Rock Mechanics
Kapoor Chand; Ved Kumar; Priyanshu Raj; Nikita Sharma; Amit Kumar Mankar; Radhakanta Koner
Abstract
Failure of tailings dams is a major issue in the mining industry as it critically impacts the environment and life. A major cause of the failure of tailings dams is the unplanned depositing of tailings and the increase in saturation due to rainfall events. This study using numerical modelling and artificial ...
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Failure of tailings dams is a major issue in the mining industry as it critically impacts the environment and life. A major cause of the failure of tailings dams is the unplanned depositing of tailings and the increase in saturation due to rainfall events. This study using numerical modelling and artificial intelligence techniques (like MLR, SVR, DT, RF, and XGB) aims to predict the slope stability of tailings dams to avoid failure. The stability of tailings dams is analysed using the finite difference method (FDM), which computes the factor of safety (FoS) using the shear strength reduction (SSR) technique. This investigation mainly focuses on the geotechnical and geometric parameters of the tailings dam, such as density, cohesion, friction angle, saturation, embankment height, slope angle and haul road width. Results of numerical modelling have been used for developing ML models and predicting slope stability. The efficiency of ML models was analysed based on the R2 and root mean square error (RMSE), mean squared errors (MSE), and mean absolute error (MAE). The XGB algorithm proved to be the most effective as it gave the highest accuracy and lowest RMSE value compared to other ML models. AI tool was developed based on the ML model results for dam slope stability prediction. The developed AI tool will help understand the role of saturation and geometry parameters in embankment stability at the initial level of investigation.
Rock Mechanics
Jitendra Singh Yadav; Poonam Shekhawat; Sreekeshava K S
Abstract
The present work aims to assess the pressure-settlement behaviour of sand beds under a square footing reinforced with coir geotextile using the PLAXIS 3D software. The angle of internal friction of sand was varied from 28° to 38°. The effect of length of coir geotextile (1B, 2B, 3B, 4B, and 5B; ...
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The present work aims to assess the pressure-settlement behaviour of sand beds under a square footing reinforced with coir geotextile using the PLAXIS 3D software. The angle of internal friction of sand was varied from 28° to 38°. The effect of length of coir geotextile (1B, 2B, 3B, 4B, and 5B; B is width of footing) and position of coir geotextile (0.2B, 0.4B, 0.6B, 0.8B, and 1B) to ultimate bearing capacity of sand were examined. A remarkable improvement in ultimate bearing capacity of sand beds was obtained with provision of coir geotextiles. It was observed that the bearing capacity of sand increases by placing coir geotextiles up to a depth of 0.4B from base of footing, thereafter it starts decreasing. The optimum length of coir geotextile was found as 4B-5B. An insignificant improvement in the bearing capacity ratio of sand reinforced with coir geotextile was observed at higher values of angle of internal friction.
Hafeezur Rehman; Wahid Ali; Kausar sultan Shah; Mohd Hazizan bin Mohd Hashim; Naseer Muhammad Khan; Muhammad Ali; Muhammad Kamran; Muhammad Junaid
Abstract
Support design is the main goal of the Q and rock mass rating (RMR) systems. An assessment of the Q and RMR system application in tunnelling involving high-stress ground conditions shows that the first system is more appropriate due to the stress reduction factor. Recently, these two systems have been ...
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Support design is the main goal of the Q and rock mass rating (RMR) systems. An assessment of the Q and RMR system application in tunnelling involving high-stress ground conditions shows that the first system is more appropriate due to the stress reduction factor. Recently, these two systems have been empirically modified for designing the excavation support pattern in jointed and highly stressed rock-mass conditions. This research work aims to highlight the significance of the numerical modelling, and numerically evaluate the empirically suggested support design for tunnelling in such an environment. A typical horse-shoe-shaped headrace tunnel at the Bunji hydropower project site is selected for this work. The borehole coring data reveal that amphibolite and Iskere Gneiss are the main rock mass units along the tunnel route. An evaluation of the proposed support based on the modified empirical systems indicate that the modified systems suggest heavy support compared to the original empirical systems. The intact and mass rock properties of the rock units are used as the input for numerical modelling. From numerical modelling, the axial stresses on rock bolts, thrust bending moment of shotcrete, and rock load from modified RMR and Q-systems are compared with the previous studies. The results obtained indicate that the support system designed based on modified version of the empirical systems produce better results in terms of tunnel stability in high-stress fractured rock mass conditions.
H. Sarfaraz; M. H. Khosravi; T. Pipatpongsa
Abstract
One of the most important tasks in designing the undercut slopes is to determine the maximum stable undercut span to which various parameters such as the shear strength of the soil and the geometrical properties of the slope are related. Based on the arching phenomenon, by undercutting a slope, the weight ...
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One of the most important tasks in designing the undercut slopes is to determine the maximum stable undercut span to which various parameters such as the shear strength of the soil and the geometrical properties of the slope are related. Based on the arching phenomenon, by undercutting a slope, the weight load of the slope is transferred to the adjacent parts, leading to an increase in the stability of the slope. However, it may also lead to a ploughing failure on the adjacent parts. The application of counterweight on the adjacent parts of an undercut slope is a useful technique to prevent the ploughing failure. In other words, the slopes become stronger as an additional weight is put to the legs; hence, the excavated area can be increased to a wider span before the failure of the slope. This technique could be applied in order to stabilize the temporary slopes. In this work, determination of the maximum width of an undercut span is evaluated under both the static and pseudo-static conditions using numerical analyses. A series of tests are conducted with 120 numerical models using various values for the slope angles, the pseudo-static seismic loads, and the counterweight widths. The numerical results obtained are examined with a statistical method using the response surface methodology. An analysis of variance is carried out in order to investigate the influence of each input variable on the response parameter, and a new equation is derived for computation of the maximum stable undercut span in terms of the input parameters.
M.A. Chamanzad; M. Nikkhah
Abstract
Drilling and blasting have numerous applications in the civil and mining engineering. Due to the two major components of rock masses, namely the intact rock matrix and the discontinuities, their behavior is a complicated process to be analyzed. The purpose of this work is to investigate the effects of ...
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Drilling and blasting have numerous applications in the civil and mining engineering. Due to the two major components of rock masses, namely the intact rock matrix and the discontinuities, their behavior is a complicated process to be analyzed. The purpose of this work is to investigate the effects of the geomechanical and geometrical parameters of rock and discontinuities on the rock mass blasting using the UDEC software. To this end, a 2D distinct element code (DEM) code is used to simulate the stress distribution around three blast holes in some points and propagation of the radial cracks caused by blasting. The critical parameters analyzed for this aim include the normal stiffness (JKN) and shear stiffness (JKS), spacing, angle and persistence of joint, shear and bulk modulus, density of rock, and borehole spacing. The results obtained show that the joint parameters and rock modulus have very significant effects, while the rock density has less a effect on the rock mass blasting. Also the stress level has a direct relationship with JKN, JKS, bulk modulus, and the shear modulus has an inverse relationship with the rock density. Moreover, the stress variation in terms of spacing and joint angle indicates sinusoidal and repetitive changes with the place of target point with respect to the blast hole and joint set. Also with a decrease in the JKN and JKS values, the radial cracked and plastic zones around a blast hole show more development. With increase in the joint persistence, the plastic zones decrease around a blast hole.
Rock Mechanics
R. Shafiei Ganjeh; H. Memarian; M. H. Khosravi; M. Mojarab
Abstract
Dynamic slope stability in open-pit mines still remains a challenging task in the computational mining design. Earthquake and blasting are two significant sources of dynamic loads that can cause many damages to open-pit mines in active seismic areas and during exploitation cycles. In this work, the effects ...
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Dynamic slope stability in open-pit mines still remains a challenging task in the computational mining design. Earthquake and blasting are two significant sources of dynamic loads that can cause many damages to open-pit mines in active seismic areas and during exploitation cycles. In this work, the effects of earthquake and blasting on the stability of the NW slope of Chadormalu mine are compared by a numerical modeling method. The dynamic results show that the maximum displacement under earthquake and blasting loads within the slope are 844 mm and 146 mm, respectively. According to the shear strain results, both the earthquake and blasting waveforms are destructive, while the earthquake waveforms cause more damages to the slope. Moreover, the deterministic and probabilistic seismic hazard analyses are carried out to assess the seismicity of the mine area. The experimental results indicate that the maximum values for the vertical and horizontal accelerations are 0.55 g and 0.75 g, respectively. The maximum calculated acceleration is then scaled to the selected earthquake accelerograms. In order to show the effective impact of the established scale, the model is executed using the original accelerograms. The results obtained show that the established scale prevents overestimation and underestimation of the displacement and strain. Therefore, applying scaled accelerograms in a dynamic slope stability analysis in mine slopes leads to more reliable and robust results. The overall results show that a strong earthquake causes plenty of damages to the slope, and consequently, interrupts the mining cycle. Hence, the seismic study and dynamic slope stability should be considered as a part of the computational mining design.
Rock Mechanics
M. Zoorabadi
Abstract
Numerical modelling techniques are not new for mining industry and civil engineering projects anymore. These techniques have been widely used for rock engineering problems such as stability analysis and support design of roadways and tunnels, caving and subsidence prediction, and stability analysis of ...
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Numerical modelling techniques are not new for mining industry and civil engineering projects anymore. These techniques have been widely used for rock engineering problems such as stability analysis and support design of roadways and tunnels, caving and subsidence prediction, and stability analysis of rock slopes. Despite the significant advancement in the computational mechanics and availability of high speed computing hardware, the input data and constitutive models remain the main source of errors affecting the reliability of numerical simulations. The problem with the input data has been deepened more by introducing empirical-based methods such as GSI classification to downgrade the rock properties from laboratory scale to field scale. The deformability modulus and strength parameters are the main outputs of these downgrading techniques. Numerical modelling users simply apply these downgrading methods and run the model without considering the real mechanics behind the stress induced failure and deformation around the underground excavations. While to the contrary to the commonly used downgrading methods that produce a constant modulus for rock at all depths, the rock modulus is stress dependent and varies with depth. In addition to this, the mechanism of stress induced displacement is not similar to the deformation of a continuum model simulated with equivalent rock properties. Apart from the mechanical characteristics of rocks, the magnitude and orientation of in-situ stresses are two other important parameters that have significant impacts on stress induced rock fracturing. The impacts of these two parameters have also been neglected in many practical cases. This paper discuss this old fashioned topic in more details with presenting the known facts and mechanics which numerical modelling users ignore them due to the unquestioning acceptance of downgrading methods. It also covers the influence of the stress magnitude and orientation on stress induced rock fracturing.
