geotechnical instrumentation research leads into development of improved mine designs

SME Annual Meeting Feb. 28-Mar. 03, 2010, Phoenix, AZ

1 Copyright 2010 by SME

Preprint 10-010

GEOTECHNICAL INSTRUMENTATION RESEARCH LEADS INTO DEVELOPMENT OF IMPROVED MINE DESIGNS

H. Maleki, Maleki Technologies, Inc., Spokane, WA

ABSTRACT

This paper recognizes the 100th year of research accomplishments by the mining industry resulting from close cooperation among mining companies, geotechnical consultants and researchers from the former United States Bureau of Mines in stability assessments and development of improved and productive mine designs. The paper provides an overview of dramatic developments in geotechnical instrumentation, and its applications in the U.S. and around the world for enhancing the understanding of ground response to mining-induced changes in stress, resulting in improvements in ground control technologies.

INTRODUCTION

Development of improved mine designs has become an achievable goal in the 21st Century thanks to significant improvements in geotechnical measurement techniques, great advances in numerical modeling and decades of experience and evaluations in mining projects all across the world. This success is attributed to close cooperation between mining companies, academia and research organizations. The researchers from the USBM made significant contributions by implementing comprehensive underground measurement programs, and making the results available to the mining community to enhance our understanding of strata mechanics in complex mining methods involving caving, load transfer, seismicity and placement of backfill (Maleki 2006, Seymour et al 1988).

The focus of this paper is on pioneering research on development and utilization of geotechnical instrumentation and measurement techniques that have enhanced our understanding of strata mechanics and ground response to mining-induced changes in the stress field. The focus is on four fundamental areas: (1) Development of stress measurement hardware and procedures highlighting the impact of far-field stress on mine stability and its control through prudent mine orientation, among other options (2) Application of geophysical instrumentation for monitoring mining-induced seismicity and changes in material properties through adoption of microseismic and tomographic techniques (3) Development and utilization of ground support instruments for enhancing the understanding of reinforcement mechanisms and (4) Mine-wide utilization of other instruments for monitoring ground deformations and stability.

APPLICATION OF STRESS MEASUREMENT TECHNIQUES Significant mining took place during the first half of the 20th

Century using common engineering principles and the experience in neighboring operations, the trial and error approach and observation techniques. Historically, the mining industry depended on quick observations of strata deformation and failure, bolt failure, and lithologic changes (Hilbert, 1978) to evaluate stability; these preliminary evaluations sometimes were used to make significant changes in mine layout and support systems, which affected the economics of a mine. Because of the subjectivity involved in these visual observations, the mining process was slowed, creating inefficiencies and safety concerns during the time a decision is being reached.

It was not; however, until the 1970s that the impact of in-situ stress field on stability of the mine openings was fully recognized. Central to the understanding of stress-induced failure mechanism (Aggson 1978, 1979, Maleki et al 1991) was the development of stress

measurement techniques including both hydraulic fracturing and overcoring stress measurements (Bickle 1993, Doliner 2004).

The USBM's borehole deformation gauge and the over-coring technique used in many US mines have been shown to be accurate and reliable in determining in situ stress when adequate numbers of measurements are taken with complete relief. The method basically consists of (1) Drilling a 38-mm-diam pilot hole at the bottom of a 153-mm-diam borehole. (2) Setting the borehole deformation gauge in the pilot hole. And (3) Overcoring the gauge with a 153-mm-diam bit. Instead of the reusable USBM deformation gauges, the Hi-Cell has found applications particularly in Australian Collieries and around the world. Instead of completing the measurements through the underground openings, during the 1990s, Gray (2003) developed the SIGRA tool so that the overcoring measurements can be completed from the surface using a downhole overcoring technique.

The gravity component of the in situ stress field can be generally estimated from the depth of cover (vertical stress) and Poisson's ratio. Horizontal stresses are influenced by gravity through Poisson's effect and by the tectonic regime. For a typical Poisson's ratio of 0.15, gravitational horizontal stress is 18 pct of vertical stress. Therefore, for a 600-m-deep U.S. coal mine, the stress is approximately 15 MPa vertically and only 2.6 MPa horizontally. Tectonic processes and global plate movements also contribute to horizontal stress, and in some underground mines, horizontal stress can be three to five times higher than vertical stresses. In this situation, the horizontal stresses induce stability problems. A discussion of stability problems caused by depositional setting of mines is excluded from this paper but the interested reader is referred to Maleki 1988b.

In mines operating under biaxial horizontal stress field, mine orientation is the best approach to control the stress-induced stability problems including compressive type of failure commonly called cutters in the coal mines (Maleki et al 1991, 1993). If horizontal stresses are high in comparison to rock strength, mines can experience ground control problems. In general, the experience is that the longwall panels should be oriented parallel to maximum horizontal stress to minimize roof stability problems along the gateroads. To optimize headgate stability during the retreat, Su and Hasenfus (1995) suggest that the best panel orientation occurs when maximum horizontal stress is at small deviation (20o to 25o degrees) from the gateroad alignment and the headgate is in the stress shadow of the gob. To increase awareness of horizontal stress problems, NIOSH (Mark 2003) has expanded on the three-dimensional, finite-element modeling completed by Su and Hasenfus. Maleki Technologies Inc. (MTI, 2003) under contract from NIOSH implemented these guidelines on Windows platform (Figure 1) while addressing the rock mechanics aspect of stress concentration and relief using field measurements and three-dimensional stress analyses (Maleki et al 2003).

To illustrate the use of stress measurements in mine designs and stability assessments, results from a case study (Maleki et al 2003) are summarized here addressing variations in measured secondary principal horizontal stress across a western US coal mine; distance between farthest profiles is approximately 1.5 km. In addition to three sets of overcoring measurements at the mine level, Haimson (1972) completed a set of hydraulic fractures within a 1,830-m-deep vertical borehole to calculate the stress field and its relationship to earthquakes in the area. These measurements were integrated for this paper to address variations in both magnitude and orientation of premining



horizontal stresses. Stress profiles were also used to identify arching mechanisms in the mine roof.

Figure 1. Horizontal stress concentration results, softer developed by MTI.

Maximum (P) and minimum (Q) measured stress profiles are presented in Figures 2 through 5 in conjunction with biaxial measurements of Young's modulus (top). In situ horizontal stress can be calculated from these measurements by a statistical treatment of those measurements far enough from the entry to exclude stress concentrations, in conjunction with Youngs modulus obtained from measurements using the biaxial chamber.

60

50

50

40

10

Maximum P

Minimum Q

KEY

Disking

30

20

0 1 2 3 4

DISTANCE , m

B

A

5

Figure 2. Horizontal stress and deformation modulus profile in mine roof at site 1.

The stress measurements from site 1 (Fig. 2) were obtained from a thin, massive limestone; because of the very high stiffness of the limestone, which is equal to 58 Mpa, there is significantly greater stress concentration in this unit than in any other unit in the mine roof. The in situ stress field agrees with the stress trend measured at sites 2 and site 3 (Figs. 3 and 4 and Table 1), confirming a stress field oriented approximately N76oE with a maximum horizontal stress of 14.7 MPa and a minimum stress of 9.9 MPa. Maximum stress is oriented parallel to the cleat and increases slightly from site 2 toward site 3.

60

50

40

30

20

10

0

20

10

0 1 2 3 4

DISTANCE , m5

A

B

Maximum P

Minimum Q

KEY

Figure 3. Horizontal stress and deformation modulus profile in mine roof, site 2.

20

10

0

30

20

0 1 2 3 4 5

DISTANCE , m

A

B

Maximum P

Minimum Q

KEY

10

Disking


Table 1. Maximum and minimum horizontal stress field and depth of cover.

Location Maximum, MPa Minimum,

MPa Direction

of maximum

Depth of

cover, m

AOvercoring site: Site 1 Site 2 Site 3

42 13.6 14.7

21 9.6 9.9

N88W N79E N73E

210 184 350

BHydraulic fracture, Rangely oilfield

59

31

N70oE

1,830

Stress measurements were also obtained at site 4 (located at 30 m distance to site 3) (Fig. 5), suggesting a shallow pressure arch in the mine roof at this location. The roof consisted of a massive, stiff



siltstone that concentrated stresses as high as 28 MPa at a distance of 1.5 m above the roof line. This stress magnitude and distance relate to the stress concentration (arch) in the mine roof and agree with the moderate stress field of 13.6 to 14.7 MPa measured at site 2 and site 3 when differences in roof rock properties are taken into account using numerical modeling techniques. Stress concentration and roof arch appear to have shifted approximately 3 m into the roof at site 3 (Fig. 4); this shifting was probably influenced by inelastic behavior of carbonaceous rocks in the immediate roof and formation of crack (cutter type) failure by the ribs.

40

30

20

10

20

10

0 1 2 3 4

DISTANCE , m5

A

B

Maximum P

Minimum Q

KEY

30


The relationship between the horizontal stress field and cover depth is illustrated in Figure 6. A slight stress in-crease occurs over a distance of approximately 1.5 km, as one moves from site 2 toward site 3, which is both at greater depth and closer to a syncline. (The syncline is located about 1-1/2 km to the north of site 3.) The stress increase can reasonably be accounted for by the increase in cover and/or measurement accuracy. Thus, the in situ stress field is not influenced by the syncline at this location. However, hydraulic fracture measurements (Table 1) suggest significant stress increase with depth and distance to the N50oE regional strike-slip fault. This increase is more than would be expected from overburden weight and so is likely to be the result of stress concentrations near the fault or the method of measurement overestimating the stress. These hydraulic fracture measurements were taken 7 km from the mine site and are presented in Figure 6 to compare with mine site measurements.

APPLICATION OF GEOPHYSICAL MEASUREMENT TECHNIQUES Geophysical monitoring techniques have also evolved

considerably during the last century, benefiting from great improvements in the area of data acquisition and processing. These measurements have mostly been used to monitor changes in stability of rock masses by measuring rock deformation and fracturing. Stability assessment is generally carried out by comparing measurements between initial (or expected) conditions and predetermined critical levels for the material under investigation. For instance, if changes in seismic velocities reach a critical level of 17 percent, the roof is expected to behave inelastically, requiring supplementary support (Maleki, 1993b).

On the large scale, geophysical techniques have become increasing popular in monitoring mining-induced seismicity, in view of the Crandall Canyon Mine disaster (Arabasz and others 2005), more recently. These techniques listen and record rock noise: under stress.

For a good discussion of these techniques, the interested reader is referred to publications by Swanson 1995, Estey 1995 and Maleki 1995. The focus here is on the application of tomographic techniques to stability evaluations and improvements in designs.

N

10 MPa15 Mpa

Holes 1 and 3Sites 3 and 4

21 MPa

42 MPaSite 1

10 MPa14 MPa

Site 2

59 MPa

31 MPa

N 70 E

N 79 E

N 88 W

N 73 E

30

150

0 300

Scale , m

150

Rangely O il F ield

Overburdencontours, m

Stress e llipse

KEY

Figure 6. Measurement location and stress ellipsoid.

Fracture initiation has been monitored using microseismic emissions (Repsher and Steblay, 1985), and fracture density has been recently studied using a controlled source and tomographic techniques (Maleki, et al, 1993). The latter approach has allowed changes in roof stability to be compared to roof deformation and fracturing, which can significantly influence dynamic properties such as velocity, attenuation, Young's modulus, and Poisson's ratio.

Controlled source techniques have also been used to identify fracture zones and zones of high stress in pillars (Jung, Ibrahim, and Born, 1991; Friedel, Jackson, Tweeton, and Olson, 1993; Westman, 1993, Maleki and Hollberg 1995); the latter is based on experimental evidence showing an increase in seismic velocity with stress increase.

Figure 7 illustrates typical geometries used in a controlled-source tomographic survey to evaluate roof and floor stability. Pillar surveys are similar except that there is usually access to all sides of a pillar, while contact is limited to three sides (two source boreholes and the roof) for roof or floor surveys. The impact source is generally some sort of hammer that transmits energy to the rock within source boreholes or directly to the pillar. Accelerometers are attached to the mine roof, floor, or pillar to receive the signal generated by the impact. These techniques quantify wave velocity within a volume of rock (area of interest, Figure 7) in contrast to obtaining point measurements when using sagmeters. Using the crosshole seismic tomography today it is possible to study the structure under investigation by a high density of waves traveling between sources and receivers. This is analogous to the use of a computerized axial tomography (CAT) scan of the human body, but sound waves are generated instead of X-rays.

The components of a crosshole seismic survey developed for routine underground applications are described by Maleki and others 1992. A PC-based, 13-channel system with programmable gains



capable of acquiring up to 150,000 samples a second was used for data acquisition. This system utilized an analog digital board, amplifiers, filters, and a trigger unit (Maleki, Ibrahim, Jung, and Edminster, 1992).

Entry

Source

Receiver

Area of interest

KEY

B B'

A'

A

Entry

Borehole

E ntry

Figure 7. Typical three-sided seismic surveys of roofs and floors. top, Plan view; middle, cross section along A-A1; bottom, cross section along B-B1.

Selected results from a comprehensive case study is presented here to show the application of the tomographic techniques for assessing changes in structural stability in a trona mine where tomographic surveys were complimented with deformation measurements and fracture mapping. Significant changes in seismic wave velocities, amount of roof separation, and number of roof fractures occurred during the monitoring period. Supplementary support, consisting of 2.4-m-long fully grouted resin bolts and straps to control roof deformation, was installed on day 80. Roof and floor movements and pillar dilation were all large, indicating rock failure and complex strata interactions around the mining excavation.

Wave velocities were reduced by 22 percent at the intersection and by 15 percent at mid-pillar during the 120-day monitoring period. Figure 8 and table 2 demonstrate the calculated changes in dynamic properties.

Table 2. Calculated changes in dynamic properties. Factor Percentage of change Seismic velocity 21 Maximum amplitude 22 Dominant frequency 48 Deformation modulus 51 Poisson's ratio 125

A preliminary relationship between velocity changes and supplementary support requirements was established for the case study using these measurements. On day 80, wave velocities were reduced by approximately 17 percent at the intersection because of increased fracturing and possible roof destressing (Maleki, 1993b). The roof had deformed 60 mm since development. Changes in wave

velocity greater or equal to 15 percent may be used as a criterion for allocating supplementary support.

0123

0

3

6

912

15

2,100

2,400

2,700

3,000

3,300

May

F ebruary

3,600

TronaMarlstone

KEY

15 days30 days

45 days

60 days80 days

120 days

04 0 4 128 16

DISTANCE , m

KEY200

100

Figure 8. History of roof velocity and deformation along B-B1 during 4-month monitoring period. top, Velocity; bottom, deformation.

APPLICATION OF DEFORMATION MEASUREMENT TECHNIQUES Deformation measurements have found wide applications for

assessment of roof stability in coal, hard rock and evaporates within the last three decades (Maleki, 1988a, 1993a) and yield pillar performance; these measurements are inexpensive, simple, and reliable but require some protection of the instruments and cables from fly rock and mining equipment. Development of failure zones in mine pillars is similarly associated with rib fracturing and dilation, which is measured using dilation points and tape extensometers (Maleki, et al 2003).

Deformation measurements are obtained with extensometers and roof-to-floor convergence meters. These measurements have become very popular for assessment of roof or pillar stability in underground mines, nuclear waste repositories (Maleki 2009), and strategic storage facilities for oil and liquid gas, among others.

Measurements of amount and rate of roof deformation have been extensively used in both longwall and room-and-pillar retreat mining systems (Maleki, 1993a, 1988a). Because there is a large database involving a variety of geologic and stress conditions, there are preliminary criteria for the routine use of this technique in underground coal mines.

Measurements in seven U.S. mines have indicated that most coal-measure rocks deform 2 to 7 cm prior to roof collapse and then accelerate to a critical velocity called the "critical rate." Between the initiation of the critical rate and the actual roof fall, a period of time elapses that is called time to caving.

The rate of roof movement is a good indicator of roof stability, because it indicates a change in the stability of the whole mining system. A change from a steady rate of movement to a high (critical) rate can indicate roof block detachment, loss of support effectiveness, roof shearing by pillars, or impending pillar failure.

A history of relative roof deformation is presented in figure 9 for the middle entry of a longwall headgate, to illustrate the application of these simple, inexpensive measurements. The span was 6 m; the immediate roof consisted of mudstone overlain by a fluvial sandstone, and the primary support was 1.2-m-long mechanical bolts and straps.



Sandstone

Mudstone

Coal

Sandstone

Mudstone

Coal

Sandstone

Mudstone

Coal

Bed separation

Bed separationShearing

Roof caved

0.1 cm/day

0.12 cm/day

0.05 cm/day 1-m anchor no. 2

2.5-m anchor no. 1

1.25

1.00

.75

.50

.25

0 10 20 30 40 50 6

150 100 25 0

FACE POSITION, m

TIME , days0

5075125175

Figure 9. History of roof deformation for a longwall gate road.

Roof deformation was monitored by two wire sagmeters (Maleki 1981), with anchors 1 (at 2.5 m) and 2 (at 1 m) positioned above and below the bolt anchor horizon, respectively. Relative roof movements were calculated (fig. 9). The difference between the two curves reflects bed separation in the mine roof.

After development, roof movements reached 0.05 cm/day, which resulted in roof flaking. Installation of posts reduced the rate of roof movement, but did not stop bed separation at the sandstone contact. As the longwall face approached within 100 m of the instrument, bed separation and rate of roof movement significantly increased. Installation of additional support slowed movement, but another accelerated roof movement and bed separation cycle occurred as the longwall face came within 30 m of the instrument. This was followed by a roof fall that occurred approximately 4 days after the last accelerated roof movement. The critical rate of roof movement and the time-to-caving are best established by actually monitoring unstable roofs.

A critical rate of roof sag of 0.05 cm/day was determined on the basis of 142 measurements in three mines with 6-m spans and stable pillars (table 3). This critical rate was based on roof sag measurements along the gate roads of five panels where no secondary support had been installed. Installation of secondary supports generally stabilized roof movements; therefore, measurements taken near secondary supports were excluded from the study. Maleki (1993a) has addressed these same caving parameters for another four pillar-pulling operations where spans are gradually increased to 15 m and pillars crush to residual strength.

Table 3. Critical rate of sag for different mines using 6-m spans.

Mine Critical

rate, cm/day

Time to caving,

days Number of measurements

1 0.06 2 to 8 115

2 0.08* NA 24

3 0.05 9 to 11 3

*No failure was recorded but area was unstable.

REINFORCEMENT MEASUREMENT TECHNIQUES The aforementioned methods are most suitable for monitoring

changes in rock mass deformation and stress conditions. There are other useful measurements developed within the last three decades that indicate strata movement, but that do not directly lend themselves

to evaluations of structural stability. Measurements of bolt tension (Maleki, 1992b), for instance, provide indirect indications of roof movements, but also valuable insight into reinforcement integrity and load transfer from the rock to bolts and loss of anchorage leading into roof stability problems.

To illustrate the use of instrumented fully grouted rebars, results from a case study is presented (Maleki et al 1994). Six to eight instrumented bolts were installed in the two test sections, conventional bolting and an alternative pattern using higher density of roof bolts. The mine used 2.5-m-long No.7 Dwyidag bolts with 1.5 m of resin grout in a 1.2- 1.2-m bolting pattern. These bolts had a yield and ultimate capacity of 182 and 260 kN, respectively. The yield strength of six slotted bolts was 977 kN, which was 7.1 pct higher than the 912-kN strength of five regular bolts; therefore, the measured loads on the instrumented bolts in the conventional section would be lower than the loads on the regular bolts. These bolts contained six pairs of strain gauges mounted at 0.6, 1, 1.4, 1.7, 2, and 2.2 m along the entire bolt length; special procedures were developed to protect the strain gauge wires, and the bolts were carefully oriented with respect to the entry to allow an evaluation of bending moments.

reduced the yield strength to 163 kN and the ductility to between 9 and 12 pct. When data from the instrumented roof bolts were reduced, the correlation coefficients from the axial calibrations were used to convert voltage changes to load changes. This process was accurate to 0.4 kN. The strain gauges used are functional to approximately 50,000 microstrain, and the slotted bolt will yield at approximately 2,300 microstrain.

The axial and bending loads at the locations with strain gauges were compared to assess the support-rock interaction. The axial bolt loads measured immediately after installation were generally less than 20 kN and increased rapidly in response to mining and strata deformations. Figure 10 compares mean axial bolt load history for the conventional test section.

-25 to 0 kN

0 to 25 kN

25 to 50 kN

50 to 75 kN

75 to 100 kN

100 to 125 kN

125 to 150 kN

150 to yie ld

12 3

4 5 6KEY

A

B

C

Figure 10. Roof bolt load profiles after bolt installation for conventional section A: 1 day, B: 1 week, C: 2 months.



The bending moments generated by the bolts at 1 week after installation (fig. 11) confirmed that there was significant lateral movement toward the entry within both the immediate and intermediate roof. In the first test section, the moments were maximum within and near the rider-seam-coal interface. In the alternative section with increased reinforcement density, the moments were more uniformly distributed along the entire length of the bolts. The first, conventional section allowed more rock movement. This could produce failure planes or surfaces that would create higher localized bending stresses in the bolts. If the rock remained intact, then the bending stresses would be more evenly distributed (Fig. 11B).

12 3

4 5 6

+ + +

+ + +

No gauge

0 to 50 N-m

50 to 100 N-m

100 to 150 N-m

150 to 200 N-m

200 to 250 N-m

250 to 300 N-m

300 to 350 N-m

350 to 400 N-m

400 to 450 N-m

KEY

137 11 8 12 9

1014

+

+ +

+ + + + +

A

B

Figure 11. Bolt bending moment profile one week after installation A: Conventional section, B: Alternative section.

CONCLUSIONS

Development of improved mine designs has become an achievable goal in the 21st Century thanks to significant improvements in geotechnical measurement techniques, great advances in numerical modeling and decades of experience and evaluations in mining

projects all across the world. This success is attributed to close cooperation between mining companies, academia and research organizations.

Using in-mine measurements of stress, deformation and seismic wave velocity in coal and trona mines, the importance of pioneering research on development and utilization of geotechnical instrumentation and enhancing our understanding of strata mechanics and ground response is demonstrated. Case studies demonstrate the successful, practical use of the following techniques in stability evaluation and mine designs: (1) stress measurements influencing roof stability and bolt loading depending on the magnitude of stress, rock strength and orientation of stress in respect to the workings (2) geophysical instrumentation showing significant changes in roof stability during development particularly at an intersection in one study mine in evaporates (3) mechanical sag meters for predicting potential for roof falls and (4) instrumented fully grouted bolts enhancing the understanding of reinforcement mechanisms depending on support density.

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ABSTRACTINTRODUCTIONAPPLICATION OF STRESS MEASUREMENT TECHNIQUESAPPLICATION OF GEOPHYSICAL MEASUREMENT TECHNIQUESAPPLICATION OF DEFORMATION MEASUREMENT TECHNIQUESREINFORCEMENT MEASUREMENT TECHNIQUESCONCLUSIONSREFERENCES

geotechnical instrumentation research leads into development of improved mine designs

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