is similar in composition to the Quartz Monzonite Porphyry, though it is slightly coarser grained (Swenson, 1975). The dike boundary is clearly outlined by an increase in phenocrysts (Tooker and Roberts, 1988).

There are pyrite veins in localized zones across the Ohio Copper Dike

(Lanier, 1975).

Latite (Bear Gulch Latite) (Oligocene) – The Bear Gulch Latite is a small intrusion located near the southeast edge of Bingham Mine and is approximately 400- by 600-m (Tooker and Roberts, 1988; Lanier et al.,

1978). The Bear Gulch Latite is buff to light-gray and porphyrytic with an aphanitic groundmass with local microcrystalline zones (Lanier et al. 1978; Lanier, 1975). Near the perimeter of the intrusive, sub- rounded quartzite fragments (1 to 7-cm) from the surrounding Bingham

Mine Formation form breccia zones (Lanier et al., 1978).

The Bear Gulch Latite is highly altered and contains quartz grains with an average diameter of 0.7-mm (10-15 percent), phlogopite

(5 percent), and aggregates of clay, sericite, quartz, and pyrite from altered ferromagnesian and feldspar (probably Amphibole and Ca- feldspar) minerals (10 percent) (Tooker and Roberts, 1988; Lanier,

1975; Lanier et al., 1978). Orthoclase phenocrysts are generally more abundant than plagioclase phenocrysts and are subhedral (Tooker and

Roberts, 1988). Remnant outlines of feldspar and ferromagnesian mineral phenocrysts are present (5 percent) (Bray et al., 1975). 38

Pyrite is observed as disseminated grains and veinlets throughout the intrusion (Bray et al., 1975; Lanier, 1975).

The relative occurrence of the Bear Gulch Latite, in relation to the other intrusions, is unclear due to a lack of cross-cutting relationships and accurate radiometric dating. Bray et al. (1975) and

Lanier’s Geologic Map of the Bingham Mine (1975) show the Bear Gulch

Latite occurring after the Quartz Monzonite Porphyry and the Latite

Porphyry; however, Lanier et al. (1978) report the Bear Gulch Latite occurring before these intrusions and Tooker and Roberts (1988) have the Bear Gulch Latite emplaced between these intrusions.

Quartz Monzonite Porphyry (Oligocene) – The Quartz Monzonite Porphyry is the most abundant porphyry intrusion and makes up five percent of the pit exposure (Bray et al., 1975). The Quartz Monzonite Porphyry is the center of mineralization at the Bingham Canyon Mine as it is highly altered. The Quartz Monzonite Porphyry is a northeast trending dike located at the center of the mine and has a variable degree of dip to the northwest and a maximum width of 400-m (Lanier et al.,

1978).

The Quartz Monzonite Porphyry is light-gray and is porphyritic with an altered aphanitic groundmass (Bray et al., 1975). The porphyry

(50 percent of the rock) contains orthoclase, andesine feldspar, and phlogopitized amphibole phenocrysts with an average diameter of 1.7-mm

(Tooker and Roberts, 1988; Lanier et al., 1978). The aphanitic groundmass of the Quartz Monzonite Porphyry is 0.05-mm orthoclase and quartz grains (51 percent of ground-mass), microscopic aggregates of 39

phlogopite shreds that have replaced hornblende-augite grains (6-7

percent), biotite (3 percent), pyroxene (3 percent), amphibole (11

percent), and plagioclase (24 percent) (Bray et al., 1975; Tooker and

Roberts, 1988). Accessory quartz veins, sulfide minerals (ore grade

chalcopyrite and associated pyrite, disseminated and in veins),

rutile, and zircon make up the rest of the rock (Bray et al., 1975;

Lanier, 1975). The feldspar phenocrysts are locally altered to clay

minerals and sericite (Bray et al., 1975).

The pre-alteration modal composition of the porphyry is estimated

to be that of a hornblende-biotite quartz monzonite with plagioclase

(32 percent), quartz (23 percent), orthoclase (32 percent), and mafic

Figure 9. The zonal distribution of igneous rocks are shown at the Bingham Canyon Mine (Phillips et al., 1997). 40 and accessory minerals (14 percent) (Lanier et al., 1978). It is likely that the intrusion of the Quartz Monzonite Porphyry was the thermal source for hydrothermal fluids that mineralized the Bingham

Stock (Lanier, 1975).

Latite Porphyry (Oligocene) – The Latite Porphyry makes up three percent of the pit exposure at the Bingham Canyon Mine and is part of a major swarm of northeast trending dikes and sills. The Latite

Porphyry includes the Main Hill Dike, the Starless Dike, and the

Fortuna Sill that intrude the Equigranular Monzonite and Quartz

Monzonite Porphyry (Lanier et al., 1978; Lanier, 1975; Bray et al.,

1975).

The Latite Porphyry dikes are light to medium-gray and are porphyritic with an aphanitic groundmass (Lanier, 1975; Bray et al.,

1975). The dikes contain plagioclase and minor orthoclase phenocrysts with an average diameter of 2.3-mm (30 percent), megascopic phlogopite aggregates in hornblende-augite replacement sites (12 percent), biotite phenocrysts (4 percent), and partially resorbed quartz phenocrysts with an average diameter of 0.6-mm (3 percent) (Bray et al., 1975). The remaining groundmass is composed of fine grained orthoclase, quartz, and biotite with an average diameter of 0.02- to

0.08-mm (Bray et al., 1975; Lanier et al., 1978). Local quartz veinlets, sulfide minerals (ore grade chalcopyrite and associated pyrite, disseminated and in veins), and other accessory minerals are present (Bray et al., 1975; Lanier, 1975).

Hornblende-Biotite Quartz Latite Porphyry (Oligocene) – The Quartz

Latite Porphyry are narrow (less than 10-m across), northeast trending dikes that make up less than one percent of the pit exposure (Lanier et al., 1978; Bray et al., 1975). The Andy Dike is the most extensive dike that intrudes the Equigranular Monzonite, Quartz Monzonite

Porphyry, and Latite Porphyry (Lanier, 1975).

The Quartz Latite Porphyry is medium-gray to green and is porphyrytic with an aphanitic groundmass. The Quartz Latite Porphyry contains feldspar phenocrysts with an average diameter of 2- to 30-mm

(25 percent), quartz phenocrysts with an average diameter of 3- to 6- mm (8 percent), and biotite phenocrysts (8 percent); however, much larger feldspar phenocrysts occur locally and are up to 3-cm across

(Bray et al., 1975; Tooker and Roberts, 1988; Lanier et al., 1978).

The aphanitic groundmass contains 0.05-mm orthoclase, quartz, plagioclase, and considerable hornblende (Bray et al., 1975; Lanier,

1975). Local zones of phlogopite and chlorite replacement of hornblende occurs within the pit (Lanier, 1975).

The Quartz Latite Porphyry is less altered, mineralized, and fractured, suggesting the rock type may be the youngest intrusion and postdate the most recent hydrothermal event (Lanier et al., 1978; Bray et al., 1975).

Structural Geology

The structural features that distinguish the Bingham Canyon Mine are asymmetrical compressional folds and a collection of 42 perpendicular, northeasterly and northwesterly striking fault systems.

These structural features are related to the Bingham Canyon Mine’s relative proximity to the leading edge of the Bingham nappe and compressional events related to the Mesozoic Sevier Orogeny (Smith,

1975; Babcock et al., 1997; Lanier et al., 1978). These structures contain several faults, fractures, and joints allowing episodes of igneous intrusions and, later, the introduction of hydrothermal solutions to create economic mineralization (Smith, 1975; Lanier et al., 1978).

Figure 10. Sections of the Bingham Syncline are locally overturned. 43

Bingham Syncline – The Bingham Syncline is a main, through-going fold that is associated with the mine’s proximity to the leading edge of the Bingham nappe (Tooker, 1999). As the Bingham nappe was transported to its current location, drag across the underlying Midas

Thrust Fault created several regional folds parallel to the leading edge of the nappe, including the Bingham Syncline.

The Bingham Syncline is the prominent fold at the Bingham mine and is expressed on the north edge of the open pit (Swenson, 1975).

The axis of the Bingham Syncline strikes N 60⁰W and plunges 12⁰ to the northwest, but becomes difficult to delineate as it continues to the southeast, as the limbs of the fold broaden (Smith, 1975; Lanier et al., 1978). The limbs of the Bingham Syncline are asymmetrical and locally overturned. The southwest limb of the syncline encompasses the

Bingham Canyon mine and dips moderately to the northeast, while the northeast limb of the syncline is north of the mine and dips gently to the southwest (Babcock et al., 1997; Tooker and Roberts, 1988). The fold steepens to the north and has a well-defined axial trough, before being truncated by the north-northwest trending Bear Fault (Swenson,

1975; Smith, 1975).

The Bingham Syncline is offset directly north of the mine by the left-lateral, strike-slipping Andy Fault and has an axial displacement of approximately 0.5- to 0.7-km (Swensen, 1975). The several episodes of igneous intrusions and hydrothermal mineralization occur in the south limb of the Bingham Syncline (Lanier et al., 1978).

Apex Fold – The Apex Fold is a secondary anticlinal structure found in the southwest limb of the Bingham Syncline. The Apex Fold follows the west-northwest trend of the Bingham Syncline and follows a similar plunge to the northwest. The limbs of the Apex Fold both dip moderately to the north, as the north limb is overturned. As the Apex

Fold enters the mine, it bends south due to topography and is offset by the Bingham Stock (Lanier et al., 1978).

Midas Thrust Fault – The Midas Thrust Fault (referred to as the

North Fault in older literature) is the regional, low-angle thrust fault that the Bingham nappe was transported across to its current location. In the vicinity of the Bingham Mine, the fault is east-west trending and dips to the south.

The Midas Fault generally places the upper sedimentary sequences of the Pennsylvanian Bingham Mine Formation in contact with the

Permian Curry Peak and Freeman Peak formations, but sections of the upper and lower plate formations are identical portions of the

Pennsylvanian sequences. The stratigraphic offset of the upper and lower plates varies from close to zero offset to upwards of 1.6-km of vertical throw (Smith, 1975). The Midas Thrust Fault has caused regional compressional folding, producing bedding-plane slippage throughout the sedimentary stratigraphy of the Bingham Mine Formation

(Lanier et al., 1978; Smith, 1975).

Northeast Trending Fault System – The Northeast Trending Fault

System is the collection of high-angle, mostly reverse-type faults 45 associated with economic mineralization and ore production (Smith,

1975). Lanier et al. (1978) hypothesize the Northeast Trending Fault

System is the result of compressional stress from the north, resulting in the creation of the North Oquirrh Thrust 6-km from the Bingham

Canyon Mine. The northeast trending faults occur in the southern part of Bingham Canyon Mine, extending into the Bingham Stock, and are en echelon faults and fissures that dip steeply to the west (Smith,

1975).

The relative age of the northeast trending faults is uncertain; however, the highly mineralized fault zones and cross cutting relationships suggest the stress regime occurred after the emplacement of the igneous intrusions and before the occurrence of hydrothermal fluids (Lanier et al., 1978; Smith, 1975). Although they are outside of the major ore-grade copper zone, the Northeast Trending Fault

System hosts a major portion of the Pb-Zn-Ag deposits (Smith, 1975).

Several of the northeast trending faults behave as left-lateral, strike-slip faults, like the Andy Fault that displaces the Bingham

Syncline. The northeast trending left-lateral faults roughly correlate with the Bingham Syncline’s decrease in amplitude (Lanier et al.,

1978).

Northwest Trending Fault System – The northwest trending fault system is hypothesized by Lanier et al. (1978) to be the result of compressional stress regimes from the southwest, and, later, extensional stresses from the Basin and Range fault system to the west of the Oquirrh mountains. The northwest trending faults are fewer than 46 those of the northeast trending system, but often much greater in length, and include the Occidental Fault, the Giant Chief Fault, and the Bear Fault.

The Northwest Trending Fault System contains mostly normal-type faults that dip steeply to the southwest (Lanier, 1975). The northwest trending faults form wide gouge and breccia zones in the Equigranular

Monzonite, suggesting they were caused by Cenozoic Basin and Range extension, after the emplacement of the intrusive complex (Lanier et al., 1978).

Figure 11. Structural map of the Bingham Canyon Mine. (Modified from Lanier et al., 1978) 47

Economic Geology

The central porphyry at Bingham Canyon Mine hosts copper- molybdenum-gold-silver mineralization along closely-spaced fractures in the Equigranular Monzonite stock and its several associated dikes and sills. Veins and dissemination replacement occurs in the surrounding host rock as lead-zinc-copper-silver-gold deposits. The carbonate-rich sediments close to the intrusions have been metasomatically altered to form copper-gold skarns (Tooker, 1999).

Mineralization and alteration started immediately following the crystallization of the Equigranular Monzonite Porphyry. As the subsequent intrusions provided magmatic fluids and warmed pore-fluid from the surrounding host sediments, the different igneous rocks were altered through mineral replacement. The relative degree of alteration occurred differently throughout the Bingham Stock, and was finished by the emplacement of the Quartz Latite Porphyry (Bray et al., 1975).

The Bingham ore body is roughly shaped like a “hollow-dome” with roots extending into the surrounding host sediments. The extent of mineralization is controlled by the density and number of planar fractures that occur in each rock type throughout the Bingham Stock.

The permeable fractures allowed for mineral-rich, altering hydrothermal fluids to enter the host rock. The extent of alteration in the north half of the pit is much greater than the rest due to the system of north-east trending faults and fractures in that section of the open-pit. The ore shell tilts 17˚ to the east according to Basin and Range extensional faulting. 48

Alteration in Equigranular Monzonite – Three primary alteration zones are recognized in the Equigranular Monzonite porphyry at Bingham

Canyon Mine: 1) quartz-orthoclase-phlogopite, 2) sericite-quartz, and

3) actinolite-chlorite-epidote.

In the quartz-orthoclase-phlogopite zone, ferromagnesian replacement occurs as the listed minerals and as ore-bearing chalcopyrite to a lesser extent. This type of potassic alteration causes phlogopite to be locally altered to chlorite as successive metasomatism has altered minerals through several phases. Similarly, plagioclase is replaced by orthoclase. Disseminated, copper bearing sulfides replace magnetite. This type of alteration is closely associated with the Quartz Monzonite Porphyry (Lanier et al., 1978).

Figure 12. The distribution of alteration zones throughout the Bingham Stock and surrounding host sediments (modified from Lanier et al., 1978). 49

The sericite-quartz alteration locally replaces previously altered minerals except quartz and orthoclase with quartz and sericite. The sericitic alteration occurs strongly in selvages around quartz-pyrite veins and veinlets and around gouge-filled fault zones

(Lanier et al., 1978).

The actinolite-chlorite-epidote zone, or the propylitic alteration zone, surrounds the potassic alteration zone. Actinolite replaces augite. Subsequently, preexisting actinolite, augite, and biotite are replaced by vein-controlled chlorite. Epidote formation is widespread are is common along veins and is disseminated throughout the host rock (Lanier et al., 1978).

Alteration in Porphyries – Four primary alteration zones are recognized in the porphyries across Bingham Canyon Mine: 1) quartz- orthoclase-phlogopite-(biotite), 2) sericite-quartz, 3) actinolite- chlorite-(epidote), and 4) calcite-chlorite-quartz.

The mineral zones in the associated porphyries are similar to those in the Equigranular Monzonite, except that 1) mafic replacement products are less of the total altered rock mass due to a lower mafic mineral content of the original rock, and 2) plagioclase is only partially altered to orthoclase as most plagioclase is altered to clay or sericite. In general, the sericitic and propylitic zones are similar to those found in the Equigranular Monzonite (Lanier et al.,

1978).

The calcite-chlorite-quartz mineralization appears exclusively in the north-west trending system of Latite dikes (Latite and Latite 50

Porphyry). The mineralization products listed replace preexisting mafic minerals. Calcite also locally replaces plagioclase. This type of alteration has been described by as a product of sulfide-associated hydrothermal fluids and CO2 bearing fluids (Lanier et al., 1978).

The igneous intrusions, including the original Equigranular

Monzonite Porphyry, are identified to have experienced metasomatic alteration during three distinct phases: 1) early hydrous, 2) Mg and K metasomatism, and 3) late hydrous. The early hydrous stage is recognized to be the formation of actinolite through the alteration of augite. Potassium is likely to have been produced as plagioclase was replaced with orthoclase. Magnesium and Potassium metasomatism occurred after the early hydrous stage, superimposing phlogopite replacement of the altered actinolite and continued replacement of plagioclase by orthoclase. Dissemination of iron and copper sulfides also occurred during the Mg and K metasomatism stage, given the close spatial relationship between phlogopite and sulfides. The late hydrous stage included alteration of phlogopite to chlorite in the potassic zone, alteration of biotite, augite, and actinolite to chlorite in the propylitic zone, local sericitic alteration near quartz-pyrite veinlets, and intermediate argillic alteration of plagioclase to montmorillonoid and sericite (Lanier et al., 1978).

Alteration in Sedimentary Rocks – The sedimentary rocks surrounding the Bingham Stock have been subject to three recognizable stages of hydrothermal alteration (Babcock et al., 1997). 51

Early stage Mg metasomatism (Lanier et al., 1978) is recognized by the development of wollastonite in cherty limestone and quartz diopside in calcareous quartzite and silty limestone in wall rocks surrounding the igneous intrusions. The alteration mineralogy changes outward from the intrusions as diopside is replaced by tremolite, talc, and dolomite

(Babcock et al., 1997).

Main stage Fe metastomatism (Lanier et al., 1978) is signified by the development of quartz-pyrite and sulfide veinlets through fracture networks that extend into the Bingham Stock. Biotite and actinolite selvages superimpose early stage alteration throughout the quartzite.

Limestones are altered by the overprinting of early stage wollastonite to garnet, diopside, quartz, magnetite, hematite, and copper sulfides

(Babcock et al., 1997).

Late stage hydrous alteration includes the replacement of previously emplaced calcsilicates by chlorite, montmorillonite, sericite, and talc (Babcock et al., 1997). The argillic alteration in sedimentary rocks may be a continuation of the clay alteration in igneous rocks and is probably centralized on bedding planes (Lanier et al., 1978).

In the quartz-rich sedimentary rocks surrounding the Bingham

Stock, mineralogy differences in the alterations of calcareous quartzite and calcareous sandstone and limestone beds is recognized.

In the calcareous quartzite beds, silicate mineral assemblages occur in sites of the calcareous cementation interstitial to the quartz grains. In general, alteration in the calcareous quartzite beds resembles the zoning patterns in igneous rocks and includes a quartz- 52 biotite-orthoclase zone, a sericite zone, and an actinolite zone. The quartz-biotite-orthoclase zones in quartzite are ore-bearing in the lower sections of the Bingham Canyon Mine (Lanier et al., 1978).

In the calcareous sandstone and thin limestone beds, the alteration stage assemblages are similar to those observed in the calcareous quartzite. However, alteration zoning telescopes inward toward the Bingham Stock. The alteration zones are identified as actinolite, diopside, and talc-tremolite (Lanier et al., 1978).

In the thick, relatively pure limestone beds, two alteration zones are recognized: 1) a garnet zone and 2) an outer wollastonite zone. The garnet alteration zone contains garnets, diopside, magnetite, clay, quartz, opal, and actinolite. Ore grade copper is found in the garnet zone. The wollastonite zone is characterized by wollastonite, diopside, idocrase, andradite, and sulfide (Lanier et al., 1978).

The pure limestone beds have been altered by three main stages of alteration: 1) early contact metamorphism, 2) Fe metasomatism, and 3) late hydrous state alteration. During the emplacement of the igneous intrusions, dispersive heat metamorphosed the pure limestone beds, resulting in wollastonite-diopside assemblages. Iron metasomatism developed garnet, magnetite, sulfides, and ore-grade copper mineralization. The late hydrous stage alteration, earlier formed assemblages were replaced by epidote, chlorite, montmorillonoids, sericite, talc, and opal (Lanier et al., 1978).

Skarn Alteration – Skarn formation occurs in the Commercial and

Jordan Limestones at North Ore Shoot and Carr Fork. Copper, molybdenum, gold, and silver are disseminated and in veins in the skarn formations (Babcock et al., 1997).

Alteration zones in the North Ore Shoot parallel the northern lobe of the Bingham Stock. The marble zone is considered to be weakly altered limestone and contains leached limestone with sooty carbonaceous material. The garnet zone contains andradite garnet, pyroxene, quartz, calcite, and magnetite. Local areas of garnet-clay skarn are present. Pyrite and chalcopyrite veinlets occur throughout the North Ore Shoot skarn formation and occur with magnetite, selvages of quartz, calcite, actinolite, epidote, biotite, and orthoclase.

Several retrograde skarn zones signify the alteration of previous skarn zones and include iron oxide, actinolite, and massive sulfide.

The massive sulfide skarn occurs along the marble and garnet zone contact, and contains the largest portion of ore grade mineralization

(Babcock et al., 1997).

The Carr Fork skarn formation differs from the higher-grade North

Ore Shoot. In the marble and garnet zones at the Carr Fork formation, wollastonite replaces chert nodules. Massive sulfide mineralization is less common than and occurs across narrower widths. Pyroxene zones are observed with garnet overprinting between the marble and garnet zones.

Maximum grade copper mineralization occurs along the pyroxene-garnet boundary (Babcock et al., 1997).

Metal Zoning – Five concentric, overlapping metal zones occur from the interior of the Bingham Stock outward: 1) barren core, 2) molybdenum, 3) copper, 4) iron, and 5) lead-zinc-silver. Molybdenum is present in molybdenite lined fractures. Copper occurs in chalcopyrite, bornite, and chalcocite as disseminations and in veinlets. Iron is mainly in pyrite, but also some magnetite and hematite. The lead-zinc- silver zones occurs as galena and sphalerite. Local manganese, barium, and fluorine are present around the outer edges of the lead-zinc- silver zone (John, 1975).

The inner most zone, or the barren core, is mostly Equigranular

Monzonite that has altered potassically. The barren core contains less than 0.5 percent sulfides as pyrite, chalcopyrite, bornite, and molybdenite. Sulfide mineralogy of the metal zones changes gradually from the center outward, as the amount of molybdenite, bornite, chalcopyrite, and pyrite increases depending on the type of alteration to effect each area (Babcock et al., 1997).

Copper Zone – The Bingham copper deposit is approximately triangular in plan view, and extends from the Bingham stock into the surrounding sedimentary rocks. The northeastern and eastern boundary of the copper zone roughly follows the interface between intrusive and sedimentary rocks. The southern boundary of the copper zone is largely within the Equigranular Monzonite. The extent of mineralization is a function of the pre-alteration fracturing of the intrusive and sedimentary host rocks. In general, quartzite was a poor host of mineralization (Babcock et al., 1997). 55

The copper mineralization is characterized by a “ring” of high grade

ore surrounding an inner core of low grade deposit. The low grade ore

is associated with the central molybdenite zone. Higher grade copper

ore occurs in the northern end of the Quartz Monzonite Porphyry and

the northern lobe of the Equigranular Monzonite, as the Quartz

Monzonite Porphyry is the most prominent host of disseminated sulfides

while the fracture-controlled mineralization occurs widely in the

Equigranular Monzonite. The higher grade zone is linked to strong clay

alteration and is associated with the bornite-chalcopyrite sulfide

zone (Babcock et al., 1997).

Figure 13. Metal zoning in cross-section showing the “dome” shaped ore body (Babcock et al., 1997).

Molybdenum Zone – The molybdenum zone follows the shape of the copper zone, but is much smaller in area. Molybdenite concentrations decrease radially from the center of the deposit. Molybdenum deposits are upwards of 0.1 percent molybdenite where the zone is most developed. Most of the molybdenite occurs in stockwork quartz veins and in along fractures surfaces (Babcock et al., 1997).

Gold Distribution – Generally, gold mineralization and distribution follows that of copper. High grade deposits of gold occur where bornite is in excess of chalcopyrite. The Quartz Monzonite

Figure 14. The economic mineral zones at Bingham Canyon Mine include molybdenum, copper, and an encircling lead-zinc-silver halo (modified from Tooker and Roberts, 1988).

Porphyry is a better host to disseminated gold than copper, as that is where gold is the most abundant (Babcock et al., 1997).

Submicroscopic particles of gold linked to pyrite crystals are in the Equigranular-Porphyritic Quartz Monzonite that is exposed in the southern part of the open-pit mine. These gold particles are associated with late stage, argillic alteration and are considered to be superimposed on the primary copper-gold system. Other gold mineralization occurs in faults and fissures in the pyrite and lead- zinc-silver halo near the outer edge of the copper deposit (Babcock et al., 1997).

Pyrite and Lead-Zinc-Silver Halo – A pyrite halo containing upwards of five percent pyrite surrounds the Bingham copper deposit.

The pyrite concentration decreases to one percent pyrite at 2,000 to

3,000 feet from the open-pit mine (Babcock et al., 1997).

A lead-zinc-silver halo encircle the Bingham copper deposit.

Historically, there have been a number of mines to produce lead, zinc, and silver near the Bingham Canyon Mine. The fault and fissure associated lead-zinc-silver mineralization post-dates the copper emplacement, and is occurs as argentiferous galena, sphalerite, and tetrahedrite-tennantite. Related alteration occurs locally as chlorite, clay, sericite, talc, and opal (Babcock et al., 1997).

Groundwater Hydrology

Bingham Canyon Mine is located over the interface between the primary and secondary recharge area for Salt Lake Valley’s (also known as the Jordan Valley) principal basin-fill aquifer. Wallace and Lowe

(2009) describe primary recharge areas as the coarse-grained, unconsolidated, and non-continuous fine-grained uplands along basin margins. Primary recharge areas are defined as having downward groundwater gradients. Secondary recharge areas are explained by

Wallace and Lowe (2009) to be mountain-front benches that have continuous fine-grained layers that are greater than 20 feet across and have a downward groundwater gradient.

In the case of Bingham Canyon Mine, groundwater is discharged to the principal basin-fill aquifer that supplies Salt Lake City, the

Jordan River, and the Great Salt Lake (Wallace and Lowe, 2009).

Groundwater contamination is discussed in the Environmental, Social, and Political Concerns section of this paper. 59

Bingham Canyon Mine

Figure 4. Bingham Canyon Mine is located on the interface between Salt Lake Valley’s primary and secondary recharge areas (modified from Wallace and Lowe, 2009). 60

Seismicity – The United States Geological Survey (USGS) 2014

National Seismic Hazard Map suggests there is a two percent probability in a 50 year period of exceeding a peak ground acceleration of 0.4 times the acceleration of gravity during a seismic event at Bingham Canyon Mine. The Oquirrh Mountains are effected by ongoing Basin and Range extensional seismicity. Similarly, the several faults across the Bingham Canyon Mine are a potential source for seismic activity.

Figure 5. The two percent probability of exceedance in 50 years at Bingham Canyon Mine is shown. (United States Geological Survey, 2014).

Geotechnical Monitoring

Before the Manefay Slide at the Bingham Canyon Mine occurred on

April 10, 2013, the mine’s geotechnical operations team used the data collected from several slope monitoring instruments to administer a preemptive response, suspending mining activities and clearing all personnel from the open pit. To accurately record millimeter-scale displacements in the rock volume of the northeastern pit wall, Rio

Tinto Kennecott utilized several slope monitoring technologies. These technologies were required to consider the geological setting, climatic conditions, and mining practices at the site (Steiakakis,

2013). In a series of publications issued in 2013, Rio Tinto Kennecott provided details on their “9 Layers of Protection” (Rio Tinto

Kennecott, 2013), including the employee training, strict adherence to

MSHA standards, and state-of-the-art technologies used to monitor and anticipate the occurrence of the Manefay slide.

Trained Eyes Everywhere – The 800+ Rio Tinto Kennecott employees working at the Bingham Canyon Mine are trained to identify safety and health related dangers, such as unstable bench slopes, rock falls, and other geotechnical hazards. Hazards identified by mine personnel are reported directly to a geotechnical supervisor or production control supervisor. A team of geotechnical engineers begins an immediate investigation into the reported incident (Rio Tinto Kennecott, 2013).

Rio Tinto Kennecott’s prescriptive “TRACK” program requires all mine 62

Figure 17. Rio Tinto Kennecott's "TRACK" program stands for T - Think through the task, R - Recognize the hazards, A - Assess the risks, C - Control the hazards, and K - Keep safety first in all tasks.

personnel to file a daily report on any hazards that are identified during their shift. Leading up to the Manefay slide, geotechnical supervisors at the Bingham Canyon Mine were implementing variable levels of a Trigger Action Response Plan (TARP). The TARP informed mine personnel on the level of anticipated risk based on slope deformation models acquired from monitoring equipment.

Regularly Documented Inspections – Different shift supervisors across all areas of mining activity conduct inspections for safety and health related dangers at the beginning of each shift. The shift supervisors are responsible for documenting, as required by the Mining 63

Safety and Health Administration (MSHA), any potential hazards identified during inspections (Rio Tinto Kennecott, 2013).

Prism Networks & Skyboxes – Rio Tinto Kennecott Utilizes 220+

Prism Networks reporting to four Robotic Total Stations (RTS) to compile real-time, slope-stability monitoring. The prisms are scanned hourly and a moving reflector (5-10 Hz) is tracked by the transmitting total station to derive a continuously updated displacement-time or acceleration-time relationship by acting as an electronic extensometer

(Rio Tinto Kennecott, 2013; Walker, 2013; Psimoulis, 2011). Prisms are installed throughout the pit surface of the Bingham Canyon Mine at

100- to 150-m intervals and communicate with total stations that can be 800- to 2,500-m away. Several meteorological sensors automatically record atmospheric conditions to adjust and correct data. Complex software (Leica Geo-systems) generates long-term deformation models through these geodetic and geotechnical sensors (Walker, 2013).

Extensometers – Several wire line cable extensometers were utilized at the Bingham Canyon Mine to measure the displacement across visible tension cracks that developed along the tops of bench slopes

(Rio Tinto Kennecott, 2013).

Time Domain Reflectometry – Downhole Time Domain Reflectometry

(TDR) cable systems are used in the Bingham Canyon Mine to detect and measure displacement along subsurface planes. TDR requires electric signals to be sent down a vertically-buried coaxial cable. If slippage 64 or some other displacement has occurred at a point along the coaxial line, the electric pulse will be reflected off of the discontinuity, recording the severity and depth of the deformation from a comparison to the signal recorded by the undamaged cable (Rio Tinto Kennecott,

2013).

Micro-seismic Array – Micro-seismic “noise” created by subsurface cracking of intact rock is detected by an array of sensitive monitoring equipment (Walker, 2013). Geophones are deployed inside boreholes drilled throughout Bingham Canyon Mine and monitor the arrivals of micro-seismic waves. Complex software calculates the

Figure 18. A Robotic Total Station (RTS) is used to measure long-term deformation of pit wall slopes (Rio Tinto Kennecott, 2013).

65 approximate location and magnitude of the seismic event based on pre- computed, 3-D velocity models for the open pit (Rio Tinto Kennecott,

2013). Application of micro-seismic arrays include monitoring the effects of mining operations and mitigating the likelihood of weakening intact rock or activating geological structures in the rock, monitoring stress changes in slope surfaces with increasing slope angle, and evaluating vibration thresholds of surfaces throughout the open pit to prevent subsequent instability (Walker, 2013).

GroundProbe Slope Stability Radar (SSR) – GroundProbe Slope

Stability Radar (SSR) uses state-of-the-art survey technology to monitor slope instability by comparing successive scans of slope surfaces (Rio Tinto Kennecott, 2013; Walker, 2013). With a high- precision accuracy of 1/10th of an inch and a range up to 10,000-ft,

Rio Tinto Kennecott utilizes five GroundProbe radars to scan slope faces in continuous cycles every four to six minutes (Rio Tinto

Kennecott, 2013). GroundProbe’s advanced analysis software monitors short and long term slope deformation and allows for specific programs that identify hazardous or accelerated surface movement (Rio Tinto

Kennecott, 2013; Walker, 2013). Also, GroundProbe’s sophisticated analysis software includes the Work Area Monitor (WAM) that advises personnel near slope surfaces with hazardous or accelerated movement through visual and audible alerts (Walker, 2013).

The GroundProbe radars are operable in adverse weather, including high humidity, rainfall, and whiteout conditions (Rio Tinto Kennecott,

2013; Walker, 2013). Bingham Canyon utilizes five GroundProbe Radars. 66

Figure 19. GroundProbe Slope Stability Radar (SSR) detects deformation in pit wall slope faces to identify hazardous trends (Kennecott, 2013).

IBIS Slope Stability Radar – IDS’s IBIS Slope Stability Radar systems is a form of ground base interferometric synthetic-aperture radar (GBinSAR). GBinSAR systems are used to measure complex, three- dimensional slope kinematics and landslides in open-pit mines across the world (Atzeni et al., 2015). IBIS systems uses interferometric radar technology to detect slope deformation at a sub-millimeter scale. IBIS is able to analyze slope stability over wide areas without using artificial reflectors. IBIS radar has a high spatial resolution, unlike most slope stability radar systems, can measure longer range working distances, and has a faster acquisition times. Also, because of the type of wavelength used by IBIS, there is a limited impact of 67 atmospheric artifacts (Farina et al., 2011). However, certain factors, like the decorrelation properties of radar waves, limit the ability of

IBIS system. The individual pixels in a radar image are a collection of returning waves that have interacted constructively or destructively. The phase contributions from each wave determine the outcome of the wave read by the IBIS system. Sophisticated algorithms built into IDS’s Guardian software work to evaluate the effective decorrelation of the each pixel to mitigate these effects through image filtering (Atzeni et al., 2015). IBIS systems utilized by

Kennecott have alarm capabilities to alert mine personnel to possible hazards (Rio Tinto Kennecott, 2013). There are four IBIS systems in use at Bingham Canyon Mine.

Figure 20. IBIS uses interferometric waves to measure sub-millimeter scale deformations in slope surfaces. (Kennecott, 2013). 68

GIS Data Display – Rio Tinto Kennecott uses a geographic information system (GIS) data display network to display and integrate information from several monitoring systems across Bingham Canyon

Mine. The GIS data display was customized by the geotechnical staff at

Bingham and captures multiple data sources through wireless technology. The system allows the geotechnical operations to quickly access information from the several monitoring instruments that are connected to the system (Rio Tinto Kennecott, 2014).

Theory

Slope Stability Analysis

The management and design of slopes requires the assessment of geologic, hydrologic, and economic factors, as well as an understanding of slope stability analysis and a recognition for the several challenges and contributory processes that effect the stability of those slopes. In this section, the many factors influencing the development of a comprehensive geologic model that contributes to the stability of a slope, the complex challenges that make the characterization of a rock mass problematic, the different failure mechanisms that can occur in a slope, and the variable methods of slope stability analysis are covered.

Contributory Processes

To develop a comprehensive geologic model that is fundamental to slope stability analysis, an understanding of the several factors that control the stability in those slopes is necessary.

Geology – A detailed understanding of the sequence and characteristics of lithological units within a rock mass and major geologic structures provides information for accurately describing the behavior of slopes (Morgenstern and Martin, 2008). It is essential that information gathered on lithological boundaries, such as bedding planes or other discontinuities, and major geologic structures, such 70

as faults and folds, is accurately collected to contribute to a

geologic model that will be used to analyze the stability of a slope.

The location and extent of second order geologic features, like

alteration and jointing, is also important (Hoek et al., 2000). At the

Bingham Canyon Mine, the copper porphyry deposit is strongly

associated with various alteration zones. The alteration types impact

the geotechnical qualities of a rock mass, ultimately contributing to

slope stability.

Geomorphology – The interplay of physical processes (e.g.

erosion) and chemical processes that affect the condition of slope

Figure 26. The relationship between uniaxial strength and alteration types in different rock masses is shown (Hoek et al, 2000).

71 surfaces and near-surface boundaries plays an important role in slope stability (Morgenstern and Martin, 2008). These processes can change in situ stresses, loading a rock mass beyond its ability to resist an external load. Generally, the upper 100+ meters of elevation in a slope are the most weathered, as they are the first surface exposed to overland flows and other terra forming processes (Meyers, 2012). As the topography of an open pit mine changes with mining activity, blast damage can extend several meters into a rock mass behind the slope face, increasing the likelihood of instability (Hoek et al., 2000).

Similarly, the probability of a potentially unstable structure to daylight out of its topography is likely to increase as the elevation of a slope from the pit floor increases as an open pit mine increases in depth (Meyers, 2012).

Hydrology/Hydrogeology – Groundwater and its ability to distribute through a rock mass contribute significantly to slope stability. Under saturated conditions (below the water table), water pressure acts equally in all directions and is at a pore-size scale in discontinuities in a rock mass. Water pressure reduces external confining stress to a new “effective” stress, which results in a lessening of the rock mass shear strength, or ability to resist sliding, as the shear strength of a discontinuity is directly proportional to the applied confining stress (Hoek et al., 2000;

Sjöberg, 1996). 72

Figure 72. A comparison of the Mohr’s circles for the same slope under fully saturated and unsaturated conditions. The height of the phreatic surface controls the overall effective stress experienced by the rock mass (modified from Sjöberg, 1996).

Groundwater pressure distribution effects the stability of slopes, as the hydrogeological characteristics of a rock mass control the phreatic surface (water table) elevation and water distribution throughout the rock mass. The depth of the water table depends on the slope geometry, drainage efforts, and recharge events surrounding the rock mass. There are several methods to control the effect of water 73 pressure on an open pit mine slope, such as horizontal and vertical drains (Sjöberg, 1996).

Different hydro-mechanical conditions exist at different elevations along a slope surface. At higher elevations, rock masses will be unaffected by pore water pressure, as the exposed surfaces lose more to evaporation and become unsaturated. Pore water pressure will become more important as elevation decreases along a slope surface. Slope surfaces at elevations which experience transient saturated conditions and partial saturation will be most affected by hydrostatic forces, until the water table is reached.

Figure 28. There are several methods to control groundwater distribution (modified from Sjöberg, 1996). 74

Other hydro-mechanical conditions require consideration when performing slope stability analysis. Surface discontinuities, such as fissures, can create ground water seepage into an unsaturated layer influencing the surrounding rock mass. Discontinuities within a rock mass control the flow paths and the patterns of water forces that can weaken a rock mass. The geometry of discontinuities is important as they shape the boundary conditions that effect potential flow paths and gradients (Hoek, 1974). As the prevalence of discontinuities increases throughout a rock mass, the water along the more permeable discontinuity surfaces increase the overall hydraulic conductivity of the rock mass. The increased hydraulic conductivity is detrimental to slope stability because the increased water pressure will lessen the strength of the rock mass (Hoek, 1974).

For these reasons, the development of an accurate groundwater model is an important component to be used in calculating slope stability.

Challenges of Slope Stability

When attempting to solve a slope stability problem, it is important to recognize the challenges that exist when attempting to characterize a rock mass. A rock mass is a complex and anisotropic system with measurable and unmeasurable qualities that contribute to slope stability (Meyers, 2012).

Spatial Variability – Rock mass components have a high degree of spatial variability in their characteristics (Meyers, 2012). Because a rock mass acts anisotropically, meaning its behavior is different in all directions, it is difficult to understand or predict how changes in a rock mass occur. In some cases, failure in a rock mass is behaves as a discontinuum and is controlled by more than one candidate failure surface or mechanism. The failure mechanism is dependent on how a rock mass varies spatially, as the failure will occur along the different weaknesses it encounters.

Figure 29. A failure in a rock mass can occur across several weaknesses and depends on the spatial variability of the rock mass (modified from Hoek et al., 2000). 76

Discontinuities – Geologic structures like faults, fissures, joints, and bedding planes are considered discontinuities in a rock mass and contribute significantly to the stability of slopes (Hoek and

Bray, 1981). Discontinuities are weak links in a rock mass and are subjected to failure when changing stress conditions occur in a rock mass (Sjöberg, 1996). For this reason, the several characteristics of discontinuities, including their inclination, spacing, and persistence, are important to understand.

The inclination of discontinuities greatly affects the stability of slopes. Sliding or toppling of a rock mass occurs when the failure plane that the rock mass is situated against is dipping towards the slope surface and reaches a critical angle, usually between 30˚ and

70˚. As a discontinuity surface daylights, or becomes present through a slope surface, it is more likely that a failure will occur along an inclined discontinuity plane (Hoek and Bray, 1981).

Because of the likelihood of failure along a discontinuity, it is essential to map, predict, or model the occurrence and frequency of discontinuities. Discontinuity or joint mapping is a useful technique that can be of substantial value when performing slope stability analysis. A discontinuity is mapped as a one-dimensional structure having length, although, in reality, joints are two-dimensional planes

(Sjöberg, 1996). Discontinuities are mapped this way because the geometry of a discontinuity is unknowable, and they are simplified for mathematical understanding. In most cases, discontinuity mapping happens at the surface level and is applied to the rock mass as a whole. Stereographic projection of the discontinuity orientations is 77

used to visualize the potential failure planes and joint sets

(Sjöberg, 1996).

Figure 210. The frequency and areal extent of discontinuities increases with the scale at which they are considered (modified from Sjöberg, 1996). 78

The frequency, or spacing, at which discontinuities occur is an important parameter when considering the stability of a slope.

Discontinuities occur at different frequencies depending on the type of discontinuities that are present, the strength properties of the rock mass, and the history of stress events that created the discontinuities. Similarly, persistence is the areal extent to which discontinuities are found. It is important to quantify the persistence of larger discontinuities as they can underlie an unstable, large volume rock mass (Meyers, 2012). In some cases persistence increases the shear strength of a failure plane, as intersecting discontinuities create rock bridges, offsetting a continuous plane and increasing the shear strength of that plane (Watters, 2015).

The scale at which one considers discontinuities in a rock mass is an important factor in slope stability analysis. Discontinuities occur at virtually every scale and include micro-cracks in the crystalline lattice of a mineral to tectonic plate boundaries

(Sjöberg, 1996). It is essential to look at discontinuities in both a gross and a local context, as there are an increasing number of weakness planes present with increasing scale.

Typical Failure Mechanisms

The failure mechanism a slope undergoes is based on the geological and in situ stress state of the rock mass, as well as the several factors detailed above. Structurally controlled failures are the most common type of failure mechanism in slopes, meaning the 79 geologic structure, the pervasiveness of discontinuities, and the geotechnical characterization of a rock mass is the most likely source of failure (Sjöberg, 1996). Instability in an open-pit mine can occur at a bench scale or as an overall pit slope failure (Hoek, 2009).

Slope failures have either a translational slip, rotational slip, or are a toppling failure (Teymur). Translational slip is a product of kinematically free rock mass shearing along one or more structural planes or discontinuities. The shear strength of the rock mass and discontinuity surface is important in these types of failures.

Rotational, or circular, slip occurs in a homogeneous rock mass or a rock mass without a critically oriented structural feature. Rotational failure is the primary failure mechanism in soils. Toppling failures are the successive breaking of rock slope associated with a concentration and subsequent unloading of slope toe stress (Sjöberg,

1996). In the occasion that a discontinuity is oriented parallel to the slope surface, slab or buckling failure can occur.

With the exception of toppling failure, the following failure mechanisms involve simple gravitational driven loading (Hoek, 2009).

Simple Plane Failure – A plane failure occurs when a rock mass is kinematically free to move along a highly geometric discontinuity

(Sjöberg, 1996). Plane failures are considered rare because they require the failure to occur along a highly ordered discontinuity, however, the two-dimensional study of plane failures is the easiest form of slope stability analysis (Hoek, 1974; Hoek and Bray, 1981). 80

To be kinematically feasible, the general conditions that must be true for a plane failure to occur: 1) the failure plane must strike within ±20˚ to parallel of the slope surface, 2) the failure surface must “daylight”, or come out on the slope surface, 3) the dip of the failure plane must be less than the slope surface angle to satisfy the previous condition (Ψf ≤ ΨP), 4) the dip of the failure surface must be greater than the friction angle of the discontinuity surface material

(Ψf ≥ ϕ), and 5) tangential (release) surfaces must provide negligible friction to sliding (Hoek and Bray, 1981). Plane failures are sensitive to changes in the shear strength properties of the materials along the discontinuity and groundwater pressure distribution (Hoek and Bray, 1981).

Plane failures are unlikely in large scale slopes as they would require a continuous geologic structural feature for the rock mass to

Figure 211. A simple plane failure requires a highly geometric discontinuity to shear along (modified from Hoek, 1974). 81

shear along (Sjöberg, 1996). However, this study hypothesizes the

Manefay Slide at Bingham Canyon Mine was a simple plane failure.

Wedge Failure – A wedge failure occurs when a rock mass is

kinematically free to move along an intersection of two highly

geometric discontinuities (Sjöberg, 1996). Unlike simple plane

failures, a wedge failure can only be interpreted in a three-

dimensional analysis. A wedge failure can be analyzed in two-

dimensions by assuming it is a simple plane failure.

To be kinematically feasible, the general conditions that must be

true for a plane failure to occur: 1) the dip of the line of

intersection must be less than the slope surface angle (Ψi ≤ ΨP), 2)

the line of intersection must “daylight” on the slope surface, 3) the

line of intersection must be steep enough to overcome the strength of

the two discontinuity surfaces, and 4) the upper most point of the

Figure 212. A wedge failure requires two highly geometric discontinuity planes to shear against (modified from Sjöberg, 1996).

82 line of intersection must intersect the top of slope or be terminated by a tension crack (Watters, 2015).

Circular Failure – Circular failures occur primarily in soil, highly weathered rock, or rock with densely spaced discontinuities

(Hoek and Bray, 1981). Unlike a simple plane failure or a wedge failure, a circular failure is not controlled by geological features such as faults, joints, bedding planes, or any other discontinuous structures. Instead, circular failures occur in homogeneous materials where a strongly defined geometric structure or critically oriented discontinuities no longer exist (Hoek and Bray, 1981; Sjöberg, 1996).

The failure surface in a circular failure is free to follow the path of least resistance and commonly follows a circular arc. For a circular failure to occur, the size of the individual particles in the

Figure 213. A circular failure occurs in a homogeneous material with no preferential failure paths (modified from Sjöberg, 1996).

83 rock or soil mass must be very small compared to the size of the slope

(Hoek and Bray, 1981).

Toppling Failure – A toppling failure is defined by the successive breakdown of a rock slope. Toppling failure starts when confining stress crushes the toe rock of a slope. After the toe rock is crushed, the load is distributed to adjacent rock which fail under the newly applied stress. Generally, the redistributed load is directed to discontinuities in the rock mass, resulting in secondary modes of failure after the slope toe collapse. Unlike the other failure types, toppling is controlled by the in situ stress conditions rather than the shear strength of the rock mass (Sjöberg, 1996).

During a toppling failure, large columns of rock formed by preexisting, steeply dipping discontinuities in the rock mass will

Figure 214. A toppling failure occurs a successive breaking of large columns of rock (modified from Hoek, 1974). 84 overturn successively. Flexural toppling is possible in bench scale failure conditions (Sjöberg, 1996).

Slab/Buckling Failure – A slab/buckling failure happens when a very thin, continuous bedding plane or joint set oriented parallel to a slope surface buckles as the slope toe collapses. The failure is initiated by slope toe crushing, but can be created by hydrostatic uplift (elevated groundwater) (Sjöberg, 1996).

Figure 15. Buckling is caused by slope toe collapse and the subsequent buckling of a discontinuity parallel to the slope surface.

Mohr-Coulomb Failure Relationship

Most rocks are assumed to behave as Mohr-Coulomb materials during shear failures (Hoek and Bray, 1981). The failure surface on a Mohr-

Coulomb material has a shear strength expressed as: 85

휏 = 푐 + 휎′ tan (∅) (1)

Where τ is shear strength, c is cohesion, σ’ is effective normal stress, and ϕ is friction angle.

The normal stress is mainly the confining pressure experienced in situ. However, the normal stress is reduced when pore water pressure acts on a potential failure surface and becomes the effective normal stress. The cohesion and friction angle of a failure surface are

Figure 16. The cohesion and friction angle of a shearing plane is derived from a Mohr-Coulomb failure envelope (modified from Hoek and Bray, 1981). 86 proportional to the effective normal stress and can be determined from a Mohr-Coulomb failure envelope derived from direct shear testing. A

Mohr-Coulomb failure envelope shows the shear stress required at increasing normal stress to cause sliding along a discontinuity (Hoek and Bray, 1981).

Slope Stability Analysis Methods

There are several methods for analyzing the stability of slopes.

The analysis technique used depends on measurable rock mass properties, site conditions, and the potential mode of failure. Slope stability analysis methods range from simple infinite slope calculations and limit equilibrium techniques to sophisticated finite- elements algorithms (Eberhardt, 2003). For the purpose of this study, only the basic methods of slope stability analysis will be discussed, and will not include probabilistic or numerical modeling.

General Kinematic Analysis – Kinematic methods of slope stability analysis are focused on the feasibility of a failure after a discontinuity “daylights” on a slope surface. Kinematic analysis relies heavily on the detailed evaluation for discontinuity geometry and an understanding for rock mass strength characteristics. This form of analysis is performed using stereographic plots of discontinuity geometries to identify is failure is possible. The stereographic projection will also tell about the mode of failure, i.e. plane, wedge, circular, or toppling. Computer programs can be used to perform 87

kinematic analysis of failures with complicated geometries (Eberhardt,

2003).

Figure 17. Kinematic analysis uses stereonets to determine the primary method of instability (Hoek et al., 2000).

Empirical Design Methods – It is important to consider the

behavior of past slope failure to predict future occurrences.

Empirical slope stability analysis is concerned with designing slopes

with precedent experience in mind. Two examples of empirical design

presented in this study includes the steepness to height relationship

in open-pit slopes by Hoek and Bray (1981) and the prediction of slope

failure by the inverse-velocity method by Rose and Hungr (2006).

Hoek and Bray (1981) use data from the highest and steepest open-

pit mine slopes from across the world to find trends in the

instability of those slopes. The slope height is plotted against the

slope angle, and a dashed line is used to extrapolate points beyond

the data they gathered. Hoek and Bray (1981) show that a higher slope

must have a lower slope angle to maintain stability. However, there

are a limited number of data points from higher slopes. Bingham Canyon 88

Mine is represented as an unstable slope with the highest slope height

(Sjöberg, 1996).

Rose and Hungr (2006) demonstrate the onset of a slope failure can be accurately predicted using the inverse-velocity method developed by Fukuzono in 1985. By recording the displacement rate of a slope surface over time, the inverse of the displacement rate plotted against time shows an approximately linear trend approaching the date

Figure 18. Hoek and Bray (1981) looked at data from open-pit mines to find a trend in instability with slope height (modified from Hoek and Bray, 1981). 89

Figure 19. A typical inverse-velocity plot will show an approximately linear trend approaching the date of failure (Rose and Hungr, 2006). of failure. Different monitoring techniques can be used to measure the displacement. Sources of attributable error from the monitoring instruments, like optical refraction and diurnal effects, can be reduced by taking readings at the same time of day (Rose and Hungr,

2006).

Limit Equilibrium Analysis – Limit equilibrium is the method of choice for routine slope stability analysis (Sjöberg, 1996). Limit equilibrium analysis uses the Mohr-Coulomb determination of rock strength characteristics to apply a force balance until a limited equilibrium condition is reached. A limited equilibrium condition 90 refers to when the resisting forces of cohesion and friction are exactly balanced by the opposing forces of gravity and water pressure

(Hoek, 1974). A simple force balance can be calculated for a sliding rock mass when a few considerations are applied. Limit equilibrium analysis does not consider the deformation of material or the failure of intact rock (Sjöberg, 1996).

The simplest form of limit equilibrium analysis uses the calculation of a factory of safety to determine slope stability. In this form of limit equilibrium analysis, the sum of mobilized, opposing forces (or moments) are expressed as a fraction of the resisting, strength forces (or moments) (Sjöberg, 1996; Aryal, 2006):

∑(푟푒푠푖푠푡푖푛푔 푓표푟푐푒푠 (표푟 푚표푚푒푛푡푠)) 퐹 = (2) ∑(표푝푝표푠푖푛푔 푓표푟푐푒푠 (표푟 푚표푚푒푛푡푠))

Janbu (1973) presents a simplified factory of safety calculation where the mobilized shear stress is expressed as a fraction of the shear strength along a sliding plane. The factory of safety calculation is expressed using the Mohr-Coulomb relationship (Aryal, 2006):

′ 휏푓 푐+ 휎 tan (∅) 퐹 = = (3) 휏 휏

Where F is the factory of safety, τf is the shear strength of the sliding rock mass, τ is the shear stress experience by the rock mass, c is cohesion, σ’ is effective stress, and ϕ is the friction angle. 91

There are several complex methods presented in the field of rock and soil mechanics to calculate a factory of safety for the different failure mechanisms using force equilibriums, moment equilibriums, stress component equilibriums, and combinations of more than of these methods. For the purpose of this study, simplified block models used to determine the factory of safety for a sliding rock mass will be presented. For analysis of the Manefay Slide at Bingham Canyon Mine, a detailed description of the Janbu Simplified Method (JSM) is included in the Methods section of this report.

Translational Failure – Plane and wedge failures are simplified as two-dimensional blocks sliding on inclined planes. It is possible to conduct a three-dimensional limit equilibrium analysis on wedge failures, however, the method is not presented in this report.

Figure 20. A plane failure is modeled as a sliding box under gravitational loading (modified from Hoek and Bray, 1981). 92

An infinitely continuous, plane failure can be modeled as a box under gravitational loading. The normal stress acting against the potential surface is expressed as (Hoek and Bray, 1981):

푊 cos(휓 ) 휎 = 푃 (4) 퐴

Where W is the weight of the block, ΨP is the slope angle, and A is base area of the block. The Mohr-Coulomb relationship applies to the shear strength of the potential failure surface (Hoek and Bray, 1981):

푊 cos(휓 ) 휏 = 푐 + 푃 tan (휙) (5) 퐴

Where τ is the shear strength of the potential failure surface and ϕ is the friction angle. The factory of safety calculation for this simplified plane failure is expressed as:

푐퐴+푊푐표푠(휓 )tan (휙) 퐹 = 푃 (6) 푊푠푖푛(휙)

In the case that the sliding block is separated by a water-filled tension crack, the water pressure will increase linearly with depth.

The water in the tension crack will create a force, V, acting against the rear end of the sliding block. Seepage along the potential failure surface will create hydrostatic uplift, U, acting against the undersurface of the sliding block. 93

Figure 21. Water pressure will affect the outcome of a sliding block model (modified from Hoek and Bray, 1981).

The factor of safety calculation in the case of added water pressure is expressed as (Hoek and Bray, 1981):

푐퐴+(푊푐표푠(휓 )−푈) tan (휙) 퐹 = 푃 (7) 푊푠푖푛(휓푃)+푉

Where U is the uplift force created by the distribution of water pressure across the potential failure surface and V is the force created by the distribution of water pressure in the tension crack.

Figure 22. A sliding block can be stabilized by installing a tensioned rock-bolt (modified from Hoek and Bray, 1981).

An effective method of stabilizing sliding blocks to a potential failure surface is by installing tensioned rock-bolts. The tensioned rock-bolts reinforce a rock mass by counteracting the pressure and uplifting forces of water. The factor of safety calculation for a sliding block with a rock-bolt is expressed as (Hoek and Bray, 1981):

푐퐴+(푊 푐표푠( 휓 )−푈+푇 푠푖푛(훽)) tan (휙) 퐹 = 푃 (8) 푊 푠푖푛( 휓푃)+푉−푇 cos (훽)

Where T is force of the tensioned rock-bolt and β is the angle of the rock-bolt with respect to the potential failure surface. 95

Rotational Failure – Despite being the most common type of failure in soil slopes, limit equilibrium analysis of rotational failures is more complicated than translational failures. Generally, a rock mass on a potential circular failure surface is divided into several rigid blocks bounded by the slope face and failure surface. This is known as the method of slices. There are certain criteria that need to be known to conduct a two-dimensional limit equilibrium analysis of circular failures: 1) the shear and normal forces along the failure path, 2) the interslice (shear and normal) forces acting between each slice, 3) the weight of each slide, and 4) the external conditions effecting the slope, like external loading and groundwater distribution (Sjöberg,

1996.

Figure 23. The method of slices can be used for analyzing a circular failure (modified from Sjöberg, 1996). 96

Open Pit Mine Design & Slope Stability

There has been a declining trend in the grade of most mined ores

since Daniel C. Jackling invented the open pit mining technique at

Bingham Canyon Mine. On average, the grade of metal ores mined at an

economically feasible limit has decreased from a minimum of 15 percent

to less than 10 percent (Meyers, 2012). While technological

advancements have had a tremendous effect on the ability to mine lower

grade ores, the open pit mining technique, as compared to other mining

practices, is the most significant reason behind the exploitation and

profitability of low grade ores.

The reduction of ore grades has followed the heightened demand

for mined resources, as the global population increases. To keep up

with these demands, both the volume of material that has to be mined

and the scale at which open pit operations occur have greatly

increased. Steepening of open pit slopes has the potential to add an

additional five or more years to the mine life of a pit. Similarly, a

1˚ increase in the overall slope angle can free hundreds to millions

Figure 39. How economically feasible is the scale of Bingham Canyon Mine? 97 of dollars-worth of ore (Meyers, 2012). The financial pressure to produce more at a larger scale has forced slope stability engineers to question: how much larger?

Economics of Open Pit Mining

The scale of an open pit mine is determined by the extent and geometry of the ore body, the type of mineralization being mined, and the mining program implemented to exploit the ore body. The planning and design process of an open pit mine includes two major phases: 1) pit optimization and 2) process and production design. The final pit outline, the decision of where to mine based on the economic feasibility of the ore grade, and plans for pit expansion are considered part of the pit optimization phase. Process and production design includes the day-to-day planning and evaluation of mining activity. Mining sequence is determined by the current economic environment and usually requires the continuous update of the pit design (Sjöberg, 1996).

Open pit mining requires the removal of non-profitable waste rock, or overburden. The stripping of waste rock is intrinsic to the open pit mining process, as the geometry of ore bodies seldom coincides with the design of a stable pit slope. The stripping ratio at an open pit mine refers to the ratio between the amount of overburden removed and the ore rock mined at a profit. It is economically advantageous to have a low stripping ratio. During the optimization phase, a favorable, overall stripping ratio might be 98 approximated. However, the day-to-day, instantaneous stripping ratio may be considered economically unfeasible as certain periods of mining activity are strictly for the removal of waste rock (Sjöberg, 1996).

The cutoff grade is the breakeven point where the costs of mining, processing, and marketing of the mined ore are covered by its profitability. The cutoff grade does not consider the stripping of waste rock to reach the ore. Therefore, the stripping ratio that corresponds to the cutoff grade is deemed the cutoff stripping ratio

(Sjöberg, 1996).

The final extent of an open pit mine changes according to the ore grade that is economically feasible to mine given the market demand for the mined resource and time value of money. Several techniques are

Figure 40. The present value of a mined resource plays an important role in determining the extent of an open pit mine. 99 used to include interest and inflation rates, and future revenues and expenditures in the calculations made to determine the final pit limit. However, before any financial analyses are conducted, it is important to study the mechanical behavior of rock slopes as they relate to the intended mining program (Sjöberg, 1996).

Architecture of Open Pit Mining

The economics of an open pit mine are directly controlled by the pit slope geometry. The smallest, managed slope in an open pit mine is referred to as a “bench”, and an operating pit is mined concurrently at several different benches. A collection of mined benches make up the working slope. The geometry of a working slope is determined by the financial benefit of mining in that location and the stability required from the interim slope. Unlike a designed slope commonly found in an urban environment, instability in an interim working slope may be beneficial to reducing future mining costs.

The maximum achievable geometry of a rock slope is governed by the mechanical stability of the rock mass and the corresponding, intermediate geometries used to make up the overall slope. A pit slope is made of benches, inter-ramp areas (catch benches), and haulage roads. The inter-ramp angle is a function of the bench angle, the bench height, and the width of the bench. The overall slope angle is defined by the inter-ramp angle and the number of access roads crossing the slope. Varying slope geometries may be required in different locations throughout the mine, as changing conditions in the 100 geologic structure, groundwater distribution, presence of discontinuities, climate, seismicity, and geomechanical properties of the rock mass occur (Sjöberg, 1996).

Figure 41. The geometry of an open pit slope is controlled by a number of factors, including economics and geomechanics.

101

Methods

Slide by Rocscience, Inc.

Slide will be used to conduct a limit equilibrium analysis of the planar slope failure that led to the Manefay Slide at Bingham Canyon

Mine. Slide is a two-dimensional, limit equilibrium analysis program developed by Rocscience, Inc. Slide is used to model and analyze both natural and engineered slopes, including soil and rock slopes, embankments, earthen dams, and retaining walls. Slide’s graphical interface (CAD) capabilities allow the user to create and edit complex, user-defined circular and non-circular failure surfaces. The program uses a number of widely used limit equilibrium analysis methods to calculate a factor of safety for circular and non-circular failure surfaces (Rocscience).

Table 1. The types of limit equilibrium and their corresponding applications are shown (modified from Aryal, 2006).

Methods Circular Non-Circ. ΣM = 0 ΣF = 0 Assumptions for T§ and E# Ordinary X - X - Neglects both E and T Bishop (Simp.) X X X (†) Considers E, but neglects T Janbu (Simp.) X X - X Considers E, but neglects T Janbu GPS X X (*) X Considers both Lowe-Karafiath - X - X Inclines at, θ = 1/2(α+β)

Corps of Engrs. - X - X Inclines at, θ = 1/2(α1+α2) Sarma X X X X Interslice shear, T = ch+Etanφ Spencer X X X X Constant Incline, T=tanθ E Morgenst-Price X X X X f(x), T=f(x)λE * Satisfies moment equilibrium for intermediate thin slices † Satisfies vertical force equilibrium for base normal force §Interslice shear forces #Interslice normal forces 102

The general goals integrated into Slide as they relate to the analysis of the Manefay Slide include: 1) the assessment of the stability of slopes under specified conditions, 2) the evaluation of the possibilities of the failure of slopes, and 3) sensitivity analyses for evaluating the influence of variations in critical parameters such as geometry, material properties, and groundwater conditions on the stability of slopes (Rocscience).

For the purpose of this study, the Janbu Simplified limit equilibrium Method (JSM) will be used to analyze the Manefay Slide.

Janbu (1954) uses a simplified method for non-circular failure

Figure 42. The Janbu Simplified Method (JSM) uses several vertical slices to calculate a factor of safety (Aryal, 2006).

103 surfaces where a sliding rock mass is divided into several vertical slices. In JSM, both the horizontal and vertical force equilibriums are satisfied, but the moment equilibrium is not satisfied. A factor of safety is determined by conducting a horizontal force equilibrium, so JSM does not consider the interslice shear forces between the vertical slices. The factor of safety calculation for the Janbu

Simplified Method (1954):

∑ (푐 푙+(푁−푢 푙) tan(∅′)) sec (훼) 퐹 = (9) 푓 ∑ 푊 tan(훼)+ ∑ ∆퐸

Where α is the slope angle, c is the cohesion, u is the pore water pressure, ϕ is the friction angle, l is the length of the slice, W is the weight of the slice, ΣΔE = E2 –E1 = net interslice normal forces, and N is the base normal force:

1 푐′푙 sin(훼) 푁 = ∑(푊 − − 푢푙 cos (훼)) (10) 푚훼 퐹

Where,

tan(∅′) 푚 = cos(훼) (1 + tan(훼) ) (11) 훼 퐹

Where F is the dynamic force.

104

Janbu’s (1954) original limit equilibrium calculations were constructed in stress terms (Aryal, 2006). The JSM in stress terms:

푏 (푐′+(푝−푢) tan(∅′) ∑{ } 푛 퐹 = 훼 (12) 표 ∑ 푝 푏 tan (훼)

Where p = W/b, b is the width of each slice, and nα:

tan(∅′) 푛 = 푐표푠2(훼(1 + tan(훼) ) (13) 훼 퐹

Determining Rock Mass Properties

To calculate a factory of safety using the JSM, Slide requires the user to define certain inputs including the slope geometry, an approximate location of the phreatic surface, and rock mass properties, like cohesion (c) and friction angle (ϕ). To determine the rock mass properties that controlled the Manefay Slide at Bingham

Canyon Mine, a shear test will be performed on several representative samples of the different rock types that occur along the slide plane.

The tests will be done using a hand-operated, direct shear apparatus. At different normal loads, a rock sample will be sheared to generate a displacement vs shear stress plot. The peak and residual shear values from each resulting graph will be identified and plotted against the tested normal stresses. A Mohr-Coulomb failure envelope will be made by applying a linear best fit line to the plot of normal 105

and shear stresses. The cohesion and friction angle will be calculated

from the resulting line equation. The cohesion and friction angle

values will be compared to those presented in Styles et al. (2011) to

ensure the rock strength properties returned from the shear tests are

roughly close to those determined by others.

Slide Testing

Slide tests will be performed using the rock mass properties

determined by the shear tests. Trials will be conducted by simulating

different fault plane scenarios (bottom layer-top layer): 1)

quartzite-quartzite contact, 2) quartzite-argillic clay contact, 3)

quartzite-altered limestone (limestone rock strength properties from

Styles et al. (2011)), 4) argillic clay-quartzite contact, 5) failure

plane in argillic clay, and 6) failure plane in altered limestone.

Different trials will be performed under each scenario by varying the

Figure 43. Slide tests will be performed at three different failure slope angles. 106 cohesion values for each rock mass within a ± 10-15 lb/ft2 range. The friction angles will be kept static throughout the tests.

The variable slide plane contacts and cohesion value tests will be performed for three failure slope angles: 1) 15˚, 2) 20˚, and 3)

25˚. It is unclear how the slide plane dipped into the open-pit, as the bedding planes dip roughly north. The apparent dip into the pit was estimated to be 17˚-20˚ using Lanier’s (1975) Geologic Map of the

Bingham Mine.

Finally, tests will be performed under both unsaturated and partially saturated conditions. The phreatic surface will be modelled in to Slide so that it is roughly parallel to the slope surface, but set back approximately ±200-300 ft. Also, under the conditions applied before the addition of the water table, the upper 100, 200, and 300 feet of overburden will be removed. A total of 180 Slide trials will be performed using these criteria.

Figure 44. The phreatic surface will be modeled in to Slide and will roughly follow the open-pit surface. 107

Analysis

The Manefay Slide at Bingham Canyon Mine was analyzed using

Slide, a two-dimensional, limit equilibrium analysis program by

Rocscience, Inc. The testing procedure is outlined in the Methods

section of this report. Recommendations as to mitigating future slope

failure are included in the Discussion section of this report.

The Manefay Slide occurred along the Manefay bedding planes,

three or four thin layers of arenaceous limestone and calcareous

sandstone. The Manefay beds are surrounded by quartzite and calcareous

quartzite (Lanier et al., 1978). To find the rock strength properties

of the quartzite, a direct shear test was performed on a

Mohr-Coulomb Failure Envelope: Quartzite

Peak Shear Stress Residual Shear Stress Linear (Peak Shear Stress) Linear (Residual Shear Stress) 700

600

500 (psi) τ 400

300

Shear Stress, ShearStress, 200

100

0 0 100 200 300 400 500 600 700 800 900 Normal Stress, σ' (psi)

Figure 45. A Mohr-Coulomb failure envelope was derived from the peak and residual shear strengths to find the strength properties quartzite. 108 representative sample of quartzite collected near the base of the

Manefay bedding planes at Bingham Canyon Mine. The peak and residual shear strengths were plotted against normal stress to derive a Mohr-

Coulomb failure envelope. It was determined from the direct shear tests that the representative sample of quartzite has a cohesion, c, of approximately 81 lb/in2 and a friction angle, ϕ, of 35˚. Tabled results from the shear tests are in Appendix A of this report.

Lanier et al. (1978) postulate that argillic alteration in sedimentary rocks is probably centralized on bedding planes. For this reason, a second laboratory shear test was conducted to find the

Mohr-Coulomb Failure Envelope: Quartzite-Argillic Clay

Residual Strength Values Linear (Residual Strength Values)

700

600

500 (psi)

τ 400

300 Shear Stress, ShearStress,

200

100

0 0 100 200 300 400 500 600 700 800 900 1000 Normal Stress, σ' (psi)

Figure 46. A Mohr-Coulomb failure envelope was derived for the specimen simulating argillic alteration 109 strength properties of the argillic bedding. A representative clay specimen, likely containing kaolinite and montmorillonite, was tested on the quartzite shear block. The clay specimen was brought to full saturation before testing to simulate the condition an argillically altered bedding plane would experience at a point below the phreatic surface. Shear tests on the representative sample of argillic clays determined the clay specimen to have no cohesion and a friction angle of 22.7˚. Tabled results from the shear tests performed are in

Appendix A of this report.

Styles et al. (2011) presents rock strength values of quartzite found throughout the pit at Bingham Canyon Mine. The numbers reported in Styles et al. (2011) are roughly the same as the rock strength properties determined by the laboratory shear test performed on the representative sample of quartzite. The friction angle was found to be the same in both tests, while the cohesion calculated by Styles et al.

(2011) is 10 lb/in2 more than the value returned by the laboratory shear tests conducted by the authors of this study. Styles et al.

Table 2. The strength properties from Styles et al. 2011.

Clay Altered Parameter Monzonite Limestone Quartzite Monzonite Limestone

Density, ρ (g/cm3) 2.88 2.64 2.8 2.78 2.6

Cohesion, c (lb/in2) 261.1 85.6 420.7 154.5 94.3

Friction Angle, φ (˚) 37 35 40 32 35 110

(2011) also include strength properties for “altered limestone”, a possible rock type found on the Manefay beds.

Using the rock strength properties from the laboratory shear tests and Styles et al. (2011), 180 limit equilibrium analyses were done in Slide. Analyses were performed for several failure plane scenarios: 1) “failure plane in quartzite along a preexisting discontinuity”, 2) “argillic clay sliding on quartzite”, 3) “quartzite sliding on argillic clay”, 4) “failure plane in saturated argillic clay”, 5) “altered limestone sliding on quartzite”, and 6) “failure

Figure 47. Trials were performed for a “preexisting discontinuity in quartzite”. 111

plane in altered limestone along a preexisting discontinuity”. The

cohesion values for each rock type were varied to find the limited

equilibrium condition in each case. In half of the trials, the water

table was elevated to create partially saturated conditions along the

failure plane, as the water table was made to follow the slope

surface. Each trial was performed at a failure plane angle of 15˚,

20˚, and 25˚. In a few trials, the rock strength properties were kept

constant while the upper 100, 200, and 300 feet of waste rock were

removed from the slope crest.

Figure 248. The phreatic surface was modeled into half of the trials to create partially saturated conditions across the failure surface. 112

The factor of safety returned from each trial was ranked according to the risk of slope instability it conveyed:

Factor of Safety: Factor of Safety: Factor of Safety: Factor of Safety: ≥ 1.500 1.250 - 1.499 1.000 - 1.249 < 1.000

Figure 49. The factor of safety values were ranked according the risk of slope instability.

Most of the trials returned a factor of safety well above the recommended safety limit of 1.3 to 1.5 (Watters, 2015).

The trials done with a “preexisting discontinuity in quartzite” were run as a control variable to observe how the addition of new

Limit Equilibrium Analysis: Quartzite-Quartzite Contact 4

2 Factor of Safety of Factor 1

0 65 70 75 80 85 90 95 Cohesion (lb/in2)

Q-Q 15 DRY Q-Q 20 DRY Q-Q 25 DRY Q-Q 15 SAT Q-Q 20 SAT Q-Q 25 SAT

Figure 50. Trials done with a “preexisting failure plane in quartzite” were all above the recommended factor of safety of 1.3 to 1.5. 113 materials effected the outcome of the limit equilibrium analysis. The quartzite-quartzite contact trials, for both unsaturated and partially saturated conditions, were well above a reasonable factor of safety.

Similarly, the limit equilibrium analyses done on the “argillic clay sliding on quartzite”, the “altered limestone sliding on quartzite”, and the “altered limestone with a preexisting failure plane” had a factor of safety well above the recommended limit.

For all trials, increasing the slope of the failure plane and adding the water table had the effect of reducing the factor of safety by approximately 0.15 to 0.5. Saturating the failure plane had less of an effect than increasing the slope of the failure surface. Increasing

Average Factor of Safety for Five Cohesion Values (Bottom Contact/Top Contact/Dry or Saturated) 4.5

3.5

2.5

FactorSafety of 1.5

0.5

Trial

15˚ 20˚ 25˚

Figure 51. The average factor of safety for five cohesion values, under dry and partially saturated conditions. 114 the slope of the failure surface from 15˚ to 20˚ had more of an effect than increasing it from 20˚ to 25˚.

Removing the upper 100, 200, and 300 feet of the slope had little effect on the factor of safety. It is likely the conditions modeled into slide did not accurately reflect those of the waste rock on top of the slope at Bingham Canyon Mine. Also, the upper slope removal trials were done independently of the other controlled variables, meaning the limited factor of safety reduction could be a product of another variable overriding the effects of the slope removal.

The “failure plane in argillic clay” and the “quartzite sliding on argillic clay” failure plane scenarios were the only trials to return factors of safety under the recommended limit. The trials had

Limit Equilibrium Analysis: Failure Surface in Argillic Clay 2

1.5

1 Factor of Safety of Factor 0.5

0 0 5 10 15 20 25 Cohesion (lb/in2)

AC-AC 15 DRY AC-AC 20 DRY AC-AC 25 DRY AC-AC 15 SAT AC-AC 20 SAT AC-AC 25 SAT

Figure 52. Slope instability occurred in several trials along a “failure surface in saturated argillic clay”. 115 factors of safety in the 1.000 to 1.500 range for a dry and partially saturated failure slope of 15˚. For a dry failure slope of 20˚, at least one trial of each failure plane scenario returned a factor of safety below 1.000. For a dry failure slope of 25˚, four of the five cohesion trials of each failure plane scenario had a factor of safety below 1.000. Adding the phreatic surface reduced the average factor of safety and increased the number of slopes with a factor of safety below 1.000.

The most unstable slope scenario was the “quartzite sliding on argillic clay” with a failure slope of 20˚ or more.

Limit Equilibrium Analysis: Quartzite Sliding on Argillic Clay 2

1.5

1 Factor of Safety of Factor 0.5

0 0 5 10 15 20 25 Cohesion (lb/in2)

AC-Q 15 DRY AC-Q 20 DRY AC-Q 25 DRY AC-Q 15 SAT AC-Q 20 SAT AC-Q 25 SAT

Figure 53. Slope instability occurred in several trials of “quartzite sliding on argillic clay”.

116

Discussion

It was determined in the Analysis section of this report that the most unstable slope conditions were when a failure plane developed in argillic clay centralized on a bedding plane, or when a quartzite block slid along an argillically altered bedding plane. The slope became increasingly unstable as the angle of the failure surface increased and with the addition of the water table.

From these results, it is likely the Manefay Slide happened when the argillically centralized Manefay bedding fault became mobilized.

The limit equilibrium trials performed in Slide confirm that an unstable slope could develop along a dry failure surface, but would more likely occur along a saturated failure surface. The mine operations team at Bingham Canyon Mine communicated that they drain the slopes of the open pit below the toe scarp of the slide area. It is likely the excess water pressure was introduced as water seeped through a tension fracture developed at the top of the slope and progressively flowed across the failure plane.

Climate records from the National Oceanic and Atmospheric

Administration (NOAA) show areas near Bingham Canyon Mine received upwards of 3.0 inches of rain on April 9, 2013, the day before the

Manefay Slide. Considering the elevation of Bingham Canyon Mine, the diurnal temperature variation is likely to cause stress fluctuations 117

Figure 54. Areas near Bingham Canyon Mine received upwards of 3.0 inches of precipitation on April 9, 2013, the day before the Manefay Slide. in the rock mass near the toe scarp, weakening the base of the failure surface.

A number of questions are left to be considered. The sedimentary units at Bingham Canyon Mine dip roughly north. Lanier’s (1975)

Geologic Map of Bingham Mine was used to approximate the apparent dip at which the Manefay beds enter the pit. However, the pit outline presented by Lanier (1975) is considerably less extensive than the current pit limit, meaning the dip indicators provided on the map may not be representative. This point is relevant because the mobilized rock mass in the Manefay Slide did not move along a failure surface that dipped toward the slope surface. The sliding rock mass curved into the pit from a failure surface striking approximately 30˚ to 45˚ from the strike of the pit slope surface. 118

Side Scarp

Figure 55. The Manefay failure surface did not dip in the same direction as the slope surface, and left a sizable side scarp.

In a traditional plane failure mechanism, the resisting forces provided by the scarps in line with the sliding plane are neglected.

In the case of the Manefay Slide, a notably large scarp is parallel to the sliding direction. It seems unreasonable that the resisting force provided by this side scarp should be neglected. While the limit equilibrium analysis performed in Slide used a number of assumptions, it becomes clear that it may not serve the best use in analyzing a unique failure condition, like the Manefay Slide.

119

Recommendations for Mitigating Future Slope Failure

Drainage of Slope – The mine operations team at Bingham Canyon

Mine shared that the regional water table is typically drained below the toe scarp of the Manefay failure surface. It is possible that local, hydrogeologic conditions exist in parts of the mine that allow the water table to extend beyond the level of the toe scarp.

Continuously monitoring drainage efforts in vertical drainage wells is important as mining activity alters the hydrogeological conditions of near surface layers. Blast fracturing can extend into the rock mass behind a bench and alter water flow conditions beyond the near surface layers. Pumping rates, depth of pumping, and number of vertical drainage wells can be optimized as the hydrogeological environment changes. In addition to vertical drainage wells, horizontal drains drilled into a slope surface can improve drainage efforts, although they can be sheared in the event that a slope is moving. A more expensive alternative is a drainage gallery. Drainage galleries are excavated tunnels that have radially extending drain pipes (Sjöberg, 1996).

While there is no exact solution to slope stabilization through the drainage of slopes, it has shown to be an effective means of slope stabilization, especially of large failure structures (Sjöberg, 1996).

Unload Overlying Waste Rock – The stability of slope surfaces at

Bingham Canyon Mine may be improved by removing the waste rock at the top of the pit slopes. While the weight of the overlying rock will 120 increase the normal stress of rock below, it does not weigh as much as the intact rock. The added weight of the waste rock is negligible compared to the effect of keeping it as part of the overall slope. As a slope increases in elevation, there are more opportunities for failure.

Reinforced Slopes – Rock support and reinforcements are becoming increasingly more popular to provide slope stabilization in open pit mines. Rock bolts, wire mesh, buttresses, and engineered retaining walls are examples of reinforcements techniques used to mitigate certain failure types. The likelihood of these methods preventing slope failure decreases as the scale of a slope increases (Sjöberg,

1996). At Bingham Canyon Mine, the stability of slopes at bench scale, or even inter-ramp scale, can be improved implementing one or more of these methods.

Continued Monitoring – Improving the methods of slope stability monitoring applied and the number of instruments tracking slope deformation can be advantageous to mitigating slope failure at Bingham

Canyon Mine. Improved in field surveys looking for tension cracks along slope crests will increase the likelihood that seepage along a failure surface is prevented before a critical stress condition is reached.

Mine Out Moving Zone – Upon detection of a moving slope face, a surface can be mined out to correct the slope geometry to a more 121 stable condition. The overall slope angle can be decreased or the individual bench dimensions can be altered to provide slope stabilization.

Improved Slope Stability Analysis Methods – The Manefay Slide demonstrates the inability of two-dimensional, limit equilibrium methods in analyzing a unique slope stability problem. To better mitigate slope failure at Bingham Canyon Mine, a move toward more advanced three-dimensional modeling techniques, numerical modeling methods, and the use of finite-element, shear strength reduction factor methods (SSR) to find a strength reduction factor (SRF) can be used.

Run Out Prediction: Volume-Fahrböschung Relationship

An empirical method commonly used in predicting run out distance is based on the relationship between the volume of material that makes up a landslide and the fahrböschung, or the angle of reach. A fahrböschung is the angle formed by a line connecting the crest of a source slope and the toe of a slide deposit. A fahrböschung is considered similar to the average friction angle, but is not the exact same, and is also a measure of the efficiency of energy dissipation. A lower fahrböschung is associated with a longer run out distance

(Clague and Stead, 2012). 122

Figure 56. The geotechnical operations at Bingham Canyon Mine grossly underestimated the extent of run out from the Manefay Slide.

The geotechnical operations team at Bingham Canyon Mine may have estimated the fahrböschung to find the run out distance from volume calculations of the rock mass over the failure surface. In reality, it is tremendously complicated to approximate the run out distance of a landslide. Consideration of the rock strength properties, slope geometry, slope failure mechanism, and several other factors needs to be included in the analysis method used to determine the run out distance.

123

Future Work

There are a number of ways in which the limit equilibrium analyses performed in this study can be improved. Additional limit equilibrium trials can be done including more variability in the rock strength properties. A better understanding as to where the regional water table is located will help more accurately model the hydrogeologic conditions along the failure plane. While an attempt to model the slope geometry in Slide was made, exact slope specifications as they are along the slope surface perpendicular to the failure plane will improve results. Implementation of other limit equilibrium methods may better the results of the trials performed. Finally, exploration into using advanced analysis techniques, like finite- element or finite-difference models, may improve the results of this study.

124

References

Arrington, L.J. and Hansen, G.B., 1963, The Richest Hole on Earth: A History of the Bingham Copper Mine: Logan, Utah State University Press, 103 p.

Aryal, K.P., 2006, Slope Stability Evaluations by Limit Equilibrium and Finite Element Methods [Ph.D. thesis]: Norwegian University of Science and Technology, 124 p.

Atenzi, C., Barla, M., Pieraccini, M., Antolini, F., 2015, Early Warning Monitoring of Natural and Engineering Slopes with Ground- Based Synthetic-Aperture Radar: Rock Mechanics and Rock Engineering, v. 48, p. 235-246.

Babcock, R.C., Ballantyne, G.H., and Phillips, C.H., 1997, Summary of the Geology of the Bingham District, Utah, in, John, D.A. and Ballantyne, G.H., eds., Geology and Ore Deposits of the Oquirrh and Wasatch Mountains, Utah: Society of Economic Geologists Guidebook Series, v. 29, p. 159-178.

Bray, R.E., Lanier, G., John, E.C., 1975, General Geology of the Open- Pit Mine, in Guidebook to the Bingham Mining District: Society of Economic Geologists, p. 49-58.

Carter, R.A., 2014, Recovering from Bingham Canyon’s Record-Setting 2013 Slide: Engineering & Mining Journal, v. 215, no. 5, p. 48- 49.

Christensen, J., 1994, Can a Copper Firm Restore a Blasted Ecosystem?: High Country News: http://www.hcn.org/issues/12/368/ (accessed February 2015).

Clague, J.J., Stead, D., 2012, Landslides: Types, Mechanisms and Modeling, New York City, Cambridge University Press, 296 p.

Davidson, L., 1997, Pollution Report Slams Kennecott over Groundwater: Deseret News: http://www.deseretnews.com/article/583351/Pollution-report-slams- Kennecott-over-groundwater.html?pg=all (accessed February 2015).

Eberhardt, E., 2003, Rock Slope Stability Analysis – Utilization of Advanced Numerical Techniques: http://www.eos.ubc.ca/personal/erik/e-papers/EE- SlopeStabilityAnalysis.pdf

U.S. Environmental Protection Agency, 2015, Kennecott South Zone/Bingham NPL Site: http://www2.epa.gov/sites/production/files/2015- 02/documents/kennecott-south-site-maps-1-22-2015.pd 125

U.S. Environmental Protection Agency, 2006, Abandon Mine Lands Case Study: Kennecott Mining Site: http://www.epa.gov/aml/tech/kennecott.pdf (accessed February 2015).

Farina, P., Leoni, L., Babboni, F., Coppi, F., Mayer, L., Ricci, P., 2011, IBIS-M, an Innovative Radar for Monitoring Slopes in Open- Pit Mines, in Proceedings, Slope Stability 2011: International Symposium on Rock Slope Stability in Open Pit Mining and Civil Engineering: Vancouver, Canada, Sept. 18-21, 2011.

Farina, P. et al., 2014, Recent Advances in Slope Monitoring Radar for Open-pit Mines, in Proceedings, Mine Closure Solutions: Ouro Preto, Minas Gerais, Brazil, April 26-30.

Gestring, B., 2011, Factsheet on Rio Tinto’s Bingham Canyon Copper- Gold-Silver-Molybdenum Mine in Utah: http://www.earthworksaction.org/files/publications/FS_Problems_Bi nghamCanyon_2011_low.pdf (accessed February 2015).

Goin, P. and C.E. Raymond, 2004, World’s Biggest Hole: Creating and Re-creating Bingham Canyon, Utah, in Changing Mines in America: Chicago, University of Chicago Press, p. 97-116.

Gorrell, M., 2013, Kennecott Slide Causes Drop in Sales Tax Revenue: The Salt Lake City Tribune, September 11.

Gorrell, M., Fahys, J., 2013, Kennecott Announces Layoffs after Bingham Canyon Slide: The Salt Lake City Tribune, May 3.

Hoek, E., Londe, P., 1974, The Design of Rock Slopes and Foundations, in proceedings, Third Congress of the International Society of Rock Mechanics: Denver, Colorado, Sept. 1974, 78 p.

Hoek, E., Bray, J.W., 1981, Rock Slope Engineering: Taylor & Francis, New York, New York, 361 p.

Hoek, E., Read, J., Karzulovic, A., Chen Z.Y., 2000, Rock Slopes in Civil and Mining Engineering, in proceedings, International Conference on Geotechnical and Geological Engineering, GeoEng2000: Melbourne, Australia, 19-24 Nov. 2000, 16 p.

Hoek, E., 2009, Fundamentals of Slope Design, in proceedings, Slope Stability 2009: Santiago, Chile, 9-11 Nov. 2009, 26 p.

Janbu, N., 1954, Stability analysis of Slopes with Dimensionless Parameters. Thesis for the Doctor of Science in the Field of Civil Engineering, Harvard University Soil Mechanics Series, No. 46.

Janbu, N., 1973, Slope Stability Computations, in Embankment Dan Engineering, Casagrande Volume, p. 47-86. 126

John, E.C., 1975, Mineral Zones of the Bingham District, in Guidebook to the Bingham Mining District: Society of Economic Geologists, p. 59-72.

Kosich, D., 2013, Inside Bingham Canyon – N. American Mining’s Largest, Most Successful Landslide: http://www.mineweb.com

Layton, S., 2014, This is Utah: http://historytogo.utah.gov/facts/brief_history/

Lanier, G., compiler, 1975, Geologic Map of the Bingham Mine, Guidebook to the Bingham Mining District, Plate 2, scale 1:4,800.

Lanier, G., John, E.C., Swensen, A.J., Reid, J., Bard, C.E., Caddey, S.W., Wilson, J.C., 1978, General Geology of the Bingham Mine, Bingham Canyon, Utah: Economic Geology, v. 73, p. 1228-1241.

Meyers, A., 2012, Fundamental Techniques for Reducing Risk Associated with Instabilities in Mining Slopes: I.J.M.G.E. University of Tehran, v. 46, no. 1, p.43-55.

Morgenstern, N.R., Matin, C.D., 2008, Lanslides: Seeing the Ground, in Chen et al. (eds), Landslides and Engineered Slopes: Taylor & Francis Group, London, p. 3-23.

Notarianni, P.F., 2014, Mining: http://historytogo.utah.gov/utah_chapters/mining_and_railroads/mi ning.html

Pankow, K.L., Moore, J., Hale, J.M., Koper, K.D., Kubacki, T., Whidden, K.M., McCarter, M.K., 2014, Massive Landslide at Utah Copper Mine Generates Wealth of Geophysical Data: GSA Today, v. 24, no. 1, p 4-9.

Pankow, K.L., Moore, J., 2014, Mine Landslide Triggered Quakes: http://www.unews.utah.edu/news_releases/mine-landslide-triggered- quakes/

Petley, D., 2013, Is the Binghcam Canyon Copper Mine Landslide the Most Expensive Single Mass Movement in History?: http://blogs.agu.org/landslideblog

Phillips, C.H., Smith, T.W., and Harrison, E.D., 1997, Alteration, Metal Zoning, and Ore Controls in the Bingham Canyon Porphyry Deposits, Utah, in, John, D.A. and Ballantyne, G.H., eds., Geology and Ore Deposits of the Oquirrh and Wasatch Mountains, Utah: Society of Economic Geologists Guidebook Series, v. 29, p. 136-149.

Powell, A.K., 1992, Utah Territory, in Encyclopedia of Mormonism: New York, Macmillan, p. 1503-1506.

127

Psimoulis, P. and Stiros, S., 2011, Robotic Theodolites (RTS): http://www.gim-international.com/issues/articles/id1693- Robotic_Theodolites_RTS.html

Rio Tinto Kennecott Utah Copper, 2013a, Bingham Canyon Mine Slide Fact Sheet:http://riotintokennecott.com/sites/kennecott.com/files/slid e_fact_sheet_fnl4_15_13_315pm.pdf (accessed February 2015).

Rio Tinto Kennecott Utah Copper, 2013, Bingham Canyon Mine Monitoring System: 9 Layers of Protection: http://riotintokennecott.com/sites/kennecott.com/files/9layers_pr otection_bcmmonitoringsystem.pdf (accessed February 2015).

Rocscience, Slide (4.0): White Paper, http://rocscience.com/assets/files/uploads/7772.pdf

Rodd, R., and Thatcher, L., 2014, Brief History of Utah: Mines & Minorities:http://historytogo.utah.gov/facts/brief_history/minesa ndminorities.html

Romboy, D., 2013, Kennecott Landslide Impacts Global Copper Industry, Utah Economy: Deseret News, April 15.

Rose, N.D. and Hungr, O., 2006, Forecasting Potential Rock Slope Failure in Open Pit Mines Using the Invserse-Velocity Method: International Journal of Rock Mechanics & Mining Sciences, v. 44, p. 308-320.

Sjöberg, J., 1996, Large Scale Slope Stability in Open Pit Mining – A Review: Technical Report 1996:10T, Lulea University of Technology, 229 p.

Smith, W.H., 1975, General Structural Geology of the Bingham Mining District, in Guidebook to the Bingham Mining District: Society of Economic Geologists, p. 41-48.

Smith, W.H., 1975, A Short History of the Bingham Canyon Mining District, in Guidebook to the Bingham Mining District: Society of Economic Geologists, p. 1-15.

Steiakakis, C., 2013, Slope Failures, Landslides and Mines: http://blog.geotechpedia.com/index.php/2013/04/slope-failures- landslides-and-mines/

Styles, T.D., Stead, D., Eberhardt, E., Rabus, B., Gaida, M., Bloom, J., 2011, Integrated Numerical Modelling and Insar Monitoring of a Slow Moving Slope Instability at Bingham Canyon Mine, in Proceedings, Slope Stability 2011: International Symposium on Rock Slope Stability in Open Pit Mining and Civil Engineering, Vancouver, Canada, September 18-21, 2011.

128

Swensen, A.J., compiler, 1975, Geologic Map of the Bingham District, Guidebook to the Bingham Mining District, Plate 1, scale 1:24,000.

Swensen, A.J., 1975, Sedimentary and Igneous Rocks of the Bingham District, in Guidebook to the Bingham Mining District: Society of Economic Geologist, p. 21-40

Teymur, Slope Stability Presentation: http://web.itu.edu.tr/~teymurb/fe1lecture3.pdf

Tooker, E.W. and Roberts, R.J., 1988, Preliminary Geologic Map, Cross- Sections, and Explanation Pamphlet for the Bingham Canyon 7 ½- Minute Quadrangle, Salt Lake and Tooele Counties, Utah: U.S. Geological Survey Open-File Report 88-699, 33 p.

Tooker, E.W., 1999, Geology of the Oquirrh Mountains, Utah: U.S. Geological Survey Open-File Report 99-571, 150 p.

United States Geological Survey, 2014, 2014 National Seismic Hazard Map:http://earthquake.usgs.gov/hazards/products/conterminous/2014 /2014pga2pct.pdf

University of Utah News, 2014, Mine Landslide Triggered Quakes: http://unew.utah.edu/news_releases/mine-landslide-triggered- quakes/ (accessed February 2015).

Wallace, J. and Lowe, M., 2009, Ground-Water Quality Classification for the Principal Basin-Fill Aquifer, Salt Lake Valley, Salt Lake County, Utah: Utah Geological Survey Open-File Report 560, 12 p., http://files.geology.utah.gov/online/ofr/ofr-560/ofr-560.pdf (accessed February 2015).

Walker, S., 2013, Careful Monitoring: The Key to Pit-wall Safety: http://www.e-mj.com/features/2836-careful-monitoring-the-key-to- pit-wall-safety.html#.VRuCBvnF9ps (accessed February 2015).

Watters, B., 2015, GE 487 Slope Stability Analysis Presentations, University of Nevada, Reno.

Whitehead, B.D. and Rampton R.E., 2006, Bingham Canyon, in Whitley, C., ed., From the Ground Up: The History of Mining in Utah: Logan, Utah State University Press, p. 220-249.

129

Appendix A

Tabled Data from Direct Shear Tests

Table 3. Quartzite: 500 psi Normal Load.

Shear Stress, Displacement, Cumulative Disp., Strain, τ δ δT ε (psi) (in) (in)

135 0.006 0.006 0.120% 157.5 0.014 0.014 0.280% 180 0.035 0.035 0.700% 180 0.073 0.073 1.460% 180 0.107 0.107 2.140%

Table 4. Quartzite: 1000 psi Normal Load.

Shear Stress, Displacement, Cumulative Disp., Strain, τ δ δT ε (psi) (in) (in) 180 0 0.107 0.000% 202.5 0 0.107 0.000% 225 0 0.107 0.000% 247.5 0.001 0.108 0.020% 270 0.005 0.112 0.100% 292.5 0.017 0.124 0.340% 292.5 0.026 0.133 0.520% 315 0.047 0.154 0.940% 315 0.067 0.174 1.340%

130

Table 5. Quartzite: 1500 psi Normal Load.

Shear Stress, Displacement, Cumulative Disp., Strain, τ δ δT ε (psi) (in) (in) 337.5 0.002 0.176 0.040% 360 0.002 0.176 0.040% 382.5 0.003 0.177 0.060% 405 0.01 0.184 0.200% 405 0.041 0.215 0.820% 405 0.068 0.242 1.360%

Table 6. Quartzite: 2000 psi Normal Load.

Shear Stress, Displacement, Cumulative Disp., Strain, τ δ δT ε (psi) (in) (in)

450 0 0.242 0.000% 472.5 0.001 0.243 0.020% 495 0.008 0.250 0.160% 495 0.027 0.269 0.540% 506.25 0.064 0.306 1.280%

Table 7. Quartzite: 2500 psi Normal Load.

Shear Stress, Displacement, Cumulative Disp., Strain, τ δ δT ε (psi) (in) (in) 562.5 0.002 0.308 0.040% 585 0.004 0.310 0.080% 607.5 0.033 0.339 0.660% 630 0.062 0.368 1.240% 630 0.115 0.421 2.300%

131

Table 8. Selected peak and residual strength values from quartzite shear test.

Normal, σ' Peak, Ƭ Residual, Ƭ (psi) (psi) (psi) 157 182 180 314 317 315 471 407 405 628 506 495 785 632 630

Table 9. Argillic Clay-Quartzite: 500 psi Normal Load.

Cumulative Disp., Shear Stress, τ Displacement, δ δT Strain, ε (psi) (in) (in) 45 0.155 0.155 3.100%

Table 10. Argillic Clay-Quartzite: 1000 psi Normal Load.

Cumulative Disp., Shear Stress, τ Displacement, δ δT Strain, ε (psi) (in) (in) 135 0.050 0.205 1.000%

Table 11. Argillic Clay-Quartzite: 1500 psi Normal Load.

Cumulative Disp., Shear Stress, τ Displacement, δ δT Strain, ε (psi) (in) (in) 180 0.142 0.347 2.840%

132

Table 12. Argillic Clay-Quartzite: 2000 psi Normal Load.

Cumulative Disp., Shear Stress, τ Displacement, δ δT Strain, ε (psi) (in) (in) 247.5 0.053 0.400 1.060%

Table 13. Argillic Clay-Quartzite: 2500 psi Normal Load.

Cumulative Disp., Shear Stress, τ Displacement, δ δT Strain, ε (psi) (in) (in) 315.0 0.024 0.424 0.480%

Table 14. Argillic Clay-Quartzite: 3000 psi Normal Load.

Cumulative Disp., Shear Stress, τ Displacement, δ δT Strain, ε (psi) (in) (in) 382.5 0.014 0.438 0.280%

133

Appendix B

Tabled Data from Slide

Table 2. Failure Plane in Quartzite; Failure Plane: 15˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 70 10080 35 70 10080 35 3.301 75 10800 35 75 10800 35 3.352 80 11520 35 80 11520 35 3.403 85 12240 35 85 12240 35 3.453 90 12960 35 90 12960 35 3.504

Table 3. Argillic Clay Sliding on Quartzite; Failure Plane: 15˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 0 0 20 3.645 80 11520 35 5 720 20 3.645 80 11520 35 10 1440 20 3.645 80 11520 35 15 2160 20 3.645 80 11520 35 20 2880 20 3.645

Table 17. Failure Plane in Argillic Clay; Failure Plane: 15˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 0 0 20 1.348 5 720 20 5 720 20 1.400 10 1440 20 10 1440 20 1.452 15 2160 20 15 2160 20 1.504 20 2880 20 20 2880 20 1.556

134

Table 4. Altered Limestone Sliding on Quartzite; Failure Plane: 15˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 150 21600 32 3.427 80 11520 35 155 22320 32 3.427 80 11520 35 160 23040 32 3.427 80 11520 35 165 23760 32 3.427 80 11520 35 170 24480 32 3.427

Table 5. Failure Plane in Altered Limestone; Failure Plane: 15˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle

150 21600 32 150 21600 32 3.838 155 22320 32 155 22320 32 3.888 160 23040 32 160 23040 32 3.939 165 23760 32 165 23760 32 3.990 170 24480 32 170 24480 32 4.041

Table 20. Quartzite Sliding on Argillic Clay; Failure Plane: 15˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 80 11520 35 1.348 5 720 20 80 11520 35 1.398 10 1440 20 80 11520 35 1.449 15 2160 20 80 11520 35 1.500 20 2880 20 80 11520 35 1.550

135

Table 21. Failure Plane in Quartzite; Failure Plane: 15˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 70 10080 35 70 10080 35 3.018 75 10800 35 75 10800 35 3.069 80 11520 35 80 11520 35 3.120 85 12240 35 85 12240 35 3.170 90 12960 35 90 12960 35 3.221

Table 6. Argillic Clay Sliding on Quartzite; Failure Plane: 15˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 0 0 20 3.277 80 11520 35 5 720 20 3.277 80 11520 35 10 1440 20 3.277 80 11520 35 15 2160 20 3.277 80 11520 35 20 2880 20 3.277

Table 7. Failure Plane in Argillic Clay; Failure Plane: 15˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 0 0 20 1.197 5 720 20 5 720 20 1.249 10 1440 20 10 1440 20 1.301 15 2160 20 15 2160 20 1.353 20 2880 20 20 2880 20 1.405

136

Table 8. Altered Limestone Sliding on Quartzite; Failure Plane: 15˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 150 21600 32 3.136 80 11520 35 155 22320 32 3.136 80 11520 35 160 23040 32 3.136 80 11520 35 165 23760 32 3.136 80 11520 35 170 24480 32 3.136

Table 9. Failure Plane in Altered Limestone; Failure Plane: 15˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 150 21600 32 150 21600 32 3.584 155 22320 32 155 22320 32 3.635 160 23040 32 160 23040 32 3.686 165 23760 32 165 23760 32 3.737 170 24480 32 170 24480 32 3.787

Table 10. Quartzite Sliding on Argillic Clay; Failure Plane: 15˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 80 11520 35 1.040 5 720 20 80 11520 35 1.090 10 1440 20 80 11520 35 1.141 15 2160 20 80 11520 35 1.192 20 2880 20 80 11520 35 1.242

137

Table 11. Failure Plane in Quartzite; Failure Plane: 20˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 70 10080 35 70 10080 35 2.649 75 10800 35 75 10800 35 2.702 80 11520 35 80 11520 35 2.756 85 12240 35 85 12240 35 2.809 90 12960 35 90 12960 35 2.862

Table 12. Argillic Clay Sliding on Quartzite; Failure Plane: 20˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 0 0 20 3.15 80 11520 35 5 720 20 3.15 80 11520 35 10 1440 20 3.15 80 11520 35 15 2160 20 3.15 80 11520 35 20 2880 20 3.15

Table 13. Failure Plane in Argillic Clay; Failure Plane: 20˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 0 0 20 0.989 5 720 20 5 720 20 1.045 10 1440 20 10 1440 20 1.100 15 2160 20 15 2160 20 1.156 20 2880 20 20 2880 20 1.212

138

Table 14. Altered Limestone Sliding on Quartzite; Failure Plane: 20˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 150 21600 32 2.892 80 11520 35 155 22320 32 2.892 80 11520 35 160 23040 32 2.892 80 11520 35 165 23760 32 2.892 80 11520 35 170 24480 32 2.892

Table 31. Failure Plane in Altered Limestone; Failure Plane: 20˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 150 21600 32 150 21600 32 3.305 155 22320 32 155 22320 32 3.359 160 23040 32 160 23040 32 3.412 165 23760 32 165 23760 32 3.466 170 24480 32 170 24480 32 3.519

Table 32. Quartzite Sliding on Argillic Clay; Failure Plane: 20˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 80 11520 35 0.989 5 720 20 80 11520 35 1.043 10 1440 20 80 11520 35 1.096 15 2160 20 80 11520 35 1.149 20 2880 20 80 11520 35 1.202

139

Table 15. Failure Plane in Quartzite; Failure Plane: 20˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 70 10080 35 70 10080 35 2.447 75 10800 35 75 10800 35 2.500 80 11520 35 80 11520 35 2.553 85 12240 35 85 12240 35 2.607 90 12960 35 90 12960 35 2.660

Table 34. Argillic Clay Sliding on Quartzite; Failure Plane: 20˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 0 0 20 2.748 80 11520 35 5 720 20 2.748 80 11520 35 10 1440 20 2.748 80 11520 35 15 2160 20 2.748 80 11520 35 20 2880 20 2.748

Table 16. Failure Plane in Argillic Clay; Failure Plane: 20˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 0 0 20 0.880 5 720 20 5 720 20 0.935 10 1440 20 10 1440 20 0.991 15 2160 20 15 2160 20 1.046 20 2880 20 20 2880 20 1.102

140

Table 17. Altered Limestone Sliding on Quartzite; Failure Plane: 20˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 150 21600 32 2.573 80 11520 35 155 22320 32 2.573 80 11520 35 160 23040 32 2.573 80 11520 35 165 23760 32 2.573 80 11520 35 170 24480 32 2.573

Table 18. Failure Plane in Altered Limestone; Failure Plane: 20˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 150 21600 32 150 21600 32 3.124 155 22320 32 155 22320 32 3.177 160 23040 32 160 23040 32 3.231 165 23760 32 165 23760 32 3.285 170 24480 32 170 24480 32 3.338

Table 19. Quartzite Sliding on Argillic Clay; Failure Plane: 20˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 0 0 20 0.841 80 11520 35 5 720 20 0.894 80 11520 35 10 1440 20 0.947 80 11520 35 15 2160 20 1.000 80 11520 35 20 2880 20 1.054

141

Table 20. Failure Plane in Quartzite; Failure Plane: 25˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 70 10080 35 70 10080 35 2.491 75 10800 35 75 10800 35 2.563 80 11520 35 80 11520 35 2.635 85 12240 35 85 12240 35 2.707 90 12960 35 90 12960 35 2.779

Table 40. Argillic Clay Sliding on Quartzite; Failure Plane: 25˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 0 0 20 2.981 80 11520 35 5 720 20 2.981 80 11520 35 10 1440 20 2.981 80 11520 35 15 2160 20 2.981 80 11520 35 20 2880 20 2.981

Table 21. Failure Plane in Argillic Clay; Failure Plane: 25˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 0 0 20 0.77 5 720 20 5 720 20 0.846 10 1440 20 10 1440 20 0.923 15 2160 20 15 2160 20 0.999 20 2880 20 20 2880 20 1.075

142

Table 22. Altered Limestone Sliding on Quartzite; Failure Plane: 25˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 150 21600 32 2.670 80 11520 35 155 22320 32 2.670 80 11520 35 160 23040 32 2.670 80 11520 35 165 23760 32 2.670 80 11520 35 170 24480 32 2.670

Table 23. Failure Plane in Altered Limestone; Failure Plane: 25˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 150 21600 32 150 21600 32 3.501 155 22320 32 155 22320 32 3.574 160 23040 32 160 23040 32 3.646 165 23760 32 165 23760 32 3.719 170 24480 32 170 24480 32 3.791

Table 24. Quartzite Sliding on Argillic Clay; Failure Plane: 25˚. DRY.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 80 11520 35 0.770 5 720 20 80 11520 35 0.842 10 1440 20 80 11520 35 0.914 15 2160 20 80 11520 35 0.986 20 2880 20 80 11520 35 1.058

143

Table 25. Failure Plane in Quartzite; Failure Plane: 25˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 70 10080 35 70 10080 35 2.286 75 10800 35 75 10800 35 2.358 80 11520 35 80 11520 35 2.430 85 12240 35 85 12240 35 2.502 90 12960 35 90 12960 35 2.574

Table 26. Argillic Clay Sliding on Quartzite; Failure Plane: 25˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 0 0 20 2.714 80 11520 35 5 720 20 2.714 80 11520 35 10 1440 20 2.714 80 11520 35 15 2160 20 2.714 80 11520 35 20 2880 20 2.714

Table 27. Failure Plane in Argillic Clay; Failure Plane: 25˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 0 0 20 0.657 5 720 20 5 720 20 0.733 10 1440 20 10 1440 20 0.810 15 2160 20 15 2160 20 0.886 20 2880 20 20 2880 20 0.962

144

Table 48. Altered Limestone Sliding on Quartzite; Failure Plane: 25˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 80 11520 35 150 21600 32 2.459 80 11520 35 155 22320 32 2.459 80 11520 35 160 23040 32 2.459 80 11520 35 165 23760 32 2.459 80 11520 35 170 24480 32 2.459

Table 28. Failure Plane in Altered Limestone; Failure Plane: 25˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 150 21600 32 150 21600 32 3.316 155 22320 32 155 22320 32 3.389 160 23040 32 160 23040 32 3.462 165 23760 32 165 23760 32 3.534 170 24480 32 170 24480 32 3.607

Table 29. Quartzite Sliding on Argillic Clay; Failure Plane: 25˚. SAT.

Material Contact #1 Material Contact #2 cohesion cohesion friction cohesion cohesion friction Factory of Safety (psi) (psf) angle (psi) (psf) angle 0 0 20 80 11520 35 0.663 5 720 20 80 11520 35 0.735 10 1440 20 80 11520 35 0.807 15 2160 20 80 11520 35 0.880 20 2880 20 80 11520 35 0.952

145

Appendix C

Slope Geometry Input for Slide

X Y 6215.37 2950 4210.337 1700 0 500 6155.37 2950 4150.337 1700 0 0 6135.169 2900 4130.136 1650 15000 0 6075.169 2900 4070.136 1650 15000 4100 6054.967 2850 4049.934 1600 8000 4100 5994.967 2850 3989.934 1600 7979.799 4050 5974.766 2800 3969.733 1550 7919.799 4050 5914.766 2800 3909.733 1550 7899.597 4000 5894.565 2750 3889.532 1500 7839.597 4000 5834.565 2750 3829.532 1500 7819.396 3950 5814.363 2700 3809.331 1450 7759.396 3950 5754.363 2700 3749.331 1450 7739.195 3900 5734.162 2650 3729.129 1400 7679.195 3900 5674.162 2650 3669.129 1400 7658.993 3850 5653.961 2600 3648.928 1350 7598.993 3850 5593.961 2600 3588.928 1350 7578.792 3800 5573.759 2550 3568.727 1300 7518.792 3800 5513.759 2550 3508.727 1300 7498.591 3750 5493.558 2500 3488.525 1250 7438.591 3750 5433.558 2500 3428.525 1250 7418.39 3700 5413.357 2450 3408.324 1200 7358.39 3700 5353.357 2450 3348.324 1200 7338.188 3650 5333.155 2400 3328.123 1150 7278.188 3650 5273.155 2400 3268.123 1150 7257.987 3600 5252.954 2350 3247.921 1100 7197.987 3600 5192.954 2350 3187.921 1100 7177.786 3550 5172.753 2300 3167.72 1050 7117.786 3550 5112.753 2300 3107.72 1050 7097.584 3500 5092.551 2250 3087.519 1000 7037.584 3500 5032.551 2250 3027.519 1000 7017.383 3450 5012.35 2200 3007.317 950 6957.383 3450 4952.35 2200 2947.317 950 6937.182 3400 4932.149 2150 2927.116 900 6877.182 3400 4872.149 2150 2867.116 900 6856.98 3350 4851.948 2100 2846.915 850 6796.98 3350 4791.948 2100 2786.915 850 6776.779 3300 4771.746 2050 2766.713 800 146

6716.779 3300 4711.746 2050 2706.713 800 6696.578 3250 4691.545 2000 2686.512 750 6636.578 3250 4631.545 2000 2626.512 750 6616.376 3200 4611.344 1950 2606.311 700 6556.376 3200 4551.344 1950 2546.311 700 6536.175 3150 4531.142 1900 2526.11 650 6476.175 3150 4471.142 1900 2466.11 650 6455.974 3100 4450.941 1850 2445.908 600 6395.974 3100 4390.941 1850 2385.908 600 6375.772 3050 4370.74 1800 2365.707 550 6315.772 3050 4310.74 1800 2305.707 550 6295.571 3000 4290.538 1750 2285.506 500 6235.571 3000 4230.538 1750 2225.506 500