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Bulking characteristics of gangue are of great significance for the stability of goafs in mining overburden in the caving zones. In this paper, a particle discrete element method with clusters to represent gangue was adopted to explore the bulking coefficient time effect of the broken rock in the caving zone under three-dimensional triaxial compression condition. The phenomena of stress corrosion, deformation, and failure of rock blocks were simulated in the numerical model. Meanwhile, a new criterion of rock fragments damage was put forward. It was found that the broken rock has obvious viscoelastic properties. A new equation based on the Burgers creep model was proposed to predict the bulking coefficient of broken rock. A deformation characteristic parameter of the prediction equation was analyzed, which can be set as a fixed value in the mid- and long-term prediction of the bulking coefficient. There are quadratic function relationships between the deformation characteristic parameter value and Talbot gradation index, axial pressure and confining pressure.

Slope stability assessment is an essential aspect of mining and civil engineering. In this study, Songwe open-pit mine in Malawi was investigated to establish possible pit slope instability. In performing the analysis, an integrated approach entailing rock mass characterisation, kinematic and numerical methods were applied. Based on rock mass classification system, Songwe Hill carbonatite rock mass is characterised as a good rock but still it possesses numerous random discontinuities that present a complex challenge in geotechnical engineering. Dip 6.0 software was used in carrying out kinematic analysis based on the attributes of discontinuities. The results show that there is a 16% likelihood of planar failure in the divided slope sections of the planned pit. Thus, slope angle optimisation to 41° has been proposed as a counter-measure to minimise the potential risk of planar failure. At the optimised angle, the risk of planar failure could be reduced by 44%. On the other hand, wedge failure was found to be improbable since no joint intersections were found in the critical zone of potential failure. For numerical analysis, finite element code was applied using FLAC3D 5.0 application. The results demonstrate that overall slope angle of 41° would offer a favourable balance between safety and mining economics as mining operations progress to deeper horizons thereby avoiding a costly push back solution due to instability.

Background:The demand for underground space utilization using pipe jacking method is increased especially in the congested urban area.Methods:In the method, the over-cutting area is formed between pipes and surrounding soils in order to smoothly construct underground pipelines. The backfilling materials as a lubricant are injected into the over-cutting area during the constructions in order to reduce the friction resistance and to sustain the surrounding ground against the overburden pressure. Therefore, the performance of lubrication for backfilling material is significant. In this study, the fatty acids are added into the backfilling material to improve the lubrication performance.Conclusion:Laboratory studies have shown that the bleeding rate is reduced with the increase in addition ratio of fatty acids, however, the excess addition may cause the shrinkage of the backfilling material due to the effect of water absorption from the pipe. Additionally, 3% addition of fatty acids into the conventional backfilling material is the most effective to reduce the thrust for pipe jacking based on the results of the direct shear tests.

The demand for specific earth retaining wall method is increasing, along with the advancement and overcrowding of underground space use, such as the presence of close structures, in an urban area. The method is named stepped-twin retaining wall. The feature of this method is to have inner and outer retaining walls and excavate the ground by two-step in order to minimize the effect of the excavation on neighboring existing structures. However, the design of the earth retaining wall is currently carried out by individual engineers based on their own experience. Therefore, it is crucial that the standard method of re-taining wall using a two-step construction is established. As the first step to-ward the standardization, evaluation of factors affecting the ground behavior of the earth retaining wall was carried out. In particular, we picked up four major considerable factors, which are horizontal distance between the outer and inner walls, depth of outer wall embedment, mechanical properties of soil. The evaluation was done by using two dimensional FEM analysis and the results were summarized to make clear the effect of each factor.

In this paper, the behaviour of the pipe roof method with different pipe roofs disposition schemes including the pipe roof gate system and L-shaped system is discussed in terms of ground settlement reduction in the condition of underground space construction adjacent to the existing structure by using numerical simulation method. Considering the construction sequences of parallel underground tunnels with rectangular cross sections, the appropriate parameters and ground response characteristics of the corresponding pipe roof methods are evaluated by the ground surface displacement. The results indicated that the construction impacts from new buildings can be reduced by the pipe roof method. However, the theoretical performance of pipe roofs is highly dependent on the spacing between two structures. Generally, the L-shaped system is more effective in case of the distance (D) that is less than or equal to 1/2 width of box culvert (R) under the condition of a new structure constructed nearby the existing building. When two structures are constructed simultaneously, the excavation and the construction of structures influence each other reciprocally. The conclusions could provide a reference for the pipe roof method with the application scenario of adjacent building protection.

In medical and welfare facilities, the staff visually checks the acquired amount of foods to manage resident health. To reduce the load of the task of the staff, we propose a method that automatically estimates the amount of these meal intakes using images of the meal on the tray. To estimate the amount of change in the food on the tray, we used color as well as depth images of the meals. Using color images, we identified 10 types of foods using machine learning. Consequently, we obtained a food identification accuracy of 85%. Furthermore, we estimated the amount of change in the food on the tray before and after eating. To estimate the change amount of foods, we classified the depth images of each food into six categories and identified the food amounts based on these six categories using the ResNet50 network. As a result, we achieved a 90% accuracy of categorizing for each food type. Using the estimated values, we also proposed a calculation method of the acquired number of calories. We also confirmed that the proposed calorie estimation method was acceptable for use in medical institutions.

The present research primarily focuses on the investigation of gate-entry stability of longwall trial panel under weak geological condition in Indonesia coal mine by means of numerical analysis. This work aims at identifying appropriate roof support at 100 m and 150 m of depth during gate development. Due to depth depending competency of dominant rock, the stability of gate-entry at 100 m of depth can be optimized by leaving at least 1 m of remaining coal thickness (RCT) above and below the gate-entry. The appropriate support for the trial panel gate-entry is steel arch SS540 with 1 m and 0.5 m spacing for 100 m and 150 m of depth, respectively. The influence of panel excavation on gate-entry is also discussed. Regarding the aforementioned influence, the utilization of additional gate mobile support is recommended at least 10 m from the longwall face.

Rock burst prediction is a worldwide challenge that we have long tried to overcome. This study tentatively proposed a method to regionally predict rock burst hazards using microseismic energy attenuation. To verify the feasibility of the proposal, first, the mechanism of microseismic energy propagation and attenuation in rock medium was explored, and dominant attenuation characteristics of microseismic waves were analyzed. Second, a spatial attenuation model of microseismic energy was established, and the average energy attenuation coefficient for each wave path was defined. A 3D seismic energy attenuation inversion algorithm was put forward, and the corresponding computation matrix was developed. Third, a continuous microseismic field investigation was carried out in a deep coal mine. Seismic energy attenuation coefficient was confirmed using the calibrated focus position and energy determination. Based on data discretization processing, energy attenuation inversion and tomography, potential rock burst hazard regions were strictly zoned in mining areas. Finally, regional prediction results obtained from the microseismic energy attenuation were compared with the direct measurement results obtained from the classical drilling dust method to verify the reliability of proposed approach. It turns out that rock burst hazard regions predicted by the microseismic energy attenuation agreed well with the objective hazardous situations. Seismic energy attenuation coefficient is a significant evaluation factor that directly mirrors the inelastic performance of rock medium. Energy attenuation coefficient threshold used for determining the rock burst hazard regions was 3.0 km−1. Reliability of the seismic energy attenuation inversion and tomography was closely related to the spatial distribution of microseisms in a localized region. The optimum spatial density of microseisms was 0.2 m−3. Regional rock burst prediction using microseismic energy attenuation is an effective approach for revealing potential hazardous regions in deep mining conditions. This approach improves the pertinence of geological hazard prevention and provides a beforehand reference for targeted hazard management.

Blasting remains as an economical and reliable excavation technique, but there are some environmental shortcomings such as the control of blast-induced vibration. The impacts of vibration over surrounding communities in a blast area have been investigated for decades and researchers have been using a myriad of empirical predictive attenuation equations. These models, however, may not have satisfactory accuracy, since parameters associated to geomechanical properties and geology affect the propagation of seismic waves, making vibration modeling a complex process. This study aims for application of an Artificial Neural Network (ANN) method and Geomechanical parameter relationships to simulate the blast-induced vibration for a Brazilian mining site and then compare them to the traditional approach. ANN had the best performance for this mine despite having demanded large datasets (as much as for the traditional approach), while geomechanical parameters like RQD and GSI may be used to deliver a fair approach even without seismic data. Also, ANN methods may be useful in dealing with a large amount of information to facilitate the simulation process when combined with other methods. Therefore, alternative prediction methods may be helpful for small budget mining operations in planning and controlling blast-induced vibration and helping mining in urban areas becoming a more sustainable activity.

Predictive model and formula of jacking force has been well-established for circular pipe jacking in the past decades. However, the frictional mechanism of pipe–soil–slurry and jacking force predic...

In next two years, the current waste dump of Narynsukhait coal mine is predicted insufficient to accommodate the overburden as limited of the waste dump capacity. Thus, redesigning waste dump is paramount to increase capacity of the dump in future. This paper describes current condition of waste dump of Narynsukhait coal mine and then discusses the optimization of waste dump geometry by analyzing the effect of different waste dump’s bench configuration on slope performance. Optimization of the geometry is carried out by investigating and comparing the performance of geometrical combinations of bench height, bench angles and number of safety berm by means of numerical modeling. The model shows that increasing height of bench is able to induce shear stress in the bench and may initiate bench instable. However, the shear stress can be limited by having safety berm and/or reducing bench angle to satisfy the stability criteria.

Stability assessment is one of the most important issues in mining ground control. Mine development and/or production instability can cause production delay, loss of reserves, as well as injury to miners. Within the scope of this study, a series of open stope’s instability under the influence of overlaying mined-out regions were carried out with different mining scenarios at Modi Taung gold mine which is operated by National Prosperity Gold Production Group Limited (NPGPGL) in Myanmar. NPGPGL has been developing stopes up to 150 m from the surface at Shwesin vein system, and the mining activities are going to continue to deeper levels to fulfill the ore mineral supply. Creating a new stope opening under overlaying mined-out regions is not easy considering the instability of mined-out regions can affect the stope. The instability of new stope opening is not only due to its own induced stress but also the strong influence by the mined-out regions situated on upper part of the stope. Therefore, the understandings of ground behaviors and failure mechanisms of new stope opening due to the influence of overlaying mined-out regions are paramount to be studied. This paper describes in detail the strength factor and failure zones under the overlaying mined-out regions with different mine conditions by using numerical simulations, 3D finite difference software (FLAC 3D).

Underground facilities play an important role in infrastructure of modern society, especially some lifeline engineering, such as subway and railway tunnels, civil air-defense engineering and so on. As we know, the entire cognitive processes of seismic response of underground openings to earthquake is transforming from taking no account of earthquake load, primordially, to adopting the theory of free-field deformation, transitively, and the employing theory of dynamic design till now. However, so far, there are no established methods which can absolutely be employed for assessing and evaluating stability and induced plastic damage of tunnel surroundings ground during earthquakes. Based on a while building subway tunnel, this paper describes the propagation of seismic wave and seismic induced plastic damage of surrounding rock with a simple harmonic wave and a field recorded wave, which have the same dominate frequency. In addition, the specific energy density (SED) is introduced to represent the energy dissipation during ground shaking. Due to the recorded time-history wave, the response of feature points along tunnel and extension of plastic zone are analyzed with the different peak ground acceleration (PGA). Results illustrate that the upper tunnel cross section is more vulnerable than the lower part. Interestingly, when the PGA reaches 0.3 and 0.4 time of gravity acceleration, the seismic response of tunnel behaves almost the same.

The basic objectives of mine roadways are to provide sufficient cross sections to accommodate equipment, transport, personnel travel and ventilation. However, many roadways become damaged to the extent of needing maintenance, generally dinting, and in some cases requiring re-ripping. Strata conditions play important effects on stability of roadways and other mining activities. Weak rocks cause excessive roadway closures, and water softens some rocks and worsens the closure problem. Therefore, study on effect of mine water on the stability of underground coal mine roadways is important for underground coal mine development. A scientific discussion on the effect of water on the stability of roadways is given on the basis of results obtained by means of field investigation and laboratory tests. Based on analysis, rock in saturated condition has only between 0.19 to 0.49 of its compressive strength and 0.17 to 0.59 of its tensile strength in dry condition. Among the coal measures rock in research area, water has the most obvious effect on the strength reduction of shale. The slaking behavior of shale has also been the worst among the other. Field investigation of roadway driven in shale shown vertical closure in the wet and dry condition area reached 40–60 and 5–15 cm respectively, in 30 days after the drivage. Among the measures, drainage is considered to be the most economical and simplest method to reduce the water content in the rocks or the rock masses.

Maintaining stability as well as optimizing recovery of crown pillar, a pillar separating surface area with the uppermost stope in overhand cut and fill underground mining method, is important. Failures in stope may lead to crown pillar failures and cause surface subsidence. Increasing crown pillar thickness will increase crown pillar stability yet reduce mining recovery because part of crown pillar is formed by ore body. Preventing stope failure is the key to maintain stability and optimize recovery of crown pillar. Therefore, it is important to study countermeasure method for stope failure especially in crown pillar area. An attempt has been made to investigate the effectiveness of various countermeasures for stope failure in crown pillar area by means of parametric study. The result shows active type support system is effective for supporting stope in high vertical stress condition while the passive one needs to be installed if the stope is opened in high horizontal stress condition. In general, more supporting capacity from both type support systems is needed if the stope is opened in more severe geological condition. Another countermeasures, sill pillar and surface pile, are introduced for stope instability in crown pillar and non-crown pillar area. Sill pillar is an abandoned slice of unstable stope based on stability analysis. Sill pillar is very effective to stabilize stope both in crown pillar and non-crown pillar area, especially for stope in high horizontal stress condition. Sill pillar application in model with stress ratio 2 can optimize 20 meter thickness of crown pillar into 5 meter. Another proposed countermeasure is surface pile. Surface pile can be installed from the surface to improve stability of crown pillar and stope. The most effective use of surface pile is found in simulation of model with stress ratio 0.75 where surface pile can optimize 15 meter thickness of crown pillar into 5 meter.

Acid Mine Drainage (AMD) which occurs when sulfide minerals are exposed to water and oxygen with an excavation is one of the serious environmental problems in the world. A dry cover system is generally constructed in waste dump for the prevention of AMD in Indonesia by virtue of low cost and availability of waste rocks for a cover layer. However, the failure of the system caused by the lack of information related to the construction of cover system in mines, which leads to AMD, has been reported recently in Indonesia. In this study, the field investigation was conducted in pit and waste dump in open cast coal mine in Indonesia with the aim of obtaining the information on the construction of a cover layer and backfilling conditions of waste rocks in the waste dump. The rock samples taken in two areas of the mine were analyzed by geochemical analysis and sequential extraction with acids. The results indicated that Net Acid Producing Potential (NAPP) of the rocks in the waste dump down to 100 cm depth in both areas was from 10 to 30 kg H2SO4/ton, suggesting that Potentially Acid Forming (PAF) was backfilled in a cover layer. The backfill of PAF was contrary to the concept of cover system, which caused the failure of constructing a cover layer. The cause of the failure was likely attributed to the shortage of cover rocks which are classified as Non Acid Forming (NAF) or the failure of proper placement of them by an operational problem in the areas. Moreover, the results of the extraction with acids suggested that the form of iron and sulfur has to be taken into account to discuss the occurrence of AMD.

Rapid excavation could mitigate the imbalance relationship between excavation speed and production needs that plays a pivotal role in the sustainable development of large underground coal mines. This paper provides a case study on Bolter Miner Rapid Excavation Technology (BMRET) in Menkeqing Coal Mine, which has a high production of 13 million tons per year in Inner Mongolia. The temporal characterization of excavation procedures is analyzed in detail based on field monitoring data. The improvement of the roadway driving process and efficiency under a new support design is introduced, and corresponding evaluation methods, including parallel operation index ( P x ) and unit drilling-hole index ( D x ), are proposed for BMRET. A field application is conducted to verify the effectiveness of the improved BMRET, which fully considers the structure characteristic of the bolter miner machine. The performance and reliability of this new support scheme are monitored in terms of roadway convergence and axial force of cable through professional instrumentation programs in the field. The results show that the average excavation speed of the BMRET is 36.15 m/day (1080 m/month), an increase of 99.72% compared with the original excavation technology, which indicates that the BMRET could provide high efficiency in roadway excavation and effectively control the stability of deep roadways. It is pivotal to apply BMRET to ensure sustainable and highly efficient coal production. This case study provides reference and guidance for rapid excavation of deep underground roadways with similar geological conditions.

A study of multiple seams longwall mining is proposed to investigate its applicability in Indonesia coal mine. The study area of this research is PT Gerbang Daya Mandiri (GDM) coal mine located in East Kalimantan Island. The study of seam interaction is crucial for developing multiple seams longwall mining especially when it comes to weak rock conditions which are usually found in most of the coal reserves in Indonesia. This paper will use numerical simulation to investigate the effect of the first mined-out seam on the development of the second coal seam gate-entry by considering a couple of key parameters including depth of the coal seam and interburden length. The simulation model consists of two main indicators for instability which include failure zone, the contour of safety factor. The results show that the effect of seam interaction on gate-entry has different intensity based on the thickness of the interburden and coal seam depth. This work also provided appropriate support configuration for maintaining the stability of gate-entry.

The shaft lining failure which occurs in deep alluvium is a sudden coal mine hazard. The strata grouting is one of the treatment methods for the shaft lining failure. To investigate the impact of grout injection pressure on the shaft lining, the field measurement and the numerical analysis of the shaft lining stress variation during the grouting were conducted. To improve the strata grouting, the underground continuous impervious curtain (UCIC) is proposed as a new method by using the chain conveyor cutter technique without the impact on the shaft lining. The effects of the new method were also analyzed by means of the numerical methods. The results show that the strata grouting in the deep alluvium needs a high injection pressure, and in the horizontal direction, the shaft lining sustains the repeat tensile and compressive force during the grouting process. The negative influence of strata grouting on the stability of the shaft lining is obvious and serious. The UCIC built around the shaft lining can restrain the stress concentration induced by the aquifer drawdown. The triangular UCIC has a similar effect as that of the vertical one, and the small angle is better for preventing the shaft lining failure.

To study the effectiveness of grouting reinforcement on a large heaving floor of the retained goaf-side gateroad (RGSG) at a 900 m depth in the Zhuji coal mine, a numerical model involving strain softening constitutive material was built. First, the simulated deformations of the RGSG were compared with field data to verify the reliability of the numerical model. Then, the dynamic distribution of the stress in the RGSG floor was examined to reveal the mechanism of floor heave. Finally, grouting reinforcement was proposed to control the RGSG floor, and the corresponding effectiveness was verified by improving the rock mechanics of the floor strata based on the results of numerical uniaxial compressive tests. The results demonstrated that a fairly good match was achieved between the field and numerical data, and the proposed FLAC3D (Fast Lagrangian Analysis of Continua) numerical model was an effective approach to study the stability of the deep RGSG. A variation of the ratio between horizontal stress to vertical stress in the floor strata was the root cause of floor heave in the deep RGSG. Ideally, the floor heave could be reduced by 41%, 62%, and 79% when the floor strata of 1 m depth were reinforced with grouting schemes I, II, and III, respectively.

Zhuji coalmine, a typical deep underground coalmine located in Huainan city, Anhui province of China, is characterised by rich gas, high in situ stress and low permeability. These geotechnical conditions cause serious problems including outburst, excessive gas emissions in the upper corner of the coal face, and low resource recovery resulting from large remaining coal pillars when using conventional mining methods. Retained gob-side gateroad (RGSG) is proposed to resolve these problems by eliminating the large coal pillars and extracting gas simultaneously. To achieve a scientific and reasonable application of the RGSG in the Zhuji coalmine, stress evolution in the surrounding rock of this roadway was studied using FLAC3D numerical simulation method. Then, a staged supporting strategy and corresponding parameters were proposed in terms of the characteristics of stress evolution. Finally, the effectiveness of the proposed method was discussed based on the field measurements and roadway maintenance ...

Faults are complex geological conditions that are commonly encountered during underground excavation. Many support schemes, such as using a single pilot heading method and 30-m-long borehole pre-grouting, have been implemented during the pilot excavation of an 800-m-deep underground opening that passes through large fault zones in East China. However, various geo-hazards, including groundwater inrush, debris flow, and roof collapse, are still occurring, which seriously threaten tunneling safety. To eliminate the geo-hazards and ensure tunneling safety, ground surface pre-grouting (GSPG) was proposed and implemented for the first time to reinforce the regional engineering rock mass for this proposed 800-m-deep underground opening passing through large fault zones. The minimum grouting pressure of GSPG at a depth of 800 m below the surface is put forward based on hydraulic fracturing theory, providing valuable guidance for GSPG engineering practice. Engineering practice demonstrates that GSPG eliminates geo-hazards, improves the objective rock mass stability, and ensures tunneling safety. Field measurements indicate that the displacement velocity of the surrounding rock shows an obvious fluctuation response under the influence of GSPG, and the impact of GSPG on the stability of the 800-m-deep underground opening that has been excavated dramatically decreases as the distance from the grouting borehole increases. Moreover, there is a strong negative exponential correlation between the maximum velocity of deformations and the distance from the grouting borehole. In addition, the safe distance underground during GSPG is greater than 137 m.

To reduce noise components from original microseismic waves, a comprehensive fine signal processing approach using the integrated decomposition analysis of the wave duration, frequency spectrum, and wavelet coefficient domain was developed and implemented. Distribution regularities of the wave component and redundant noise on the frequency spectrum and the wavelet coefficient domain were first expounded. The frequency threshold and wavelet coefficient threshold were determined for the identification and extraction of the effective wave component. The frequency components between the reconstructed microseismic wave and the original measuring signal were compared. The noise elimination effect via the scale-changed domain decomposition was evaluated. Interaction between the frequency threshold and the wavelet coefficient threshold in the time domain was discussed. The findings reveal that tri-domain decomposition analysis achieves the precise identification and extraction of the effective microseismic wave component and improves the reliability of waves by eliminating the redundant noise. The frequency threshold and the wavelet coefficient threshold on a specific time window are two critical parameters that determine the degree of precision for the identification of the extracted wave component. This research involves development of the proposed integrated domain decomposition method and provides a diverse view on the fine processing of the microseismic signal.

This paper uses a two-dimensional finite element method (FEM) numerical modeling to predict groundwater inflow from an unconfined aquifer into the open-pit coalmine in the Phulbari area of northwest Bangladesh. The rock mass of the unconfined aquifer in the Phulbari coal deposit is highly porous and permeable. Interconnected pores within rock mass are the main paths for groundwater flow. Model assumes groundwater inflow rate through the rock mass associated with interconnected pores. Eleven stages of computation associated with variable overall pit slope angle (OPSA) 25°, 26°, 27°, 28°, 29°, 30°, 31°, 32°, 33°, 34°, and 35° are presented here. Model reveals that the probable inflow rate from unconfined aquifer in the open-pit coalmine would be about 377.68, 364.93, 354.38, 340.20, 331.29, 326.36, 316.86, 315.36, 308.03, 302.86, and 298.00 m3/h with relevant slope angles of 25°, 26°, 27°, 28°, 29°, 30°, 31°, 32°, 33°, 34°, and 35°, respectively. It is reasonable to emphasize that the computed inflow rate (m3/h) in the present numerical model would be effective through unit width of rock strata up to the pit depth of about 320 m in the Phulbari coal deposit. From the simulation results, it is thought that the open-pit coalmine in the Phulbari area would be an environmentally challenging issue because of regional unconfined mega-aquifer.

Goaf-side entry driving in underground coal mines could greatly improve coal recovery rates. However, it becomes more difficult to maintain stability, especially in deep coal mines. Pillar width plays a pivotal role in the stability of goaf-side entry driving. To obtain a reasonable and appropriate narrow pillar width, theoretical calculations of the widths of mining-damaged zone and limit equilibrium zone in the pillar are derived according to limit equilibrium theory. Based on the stability issues of goaf-side entry driving in the first island longwall coal face (LCF) at a depth of 800 m below the surface in Guqiao Coal Mine in China, a numerical model is established by FLAC software to analyze the stability of the surrounding rock of goaf-side entry driving during excavation, using various coal pillar widths and support schemes. The results obtained from theoretical calculations, numerical simulation, and engineering practice indicate that an 8-m-wide coal pillar is relatively reasonable, appropriate, and feasible. Field measurements show that deformations of the surrounding rock could be efficiently controlled 31 days after the support schemes were implemented in goaf-side entry driving with an 8-m-wide narrow pillar along the adjacent goaf side with a compaction duration of 10 months. The mining influence range of the overlying LCF on the stability of goaf-side entry driving is found to be the area from 50 m ahead of the LCF to 70 m behind the LCF as it passes over the measurement point.

The increase in the number of coal-fired power plants with the increase in coal production and its consumption has caused the problem of the treatment of a large amount of coal ash in Indonesia. In the past studies, coal ash was applied to postmine land with the aim of improving soil conditions for plant growth; however, heavy rain in the tropical climate may cause soil erosion with the change in soil conditions. This study presents the effects of application of coal ash to postmine land on soil erosion by performing the artificial rainfall test as well as physical testing. The results indicate that the risk of soil erosion can be reduced significantly by applying the coal ash which consists of more than 85% of sand to topsoil in the postmine land at the mixing ratio of over 30%. Additionally, they reveal that not only fine fractions but also microporous structures in coal ash enhance water retention capacity by retaining water in the structure, leading to the prevention of soil erosion. Thus, the risk of soil erosion can be reduced by applying coal ash to topsoil in consideration of soil composition and microporous structure of coal ash.

The stability of underground openings is pivotal to sustainable safe mining in underground coal mines. To determine the stability and tunneling safety issues in 800-m-deep underground openings through large fault zones in argillaceous rocks in the Guqiao Coal Mine in East China, the pilot industrial test, laboratory experimentation, and field measurements were used to analyze the large deformations and failure characteristics of the surrounding rock, the influence factors of safe excavation and stability of underground openings, and to study the stability control countermeasures. The main factors influencing the stability and tunneling safety include large fault zones, high in situ stress, poor mechanical properties and engineering performance of the argillaceous rock mass, groundwater inrush and gas outburst. According to the field study, the anchor-ability of cables and the groutability of cement-matrix materials in the argillaceous rock in the large fault zones were extremely poor, and deformations and failure of the surrounding rock were characterized by dramatic initial deformation, high long-term creep rate, obviously asymmetric deformations and failure, rebound of roof displacements, overall loosened deformations of deep surrounding rock on a large scale, and high sensitivity to engineering disturbance and water immersion. Various geo-hazards occurred during the pilot excavation, including roof collapse, groundwater inrush, and debris flow. Control techniques are proposed and should be adopted to ensure tunneling safety and to control the stability of deep underground openings through large fault zones, including regional strata reinforcement technique such as ground surface pre-grouting, primary enhanced control measures, floor grouting reinforcement technique, and secondary enclosed support measures for long-term stability, which are critical for ensuring the sustainable development of the coal mine.

Stability of the retained goaf-side gateroad (RGSG) is influenced mainly by the movements of the roof strata near coal seam after coalface passes by. To make effective controlling technology for the stability of the RGSG, we analyze the roof structure over the RGSG to illustrate the mechanism causing the RGSG instability under different roof conditions. We then examine the dynamic evolution of the deformation and abutment stress in the rock surrounding the RGSG during coal seam mining, using the FLAC3D numerical software to reveal the instability characteristics of the RGSG under different roof conditions. Next, corresponding stability controlling technologies for the RGSGs are proposed and tested in three typical deep underground coalmines. Results show that: sink and rotation of the roof cantilever over the RGSG impose severer influence on the stability of the RGSG. The RGSG suffers disturbances three times during the coal-seam mining, and the deformation and abutment stress in the rock surrounding the RGSG increase significantly when the main roof becomes thicker and the immediate roof becomes thinner. Staged support technology involving grout cable bolts has better controlling results of the RGSG stability than that composed of conventional rock bolts, when the RGSG is beneath weak immediate roof with large thickness. Roof structure optimizing technology involving pre-split technology can improve the stability of the RGSG effectively when the RGSG is covered by hard main roof with large thickness directly.

In this paper, a series of parametric three-dimension numerical simulations were carried out to estimate the face stability and to calculate the minimum allowable slurry pressure of pipe jacking tunnel in frozen ground for the first time. In total, 5120 of simulation schemes were done with different ground temperature, diameter and cover thickness of tunnel, cohesion and friction angle of frozen soil. In order to figure out the optimal grouting pressure, 4 groups of additional simulations are computed with the uniform face support pressure of 0.5, 1.0, 1.5 and 2.0 times of horizontal stress, σH, at tunnel axis. The qualitative analysis of heading face deformation mechanism and quantitative analysis between deformation profiles and influential factors were implemented to comprehend the heading face deformation characteristics when conducting pipe jacking tunnel in frozen ground. The results show that ground temperature plays a dominant role to control the face deformation of jacking tunnel in frozen ground. And, factors of tunnel diameter and cover thickness have relatively greater influence on the deformation regulation than that of shearing parameters of frozen soil, cohesion and friction angle. Finally, the minimum allowable slurry pressure for each simulation schemes are obtained, which may be used in construction the pipe jacking tunnel in frozen ground.

Almost all the coal is produced from open cut mines in Indonesia. As a consequence of open cut mine application, a great deal of coal is left out in the highwalls of the mined-out pits. Highwall mining systems can be used to recover this coal. The use of highwall mining systems has increasingly come into play in the US and Australia. However, it is not common in Indonesia. Moreover, Indonesia coal measure is categorized as weak geological condition. Some problems are likely to arise during the application of the highwall mining system for example instability of openings and highwalls due to the roof and pillar failures. Therefore, study of highwall mining system application in Indonesia is needed in order to increase the recovery rate of coal mining in Indonesia. This paper described the characteristics of the highwall mining system and discussed the appropriate highwall mining system application in weak geological condition, Indonesia. From the results of a series of laboratory tests and numerical analyses, it can be concluded that the stability of pillars and mine openings in auger mining systems is much higher than that in CHM and an auger mining system is suitable for such as very weak/poor strata conditions. Moreover, the application of backfilling system is very effective for improvement of the stability of pillar and openings.