Archives of Mining Sciences

The paper presents a simulation analysis of four control systems of the raw coal feed to a jig: stabilization of the volumetric flow of the feed, stabilization of the feed tonnage, stabilization of the feed flow with the additional measurement of the feed bulk density or the additional measurement of ash content in the feed. Analysis has been performed for the first and second compartments of a jig. The aim of the feed control was to stabilize the mass of the bed in the zone where the material stratifies; the mass may change due to changes in the washability characteristics of the feed. Such control should result in stable conditions in which material loosens during subsequent media pulsation cycles; stabilizing conditions minimizes the dispersion of coal particles in the bed. The best results have been achieved for the system of feed control where the ash content was measured in the first compartment, and for feed tonnage control in the second compartment.

With the rise of coal mine underground reservoir engineering in the Shendong Mining Area, the space time dynamic evolution prediction of storage coefficient is becoming one of the critical technical problems for long-term reservoir operation. This coefficient directly determines the storage capacity and the comprehensive benefits of the operation of a coal mine underground reservoir. To this end, the proposed underground reservoir in Daliuta coal mine (No. 22616 working face) is selected in this study for the development and application of an experimental device to measure the storage coefficient. Rock and coal fragments from similar materials are prepared, which are filled and loaded according to the caving rock nature as well as the lumpiness and accumulation mode characteristics pertaining to No. 22616 working face. Subsequently, the measured storage coefficient under circulating water injection conditions revealed a four-dimensional spatial and temporal pattern. It followed the law of storage coefficient under joint interaction of water-rock and stress. The results showed that, prior to the experiment, rock and coal fragments made from similar materials had good water resistance when the paraffin content was set at 8%. The three stress zones were defined based on a theoretical analysis, which were applied on the corresponding loads. During the experiments, significant regional differences were found in the top surface with persisting subsidence of each stress loading zone. Hence, compared with its initial state, the maximum subsidence in the stress stability zone, the stress recovery zone, and the low-stress zone was 7.89, 5.8, and 1.83 mm, respectively. While the storage capacity and the storage coefficient gradually decreased, the former ranged from 0.2429 to 0.2397 m3, and the latter ranged from 0.270 to 0.266. The experimental results are verified by drainage engineering tests in the Shendong Mining Area. In essence, the storage coefficient had remarkable spatial distribution characteristics and a time-varying effect. In space, the storage coefficient increased with height along the vertical direction of the coal mine underground reservoir. However, it decreased with the distance from the boundary of the dam body in the horizontal direction. With time, the storage coefficient decreased dynamically. This study provides a new way of predicting the storage coefficient of a coal mine underground reservoir.

Mining the lower seams in a sequence of shallow, closely spaced coal seams causes serious air leakage in the upper goaf; this can easily aggravate spontaneous combustion in abandoned coal. Understanding the redevelopment of fractures and the changes in permeability is of great significance for controlling coal spontaneous combustion in the upper goaf. Based on actual conditions at the 22307 working face in the Bulianta coal mine, Particle Flow Code (PFC) and a corresponding physical experiment were used to study the redevelopment of fractures and changes in permeability during lower coal seam mining. The results show that after mining the lower coal seam, the upper and lower goafs become connected and form a new composite goaf. The permeability and the number of fractures in each area of the overlying strata show a pattern of „stability-rapid increase-stability“ as the lower coal seam is mined and the working face advances. Above the central area of goaf, the permeability has changed slightly, while in the open-cut and stop line areas are significant, which formed the main air leakage passage in the composite goaf.

Time-dependent behavior of rock mass is important for long-term stability analysis in rock engineering. Extensive studies have been carried out on the creep properties and rheological models for variable kinds of rocks, however, the effects of initial damage state on the time-dependent behavior of rock has not yet been taken into consideration. In the present study, the authors proposed a creep test scheme with controlled initial damage to investigate the influence of initial damage on the time-dependent behavior of sandstone. In the test scheme, the initial states of damage were first determined via unloading the specimen from various stresses. Then, the creep test was conducted under different stress levels with specific initial damage. The experimental results show that there is a stress threshold for the initial damage to influence the behavior of the rock in the uniaxial compressive creep tests, which is the stress threshold of dilatancy of rock. When the creep stress is less than the stress threshold, the effect of the initial damage seems to be insignificant. However, if the creep stress is higher than the stress threshold, the initial damage has an important influence on the time-dependent deformation, especially the lateral and volumetric deformation. Moreover, the initial damage also has great influence on the creep failure stress and long-term strength, i.e., higher initial damage leading to lower creep failure stress and long-term strength. The experimental results can provide valuable data for the construction of a creep damage model and long-term stability analysis for rock engineering.

Scientific research discussed in the present article is focused on the determination of the vertical conveyance capacity in the process of mining minerals, while applying a mathematical calculation and verification of the calculation results by simulation. Input parameters for the capacity calculation include the transport cycle time. The article presents the results of measuring a transport cycle during the operation and a calculation of the transport cycle while using known formulas. On the basis of the observed findings, two methods of increasing the hoisting machine capacity were proposed. The first method is increasing the velocity from the original value of 6 m.s–1 to the velocity of 7 m.s–1. In this case, we achieved the daily capacity increase in 2-2.5%. The second method consisted in changing the hoisting machine acceleration and deceleration modes by which we achieved as much as 9% increase in the daily capacity. The article also describes a transport cycle simulation model, with its output being the number of work cycles parameter. The obtained parameter was used again in the capacity calculation. The simulation model was used in experiments for both, the current status as well as proposed solutions. The simulation model serves also for calculation verification.

The blasting technique is currently the basic excavation method in Polish underground copper mines. Applied explosives are usually described by parameters determined on the basis of specific standards, in which the manner and conditions of the tests performance were defined. One of the factors that is commonly used to assess the thermodynamic parameters of the explosives is the velocity of detonation. The measurements of the detonation velocity are carried out according to European Standard EN ­13631-14:2003 based on a point-to-point method, which determines the average velocity of detonation over a specified distance. The disadvantage of this method is the lack of information on the detonation process along the explosive sample. The other method which provides detailed data on the propagation of the detonation wave within an explosive charge is a continuous method. It allows to analyse the VOD traces over the entire length of the charge. The examination certificates of a given explosive usually presents the average detonation velocities, but not the characteristics of their variations depending on the density or blasthole diameter. Therefore, the average VOD value is not sufficient to assess the efficiency of explosives. Analysis of the abovementioned problem shows, that the local conditions in which explosives are used differ significantly from those in which standard tests are performed. Thus, the actual detonation velocity may be different from that specified by the manufacturer. This article presents the results of VOD measurements of a bulk emulsion explosive depending on the diameter of the blastholes carried out in a selected mining panel of the Rudna copper mine, Poland. The aim of the study was to determine the optimal diameter of the blastholes in terms of detonation velocity. The research consisted of diameters which are currently used in the considered mine.

The article describes mine survey works during opening old St. Anthony of Padua water adit in Horní Město (Czech Republic) to make it accessible to visitors. The works cover the connecting survey and orientation measurement, traverse measurement of the first opened part, setting-out projection of the end of opened part to the surface to make shaft from the surface, new connecting survey and orientation measurement by shaft and traverse measurement of the rest of water adit. Non-standard aids and techniques were used during surveying. One of the tools is a suspended prism holder developed at Institute of geodesy and mine surveying, VSB – Technical university of Ostrava, registered as a utility patent.

The mining of hard coal deposits at increasingly greater depth leads to an increase in hazards related to the loss of stability of steel arch supports as a result of excessive static and dynamic loads. Camber beam reinforcement via rockbolting is often utilised in order to improve the stability of the yielding steel arch support.

This article presents the results of comparative bench tests of the ŁP10/V36-type steel arch support, tested with and without reinforcement by means of self-drilling bolts with drunken R25 threads, using short joists formed from V32 and V25 sections. It also presents the results of comparative tests of the ŁPP10/4/V29/I-type steel arch support, tested with and without reinforcement by means of rock bolts with trapezoidal Tr22/13 threads, using short joists formed from V25 sections. The obtained test results, in the form of load courses and work values of the steel arch and mixed (arches and rock bolts) support systems, demonstrate that the utilisation of mixed support may significantly improve the stability of workings, particularly immediately after they are driven. A mixed support system quickly achieves its maximum load capacity together with a significant increase in its work value. It may thus prevent the stratification of the rocks surrounding the working, and therefore better utilise the self-supporting capacity of the rock mass. As evidenced by the test results, the mixed support work may be as much as 3.5 times as great compared to the steel arch support at the beginning of the height reduction process initiated by loading – i.e. until its reduction by a presupposed value of 100 mm.

The stability of gateroads is one of the key factors for the mining process of hard coal by a longwall system. Wrong designed and applied the gateroad support at the stage of drilling, may adversely affect the functionality of the gateroad and the safety of the crew throughout its existence.

The article presents the results of the underground tests and observations such as: convergence of the gateroad, stratification and the fractured zone range in the roof rocks, carried out in four longwall gateroads at the stage of their drilling.

The obtained test results were the basis for the assessment of the possibility of using a convergence control method in the design of the gateroad support. The method is based on three interdependent relationships, such as: Ground Reaction Curve (GRC), Longitudinal Displacement Profile (LDP), and a Support Characteristic Curve (SCC). All calculations were performed using numerical modeling in the Phase2 program, based on the finite element method (FEM).

In order to provide sufficient cooling capacity for working and heading faces of the coal mine, chilled water is often transported a long distance along pipelines in deep mine, which inevitably results in its temperature rising owing to heat transfer through pipe wall and the friction heat for flow resistance. Through theoretical models for temperature increasing of the chilled water were built. It is pointed out that the temperature rising of the chilled water should be considered as a result of the synergy effects of the heat transfer and the friction heat, but theoretical analysis shows that within engineering permitting error range, the temperature increasing can be regarded as the sum caused by heat transfer and fraction heat respectively, and the calculation is simplified. The calculation analysis of the above two methods was made by taking two type of pipe whose diameters are De273 × 7 mm and De377 × 10 mm, with 15 km length in coal mine as an example, which shows that the error between the two methods is not over 0.04°C within the allowable error range. Aims at the commonly used chilled water diameter pipe, it is proposed that if the specific frictional head loss is limited between 100 Pa/m and 400 Pa/m, the proportion of the frictional temperature rising is about 24%~81% of the total, and it will increase with high flow velocity and the thin of the pipe. As a result, the friction temperature rising must not be ignored and should be paid enough attention in calculation of the chilled water temperature rising along pipe.

This study attempts to evaluate the field performance of various mining equipment used at the development galleries of coal mines. These are hand-held and jumbo rock drills, and a roadheader used in mechanical excavation. For this purpose, the penetration rates of rock drills were monitored and measured in the field. The physical, mechanical, and drillability properties were determined through the collected samples in order to understand the complex interactions between the rock and bit/pick. The abrasive mineral content was also analyzed with XRD analysis to examine the wear on the cutting/drilling tools. Besides, the specific energy of the equipment was calculated relying on the operational parameters. A comparison of the monthly advance and production rates of the drilling rigs and roadheader was made. The relations among operating power, specific energy, and design of buttons/picks were investigated. It has been found that the average advance and production rates of the mining equipment are consistent with the penetration rate. The results verified that the roadheader used in mechanical excavation and the jumbo drill used in drilling and blasting technique are the machines maximizing the advance and production rates.

The underground complicated testing environment and the fan operation instability cause large random errors and outliers of the wind speed signals. The outliers and large random errors result in distortion of mine wind speed monitoring, which possesses safety hazards in mine ventilation system. Application of Kalman filter in velocity monitoring can improve the accuracy of velocity measurement and eliminate the outliers. Adaptive Kalman Filter was built by automatically adjusting process noise covariance and measurement noise covariance depending on the differences between measured and expected speed signals. We analyzed the fluctuation of airflow flow using data of wind speed flow and distribution characteristics of the tunnel obtained by the Laser Doppler Velocimetry system (LDV) studies. A state-space model was built based on the tunnel airflow fluctuations and wind speed signal distribution. The adaptive Kalman Filter was calculated according to the actual measurement data and the Expectation Maximization (EM) algorithm. The adaptive Kalman filter was used to shield fluid pulsation while preserving system-induced fluctuations. Using the Kalman filter to treat offline wind speed signal acquired by LDV, the reliability of Kalman filter wind speed state model and the characteristics of adaptive Kalman Filter were investigated. Results showed that the adaptive Kalman filter effectively eliminated the outliers and reduced the root-mean-squares error (RMSE), and the adaptive Kalman filter had better performance than the traditional Kalman filter in eliminating outliers and reducing RMSE. Field experiments in online wind speed monitoring were conducted using the optimized adaptive Kalman Filter. Results showed that adaptive Kalman filter treatment could monitor the wind speed with smaller RMSE compared with LVD monitor. The study data demonstrated that the adaptive Kalman filter is reliable and suitable for online signal processing of mine wind speed monitor.

Influenced by the dynamic pressure of the front abutment pressure and the lateral abutment pressure, large deformation of surrounding rock occurs advancing working face in the entry heading adjacent to the active longwall mining face. Based on the cause analysis of entry large deformation, a new technology was put forward to solve the problem, and the designing method of drilling hole parameters for directional hydraulic fracturing was formed. Holes are drilled in the entry or in the high drainage entry to a certain rock layer over the adjacent working face, hydraulic cutting or slotting at the bottom of a borehole were also applied in advance to guide the hydraulic fractures extend in expected direction, through which the hard roof above the coal pillar can be cut off directionally. As a result, the stress concentration around the entry was transferred, and the entry was located in a destressing area. The field test at Majialiang coal mine indicates that the propagation length of cracks in single borehole is more than 15 m. After hydraulic fracturing, the large deformation range of the entry is reduced by 45 m, the average floor heave is reduced by 70%, and the average convergence of the entry’s two sides is reduced by 65%. Directional hydraulic fracturing has a better performance to control the large deformation of the dynamic pressure of the entry heading adjacent to the advancing coal face. Besides, it can improve the performance of the safety production.