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Case study of establishing a safe blasting criterion for the pit slopes of an open-pit coal mine
International Journal of Rock Mechanics & Mining Sciences 57 (2013) 1–10
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Technical Note
Case study of establishing a safe blasting criterion for the pit slopes of an open-pit coal mine Byung-Hee Choi a, Chang-Ha Ryu a, Debasis Deb b, Yong-Bok Jung a, Ju-Hwan Jeong a,n a b
Korea Institute of Geoscience and Mineral Resources (KIGAM), Daejeon 305-350, Korea Indian Institute of Technology (IIT), Kharagpur, India
a r t i c l e i n f o Article history: Received 1 August 2011 Received in revised form 18 April 2012 Accepted 24 July 2012 Available online 23 October 2012
1. Introduction The average depth of the pits of Pasir Coal Mine, which is located in the island of Kalimantan, Indonesia, has increased during the 17 years after the first development of the mine, and now reaches about 150 m below the ground surface. The pits are mostly composed of three rock layers of mudstone, sandstone, and coal seam. The layers of mudstone and sandstone become the host rocks of the coal seams, but they are in an unconsolidated state and thus so weak in nature. The weakness of these host rocks is believed to provide the primary source of the failure or instability of many slopes in the mine. Another important element that may affect the stability of the pit slopes is the ground vibrations that are produced from everyday blasting operations. Surface blasting with 80–130 kg/ hole of charge in pit bottom or on pit slope itself can produce large ground vibrations. Above all, since such a large-scale blasting is always conducted regardless of the stand-off distance from the blast to the slopes, repeated ground vibrations can seriously affect the stability of nearby slopes especially after rain. Moreover, the mine company has a plan to develop the pits up to over 300 m of depth in 15 years henceforth. Thus, controls of ground vibrations will become a rather critical issue to the slope stability as the pits become deeper and deeper. In this regard, the case study was launched to evaluate the propagation characteristic of the ground vibrations and to prepare a practical safe blasting guideline. One of the most important problems in establishing a safe blasting criterion will be to set up an allowable level of ground vibration for the structure concerned. Many research results were
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reviewed to get a clue to solve the imminent problem of determining the allowable level of ground vibration for the pit slopes. However, it was difficult to find an existing criterion except the Russian criterion [1], which defines the allowable level of ground vibration for two different types of slopes. That is, the allowable level for saturated sandy slopes is defined as 60 mm/s of PPV (peak particle velocity) and that for soil slopes, which are part of primary structure, is 120 mm/s. These two limits are applied to the slopes that are subjected to repetitive ground vibrations. In the case of single vibration, the corresponding allowable levels are 120 and 240 mm/s, respectively. Hence, in this study the allowable level of ground vibration was determined to be set by using a numerical analysis and the Russian criterion was used as a reference level. Necessary field measurements on the ground vibrations and data analyses are also conducted to derive a prediction equation for the vibration level in the mine area. A scaled distance equation is then derived using the allowable level and the prediction equation so that the maximum charge weight per hole can be determined for future blasting. In addition, some standard blasting patterns including the ones that can be used in the vicinity of slopes are proposed for field use through a series of test blasts.
2. Overall geology and mining method According to the previous report of Chung et al. [2], Tertiary sedimentary rocks are widely distributed on the island of Kalimantan in which the Pasir mine is located. The basement to the sedimentary rocks is composed of Paleozoic gneiss and schist of the Subda shield and Mesozoic igneous rocks containing ophiolite. The Pasir coal field is located in the Warukin Formation, which forms a large basin structure. In Pasir mine, the Tertiary
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sedimentary rocks are mainly comprised of mudstone, sandstone, and coal, each of which occupies approximately 80, 15, and 5% of the total volume of the rock of the mine, respectively. These three sedimentary rocks were all formed by fluvial deposition on a flood plain, or in a channel. Table 1 shows the laboratory test results of the representative physical properties of the mudstone, sandstone, and coal that occur in the mine. However, if these Tertiary sedimentary rocks, that is, the mudstone and sandstone, are saturated with water, they are nearly completely decomposed within two hours as can be seen in Fig. 1. The open pits are distributed along the coal seams, which extend over 16 km in the north–south direction as shown in Fig. 2. The dip angles of the coal seams varies in the range of about 751–851. The average elevations of the ground surface and the pit bottom are 170 m and 0 m respectively, so that the average depth of the pits is currently around 170 m. However, the pits will be developed up to 150 m below the sea level, so that the excavation will continue until the depth of pits reaches about 320 m. The widths of exploitable coal seams varies from a few meters to several tens of meters, and the maximum width is about 80 m. Fig. 3 shows an overview of a typical pit. In the mine, a blasting method with a single free face, which is usually the pit bottom, has been used. This is mainly because the host rocks are so weak that it is possible to exploit both the host rocks and coal seams with an excavator, after the rocks are appropriately loosened by
Table 1 Physical and mechanical properties of rock in Pasir coal mine. Rock type
Mudstone Sandstone Coal 3
Density (kg/m ) Uniaxial compressive strength (MPa) Tensile strength (MPa) Young’s modulus (MPa) Poisson’s ratio Porosity (%) Cohesion (MPa) from triaxial test Internal friction angle (1) from triaxial test Peak cohesion (MPa) from shear test Peak internal friction angle (1) from shear test Residual cohesion (MPa) from shear test Residual internal friction angle (1) from shear test
1992 6.1 0.5 550 0.26 8 1.78 26.6 0.11 23.3
2156 9.1 0.5 1384 0.25 21 2.44 36.1 0.13 32.8
1336 6.7 0.7 548 0.24 7 0.77 49.5 0.23 37.7
0.06 14.2
0.07 15.2
0.10 20.9
Fig. 2. Distribution of the coal seams.
Fig. 3. Overview of a typical pit.
Fig. 1. Decomposition phenomenon of mudstone and sandstone when saturated with water.
blasting. Fig. 4 is a schematic diagram that shows the stepwise development of a pit by vertical drilling and blasting, that is, a crater blasting, in pit bottom.
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Fig. 4. Schematic diagram of blasting operation in pit bottom.
3. Blasting operations and ground vibrations 3.1. Conventional blasting method In a typical round, 80–130 kg of ANFO (ammonium nitrates and fuel oil) is charged into a vertical hole of 200 mm diameter and 9 m depth, and the total number of holes can become up to 200 holes. Cartridge-type emulsion explosives are generally used to prime the ANFO. Because of the frequent rain in the island of Kalimantan, non-electric detonators are always used for the safety of blasting. Fig. 5(a) shows a schematic diagram of a typical conventional initiation pattern that has been used in the mine. The explosion starts from the IP (initiation point) and propagates along the main line first, and then along the branch lines. The circles represent blast holes and each number in a circle denotes the total delay time of the hole determined from the combination of the surface detonators. This conventional pattern is characterized by shorter delay time of the main line than the branch lines. In this particular example, delay times of the main and branch lines are 25 and 42 ms, respectively. The shock waves that transfer through NONEL tubes come back toward the starting position or IP since the branch lines are connected in backward echelon forms. This type of ignition pattern may cause cut-off problems. Thus, to prevent the cut-off, inner-hole detonators with 500 ms of delay time are used in all blast holes. The sequential movements of broken rocks or burdens in this pattern start from the 1st row and are indicated by thick dashed lines in Fig. 5(b). Although there is only a single free face with the crater blasting in the pit bottom, this pattern can also be thought of as a blasting with two free faces. That is, as with tunnel blasting, it can be assumed that a crater, which has been made by the blasting of the preceding hole, provides a new free face to subsequently blasted adjacent holes. To distinguish this burden from the actual burden, which is usually defined as the depth from the ground surface to the center of charge in a crater blasting, the term ‘apparent burden’, denoted by B in Fig. 5(b), is used here. The apparent burden B and spacing S in a typical round are usually about 6–7 m and 8–9 m, respectively. 3.2. Prediction of ground vibration level Field measurements were carried out to obtain the typical ground vibration data from daily blasting operations. The seismographs were installed over a broad range of distance of 50–1200 m from blasting sources. Fig. 6 shows a typical ground vibration record including an airblast record (MicL). The dominant frequencies of the ground vibration waveforms are very low, that is, about 6 Hz in this case. Actually, such a low frequency is found in almost all of the waveforms measured in the mine area. These waves are thought of as surface waves, and show dominant frequencies of less than 10 Hz. Generally, surface waves with frequencies of
Fig. 5. Conventional blasting pattern: (a) Schematic of a conventional blasting pattern in Pasir mine and (b) rock movement in the conventional initiation pattern.
4–8 Hz show higher amplitudes than those observed on solid rocks. Siskind [3] described that these higher amplitudes could be caused by constructive reinforcements of the waves generated from close blast holes. Ideally, sinusoidal waves are constructively reinforced if their phase differences are less than a fourth of their periods. Thus, these low-frequency and high-amplitude ground vibrations can induce a large amplification of slope vibrations, which, in turn, can result in the failure of slopes. Blast vibration monitoring and control usually begins with predicting ground motions induced by blasting in a concerned area. Most popular equations to predict PPV are those based on the square-root scaling (SRS) and cube-root scaling (CRS) of distance [4]. These two types of equations are given in Eqs. (1) and (2) below, respectively, and have frequently been used by many engineers and researchers. In these equations, the ground vibration level is represented by PPV, that is, the maximum amplitude between three components in a ground vibration record. Ghosh and Daemen [5] and Bhandari [6] summarized various predictive equations that had been suggested at that time.
V ¼K
V ¼K
D
n ð1Þ
1=2
L
D L1=3
n ð2Þ
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Fig. 6. Typical ground vibration record together with an airblast (MicL).
In Eqs. (1) and (2), V is PPV in mm/s, D is the distance from blast to measuring point in meters, and L is the maximum charge per delay in kg. The constants K and n are the parameters that should be determined through statistical data processing process and are dependent upon various conditions under which the specific blasting is conducted. It is important to choose a predictive model that has higher determination coefficient between the SRS and CRS models. The SRS model, however, is adopted in this study because there is no noticeable difference between the determination coefficients of the two models, and it is also a common practice to use the SRS model in a large-scale surface blasting. As Siskind [3] indicated, the traditional concept of ‘charge weight per 8 ms delay interval’ may or may not be used because there is no clear definition for the ‘delay’ in a low-frequency site, and Pasir mine can be thought of as such a site. Thus, in this study the charge weight per delay is redefined as the ‘charge weight per hole’ under the assumption that there is no noticeable change in the blasting method. The result of processing the vibration data using the SRS model is shown in Fig. 7, where PPV data are plotted on log–log scaled axes. In the graph, the solid line is the least-square regression line, and the dashed line is the ‘95% line’ below which 95% of the total data may fall. This 95% line corresponds to the probability value of P(zr1.645) in the standard normal distribution, where z is the standardized normal random variable. Total 312 field data were used for the data processing process, but the determination coefficient of the regression line was as low as r2 ¼0.57. Hence, to accomplish the safe blasting purpose, the predicting equation for the ground vibration level was chosen as the higher 95% line. Thus, the ground vibration levels for future blasting are expected to be lower than the level that is estimated by the following prediction equation.
V 95% ¼ 558:56
D
L1=2
0:95 ð3Þ
Fig. 7. Result of regression analysis for the ground vibration data in Pasir coal mine.
4. Establishment of safe blasting criterion 4.1. Strategy to determine the allowable level of ground vibration To establish a safe blasting criterion, an allowable level of ground vibration should be determined for the structure concerned. The most desirable way to set the allowable level of ground vibration for the pit slopes would be to use an experimental method, but
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Fig. 8. Analysis model of a typical section of the open pit mine.
practically it was not possible within a short period of time. Hence, a numerical approach was adopted to determine the allowable level of ground vibration with some relevant considerations on the physical properties of the rocks. Numerical analysis was first performed with the values of cohesions and internal friction angles obtained from the direct shear tests as shown in Table 1. However, it was found that the overall pit could be collapsed with these input values. Hence, the values of cohesions and internal friction angles that were obtained from the triaxial tests were used for the analysis. That is, the cohesions of 1.78, 2.44, and 0.77 MPa and the internal friction angles of 26.61, 36.11, and 49.51 were used in the analysis for the layers of mudstone, sandstone, and coal, respectively. With these input values, it appears that the average ground vibrations from everyday blasting do not affect the stability of the overall pit if the stand-off distance is larger than 50 m because the peak shear stress of 0.5 MPa occurred in that distance does not exceed the lowest cohesion of 1.78 MPa for the mudstone, which was obtained from the triaxial tests [7]. However, the surface rocks of local benches may actually have the values of cohesions and internal friction angles that were obtained from the direct shear tests although the deeper rocks may have higher strengths. That is, the mudstone layers in the local benches may have the residual cohesion of 0.06 MPa and peak cohesion of 0.11 MPa, not of 1.78 MPa. Moreover, the layers of sandstone up to 1–2 m below the bench surface can be saturated with rain, and thus there is always the possibility that part of the host rocks in the local benches can be decomposed to a certain degree especially after rain, as shown in Fig. 1. Thus, although the whole pit may remain stable from everyday blasting, the local benches or parts of pit slopes may not, especially after rain. Hence, in this study a value of cohesion, beyond which a shear failure could take place, is first defined as the critical level of shear strength of the local benches. Then, the ground vibration level that can produce a shear stress that exceeds this critical level of shear strength is to be sought for the local benches by changing the magnitude of explosion pressure in the numerical analysis. Finally, the critical vibration level obtained from the numerical analysis is adopted as the allowable limit of ground vibration for the pit slopes. 4.2. Numerical analysis on slope stability under dynamic loading condition 4.2.1. General description Stability of benches and overall pit slope is of prime concern to the mine management. In Pasir mine, coal seams have dip of about 801. The host rock is made of mostly mudstone and weak in
nature and also contains considerable amount of moisture. The uniaxial compressive strengths of rock strata vary between 6–9 MPa as in Table 1. As a result, the overall pit slope is kept 271 to improve the safety factor. Blasting is an integrated process of mining and it inevitably causes ground vibration, which may further destabilize the slopes. Moreover, as the mining progresses and coal is being excavated from deeper benches, there are more chances of slope failure problems. As a result, it has become imperative to estimate the effect of blasting or dynamic loading on the slopes as well as to investigate the slope stability problems at Pasir mine due to static loading. In this study, two dimensional numerical modeling technique is applied to analyze the slopes of a representative section for three stages of mining operation: (i) first stage of pit section, which is the pit layout after one year of current pit section, (ii) second stage of pit section, which is the pit layout after one year or 15 m of vertical excavation from the first stage of pit section, and (iii) third stage of pit section, which is one year after or 15 m of vertical excavation from the second stage of pit section. In this paper, these three stages will be termed as 1st excavation, 2nd excavation, and 3rd excavation, respectively. During the whole project, three types of numerical analyses, that is, static analysis and static equivalent of dynamic analysis for all three stages of pit sections, and dynamic analysis of the first stage of pit section were conducted, but in this paper, only the dynamic analysis will be discussed. Field data of ground vibrations are used for the verification of the results obtained from the numerical models. On the other hand, the strike of geological rock layers is more or less parallel to the strike of the coal layers, as mentioned in Section 2. The Pasir coal deposits are unique and can be thought of as inclined ore bodies having hanging and footwall rock strata extending over 1 km across the strike. Hence, the number and composition of rock strata across the strike will be the same at any vertical cross-section taken from hanging wall to the footwall. Due to this reason, 2D numerical analysis in plane strain condition has been preferred over expensive 3D numerical analysis. Since plane strain condition is assumed, out-of-plane strain along the strike direction has been neglected. This assumption generally provides accurate results in terms of displacements and stresses in long tunnels or other underground excavations made in rock mass. In this respect, 2D analysis with plane strain condition will be sufficient for Pasir site.
4.2.2. Numerical modeling of pit slopes As mentioned in Section 4.1, three finite element models of a typical section of a slope were developed for three excavation stages, but in this paper only the dynamic model is described.
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In the dynamic model, dynamic load is directly applied and the model is analyzed considering dynamic equilibrium conditions. Fig. 8 shows the analysis model of a typical section of the mine, and in the static and the static equivalent dynamic analyses all three different stages of excavations are modeled in ANSYS finite element software incorporating the three different materials, that is, sandstone, mudstone and coal. As mentioned in Section 2 and Section 4.1, the predominant rock is mudstone followed by sandstone and coal, and the angle of overall slope is around 271. The three excavations indicate 15 m depth difference or yearly advancement of the mine (three lines for profiles of slope design). Rock layers more than 2 m thickness are included as it is in the finite element models. Thinner layers (thickness less than 2 m) are combined as one layers and included in the finite element models. This procedure is adopted to generate optimum number of elements with minimum number of skewed elements. In order to apply boundary conditions, each model is extended to 500 m in each vertical side and to 100 m in the bottom horizontal side. Fig. 9 shows a part of the solid model of 1st stage of excavation in ANSYS. The entire model has 10 benches though only a half of
that is shown in Fig. 9. Apart from this excavation model, an in-situ model is also developed considering the pre-mining or pre-excavation condition. This model is required to obtain in-situ displacement, which is subtracted from that of excavation model to obtain the displacement due to excavation only. The source of dynamic loading on the pit slopes is considered to be blasting load only. In Pasir coal field, bottom 4 m is loaded with explosive and stemming is done at the top 5 m out of a 9 m blast hole. The average burden is taken to be about 7 m and 3 rows of holes are assumed to be blasted in given blast round. Hence in the numerical model, dynamic loading is applied to an area of 21 m by 4 m rectangle, thus the blast zone is defined by an area of that size located 5 m below the surface as shown in Fig. 9. The dynamic pressure loading is applied all around the boundaries of this rectangular area based on the impulse loading patterns shown in Fig. 10. These loading patterns simulate the dynamic loads generated from three blast holes and are mentioned later. Apart from this, all other boundary conditions remain the same as the model in Fig. 8. As mentioned in Section 4.1, three types of rocks mainly exist in the study area, and the properties of these
Fig. 9. Numerical analysis models: solid model of 1st stage excavation.
Fig. 10. Trapezoidal form of dynamic pressure versus time curve for 3 blast holes.
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three rocks that are used in the analysis are the same as those in Table 1 with the exception that the cohesions and internal friction angles are the ones obtained from the triaxial tests. As shown in Fig. 10, the blasting load is considered to be an impulsive load of high magnitude for a very short period of time. In this case, holes are blasted with delay of 42 ms and 67 ms respectively and have the duration of the peak impulse pressure of 1 ms. For each blast hole, a trapezoidal form of impulse loading is assumed in consideration of 0.5 ms to reach the peak pressure of 10, 50, 100, 500, or 1000 MPa followed by 1 ms of peak pressure loading and 0.5 ms to drop the pressure to zero. The peak dynamic loads of different magnitudes are applied to find the level of resulting PPV and shear stress for nearby pit slope. Thus, the analysis procedure is essentially the same as that of Deb et al. [7] except the differing explosion pressures. 4.2.3. Analysis results and discussions The dynamic model is analyzed for 1st stage of excavation considering the damping coefficients of 3% for sandstone and 5% for mudstone. Full transient analysis is performed with explicit integration scheme using Newmark method. The model is analyzed for 2 s although loading is applied until 116 ms. The results are reviewed in terms of PPV and shear stress at the locations of toes (T1–T4) and crests (C1–C4) as shown in Fig. 9. These points are located about 1–2 m below the bench surface. Table 2 PPV and shear stress with distance in the case of 1000 MPa of peak pressure. Measuring points
Toes T1 T2 T3 T4 Crests C1 C2 C3 C4
Distance (m)
PPV (mm/s)
Shear stress (kPa)
25 116 210 295
4000 77 22 10
1050 24 6 4
86 183 273 367
88 27 13 8
40 5 2 1
7
Table 2 summarizes the resulting levels of PPV and shear stress with distance in the case of 1000 MPa of peak blasting pressure, and Fig. 11 shows the obtained relationship between the shear stress (c) and PPV (V). If the units are changed properly, this relationship can be rewritten as follows: 1:31 V c ¼ 788 ð4Þ 1000 In Eq. (4), the shear stress c is in kPa and the PPV (V) is in mm/s. On the other hand, the strengths of in-situ rocks are usually lower than those of the corresponding intact rocks. According to Mohammad et al. [8], Young’s modulus and uniaxial compressive strength of field rock may be decreased up to about 0.5 and 0.3 times those of laboratory values, respectively. This means that although the obtained peak cohesion of the fresh specimen was 0.11 MPa for the mudstone in Pasir mine, the corresponding in-situ rocks of the mudstone layers, which actually comprise the local benches, can have a lower strength than this laboratory value. That is, the actual cohesion cin situ of the local benches in Pasir mine may also be as low as 0.5 times the laboratory value cintact, that is, cin situ ¼0.5cintact. In this regard, if the safety factor is taken to be k ¼ 2, the critical level ccritical of cohesion for the local benches can be 1/k ¼0.5 times the laboratory value of peak cohesion, cintact, peak ¼110 kPa, for the mudstone. Hence, in this study the critical level of cohesion is taken to be ccritical ¼0.5cintact, peak ¼55 kPa. The corresponding PPV value Vcritical that can be allowed for this critical level of shear stress (55 kPa) then becomes Vcritical ¼ 131 mm/s, according to the analysis result of Eq. (4). This result implies that the ground vibrations of less than 130 mm/s would not affect the stability of the local benches even after rain. Thus, the allowable level of ground vibration for the pit slopes is defined as Vallow ¼130 mm/s in this study. 4.3. Design criterion for safe blasting Now that both the allowable level of ground vibration (Vallow) for the pit slopes and the prediction equation (Eq. (3)) have been obtained, a safe blasting criterion can be easily established by using the concept of scaled distance equation. Just like the criterion
Fig. 11. Shear stress versus PPV in the case of 1000 MPa of peak pressure input.
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of the US Office of Surface Mining (OSM) [9], if a scaled distance equation is used, charge weight L per delay can be calculated according to the distance D from the blast. The design criterion can then be established by deriving the scaled distance factor for the structure concerned. The scaled distance factor Ds ¼D/L1/2 for the pit slopes is calculated from both the prediction equation of Eq. (3) and the allowable level of Vallow ¼130 mm/s as follows: D 0:95 ¼ 130 ð5Þ V 95% ¼ 558:56 1=2 L Then, the final design criteria for safe blasting in Pasir mine are defined by the following scaled distance inequality: D L1=2
¼ Ds Z5
ð6Þ
in Eq. (6) Ds is in the unit of m/kg1/2. Therefore, to secure the safety of the pit slopes, the maximum charge weight per delay should satisfy this inequality condition. For example, if the distance from the blast to a pit slope is D ¼50 m, the charge weight L per hole that can be used in a blast while securing the safety of the pit slope should be limited to less than 100 kg as follows: 2 2 D 50 Lr ¼ ¼ 100 ð7Þ Ds 5 5. Design of safe blasting pattern Until now, blasting operations in Pasir mine have been conducted without the consideration of the stand-off distance from blast to nearby slopes, so that there have always been some possibilities of slope failure due to ground vibrations especially after rain. According to the calculation of Eq. (7), the charge weight per hole should be maintained less than 100 kg if the blasting is to be conducted within 50 m from a pit slope especially when the blasting is conducted right after rain. However, until now, blasting with a scale of larger than this has actually been conducted near the slopes or on the slopes themselves, and this may have been direct or indirect cause of the local failures that have been occurred in the mine. Therefore, some standard blasting patterns that can be used in the vicinity of or on the slopes themselves should be prepared for field use. In this respect, a simple design method with which the blasting scale can be freely modified is shown here, and several standard blasting patterns are also presented for ease of field use.
In the particular example in Fig. 12, the delay time of the main and branch lines are 42 ms and 67 ms respectively, so that the apparent burden B and the spacing S are reversed with respect to those of Fig. 5(b). Thus, the broken rocks or burdens move backward direction. The moving direction of burden can affect the fragmentation of the broken rocks. Hence, more preferable pattern can be chosen between the patterns of Fig. 5(b) and Fig. 13. Also, it can be seen in the pattern of Fig. 12 that the shock waves passing through the NONEL tubes run away from the IP along the branch lines, thus there will be no cut-off problem in this pattern. On the other hand, in the pattern of Fig. 12, the shortest delay interval between two successively blasted holes is actually 25 ms. That is, the 2nd hole in the main line has the total delay time of 42 ms and that of a branch line has 67 ms. Thus, the difference of delay time between these two successively blasted holes is only 25 ms, and this may not be the optimum interval for the destructive wave interference according to Aldas and Bilgin [10]. Therefore, surface detonators with longer delay time, say, 109 ms, would be more useful to further decrease the ground vibrations. In that case, the shortest delay interval in main line will become 42 ms, but that of any other two sequentially blasted holes will become 67 ms. 5.1.2. Drilling pattern Generally, if a series of charged holes is simultaneously blasted in a bench blasting, the volume of broken rock will be increased. This is known as the result of wave reinforcement between blast holes. For instance, a wide space blasting can be used in bench blasting. Such an effect of simultaneous blasting, however, will decrease in a blasting with a single free face like crater blasting because of the increase of the degree of confinement. Especially, the bottom parts of the vertical holes will not be easily broken in a blasting with a single free face. Hence, it is recommended that the hole spacing ab in Fig. 13 be set to be the same as the actual burden W. This is because if the spacing gets larger than the
5.1. Changing blasting scales 5.1.1. Initiation pattern Pasir mine can be a low-frequency site as described in Section 3.2. In such a place, the traditional 8 ms delay interval may not be sufficient to effectively separate the charges of adjacent blast holes. Much longer delay interval may be suitable to minimize the ground vibrations because there are more chances of destructive wave interference in such case. Aldas and Bilgin [10] proposed the most suitable time delays to minimize the constructive interference of the surface waves produced from blasting at coal mine. They suggested optimum delay intervals of 42 ms and 67 ms for that mine. In this study, delay intervals of 42 ms and 67 ms are adopted for the main and branch lines respectively. A slightly different initiation pattern is also suggested as in Fig. 12. This pattern is different from the conventional one of Fig. 5(b) in that the branch lines are now made to be in forward echelon forms. Basically the suggested pattern is given a shorter delay time in the main line compared to the branch lines so that the apparent burden (B) can be made to be as small as possible.
Fig. 12. Suggested initiation pattern and rock movement.
Fig. 13. Suggested crater blasting pattern.
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burden, that is, if ab 4W, the bottom portion cde in Fig. 13 is not broken completely. It was found from several test blasts that the most preferable result was obtained when the apparent burden B, or hole spacing ab, was set to be the same as the actual burden W. 5.1.3. Charge calculation By using the concept of specific charge, or powder factor, which is defined as the charge weight divided by the volume of rock to be fractured, the proper charge in a crater blasting with a single free face can be calculated as follows: L ¼ CW 3
ð8Þ
where L is the charge weight and W is the actual burden, which is the distance between the ground surface and the center of charge in a crater blasting and is different from the apparent burden B. The proportionality constant C in Eq. (8) is the specific charge or powder factor, which describes the whole blasting condition. However, as the burden W increases, the required charge L calculated by Eq. (8) generally has the tendency of being overestimated. Thus, a correction factor h(W) called a blast scale function is introduced and Eq. (8) is rewritten as follows [11]: L ¼ h ðWÞ CW 3
ð9Þ
There are various blast scale functions available for crater blasting [11], but the following Lares’ formula will be sufficient: !3 rffiffiffiffiffiffiffiffiffiffiffiffiffiffi 1 ð10Þ hðWÞ ¼ 1 þ 0:41 W To properly use the blast scale function h(W) of Eq. (10), a simultaneous equation is needed. That is, if the proper charge weight L(W0) for burden W0 in a given blasting condition (C) is known, a desired charge weight L(W) for a new burden W for the same blasting condition (C) can be obtained by the following simultaneous equations: LðW 0 Þ ¼ hðW 0 Þ CW 30 LðWÞ ¼ hðWÞ CW 3
ð11Þ
Solving Eq. (11) for L(W) gives LðWÞ ¼
h ðWÞ W 3 h ðW 0 Þ W 30
L ðW 0 Þ
ð12Þ
5.2. Test blasting To verify whether the suggested method works properly, a series of test blasts should be conducted. In these tests, a situation in which the burden needs to be changed from W0 ¼8 m to W¼4 m is assumed. The apparent burden B and the spacing S are first determined according to the suggested pattern. Since the suggested pattern assumes a blasting with a single free face, or a crater blasting, the apparent burden B should be the same as the new burden W, as discussed in Section 5.1.2. Hence, B ¼W¼4 m. The hole spacing S is set to be the same as the apparent burden B for convenience, so that S¼B ¼4 m. The actual depth of hole H is set to be H¼1.1 W¼4.4 m considering the sub-drilling. In the current blasting pattern, the expected burden is W0 ¼8 m although the depth of hole is H¼9 m. However, it was found that up to now, at most 7.5–8 m of the burden has actually been excavated on average. Thus, given the standard charge weight L(W0) ¼100 kg for the current pattern, two cases of standard burden can be considered: (i) W0 ¼7.5 m, and (ii) W0 ¼8 m. If the standard charge and burden of current blasting pattern are supposed to be L(W0) ¼100 kg and W0 ¼ 7.5 m, respectively,
9
Table 3 Suggested standard blasting patterns. Burden B (m) Spacing S (m) Depth of hole H (m) Charge weight per hole L (kg)
3.0 3.0 3.3 9
4.0 4.0 4.4 18
4.5 5.5 5.5 32
5.5 6.5 6.6 51
7.0 8.0 8.5 100
then the standard charge weight L(W¼4) for a new blast with burden W¼4 m can be calculated by Eq. (12), i.e., Lð4Þ ¼
h ð4Þ 43 h ð7:5Þ 7:53
Lð7:5Þ ¼
ð0:3549Þð64Þ ð100Þ ¼ 19 kg ð0:2805Þð422Þ
ð13Þ
Therefore, in this case the standard charge weight for the new blast with the burden W¼4 m becomes L(4)¼ 19 kg per hole. In the test blast, the depth of hole was 4.4 m, charge weight per hole was 20 kg, and total number of holes was 40 holes (8 holes/row, 5 rows). The fragmented rock mass from the test blast was heaved over 2 m. This indicated that the holes were slightly overcharged. Similarly, when the standard charge and burden of current blast pattern are assumed to be L(W0) ¼100 kg and W0 ¼ 8 m, respectively, the standard charge weight L(W¼4) for a new blast with burden W¼ 4 m can again be obtained by Eq. (12) and becomes L(4)¼16 kg. To confirm the result, a test blast was conducted with the same drilling and charging pattern with the above case except that the charge weight was now changed to 15 kg per hole. As a result, the fragmented rock mass was heaved slightly over 1 m. Although long tensile cracks with widths of over 10 cm were found along the boundaries of the blast site, this round was found to be slightly undercharged. From the results of these two test blasts, when the standard burden W0 was supposed to be 7.5 m, the blasting with the target burden W¼4 m was slightly overcharged. In contrast, when W0 was 8 m, the same test blast was slightly undercharged. After conducting several additional tests, the standard burden for the current blasting pattern was determined to be W0 ¼7.7 m. Now that the standard burden W0 ¼7.7 m and the standard charge L(W0)¼ 100 kg per hole have been determined, the standard charge weights L(W) for a new blast with arbitrary size of burden W can be easily obtained by using Eq. (12). In this way, several standard patterns for a range of apparent burdens B¼3–7 m and hole depths H¼3.3– 8.5 m were designed and the results are presented in Table 3. In this table, the apparent burden B is used instead of the actual burden W for ease of field use. In addition, it is recommended that the diameter of 200 mm still be used even in the short hole of 3.3 m depth because a concentrated charge is more efficient than a distributed charge to break the bottom part in a crater blasting. 6. Conclusion The objective of this case study was to establish a safe blasting guideline that can secure the safety of pit slopes against blastinduced ground vibrations in Pasir coal mine, Indonesia. For the purpose, field measurements on the ground vibrations have been conducted to obtain PPV data throughout the investigation periods. Based on the acquired data, a general prediction equation for the ground vibration level was derived. Also, the allowable level of ground vibration for the pit slopes was set to be Vallow ¼130 mm/s of PPV based on the results of a numerical analysis. Then, using the prediction equation and the allowable level, the scaled distance inequality for the pit slopes was derived as Ds Z5 in order for the maximum charge weight per hole to be determined. In addition, a simple method of calculating the proper charge weight to meet the safe blasting criterion was illustrated using the concept of a blast scale function. An alternative initiation sequence, which can be used to control the fragmentation according to the direction of rock
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layers, was additionally suggested. Finally, several standard blasting patterns that can be used for the depth of hole of 3.3–8.5 m were presented for ease of field use. However, rock conditions encountered in everyday blasting will be different site by site. Hence, it is recommended that an optimum blasting pattern for a site of special interest be found through a series of test blasts based on the proposed standard patterns.
Acknowledgment This research was supported by the Basic Research Project of the Korea Institute of Geoscience and Mineral Resources (KIGAM) funded by the Ministry of Knowledge Economy of Korean government. References [1] Pal RP. Rock blasting: effects and operations. Rotterdam: Balkema; 2005. [2] Chung SK, Shin HS, Han KC, Sunwoo C, Song WK, Choi SO, Park C, Lee BJ Geotechnical study on the stabilization for the slopes of the Pasir coal mine. KIGAM report for KIDECO, Daejeon, Korea; 2003.
[3] Siskind DE. Vibrations from blasting. Cleveland, OH: ISEE; 2000. [4] Dowding CH. Construction vibrations. Englewood Cliffs, NJ: Prentice Hall; 1996. [5] Ghosh A, Daemen JJK. A simple new vibration predictor (based on wave propagation laws). In: proceedings of the 24th US rock mechanics symposium, college station, Texas; 1983. p. 151–57. [6] Bhandari S. Engineering rock blasting operations. Rotterdam: Balkema; 1997. [7] Deb D, Kaushik KNR, Choi BH, Ryu CH, Jung YB, Sunwoo C. Stability assessment of a pit slope under blast loading: a case study of Pasir coal mine. Geotech Geol Eng 2011;29:419–29. [8] Mohammad N, Reddish DJ, Stace LR. The relation between in situ and laboratory rock properties used in numerical modelling. Int J Rock Mech Min Sci 1997;34:289–97. [9] Atlas Power Company. Explosives and rock blasting. Atlas Power Company; 1987. [10] Aldas GGU, Bilgin HA. An interpretation of the effects of using different delay intervals in blasting at an open-cast mine in Turkey. Rock Mech Rock Eng 2003;36:409–21. [11] Makhin PA, Karchevskii VK. New theory of rock fracturing by blasting. J. Min. Sci 1965;1:218–31.
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Technical Note
Case study of establishing a safe blasting criterion for the pit slopes of an open-pit coal mine Byung-Hee Choi a, Chang-Ha Ryu a, Debasis Deb b, Yong-Bok Jung a, Ju-Hwan Jeong a,n a b
Korea Institute of Geoscience and Mineral Resources (KIGAM), Daejeon 305-350, Korea Indian Institute of Technology (IIT), Kharagpur, India
a r t i c l e i n f o Article history: Received 1 August 2011 Received in revised form 18 April 2012 Accepted 24 July 2012 Available online 23 October 2012
1. Introduction The average depth of the pits of Pasir Coal Mine, which is located in the island of Kalimantan, Indonesia, has increased during the 17 years after the first development of the mine, and now reaches about 150 m below the ground surface. The pits are mostly composed of three rock layers of mudstone, sandstone, and coal seam. The layers of mudstone and sandstone become the host rocks of the coal seams, but they are in an unconsolidated state and thus so weak in nature. The weakness of these host rocks is believed to provide the primary source of the failure or instability of many slopes in the mine. Another important element that may affect the stability of the pit slopes is the ground vibrations that are produced from everyday blasting operations. Surface blasting with 80–130 kg/ hole of charge in pit bottom or on pit slope itself can produce large ground vibrations. Above all, since such a large-scale blasting is always conducted regardless of the stand-off distance from the blast to the slopes, repeated ground vibrations can seriously affect the stability of nearby slopes especially after rain. Moreover, the mine company has a plan to develop the pits up to over 300 m of depth in 15 years henceforth. Thus, controls of ground vibrations will become a rather critical issue to the slope stability as the pits become deeper and deeper. In this regard, the case study was launched to evaluate the propagation characteristic of the ground vibrations and to prepare a practical safe blasting guideline. One of the most important problems in establishing a safe blasting criterion will be to set up an allowable level of ground vibration for the structure concerned. Many research results were
n
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reviewed to get a clue to solve the imminent problem of determining the allowable level of ground vibration for the pit slopes. However, it was difficult to find an existing criterion except the Russian criterion [1], which defines the allowable level of ground vibration for two different types of slopes. That is, the allowable level for saturated sandy slopes is defined as 60 mm/s of PPV (peak particle velocity) and that for soil slopes, which are part of primary structure, is 120 mm/s. These two limits are applied to the slopes that are subjected to repetitive ground vibrations. In the case of single vibration, the corresponding allowable levels are 120 and 240 mm/s, respectively. Hence, in this study the allowable level of ground vibration was determined to be set by using a numerical analysis and the Russian criterion was used as a reference level. Necessary field measurements on the ground vibrations and data analyses are also conducted to derive a prediction equation for the vibration level in the mine area. A scaled distance equation is then derived using the allowable level and the prediction equation so that the maximum charge weight per hole can be determined for future blasting. In addition, some standard blasting patterns including the ones that can be used in the vicinity of slopes are proposed for field use through a series of test blasts.
2. Overall geology and mining method According to the previous report of Chung et al. [2], Tertiary sedimentary rocks are widely distributed on the island of Kalimantan in which the Pasir mine is located. The basement to the sedimentary rocks is composed of Paleozoic gneiss and schist of the Subda shield and Mesozoic igneous rocks containing ophiolite. The Pasir coal field is located in the Warukin Formation, which forms a large basin structure. In Pasir mine, the Tertiary
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sedimentary rocks are mainly comprised of mudstone, sandstone, and coal, each of which occupies approximately 80, 15, and 5% of the total volume of the rock of the mine, respectively. These three sedimentary rocks were all formed by fluvial deposition on a flood plain, or in a channel. Table 1 shows the laboratory test results of the representative physical properties of the mudstone, sandstone, and coal that occur in the mine. However, if these Tertiary sedimentary rocks, that is, the mudstone and sandstone, are saturated with water, they are nearly completely decomposed within two hours as can be seen in Fig. 1. The open pits are distributed along the coal seams, which extend over 16 km in the north–south direction as shown in Fig. 2. The dip angles of the coal seams varies in the range of about 751–851. The average elevations of the ground surface and the pit bottom are 170 m and 0 m respectively, so that the average depth of the pits is currently around 170 m. However, the pits will be developed up to 150 m below the sea level, so that the excavation will continue until the depth of pits reaches about 320 m. The widths of exploitable coal seams varies from a few meters to several tens of meters, and the maximum width is about 80 m. Fig. 3 shows an overview of a typical pit. In the mine, a blasting method with a single free face, which is usually the pit bottom, has been used. This is mainly because the host rocks are so weak that it is possible to exploit both the host rocks and coal seams with an excavator, after the rocks are appropriately loosened by
Table 1 Physical and mechanical properties of rock in Pasir coal mine. Rock type
Mudstone Sandstone Coal 3
Density (kg/m ) Uniaxial compressive strength (MPa) Tensile strength (MPa) Young’s modulus (MPa) Poisson’s ratio Porosity (%) Cohesion (MPa) from triaxial test Internal friction angle (1) from triaxial test Peak cohesion (MPa) from shear test Peak internal friction angle (1) from shear test Residual cohesion (MPa) from shear test Residual internal friction angle (1) from shear test
1992 6.1 0.5 550 0.26 8 1.78 26.6 0.11 23.3
2156 9.1 0.5 1384 0.25 21 2.44 36.1 0.13 32.8
1336 6.7 0.7 548 0.24 7 0.77 49.5 0.23 37.7
0.06 14.2
0.07 15.2
0.10 20.9
Fig. 2. Distribution of the coal seams.
Fig. 3. Overview of a typical pit.
Fig. 1. Decomposition phenomenon of mudstone and sandstone when saturated with water.
blasting. Fig. 4 is a schematic diagram that shows the stepwise development of a pit by vertical drilling and blasting, that is, a crater blasting, in pit bottom.
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3
Fig. 4. Schematic diagram of blasting operation in pit bottom.
3. Blasting operations and ground vibrations 3.1. Conventional blasting method In a typical round, 80–130 kg of ANFO (ammonium nitrates and fuel oil) is charged into a vertical hole of 200 mm diameter and 9 m depth, and the total number of holes can become up to 200 holes. Cartridge-type emulsion explosives are generally used to prime the ANFO. Because of the frequent rain in the island of Kalimantan, non-electric detonators are always used for the safety of blasting. Fig. 5(a) shows a schematic diagram of a typical conventional initiation pattern that has been used in the mine. The explosion starts from the IP (initiation point) and propagates along the main line first, and then along the branch lines. The circles represent blast holes and each number in a circle denotes the total delay time of the hole determined from the combination of the surface detonators. This conventional pattern is characterized by shorter delay time of the main line than the branch lines. In this particular example, delay times of the main and branch lines are 25 and 42 ms, respectively. The shock waves that transfer through NONEL tubes come back toward the starting position or IP since the branch lines are connected in backward echelon forms. This type of ignition pattern may cause cut-off problems. Thus, to prevent the cut-off, inner-hole detonators with 500 ms of delay time are used in all blast holes. The sequential movements of broken rocks or burdens in this pattern start from the 1st row and are indicated by thick dashed lines in Fig. 5(b). Although there is only a single free face with the crater blasting in the pit bottom, this pattern can also be thought of as a blasting with two free faces. That is, as with tunnel blasting, it can be assumed that a crater, which has been made by the blasting of the preceding hole, provides a new free face to subsequently blasted adjacent holes. To distinguish this burden from the actual burden, which is usually defined as the depth from the ground surface to the center of charge in a crater blasting, the term ‘apparent burden’, denoted by B in Fig. 5(b), is used here. The apparent burden B and spacing S in a typical round are usually about 6–7 m and 8–9 m, respectively. 3.2. Prediction of ground vibration level Field measurements were carried out to obtain the typical ground vibration data from daily blasting operations. The seismographs were installed over a broad range of distance of 50–1200 m from blasting sources. Fig. 6 shows a typical ground vibration record including an airblast record (MicL). The dominant frequencies of the ground vibration waveforms are very low, that is, about 6 Hz in this case. Actually, such a low frequency is found in almost all of the waveforms measured in the mine area. These waves are thought of as surface waves, and show dominant frequencies of less than 10 Hz. Generally, surface waves with frequencies of
Fig. 5. Conventional blasting pattern: (a) Schematic of a conventional blasting pattern in Pasir mine and (b) rock movement in the conventional initiation pattern.
4–8 Hz show higher amplitudes than those observed on solid rocks. Siskind [3] described that these higher amplitudes could be caused by constructive reinforcements of the waves generated from close blast holes. Ideally, sinusoidal waves are constructively reinforced if their phase differences are less than a fourth of their periods. Thus, these low-frequency and high-amplitude ground vibrations can induce a large amplification of slope vibrations, which, in turn, can result in the failure of slopes. Blast vibration monitoring and control usually begins with predicting ground motions induced by blasting in a concerned area. Most popular equations to predict PPV are those based on the square-root scaling (SRS) and cube-root scaling (CRS) of distance [4]. These two types of equations are given in Eqs. (1) and (2) below, respectively, and have frequently been used by many engineers and researchers. In these equations, the ground vibration level is represented by PPV, that is, the maximum amplitude between three components in a ground vibration record. Ghosh and Daemen [5] and Bhandari [6] summarized various predictive equations that had been suggested at that time.
V ¼K
V ¼K
D
n ð1Þ
1=2
L
D L1=3
n ð2Þ
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Fig. 6. Typical ground vibration record together with an airblast (MicL).
In Eqs. (1) and (2), V is PPV in mm/s, D is the distance from blast to measuring point in meters, and L is the maximum charge per delay in kg. The constants K and n are the parameters that should be determined through statistical data processing process and are dependent upon various conditions under which the specific blasting is conducted. It is important to choose a predictive model that has higher determination coefficient between the SRS and CRS models. The SRS model, however, is adopted in this study because there is no noticeable difference between the determination coefficients of the two models, and it is also a common practice to use the SRS model in a large-scale surface blasting. As Siskind [3] indicated, the traditional concept of ‘charge weight per 8 ms delay interval’ may or may not be used because there is no clear definition for the ‘delay’ in a low-frequency site, and Pasir mine can be thought of as such a site. Thus, in this study the charge weight per delay is redefined as the ‘charge weight per hole’ under the assumption that there is no noticeable change in the blasting method. The result of processing the vibration data using the SRS model is shown in Fig. 7, where PPV data are plotted on log–log scaled axes. In the graph, the solid line is the least-square regression line, and the dashed line is the ‘95% line’ below which 95% of the total data may fall. This 95% line corresponds to the probability value of P(zr1.645) in the standard normal distribution, where z is the standardized normal random variable. Total 312 field data were used for the data processing process, but the determination coefficient of the regression line was as low as r2 ¼0.57. Hence, to accomplish the safe blasting purpose, the predicting equation for the ground vibration level was chosen as the higher 95% line. Thus, the ground vibration levels for future blasting are expected to be lower than the level that is estimated by the following prediction equation.
V 95% ¼ 558:56
D
L1=2
0:95 ð3Þ
Fig. 7. Result of regression analysis for the ground vibration data in Pasir coal mine.
4. Establishment of safe blasting criterion 4.1. Strategy to determine the allowable level of ground vibration To establish a safe blasting criterion, an allowable level of ground vibration should be determined for the structure concerned. The most desirable way to set the allowable level of ground vibration for the pit slopes would be to use an experimental method, but
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Fig. 8. Analysis model of a typical section of the open pit mine.
practically it was not possible within a short period of time. Hence, a numerical approach was adopted to determine the allowable level of ground vibration with some relevant considerations on the physical properties of the rocks. Numerical analysis was first performed with the values of cohesions and internal friction angles obtained from the direct shear tests as shown in Table 1. However, it was found that the overall pit could be collapsed with these input values. Hence, the values of cohesions and internal friction angles that were obtained from the triaxial tests were used for the analysis. That is, the cohesions of 1.78, 2.44, and 0.77 MPa and the internal friction angles of 26.61, 36.11, and 49.51 were used in the analysis for the layers of mudstone, sandstone, and coal, respectively. With these input values, it appears that the average ground vibrations from everyday blasting do not affect the stability of the overall pit if the stand-off distance is larger than 50 m because the peak shear stress of 0.5 MPa occurred in that distance does not exceed the lowest cohesion of 1.78 MPa for the mudstone, which was obtained from the triaxial tests [7]. However, the surface rocks of local benches may actually have the values of cohesions and internal friction angles that were obtained from the direct shear tests although the deeper rocks may have higher strengths. That is, the mudstone layers in the local benches may have the residual cohesion of 0.06 MPa and peak cohesion of 0.11 MPa, not of 1.78 MPa. Moreover, the layers of sandstone up to 1–2 m below the bench surface can be saturated with rain, and thus there is always the possibility that part of the host rocks in the local benches can be decomposed to a certain degree especially after rain, as shown in Fig. 1. Thus, although the whole pit may remain stable from everyday blasting, the local benches or parts of pit slopes may not, especially after rain. Hence, in this study a value of cohesion, beyond which a shear failure could take place, is first defined as the critical level of shear strength of the local benches. Then, the ground vibration level that can produce a shear stress that exceeds this critical level of shear strength is to be sought for the local benches by changing the magnitude of explosion pressure in the numerical analysis. Finally, the critical vibration level obtained from the numerical analysis is adopted as the allowable limit of ground vibration for the pit slopes. 4.2. Numerical analysis on slope stability under dynamic loading condition 4.2.1. General description Stability of benches and overall pit slope is of prime concern to the mine management. In Pasir mine, coal seams have dip of about 801. The host rock is made of mostly mudstone and weak in
nature and also contains considerable amount of moisture. The uniaxial compressive strengths of rock strata vary between 6–9 MPa as in Table 1. As a result, the overall pit slope is kept 271 to improve the safety factor. Blasting is an integrated process of mining and it inevitably causes ground vibration, which may further destabilize the slopes. Moreover, as the mining progresses and coal is being excavated from deeper benches, there are more chances of slope failure problems. As a result, it has become imperative to estimate the effect of blasting or dynamic loading on the slopes as well as to investigate the slope stability problems at Pasir mine due to static loading. In this study, two dimensional numerical modeling technique is applied to analyze the slopes of a representative section for three stages of mining operation: (i) first stage of pit section, which is the pit layout after one year of current pit section, (ii) second stage of pit section, which is the pit layout after one year or 15 m of vertical excavation from the first stage of pit section, and (iii) third stage of pit section, which is one year after or 15 m of vertical excavation from the second stage of pit section. In this paper, these three stages will be termed as 1st excavation, 2nd excavation, and 3rd excavation, respectively. During the whole project, three types of numerical analyses, that is, static analysis and static equivalent of dynamic analysis for all three stages of pit sections, and dynamic analysis of the first stage of pit section were conducted, but in this paper, only the dynamic analysis will be discussed. Field data of ground vibrations are used for the verification of the results obtained from the numerical models. On the other hand, the strike of geological rock layers is more or less parallel to the strike of the coal layers, as mentioned in Section 2. The Pasir coal deposits are unique and can be thought of as inclined ore bodies having hanging and footwall rock strata extending over 1 km across the strike. Hence, the number and composition of rock strata across the strike will be the same at any vertical cross-section taken from hanging wall to the footwall. Due to this reason, 2D numerical analysis in plane strain condition has been preferred over expensive 3D numerical analysis. Since plane strain condition is assumed, out-of-plane strain along the strike direction has been neglected. This assumption generally provides accurate results in terms of displacements and stresses in long tunnels or other underground excavations made in rock mass. In this respect, 2D analysis with plane strain condition will be sufficient for Pasir site.
4.2.2. Numerical modeling of pit slopes As mentioned in Section 4.1, three finite element models of a typical section of a slope were developed for three excavation stages, but in this paper only the dynamic model is described.
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In the dynamic model, dynamic load is directly applied and the model is analyzed considering dynamic equilibrium conditions. Fig. 8 shows the analysis model of a typical section of the mine, and in the static and the static equivalent dynamic analyses all three different stages of excavations are modeled in ANSYS finite element software incorporating the three different materials, that is, sandstone, mudstone and coal. As mentioned in Section 2 and Section 4.1, the predominant rock is mudstone followed by sandstone and coal, and the angle of overall slope is around 271. The three excavations indicate 15 m depth difference or yearly advancement of the mine (three lines for profiles of slope design). Rock layers more than 2 m thickness are included as it is in the finite element models. Thinner layers (thickness less than 2 m) are combined as one layers and included in the finite element models. This procedure is adopted to generate optimum number of elements with minimum number of skewed elements. In order to apply boundary conditions, each model is extended to 500 m in each vertical side and to 100 m in the bottom horizontal side. Fig. 9 shows a part of the solid model of 1st stage of excavation in ANSYS. The entire model has 10 benches though only a half of
that is shown in Fig. 9. Apart from this excavation model, an in-situ model is also developed considering the pre-mining or pre-excavation condition. This model is required to obtain in-situ displacement, which is subtracted from that of excavation model to obtain the displacement due to excavation only. The source of dynamic loading on the pit slopes is considered to be blasting load only. In Pasir coal field, bottom 4 m is loaded with explosive and stemming is done at the top 5 m out of a 9 m blast hole. The average burden is taken to be about 7 m and 3 rows of holes are assumed to be blasted in given blast round. Hence in the numerical model, dynamic loading is applied to an area of 21 m by 4 m rectangle, thus the blast zone is defined by an area of that size located 5 m below the surface as shown in Fig. 9. The dynamic pressure loading is applied all around the boundaries of this rectangular area based on the impulse loading patterns shown in Fig. 10. These loading patterns simulate the dynamic loads generated from three blast holes and are mentioned later. Apart from this, all other boundary conditions remain the same as the model in Fig. 8. As mentioned in Section 4.1, three types of rocks mainly exist in the study area, and the properties of these
Fig. 9. Numerical analysis models: solid model of 1st stage excavation.
Fig. 10. Trapezoidal form of dynamic pressure versus time curve for 3 blast holes.
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three rocks that are used in the analysis are the same as those in Table 1 with the exception that the cohesions and internal friction angles are the ones obtained from the triaxial tests. As shown in Fig. 10, the blasting load is considered to be an impulsive load of high magnitude for a very short period of time. In this case, holes are blasted with delay of 42 ms and 67 ms respectively and have the duration of the peak impulse pressure of 1 ms. For each blast hole, a trapezoidal form of impulse loading is assumed in consideration of 0.5 ms to reach the peak pressure of 10, 50, 100, 500, or 1000 MPa followed by 1 ms of peak pressure loading and 0.5 ms to drop the pressure to zero. The peak dynamic loads of different magnitudes are applied to find the level of resulting PPV and shear stress for nearby pit slope. Thus, the analysis procedure is essentially the same as that of Deb et al. [7] except the differing explosion pressures. 4.2.3. Analysis results and discussions The dynamic model is analyzed for 1st stage of excavation considering the damping coefficients of 3% for sandstone and 5% for mudstone. Full transient analysis is performed with explicit integration scheme using Newmark method. The model is analyzed for 2 s although loading is applied until 116 ms. The results are reviewed in terms of PPV and shear stress at the locations of toes (T1–T4) and crests (C1–C4) as shown in Fig. 9. These points are located about 1–2 m below the bench surface. Table 2 PPV and shear stress with distance in the case of 1000 MPa of peak pressure. Measuring points
Toes T1 T2 T3 T4 Crests C1 C2 C3 C4
Distance (m)
PPV (mm/s)
Shear stress (kPa)
25 116 210 295
4000 77 22 10
1050 24 6 4
86 183 273 367
88 27 13 8
40 5 2 1
7
Table 2 summarizes the resulting levels of PPV and shear stress with distance in the case of 1000 MPa of peak blasting pressure, and Fig. 11 shows the obtained relationship between the shear stress (c) and PPV (V). If the units are changed properly, this relationship can be rewritten as follows: 1:31 V c ¼ 788 ð4Þ 1000 In Eq. (4), the shear stress c is in kPa and the PPV (V) is in mm/s. On the other hand, the strengths of in-situ rocks are usually lower than those of the corresponding intact rocks. According to Mohammad et al. [8], Young’s modulus and uniaxial compressive strength of field rock may be decreased up to about 0.5 and 0.3 times those of laboratory values, respectively. This means that although the obtained peak cohesion of the fresh specimen was 0.11 MPa for the mudstone in Pasir mine, the corresponding in-situ rocks of the mudstone layers, which actually comprise the local benches, can have a lower strength than this laboratory value. That is, the actual cohesion cin situ of the local benches in Pasir mine may also be as low as 0.5 times the laboratory value cintact, that is, cin situ ¼0.5cintact. In this regard, if the safety factor is taken to be k ¼ 2, the critical level ccritical of cohesion for the local benches can be 1/k ¼0.5 times the laboratory value of peak cohesion, cintact, peak ¼110 kPa, for the mudstone. Hence, in this study the critical level of cohesion is taken to be ccritical ¼0.5cintact, peak ¼55 kPa. The corresponding PPV value Vcritical that can be allowed for this critical level of shear stress (55 kPa) then becomes Vcritical ¼ 131 mm/s, according to the analysis result of Eq. (4). This result implies that the ground vibrations of less than 130 mm/s would not affect the stability of the local benches even after rain. Thus, the allowable level of ground vibration for the pit slopes is defined as Vallow ¼130 mm/s in this study. 4.3. Design criterion for safe blasting Now that both the allowable level of ground vibration (Vallow) for the pit slopes and the prediction equation (Eq. (3)) have been obtained, a safe blasting criterion can be easily established by using the concept of scaled distance equation. Just like the criterion
Fig. 11. Shear stress versus PPV in the case of 1000 MPa of peak pressure input.
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of the US Office of Surface Mining (OSM) [9], if a scaled distance equation is used, charge weight L per delay can be calculated according to the distance D from the blast. The design criterion can then be established by deriving the scaled distance factor for the structure concerned. The scaled distance factor Ds ¼D/L1/2 for the pit slopes is calculated from both the prediction equation of Eq. (3) and the allowable level of Vallow ¼130 mm/s as follows: D 0:95 ¼ 130 ð5Þ V 95% ¼ 558:56 1=2 L Then, the final design criteria for safe blasting in Pasir mine are defined by the following scaled distance inequality: D L1=2
¼ Ds Z5
ð6Þ
in Eq. (6) Ds is in the unit of m/kg1/2. Therefore, to secure the safety of the pit slopes, the maximum charge weight per delay should satisfy this inequality condition. For example, if the distance from the blast to a pit slope is D ¼50 m, the charge weight L per hole that can be used in a blast while securing the safety of the pit slope should be limited to less than 100 kg as follows: 2 2 D 50 Lr ¼ ¼ 100 ð7Þ Ds 5 5. Design of safe blasting pattern Until now, blasting operations in Pasir mine have been conducted without the consideration of the stand-off distance from blast to nearby slopes, so that there have always been some possibilities of slope failure due to ground vibrations especially after rain. According to the calculation of Eq. (7), the charge weight per hole should be maintained less than 100 kg if the blasting is to be conducted within 50 m from a pit slope especially when the blasting is conducted right after rain. However, until now, blasting with a scale of larger than this has actually been conducted near the slopes or on the slopes themselves, and this may have been direct or indirect cause of the local failures that have been occurred in the mine. Therefore, some standard blasting patterns that can be used in the vicinity of or on the slopes themselves should be prepared for field use. In this respect, a simple design method with which the blasting scale can be freely modified is shown here, and several standard blasting patterns are also presented for ease of field use.
In the particular example in Fig. 12, the delay time of the main and branch lines are 42 ms and 67 ms respectively, so that the apparent burden B and the spacing S are reversed with respect to those of Fig. 5(b). Thus, the broken rocks or burdens move backward direction. The moving direction of burden can affect the fragmentation of the broken rocks. Hence, more preferable pattern can be chosen between the patterns of Fig. 5(b) and Fig. 13. Also, it can be seen in the pattern of Fig. 12 that the shock waves passing through the NONEL tubes run away from the IP along the branch lines, thus there will be no cut-off problem in this pattern. On the other hand, in the pattern of Fig. 12, the shortest delay interval between two successively blasted holes is actually 25 ms. That is, the 2nd hole in the main line has the total delay time of 42 ms and that of a branch line has 67 ms. Thus, the difference of delay time between these two successively blasted holes is only 25 ms, and this may not be the optimum interval for the destructive wave interference according to Aldas and Bilgin [10]. Therefore, surface detonators with longer delay time, say, 109 ms, would be more useful to further decrease the ground vibrations. In that case, the shortest delay interval in main line will become 42 ms, but that of any other two sequentially blasted holes will become 67 ms. 5.1.2. Drilling pattern Generally, if a series of charged holes is simultaneously blasted in a bench blasting, the volume of broken rock will be increased. This is known as the result of wave reinforcement between blast holes. For instance, a wide space blasting can be used in bench blasting. Such an effect of simultaneous blasting, however, will decrease in a blasting with a single free face like crater blasting because of the increase of the degree of confinement. Especially, the bottom parts of the vertical holes will not be easily broken in a blasting with a single free face. Hence, it is recommended that the hole spacing ab in Fig. 13 be set to be the same as the actual burden W. This is because if the spacing gets larger than the
5.1. Changing blasting scales 5.1.1. Initiation pattern Pasir mine can be a low-frequency site as described in Section 3.2. In such a place, the traditional 8 ms delay interval may not be sufficient to effectively separate the charges of adjacent blast holes. Much longer delay interval may be suitable to minimize the ground vibrations because there are more chances of destructive wave interference in such case. Aldas and Bilgin [10] proposed the most suitable time delays to minimize the constructive interference of the surface waves produced from blasting at coal mine. They suggested optimum delay intervals of 42 ms and 67 ms for that mine. In this study, delay intervals of 42 ms and 67 ms are adopted for the main and branch lines respectively. A slightly different initiation pattern is also suggested as in Fig. 12. This pattern is different from the conventional one of Fig. 5(b) in that the branch lines are now made to be in forward echelon forms. Basically the suggested pattern is given a shorter delay time in the main line compared to the branch lines so that the apparent burden (B) can be made to be as small as possible.
Fig. 12. Suggested initiation pattern and rock movement.
Fig. 13. Suggested crater blasting pattern.
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burden, that is, if ab 4W, the bottom portion cde in Fig. 13 is not broken completely. It was found from several test blasts that the most preferable result was obtained when the apparent burden B, or hole spacing ab, was set to be the same as the actual burden W. 5.1.3. Charge calculation By using the concept of specific charge, or powder factor, which is defined as the charge weight divided by the volume of rock to be fractured, the proper charge in a crater blasting with a single free face can be calculated as follows: L ¼ CW 3
ð8Þ
where L is the charge weight and W is the actual burden, which is the distance between the ground surface and the center of charge in a crater blasting and is different from the apparent burden B. The proportionality constant C in Eq. (8) is the specific charge or powder factor, which describes the whole blasting condition. However, as the burden W increases, the required charge L calculated by Eq. (8) generally has the tendency of being overestimated. Thus, a correction factor h(W) called a blast scale function is introduced and Eq. (8) is rewritten as follows [11]: L ¼ h ðWÞ CW 3
ð9Þ
There are various blast scale functions available for crater blasting [11], but the following Lares’ formula will be sufficient: !3 rffiffiffiffiffiffiffiffiffiffiffiffiffiffi 1 ð10Þ hðWÞ ¼ 1 þ 0:41 W To properly use the blast scale function h(W) of Eq. (10), a simultaneous equation is needed. That is, if the proper charge weight L(W0) for burden W0 in a given blasting condition (C) is known, a desired charge weight L(W) for a new burden W for the same blasting condition (C) can be obtained by the following simultaneous equations: LðW 0 Þ ¼ hðW 0 Þ CW 30 LðWÞ ¼ hðWÞ CW 3
ð11Þ
Solving Eq. (11) for L(W) gives LðWÞ ¼
h ðWÞ W 3 h ðW 0 Þ W 30
L ðW 0 Þ
ð12Þ
5.2. Test blasting To verify whether the suggested method works properly, a series of test blasts should be conducted. In these tests, a situation in which the burden needs to be changed from W0 ¼8 m to W¼4 m is assumed. The apparent burden B and the spacing S are first determined according to the suggested pattern. Since the suggested pattern assumes a blasting with a single free face, or a crater blasting, the apparent burden B should be the same as the new burden W, as discussed in Section 5.1.2. Hence, B ¼W¼4 m. The hole spacing S is set to be the same as the apparent burden B for convenience, so that S¼B ¼4 m. The actual depth of hole H is set to be H¼1.1 W¼4.4 m considering the sub-drilling. In the current blasting pattern, the expected burden is W0 ¼8 m although the depth of hole is H¼9 m. However, it was found that up to now, at most 7.5–8 m of the burden has actually been excavated on average. Thus, given the standard charge weight L(W0) ¼100 kg for the current pattern, two cases of standard burden can be considered: (i) W0 ¼7.5 m, and (ii) W0 ¼8 m. If the standard charge and burden of current blasting pattern are supposed to be L(W0) ¼100 kg and W0 ¼ 7.5 m, respectively,
9
Table 3 Suggested standard blasting patterns. Burden B (m) Spacing S (m) Depth of hole H (m) Charge weight per hole L (kg)
3.0 3.0 3.3 9
4.0 4.0 4.4 18
4.5 5.5 5.5 32
5.5 6.5 6.6 51
7.0 8.0 8.5 100
then the standard charge weight L(W¼4) for a new blast with burden W¼4 m can be calculated by Eq. (12), i.e., Lð4Þ ¼
h ð4Þ 43 h ð7:5Þ 7:53
Lð7:5Þ ¼
ð0:3549Þð64Þ ð100Þ ¼ 19 kg ð0:2805Þð422Þ
ð13Þ
Therefore, in this case the standard charge weight for the new blast with the burden W¼4 m becomes L(4)¼ 19 kg per hole. In the test blast, the depth of hole was 4.4 m, charge weight per hole was 20 kg, and total number of holes was 40 holes (8 holes/row, 5 rows). The fragmented rock mass from the test blast was heaved over 2 m. This indicated that the holes were slightly overcharged. Similarly, when the standard charge and burden of current blast pattern are assumed to be L(W0) ¼100 kg and W0 ¼ 8 m, respectively, the standard charge weight L(W¼4) for a new blast with burden W¼ 4 m can again be obtained by Eq. (12) and becomes L(4)¼16 kg. To confirm the result, a test blast was conducted with the same drilling and charging pattern with the above case except that the charge weight was now changed to 15 kg per hole. As a result, the fragmented rock mass was heaved slightly over 1 m. Although long tensile cracks with widths of over 10 cm were found along the boundaries of the blast site, this round was found to be slightly undercharged. From the results of these two test blasts, when the standard burden W0 was supposed to be 7.5 m, the blasting with the target burden W¼4 m was slightly overcharged. In contrast, when W0 was 8 m, the same test blast was slightly undercharged. After conducting several additional tests, the standard burden for the current blasting pattern was determined to be W0 ¼7.7 m. Now that the standard burden W0 ¼7.7 m and the standard charge L(W0)¼ 100 kg per hole have been determined, the standard charge weights L(W) for a new blast with arbitrary size of burden W can be easily obtained by using Eq. (12). In this way, several standard patterns for a range of apparent burdens B¼3–7 m and hole depths H¼3.3– 8.5 m were designed and the results are presented in Table 3. In this table, the apparent burden B is used instead of the actual burden W for ease of field use. In addition, it is recommended that the diameter of 200 mm still be used even in the short hole of 3.3 m depth because a concentrated charge is more efficient than a distributed charge to break the bottom part in a crater blasting. 6. Conclusion The objective of this case study was to establish a safe blasting guideline that can secure the safety of pit slopes against blastinduced ground vibrations in Pasir coal mine, Indonesia. For the purpose, field measurements on the ground vibrations have been conducted to obtain PPV data throughout the investigation periods. Based on the acquired data, a general prediction equation for the ground vibration level was derived. Also, the allowable level of ground vibration for the pit slopes was set to be Vallow ¼130 mm/s of PPV based on the results of a numerical analysis. Then, using the prediction equation and the allowable level, the scaled distance inequality for the pit slopes was derived as Ds Z5 in order for the maximum charge weight per hole to be determined. In addition, a simple method of calculating the proper charge weight to meet the safe blasting criterion was illustrated using the concept of a blast scale function. An alternative initiation sequence, which can be used to control the fragmentation according to the direction of rock
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layers, was additionally suggested. Finally, several standard blasting patterns that can be used for the depth of hole of 3.3–8.5 m were presented for ease of field use. However, rock conditions encountered in everyday blasting will be different site by site. Hence, it is recommended that an optimum blasting pattern for a site of special interest be found through a series of test blasts based on the proposed standard patterns.
Acknowledgment This research was supported by the Basic Research Project of the Korea Institute of Geoscience and Mineral Resources (KIGAM) funded by the Ministry of Knowledge Economy of Korean government. References [1] Pal RP. Rock blasting: effects and operations. Rotterdam: Balkema; 2005. [2] Chung SK, Shin HS, Han KC, Sunwoo C, Song WK, Choi SO, Park C, Lee BJ Geotechnical study on the stabilization for the slopes of the Pasir coal mine. KIGAM report for KIDECO, Daejeon, Korea; 2003.
[3] Siskind DE. Vibrations from blasting. Cleveland, OH: ISEE; 2000. [4] Dowding CH. Construction vibrations. Englewood Cliffs, NJ: Prentice Hall; 1996. [5] Ghosh A, Daemen JJK. A simple new vibration predictor (based on wave propagation laws). In: proceedings of the 24th US rock mechanics symposium, college station, Texas; 1983. p. 151–57. [6] Bhandari S. Engineering rock blasting operations. Rotterdam: Balkema; 1997. [7] Deb D, Kaushik KNR, Choi BH, Ryu CH, Jung YB, Sunwoo C. Stability assessment of a pit slope under blast loading: a case study of Pasir coal mine. Geotech Geol Eng 2011;29:419–29. [8] Mohammad N, Reddish DJ, Stace LR. The relation between in situ and laboratory rock properties used in numerical modelling. Int J Rock Mech Min Sci 1997;34:289–97. [9] Atlas Power Company. Explosives and rock blasting. Atlas Power Company; 1987. [10] Aldas GGU, Bilgin HA. An interpretation of the effects of using different delay intervals in blasting at an open-cast mine in Turkey. Rock Mech Rock Eng 2003;36:409–21. [11] Makhin PA, Karchevskii VK. New theory of rock fracturing by blasting. J. Min. Sci 1965;1:218–31.