2011G - Geotechnical Slope Remediation Via RC-drilling And “Controlled … · 2019-05-12 · Geotechnical Slope Remediation Via RC-drilling And “Controlled Slope Failure Blasting”

Geotechnical Slope Remediation Via RC-drilling And “Controlled Slope Failure Blasting”

Tyler Acorn

Abstract

In December of 2009 an estimated one million ton slope failure activated a 10 million ton mass

impacting use of the main ramp system in the Newmont Gold Quarry pit. The failure left behind a 400

foot (122 meters), vertical to sub vertical scarp feature and shut down production in the lower benches of

the pit.

The Geotechnical group recommended that no equipment should go within an 80 foot (24 meter) safety

zone of the scarp crest due to stability concerns. This meant that the traditional approach of load/haul

could not be used to remediate the slope failure. Following careful analysis, the drill and blast group was

then tasked with finalizing the scarp feature remediation planning. The key considerations determining

the viability of blasting for stabilizing the crest were:

Safety – Ability for equipment and personnel to perform remediation activities while

remaining outside of the 80 foot (24 meters) safety zone

Material type – The rockmass is an alluvial material that lends itself well towards this

style of blasting

Crest and face topography – Face profiles provided evidence that lent itself towards a

blasting approach

Drill Selection – Exploration RC drill rigs were selected due to their ability to drill angled

holes down to the depth required and at the recommended hole diameter

A drilling and blasting plan was successfully implemented, improving the stability of the crest by

performing a series of controlled small scale slope failures, dropping the slope back to a near 45 degree

slope face. The plan utilized seven separate blasts. Each blast involved angled holes, up to 45 degrees

and 230 feet (70 meters) long, drilled from outside the safety zone. This paper describes the methods

used to design each shot, and the challenges found when implementing these plans by reviewing results

from over a period of 4 months.

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Copyright © 2011 International Society of Explosives Engineers

2011G - Geotechnical Slope Remediation Via RC-drilling And •Controlled Slope Failure BlastingŽ

1.0 Introduction: On December 24

th, 2009, a portion of the scarp associated with the Gold Quarry Nine Points Slide,

which had culminated on April 26th

, 2009, rapidly accelerated toward failure. The rapid-brittle behavior

exhibited by the Carlin Formation was a mechanism that had not been previously observed; previous

instability was characterized by time-dependent strain-softening material deformation. The result was a

failure that contained approximately 1.25-1.5 million tons, it created a 400-foot (122 meter) tall failure

scarp 125-feet (38 meters) toward the east. Although this particular failure was only of moderate size,

the sudden change in the loading condition on the Original Nine Points Failure mass caused it to

mobilize and run-out several hundred feet into the bottom of the pit.

Figure 1: Picture 1 is a view of the failure from across the pit while Picture 2 is a view from below the

failure scarp

The Carlin Formation can be summarized as four main sections. The bedrock contact is characterized

by zones of basal gravel units intermixed and overlain by medium to high-plasticity clayey silts and

swelling clays. Clay-altered tuffaceous units, with silty sand and gravel lenses were deposited upon the

basal gravels and basal clays. The middle units are predominately composed of interbedded siltstone

and sandstone. The upper layers consist of variably calcite-cemented sand and gravel debris flows.

There is an overlying alluvium/colluvium layer in undisturbed sections of the Carlin Formation. (Harlen

et al. 1999; Regnier 1960). During deposition of the Carlin Formation, Basin and Range normal faulting

activated north-south striking normal faults and reactivated most pre-existing NNE-NNW and E-W

faults in the region. The reactivation of E-W faults during Basin and Range activity has been found to

be important for imparting a preferential fabric on the Carlin Formation. The failure scarp, which

required special blasting techniques to remediate, developed within the upper variably cemented sands

while being partially defined by fault structures. 1

The rapid brittle behavior in the 9-Points Slide and the 400 feet (122 meter) high failure scarp forced us

to reevaluate our normal slide remediation techniques. The normal technique for remediating the failure

zones in the Carlin Formation utilized track dozer’s. within the area of concern that pushed the slide

material to shovels set up on more competent ground. As a safety consideration our Geotechnical group

recommended that we set an 80 foot (24 meter) exclusion zone behind the failure scarp inside which no

equipment would be allowed.

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Figure 2: 80 ft. exclusion zone

The 9-Points Slide shut down the only access into the bottom of the pit. Production for this mine was

just under 60 million tons of material mined per year, with an annual gold production of 300,000 ounces

of gold. There was one other access into the upper portion of the east side of the pit, however this ramp

system was never intended for accessing the main ore body of the pit and could not be realigned for such

work. The location of the 9-Points Slide and the nature of the Carlin Formation also prohibited us from

re-establishing access to the main ramp system. It was determined that working below the failure zone

posed a safety hazard. The safest way to remediate the area of concern was from the top down.

Therefore a method needed to be established for stabilizing the crest of the failure scarp that would

permit it to be mined out with typical mining equipment.

It was determined that blasting would be a viable alternative as a way to improve the stability of the

crest of the 9-Points Scarp. The key considerations for this were: 2

Safe drilling and loading procedures – all drilling loading was performed outside the 80 ft. (24

meter) exclusion zone

Material type – the rockmass consisted primarily of alluvium material

Crest and face topography – face profiles were used to develop drill depths, angles, and azimuths

Drill selection – exploration drill rigs were selected due to the angle and depth requirements

Drill diameter – 9.875 inch (25 cm) diameter holes were used for the initial blast and then were

switched to 7.875 (20 cm) inch holes for the following blasts

Explosive selection – the rockmass was dry therefore ANFO and Heavy-ANFO were used

Initiation selection – Electronic detonators were used to provide vibration control

2.0 Theory Behind the “Controlled Slope Failure” Blasting The first step in the blast design process was to the break the crest into individual blast zones with

common blast baselines. These baselines were oriented parallel to the crest allowing the blastholes to be

drilled perpendicular to the crest and the baseline.

A survey of the scarp face was then performed using a laser scanner to acquire an updated topo. Sections

were drawn at 25 foot (7.62 meter) intervals and profiles were plotted to provide markers for the drill

and blast design. The following blast parameters were used as a guideline for the first blast:

A “no charge” zone was offset from the face of the scarp. This was initially set at 20 charge

diameters based on the strength of the alluvium in the Carlin Formation. This was to limit the

potential of flyrock from the face and to maintain confinement.

Two explosive decks were used to minimize vibration amplitude.

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A minimum of 14 charge diameters was used to separate the charges and prevent premature

detonations.

Two stem heights were evaluated. 25 feet (7.62 meter) and 30 feet (9.1 meter).

The initial design used an overall energy factor of 140 kcal/t (0.35lb/t) with the lower half of the

face around 250 kcal/t 3

Using the 80 foot (24 meter) exclusion zone and the “no charge” zone as guides the angle and depth of

the holes were determined.

Figure 3: Determine “no charge” zone, angle and depth of blast hole

Using these guidelines the profiles of each section were analyzed. First the area of the overall zone that

would be displaced was calculated. Then since most of the displacement would occur in the lower half

of the blast hole the area of the bottom half was also calculated.

Figure 4: Calculate area of overall wedge and the area of the bottom half of the blast hole

The remainder of the blast parameters were then used to determine the spacing of the holes and the

charge diameter. An Excel sheet was utilized for calculating the blast design.

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Figure 5: Blast design worksheet for Shot One

4

Design Choices:

Blasting Product – Reasons for ANFO

1. Water was not found in any of the holes

2. There was some concern about getting the product to flow down a -61 degree hole (horizontal

being 0 degrees). The looser aspect of ANFO versus a Heavy ANFO or an emulsion was

predicted to flow easier and be less likely to plug up.

3. We could achieve the desired energy factors we wanted with ANFO

Charge Diameter – 9.875 inches (25 cm)

1. The exploration rigs were pre-tooled for 9 .875 inch (25 cm) holes

2. It allowed us to spread out the spacing to 22 feet (6.7 meters)

Delay Intervals – 40 ms between holes and 20 ms between explosive decks

To protect the slope from excessive damage it was imperative that blast vibration amplitudes and

frequencies be controlled. The natural frequency of the Carlin formation has a narrow range

between 3 and 5 Hz. Accordingly, it is not possible to effectively shift blast energy to higher,

less potentially damaging frequencies. As a result, the blast vibrations were primarily controlled

by using relatively long delay intervals (40 ms) between blastholes and employing two separate

charges per hole with 20 ms of separation.

Charge Weight – higher charge weight in bottom charge vs. top charge

The intent was not to cast the material as in a “cast blast” but to create a failure plane and force

the material to fail. The geotechnical group suggested sending as little as possible energy into the

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already failed mass below the blasts. One potential concern was that we would reactivate the

material below the blasts causing it push out into the pit even more.

The bottom charge was intended to create the greatest displacement of material. The intent was

to “kick out” the toe of the wedge thereby destabilizing the top half of the wedge and forcing it

to fail

The top charge was meant to create a crack between the holes (similar to what you would do

with a pre-split design) and create a failure plain for the wedge to fail along. Secondarily it

would also give the top half of the wedge a “nudge” encouraging it to fail.

Initiation – Electronic Detonators

To help control the blast vibration amplitudes the accuracy of electronic detonators was

preferred.

To rectify any issues with hole cutoffs during the loading of the blast holes. In addition two caps

and two boosters per charge deck were utilized.

3.0 Drilling Shot One

Accuracy in the drilling of the blastholes was important and therefore tightly controlled. Using standard

survey procedures and equipment the blasthole collars were laid out in the field. Each blasthole collar

was staked with the hole number, azimuth, drill hole angle, and drill hole depth labeled on the stake.

Figure 6: Survey Techniques. Note: survey stake, and orange painted lines for the azimuth offsets

The pit surveyors then painted an orange offset line four feet to the right of each blast hole collar and

parallel to the design azimuth. This allowed the drillers to line up the tires of their RC Rigs on the

painted line ensuring the holes were drilled along the correct azimuth.

At a mid point during the drilling of each drill hole a down the hole survey was conducted. This was

then plotted in AutoCAD and analyzed to determine if any changes needed to be made to the drill depth.

The key parameter analyzed was drill hole deviation towards the face or away from the face. Hole

deviations towards the face prompted recalculation of the target depth to curtail the drillers from

punching the drill hole out of the face. If the holes deviated away from the face, depth of the hole was

increased to get as close to the “no charge” zone as possible for the first shot. After each hole was drilled

to completion a “end hole” survey was conducted to provide us with an as-built of each hole. This

allowed a more accurate determination of the wedge area that each blast hole was intending to dislodge

and also to ensure as-drilled blast holes separation to arrest premature detonation. The spacing of the

drill holes for Shot One were far enough apart that we did not see any issues with drill hole deviation

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causing the drill holes to get to close together. However in some of the subsequent shots we did see

issues with hole deviation.

Figure 7: Down the hole Survey

4.0 Blast design and loading Shot One

“End hole” surveys were conducted for each drill hole utilizing down the hole survey techniques. This

data was then plotted in AutoCAD and as-built profiles were generated and compared to the original

blast design. This was analyzed to see if any changes needed to be made to the original design. Key

design features that were analyzed for change were:

Hole deviation – If any of the holes deviated too close to the face of the scarp thus entering the

“no charge” zone, they were backfilled to an appropriate depth. Drill holes deviating too close to

each other (within 14 charge diameters) were watched for but never occurred.

As-built areas for the overall wedges of each blast hole and the bottom half wedges of each blast

hole were determined and used to tweak the charge weights for each hole.

Figure 8: Designed profile for hole one in shot one (blue line is the “no charge” zone boundary)

Figure 9: As-built profile for hole 1 in shot 1

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As shown in Figures 8-9 hole 1 was designed at a -60 degrees but was drilled at a –64 degrees. This

changed the depth of the hole, and the areas of the wedges that this hole was “trying to push”. This was

all taken into account in the final blast design.

Since 10-foot (3 meter) steel surface casing was used when drilling the blast holes in shot one, each steel

casing was lined with a pvc pipe to guard against static charge.

Figure 10: 6 inch PVC pipe guarding against static charge.

After each hole was loaded the steel surface casing was then removed to prevent blast projectiles. This

was done after the each hole was loaded to ensure the top of the holes did not cave in before they were

loaded.

Figure 11: Pulling the steel surface casing

The detonator cords were also stowed in a PVC pipe and covered with a plastic bag to protect against

static charge and from the cutting the cords.

Figure 12: PVC pipe and plastic bag protecting the Detonator cords.

Seismographs were installed in five locations to record the blast vibrations. Four of the locations were

arrayed behind the scarp towards a crusher at roughly 200 feet (60 meters), 400 feet (122 meters), 700

feet (213 meters), and 1700 feet (518 meters). This was to record vibrations along the direction that the

9-Points failure had propagated and the direction it was predicted to propagate if it did fail again. They

were spaced at regular intervals to gather both near-field and far-field vibration attenuation data. A fifth

seismograph was located below the scarp near the location of a proposed buttress at roughly 1300 feet

(396 meters).

5.0 Shot One Performance

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Shot one performed as predicted. The crest broke back to the baseline with no back break behind it. The

face of the scarp was laid to the angle of the blast holes (-60 to -65 degrees). To monitor shot one the

following equipment was used:

5 seismographs

High speed video camera

Normal speed video camera

Digital photo camera

Pre and post blast survey’s utilizing laser scans for topographic images

Below is a sequence of images showing the actual blast.

Figure 13: shot one detonation sequence

The desired effect can be seen by looking at pre and post blast plan views and profile views

Figure 14: Topographic plan view showing pre-blast and post-blast

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Figure 15: Face profiles. Green line represents pre-blast profile, red line represents post-blast, and solid

black line represents drill hole survey.

Penetration rate and hole deviation data was analyzed for shot one. This provided insight into the

efficiency and accuracy of drilling with RC rigs.

The penetration rate for 9.875 inch (25 cm) holes was 44.1 ft/hr (13.4 m/hr).

Surface casing with steel casing was 9.7 ft/hr (3 m/hr)

Average deviation – Inclination – 1.424 degrees

Average deviation – Azimuth – 1.417 degrees

6.0 Conclusions from first three shots 6.1 Shot One The performance of shot number one exceeded expectations. As seen in the blast sequence images in

Figure #13, the toe of the shot blew out an excessive amount. Therefore the conclusion was made that

the 140 kcal/ton overall energy level, as well as the 240 kcal/ton energy level in the lower section could

be scaled back without endangering the blast performance.

At -60 to -65 degrees the augured ANFO loaded without any problem. The initial concerns of air-blocks

occurring proved insubstantial. Therefore it was recommended that blast holes be drilled at shallower

angles. With more shallow angled holes, it would allow us to stabilize the crest without exposing

drilling crews working directly on the crest of the scarp.

The 30 foot (9.1 meters) crushed rock stemming contained the energy better then the 25 foot crushed

rock. It is recommended that future blasts using 9.875 inch (25 cm) charge diameter holes be stemmed

with 30 foot of crushed rock.

The penetration rate and drill hole deviation for 9.875 inch (25 cm) holes indicated some potential gains

by using 7.875 inch (20 cm) holes. The 7.875 inch (20 cm) drill holes utilized the same drill rod

diameter but with a smaller drill bit. This would allow for higher penetration rates. The smaller sized

hole would also allow for less flex between the drill rods and the side walls which should produce less

deviation thus providing more accurate blast holes.

6.1 Shot Two For Shot two the charge diameter was decreased to 7.875 inches (20 cm) and the spacing to 22 feet (6.7

meters) for holes 1-6. The drill hole angles were similar to angles in shot one. Holes 7-13 went through

an area that had originally been a valley that was previously filled in with loose unconsolidated material.

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Figure 16: Fill Material as seen in the face of the Scarp

Because of the known issues of drilling through loose, unconsolidated material it was decided to use

symetrix casing to line the holes. This would allow the hole to be drilled, then push 6 inch (15.2 cm)

PVC pipe down the hole inside the symetrix casing, and then finally, pull the casing out. This cased hole

could then be loaded with explosives without fear of collapsing. Because the charge diameter dropped

down to 6 inches (15.2 cm) the spacing was also decreased to 12 feet (3.7 meters). The last couple of

holes in this shot were drilled at shallower angles to test performance. The shallowest angle tested was -

50 degrees. The down-hole survey also showed that a section of the holes had a toe burden (“no charge”

zone) of 27 feet (8.2 meters) and 35 feet (10.7 meters). This tested the theory that the “no charge” zone

could be increased to 41 charge diameters.

Shot two performed similarly to shot one. There was no noticeable adverse effect from the shallower

angled holes through the fill material. The toe was successfully destabilized by kicking it out with the

bottom charge and a failure plane was created by propagating a crack between the drill holes. This test

shot created a “controlled slope failure.”

Figure 17: Shot 2 detonation Sequence. Notice blast hole traces in the fourth image

No back break occurred behind the baseline.

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Figure 18: Post blast view of the baseline.

6.2 Shot Three

Shot two’s performance showed that shallower angle drill holes and an increased toe burden could yield

adequate results in achieving the design criteria. Therefore shot three tested those conclusions further.

Nine holes were drilled at a spacing of 13 feet (3.9 meters), and an angle of -45 degrees (the shallowest

that the RC rigs could drill), with toe burdens (“no charge” zones) ranging from 56 to 110 times the

charge diameter. Four extra holes were drilled with the larger toe burden designs to test where the

confinement would become too much to kick out the toe. Heavy ANFO was also utilized in this shot to

test how well it could be loaded in the shallower holes and to increase the energy factor. The energy

factor for this shot ranged from 149 kcal/ton to 226 kcal/ton in the bottom half and 131 kcal/ton to 159

kcal/ton overall.

This shot also performed as predicted, and met the design criteria goals. By destabilizing the toe and

creating a failure plane a “controlled slope failure” occurred.

Figure 19: Shot 3 detonation sequence

Evaluating the fourth image in figure #19 and the pre-blast and post-blast profiles in figure #19 below, it

was determined that the largest toe burdens were too confined to adequately kick out the toe. Instead of

kicking out the toe towards the face it pushed it towards the left which became the path of least

resistance.

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Figure 20: Green line represents pre-blast topo, black line represents post blast topo, red line represents

drill hole

After reviewing the survey data we determined that the confinement likely became too high around a toe

burden of 92 to 98 charge diameters.

7.0 Conclusion

In conclusion shots one through three performed as predicted in the acceptable range. The following

predictions fit the resulting data from each shot.

7.875 inch (20 cm) holes provided a faster penetration rate and more accurate drill hole

Symetrix casing further increased the accuracy of the drill holes

The energy factor could be decreased to roughly 180 kcal/ton in the bottom half of the hole and

134 kcal/ton overall.

Destabilizing the toe by “kicking it out” and creating a failure plane by propagating a crack

between the drill holes successfully created a “controlled slope failure”

The process developed for creating “controlled slope failures” through blasting could be beneficial for

future blasts. The step by step process for designing and implementing this drill and blast program were

successful in stabilizing the crest of the 9-points scarp.

Figure 21: View of the partially remediated Scarp feature from across the pit - before the eighth shot

Acknowledgements

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I would like to acknowledge :

Mr. John Floyd from Blast Dynamics, Inc. He was instrumental in working through the initial design

and theory for shot 1 which was the basis for all preceding shots. The initial idea and theory behind the

“controlled slope failure” through blasting should be accredited to Mr. John Floyd.

Mr. Russell Sheets for provided most of the geotechnical background for this paper. Mr. Russell Sheets

is the Senior Geotechnical Engineer at the Gold Quarry pit who was instrumental in understanding the

mechanics behind the December 25th

9-Points escarpment failure.

Mr. Paul Keim (blast foreman) and the rest of the blasting crew for provide the practical expertise in

working through the many small loading issues and concerns that we ran across

Mr. Bryan and Sam from Boart Longyear for providing feedback and drilling expertise that was

instrumental in working through the survey issues and drilling issues that was instrumental in providing

accurately drilled holes.

For everyone else on site that put time, effort, and ideas forward ensuring that the crest stabilization

through “controlled slope failure” blasting was a success. Terry Martin, Mario Gallegos, Bob St. Louis,

Steve Cook and everyone else involved in the crest stabilization efforts.

References 1 Russell Sheets, 2010, Geotechnical Outline of the December 25

th 9-points failure via email

2 John Floyd, 2010, “Nine Points Slide Stabilization”, Report to Newmont Mining Gold Quarry Mine,

pg 1 3 John Floyd, 2010, “Nine Points Slide Stabilization”, Report to Newmont Mining Gold Quarry Mine,

pg 8 4 John Floyd, 2010, Loading Plan for Blast 1 – Nine Points Slide Stabilization – 9.875 inch blastholes,

ANFO, Excel file

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2011G - Geotechnical Slope Remediation Via RC-drilling And “Controlled … · 2019-05-12 · Geotechnical Slope Remediation Via RC-drilling And “Controlled Slope Failure Blasting”