Royal Dutch Shell is changing its tune on carbon, saying it will tie executive pay to shorter-term reductions in emissions. Shareholders will vote on the revisions in 2020. Shell's announcement marks a change in stance by Chief Executive Officer Ben van Beurden, who for years rejected investor demands that Shell detail its plans to curtail emissions, saying it would make the company more vulnerable to lawsuits. "Meeting the challenge of tackling climate change requires unprecedented collaboration, and this is demonstrated by our engagements with investors," the Shell chief said in a statement. "This joint statement is the first of its kind, sets a benchmark for the rest of the oil and gas sector, and shows the benefit of engagement—aligning institutional investors' long-term interests with Shell's desire to be at the forefront of the energy transition," Matthews said.
Naka, Ryosuke (Hokkaido University) | Tatekawa, Takuto (Hokkaido University) | Kodama, Jun-ichi (Hokkaido University) | Sugawara, Takayuki (Hokkaido University) | Itakura, Ken-ichi (Muroran Institute of Technology) | Hamanaka, Akihiro (Kyushu University) | Deguchi, Gota (NPO Underground Resources Innovation Network)
Underground Coal Gasification is expected to be efficient technique for coal energy recovery from deep or complex coal seam since directional drilling technique is advancing in these days. Authors have been performing small-scale UCG model tests to clear gasification and combustion process in UCG. Then, we found that radial cracks were initiated from the cavity formed in the artificial coal seam. Understanding mechanism of the crack initiation is important for clarification of the detail process of combustion and gasification and assessment for environmental risks. In this study, thermal stress analysis was performed on the small-scale UCG model tests to consider the initiation mechanism of the cracks by assuming that combustion and gasification of coal were progressing through the following three processes which are often observed in coal carbonization: (A) thermal expansion, (B) softening and melting and (C) thermal contraction. It was found that tensile stress was induced in the vicinity of the cavity in the tangential direction in process C. Direction of principal stress in the coal was almost parallel to tangential or radial direction of the cavity and the magnitude of it exceeded coal tensile strength. It was also found that tensile stress zone was extended into deeper coal seam with increase in temperature and time and compressive stress zone was formed outside of the tensile stress zone. It can be considered that the radial cracks initiated at the surface of the cavity since tangential tensile stress exceeded tensile strength of coal. Then, radial cracks were arrested at the boundary of tensile stress zone and compressive stress zone after they were propagating in coal seam.
Underground Coal Gasification (UCG) is a technique to use coal energy more efficiently and cheaply. In UCG, oxidant is injected into underground through an injection well to gasify coal seam, and syngas is recovered from a production well (Fig. 1). It is expected that UCG increases available amount of coal energy because even low-grade, complex and deep coal can be used by UCG.
It is pointed out that UCG has risks of surface subsidence and groundwater pollution because cracks are likely to initiate in coal seam by combustion and gasification. Therefore, clarification of initiation and growth mechanisms of the cracks is significant for stability assessment of ground as well as assessing environmental risks.
We performed small-scale UCG model tests on massive coal and crushed coal samples to clear gasification and combustion process in UCG. It was found that radical cracks were initiated in an artificial coal seam made by massive coal as well as crushed coal (Fig. 2 (Kodama et al., 2016)). Similar radial cracks were also observed in large-scale UCG model test (NPO Underground Resources Innovation Network, 2016).
3D laser scanning is a unique technology used for the description and subsequent modeling of real shape of spatially complex underground mining environment. Groundbreaking was its application in the pilot deployment of the Room and Pillar method at the CSM mine, where this method was used for the first time within the Upper Silesian Coal Basin, also known to have one of the most difficult mining and geomechanical conditions in the world. Very difficult mining conditions at depth over 800 m warranted searching for complex geotechnical tool or method that would capture all changes without distortion. Despite some shortcomings, 3D laser scanning was selected, although there is still no suitable device for dusty and humid mining environment. During the pillar development phase, comprehensive geotechnical monitoring was undertaken including the frequent scanning of pillar movement using 3D laser scanning technology. Based on repeated time-separated measurements, spatio-temporal analyses of deformation changes during ongoing mining were carried out. These analyses captured dynamic changes in coal rib, roof and floor movements of designated roadways while developing the pillar panel. In addition, time dependent long term post-mining measurements quantified additional strata movements within the panel enabling assessment of the long term pillar and mine roadway stability. The time-lapse scanning indicated variable pillar rib movement with maximum measured displacements of 60 cm. The scans indicated that in most cases, the bottom of the seam displaced more than the top of the rib side due to low floor strength causing large floor heave of up to 100 cm. During the 3-year monitoring, more than 2 billion spatial points were captured that can be used for further analysis.
A considerable amount of coal reserves are located in protection pillars that lie under built-up region in active Czech mining areas of the Upper Silesian Coal Basin. The commonly used controlled caving longwall mining method is not applicable in these areas because significant deformation of the surface is not permitted. For this reason the modified room and pillar method with stable coal pillars has been tested in order to minimise strata convergence. The trial operation of room and pillar method has been implemented at the shaft protective pillar where no mining was carried out in the past. Mining depth of room and pillar trial ranged from 700 m to 900 m. It is perhaps the deepest room and pillar coal mining in the world.
The post peak behavior of rocks has a significant influence on its strength and deformational characteristics and is crucial in many applications like underground coal mining. A clear understanding of post peak behavior of coal is necessary for economical and safe underground coal extraction. In this study, laboratory experiments and corresponding numerical modeling is done to understand the post peak behavior of a coal and coal pillar. A series of laboratory tests are conducted using coal samples and also artificially prepared gypsum samples with varying material strength. A numerical model is developed in FLAC3D and simulations are run with strain softening model to capture the post peak responses of the material. This model allows to attain the peak strength following the Mohr-Coulomb behavior and once it attains the peak, the strength parameters are softened with respect to plastic strain that the material experienced using a piecewise linear functions. The softening parameters are selected in such a way that a realistic behavior could be achieved. With the validated model, several parametric studies such as influence of dilation, confinement, and friction angle are performed. Understanding the influence of the post peak parameters gave a full extent of the usage of material performance in the numerical model for better design of underground excavations. The numerical model is subsequently extended to design coal pillar for a underground mines is briefly discussed.
Design of supporting rock pillars in underground excavations specially applications like mining is based on the maximum pillar strength as well on the post-failure behavior. The complete stress-strain behavior of the pillars play an important role for those pillar stability. Around the underground structures there is possibility of formation of plastic regions. Hence, the design steps together with support systems expected to be accommodative in accordance with the existence of such regions. To estimate the parameters related to the post-failure of supporting pillars, large scale in-situ compression tests are needed to be conducted, which is quite difficult and expensive. In this study, post peak characteristics of coal samples and artificially prepared gypsum samples are obtained using MTS servo controlled testing machine available in the Department of Civil Engineering, IIT Madras. The same experiments are also numerically modeled in FLAC3D and attempt was made to capture the post peak response. Simulation were done by moderating various parameters namely cohesion, friction angle and dilation. Mohr-Coulomb Strain Softening model is used as the decay of strength parameters with respect to the plastic strain to obtain the softening behavior (Itasca FLAC3D manuals, 2008). With the validated model, parametric studies are done in order to understand the influence of various parameters on post peak responses.
Drill and blast is the excavation method adopted to remove overburden material at the open pit coal mine of PT Buma Job Site Lati. Recently, the company applied deep hole drilling for blasting with double rods to reach the depth of 10-18 m. The main explosive was an emulsion explosive with target explosive consumption of 0.23 kg/m3. The blast hole was not fully charged but vertically decoupled using air decks at the middle and bottom of each blasting hole. Blasted rocks were then measured by digital photograph, and the size of P80 was found to range from 200 to 800 mm. The evaluation results indicate there is close relationship between the explosive consumption and the fragment size as well as the digging time. The air deck technique adopted in this study has been giving good results in terms of fragmentation size and explosive consumption.
In mining operations, drilling and blasting with a deep hole is a preferred method to reduce lost production time caused by the delays in blasting, to increase blasting inventory, and to minimize the number of drill pads so the drilling deviation can be minimized as well. Recently, PT BUMA Job Site Lati (called BUMA from now on) applied drilling for blasting operations with double rods to reach the depth of 10-18 m. Consequently, explosive consumption was high in the lower part of the blast hole, then an air deck was used to distribute the explosive along the blast hole. Another benefit of the air deck was that it cut the waiting time of the on-site sensitized explosive expansion to accommodate gassing.
In this study, the performance of blasting operation using the air deck is evaluated in terms of fragmentation. This study aims to check the effectiveness of using an air deck with the productivity target at the particular mine and to develop a blasting-fragmentation model that can be used to predict the size of the fragmented rock.
2. Literature Review
The air deck method is well-known in blasting operations to improve the quality of the blasting results. In the early 1940s, Russian scientists first came up with the idea of using an air gap between explosive columns. This method reduced explosive consumption in blasting activity. Melnikov et al. (1979) mentioned that an air deck can act as an energy accumulator. Marchenko (1982) found that pressure in an air deck would expand micro fractures that were previously generated by the main shock wave during blasting. Pompanna and Chikkareddy (1993) concluded that the presence of an air gap in the blast hole can reduce ground vibration and back break at the Kudrremukh iron mines. Jhanwar et al. (1996) revealed that the mechanism of air deck can reduce 25-30% of explosive consumption. Chiapppeta (2004) conducted experiments in the field and found that the air deck technique could remove the sub-drill which in turn reduced the explosive consumption by 16-25%, decreased vibration due to blasting by 33%, and improved fragmentation by 25%.
Utilization of an air deck will increase the fracture network due to the secondary shock waves formed as the result of wave reflection in the air gap. The fracture degree increases as a result of secondary shock waves as the duration of the shock wave effect on the rock mass around the blast hole becomes longer. The pressure reflections from the upper and lower explosive columns will collide in the middle of the air deck and is expected to interact with the surrounding rock mass to form additional radial fractures (Moxon et al., 1993; Zhang, 2016; see Fig.1). Air deck methods have been used in some open pit mines to reduce the consumption of explosive and to improve fragmentation (Chiapetta, 2004).
The contribution deals with the development of a new monitoring device for the measurement of the whole stress tensor in rock masses. The stress field is one of the principal factors which in a decisive way together with mechanical and deformation characteristics of rocks affect behaviour of rock masses. This issue has been investigated at the Institute of Geonics for a number of years, particularly in the context of assessing the impact of changes of the original stress field caused by anthropogenic interventions in masses (mainly in connection with dynamically changing the geomechanical situation as a result of mining). The device is based on the measurement of local deformations at the conical bottom of the well (Compact Conical-ended Borehole Overcoring and Monitoring techniques), from which it can be calculated the principal components of the stress field, or changes to it, and their orientation in space. The previously developed device can participate in repeated measuring. This significantly affects the frequency of read data especially in areas with a rapidly changing geomechanical situation. The new equipment is developed for use in an environment with a CH4 explosion hazard too. It is based on the already proven cited techniques in an ordinary environment. The devices allow monitoring and recording data about changes in stress in several places at the same time. In the case of system failure, each of the probes goes into autonomous mode data reading. These data will then be transferred from the backup memory of probes into the central data repository after resuming the system.
Knowledge, as accurate as possible, of the stress-strain state and its changes in rock mass is the determiningfactor for the proper planning of roadway support and for the correct design of underground mining as well as rockburst measures.
Stresses as well as stress changes cannot be measured directly. Stress determination is made indirectly, e.g. by the measurement of strain (e.g. Zang and Stephansson, 2010). A wide overview of these stress measurement methods has been presented (e.g. Zang and Stephansson, 2010). Deformation values obtained from an unbalanced body approaching equilibrium in combination with theoretical knowledge about constitutive behaviour (stress-strain relationship) allows us to evaluate the state of stress existing in any deformable body. Cells with strain gauges belong to physical methods for determining stress and stress changes in deformable materials with application to rock mass (e.g. Zang and Stephansson, 2010). Compact Conical-ended Borehole Overcoring (CCBO) methods are one possible method for in-situ stress measurement, which was established by Sugawara and Obara (1999). A modified version of the CCBO method called the Compact Conical-ended Borehole Monitoring (CCBM) method has been developed in the Czech Academy of Sciences, Institute of Geonics by Stas (Stas et al. 2004, 2005, 2007, 2011) and was tested for in-situ stress changes measurement in different rock masses (e.g. Konicek et al. 2012, 2013; Ptacek et al., 2015; Soucek et al., 2017).
A quick, simple and quantitative method for the estimation of surface subsidence susceptibility in mined areas with a lack of detailed geological and geometrical information in underground is presented in this paper. In the method only gangway depth from the surface and the attitude (dip and dip direction) of main geological features are used as input data based on the degree of availability and reliability. Underground gangways are represented as a series of points instead of closed polygons for easy calculation. The core assumption in this method is that the susceptibility to subsidence within a unit area increases both as the depth of the gangway from the surface decreases and as the number of gangways below the unit area increases. In spite of the simplicity of the proposed method, it gave satisfactory results when applied to a virtual excavation model and a closed coal mine where subsidence occurred actually.
Several methods for predicting ground subsidence due to mining excavation such as the profile method and the influence function method have been proposed (Whittaker and Reddish, 1989; Sheory, 2000). The National Coal Board (1975) has presented a basic technique to determine the surface area affected by coal mining based on the height and width of mined areas and the angle of inclination of coal seams. All these methods were developed and verified for conditions involving horizontal coal seams and long wall mining, which are the common mining conditions in Europe. However, coal-associated geological structures in Korea are very complicated, and coal seams have various widths and irregular dip angles. Consequently, the slant chute block caving method has been widely used in Korea, and sinkhole type subsidence is more common than trough type. As a result, the conventional prediction methods must be adapted to the Korean geology and mining conditions, or new subsidence estimation methods must be developed.
The goal of this study is to develop a simple, general, quantitative and reliable method for identifying subsidence susceptibility of the closed or abandoned coal mines, which is proper to be employed in geologically complicated areas. The proposed method in this paper considers only gangway depths and attitude of geological features like dip and dip direction is an optional parameter, because these data are relatively easy to acquire and generally reliable.
2. Estimation of subsidence susceptibility
2.1 Basic assumption
The depth of gangways is selected as an input data of this study after surveying the availability and effectiveness of data because it is reliable and can be easily acquired. In fact, several researchers revealed that the magnitude (volume) and depth of excavation are the principal factors influencing on the subsidence (Whittaker and Reddish, 1989; Singh and Dhar, 1997; MIRECO, 2008).
The method proposed in this study is based on the fact that the excavation volume and shape (or distribution of coal seams) are closely related to the gangway distribution. Two basic assumptions considered in the method are that the susceptibility to subsidence within a unit surface area increases as the depth of a gangway from the surface decreases and the number of gangways below the unit area increases. The first assumption is based on the bulking of failed rock mass which can fill the excavation and prohibit the propagation of roof failure. The second assumption comes from the fact that the rock mass around the excavation is damaged due to blasting and induced stresses.
The susceptibility related to the depth of a gangway is quantified using a negative exponential equation based on the results of numerical analyses (Park et al., 2005) and statistical data of subsidence occurrences in Korean coal mines as shown in Fig. 1 (MIRECO, 2008). Park et al. investigated the influence of the depth and width of excavation and of the spacing and dip of discontinuity on ground subsidence using PFC2D capable of modeling the bulking effect and showed that the overburden remains undamaged as the mining depth increases. Fig. 1 shows that most of subsidence occurred within a depth of 100 m from the surface. The number of subsidence events decreases exponentially as the gangway depth increases.
Singh, V. K. (CSIR-Central Institute of Mining and Fuel Research) | Singh, J. K. (CSIR-Central Institute of Mining and Fuel Research) | Kumar, A. (CSIR-Central Institute of Mining and Fuel Research) | Roy, S. K. (CSIR-Central Institute of Mining and Fuel Research) | Kumar, R. (CSIR-Central Institute of Mining and Fuel Research) | Singh, R. K. (CSIR-Central Institute of Mining and Fuel Research) | Kumar, M. (CSIR-Central Institute of Mining and Fuel Research)
The paper deals with geotechnical study and slope stability of the Coal and overburden benches in the hanging wall of an up-throw strike fault in the dip side of NMOC-II open cast coal mine, WCL. The opencast mine is located in the Maharashtra state of India. The detailed slope stability analysis is carried out by limit equilibrium method. The overburden slope of the quarry is mainly characterized by sandstone which is fractured and weathered. The mine is mostly covered by black cotton soil. Well fractured rock mass, existing open cast working have made the geo-mining condition of the quarry to drained condition for all practical purposes after implementing an effective drainage system. The relevant strength properties were determined in the soil and rock mechanics laboratory of CIMFR and subsequently used for slope stability analyses. The rock mass rating was also used to estimate the strength properties. The detailed slope stability analysis was carried out by GALENA software based on limit equilibrium method. The optimum design parameters for final pit slopes have been recommended.
The geotechnical study was conducted for slope stability study of the Coal and overburden benches in the hanging wall of an up-throw strike fault in the dip side of NMOC-II open cast coal mine of Majri Area. The mine belongs to Western Coalfields Ltd, a subsidiary of Coal India Ltd. The opencast mine is located in the Maharashtra state of India. Total overburden and coal have been planned to be excavated by Shovel Dumper combination. The OB bench height is kept at 10m. The shovel and dumper up to 5m3 and 60 tonne have been deployed at the project. The coal and overburden waste production are 2 million tones and 12 million m3 respectively. The existing maximum depth of the pit would be 150m. The study was conducted to optimum slope design with special reference to F1 fault without sacrificing the safety of the mine operation (CSIR-CIMFR Report, 2017).
The top soil of about 10m thickness forms the top most lithology in the project area. Soil is underlain by main overburden constituted sandstone with coal seam. Floor of the seam is generally composed of sandstone. The dip of the seam in this area is about 14 degree.
The importance of safe, properly designed and scientifically engineered slope is well known. The benefit of an openpit operation largely depends on the use of the steepest slopes possible, which should not fail during the life of the mine. So, the design engineer is faced with the two opposite requirements, stability and steepness, in designing the deep openpit slopes. Steepening the slopes, thereby reducing the amount of material to be excavated, can save a vast sum of money (Hustrulid et al., 2000). At the same time excessive steepening may result into slope failure leading to loss of production, extra stripping costs to remove failed material, reforming of benches, rerouting of haul roads and production delays. The Directorate of Mines Safety, the highest statutory body, may even close the mine, in case unsafe conditions are created. Therefore, it is necessary that a balance between economics and safety should be achieved.
The mine is mostly covered by black cotton soil. The information regarding geological sequence is therefore available only from the borehole data. Table 1 furnishes summarized statement of lithological formations encountered in the boreholes drilled in the mine quarriable area. The area is structurally disturbed by fault. F1 fault is present in the present mining area.
The failing process of coal sample under loading conditions is investigated by using both laboratory experiment and 3D finite-discrete element method in the present paper. The cohesive zone model was used to characterize nucleation, growth and propagation of cracks, while the potential contact detection and interaction of fractured solids were examined by means of the penalty method in ABAQUS software, where the parallel computation was employed to accelerate the calculations. Uniaxial and Brazilian tests were performed in the laboratory to obtain the mechanical properties of the coal such as Young’s modulus, fracture energy, cohesive strength, friction angle, uniaxial compression and shear strength. Further, these properties were carefully calibrated prior to being taken as input arguments in the continuous-discontinuous modelling. All the simulating results were basically in agreement with that obtained from the tensile tests in laboratory. This study shows that such computational mechanics of discontinua can be employed to gain powerful insight into the failure mechanism of coal, which could also be a useful tool to clarify the collapse mechanism of coal block caving in mining engineering design and rock test scheme optimization.
As a special kind of rock, coal is generally at a complex stress state under mining conditions. Thus, understanding for the mechanical behavior of coal plays a very important role in designing rock structures such as coal mining, underground excavation. In the literature, numerical methods such as continuum and discontinuum are often used to describe the failure mechanism of rock (Bobet et al., 2009; Li et al., 2015; Lisjak and Grasselli, 2014). For example, plastic deformation and damage softening are perhaps the most studied problems in the continuum method while the internal length of geomaterial is usually not considered in its formulation, which is the most serious drawback because of its predictions significantly depended upon mesh size. To bypass the shortcomings mentioned above, an enriched or higher-order continuum formulation for the softening was developed (de Borst and Pamin, 1996; de Borst, 2002) and nonlocal continuum was also introduced (Bazant and Planas, 1998). However, interaction between fragments during the evolution of multi-cracks cannot still be taken into account in the enhanced continuum methods. On the other hand, discontinuous modeling techniques that are known as discrete element methods (DEM) treat the material directly as an assembly of separate blocks or particles, which was originally proposed by Cundell (1971) from the viewpoint of analogous molecular dynamics simulation to better account for and understand the interaction between the blocks. In discontinuous methods, the length scale can be automatically incorporated into the modelling, which naturally accommodates the real size of elements or particles to capture failure zone of the shear process.
Yang, Yiran (Xi’an University of Science and Technology / Ministry of Education of China) | Lai, Xingping (Xi’an University of Science and Technology / Ministry of Education of China) | Shan, Pengfei (Xi’an University of Science and Technology / Ministry of Education of China)
Occurrence and stress setting in steeply inclined and thick coal seams are totally distinct from the ones in gently inclined coal seams. The difference would result in different roof structures when the roof is unstable. Furthermore, it would lead to inappropriate selection of supports without correct understanding of dynamic evolution of roof structure. This paper adopted comprehensive methods including theoretical analysis, 3DEC numerical simulation, field data comparisons and analysis, and aimed at the appropriate selection of supports and safe mining of Urumchi coal field. The results indicated that parameters including length, position of crest section of the asymmetric flat-topped arch structure(AFAS) and its height kept decreasing with increment of coal seam angle, showing extremely asymmetrical characteristic. simultaneously, the total weight of caving coal and rock mass and the average loading of supports were reducing obviously, while the reduced rate tended to be slow. As a result, the safe mining was improved dramatically and it is very crucial to do intense research and master the dynamic evolution.
The appropriate selection of supports has always been a vital ingredient for safe mining (Liu Changyou, 2015, Liu Jinhai, 2012, Zhang Dongsheng, 2013, Wang Jiachen, 2009), while the empirical method that is employed by most coal mines proves to be unreasonable. Unlike the result that is inferred by empirical method, the field monitoring data of ground pressure suggests that the working resistance of supports do not increase obviously with increment of mining depth. For example, both the mean value of measured maximum working resistance and time-weighted average resistance of working face of different depth in Wudong coal mine are less than 8000KN/support. Data and examples indicate that the supports are protected by a temporary roof structure. Many researchers have successively tried to describe the roof structure and determine the average loading of supports with various methods (Lai Xingping, 2010 and 2013, Shao Xiaoping, 2006, 2008 and 2009, Yang Fan, 2006). Shao Xiaoping (2007) proves that the roof structure would change with coal seam angle and supports only bear the weight within roof structure. Niu Shaoqing (2014) believes that roof instability is caused by relative sliding at roof sliding interface and provides support method. In addition, a fitting logarithm function between safety factor of roof and thickness to span ratio and span of roof with strength reduction method has been achieved by Zhao Yanlin (2010). Surely, all of these researches have made great contribution to the study of roof structure. However, seldom study have focused on the dynamic evolution of roof structure. To solve the above mentioned problem, this paper adopted comprehensive methods including theoretical analysis, 3DEC numerical simulation, field data comparisons and analysis. And the results indicated that there do exist a AFAS in the roof of working face. It is precisely the AFAS that allow supports only to bear the weight of caving coal and rock mass within AFAS. It is necessary for appropriate support selection to master the dynamic evolution of roof structure.