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Ventilation plays a key role in ensuring safe exploitation in underground gassy mines. Over the years, the structure of a mine’s ventilation network changes. Therefore, it becomes necessary to construct new excavations, while some existing excavations lose their potential for future mining activity. Constructing new excavations, especially shafts, is very expensive. Therefore, mine operators are looking for solutions to ensure appropriate ventilation by reorganizing the ventilation network and using existing infrastructure, including shafts. This article presents the example of a coal mine located in the Upper Silesian Coal Basin in Poland to discuss the factors relevant to switching the function of one of the central shafts from a downcast function to an upcast one. This change is accompanied by the closure of a peripheral upcast shaft. The main aim of this change is to assess the possibility of further safe operation without the construction of new shafts. This action also results in the release of the coal currently closed in the pillar of the shaft being closed. Using a numerical model of the mine ventilation network allowed for the comparison of the considered solutions before making final decisions and implementing changes in the network. The calculations showed that it is possible to provide appropriate ventilation in the mine, but it would need to take into account certain technological assumptions, like the additional technical function of the changed shaft. This article discusses the advantages and disadvantages of modifications to the mine ventilation network, as well as their guiding principles, in the context of existing methane hazards. The procedure presented in this article can be adopted in other mine ventilation networks in which analogous modifications are considered.

The determination of natural gas content in coal seams is essential to safe mining activities and to estimate gas resources. It is used to estimate methane emissions into workings, which is very important when using preventive measures of methane ignition and explosion. Methane content in a coal seam is mostly determined using a direct method. However, a problem with estimation of gas loss during collecting a sample in a mine occurs in each method. That gas loss is estimated using different procedures. In the Polish mining industry, methane content in a coal seam is mostly determined during roadway advancing faces in coal seams. The differences in procedures for determining methane content in coal in various laboratories gave rise to starting research into developing a new method. This article presents results of study of gas loss unification for the determination of methane content in coal seams. The research focused mainly on determining a uniform method for estimating gas loss using the early adsorption diffusion process. The method of coal sample saturation with methane was used to measure gas desorption rate. Saturation pressures of samples were determined for coal seams in the Upper-Silesian Coal Basin. The analysis of desorption results allowed to determine the dependence between methane content in a coal sample and the initial methane content in a coal seam. It was concluded that the desorption rate was highest within two minutes. For the procedure of coal sampling, gas loss was determined at 12% of both desorbed and residual gases. The presented research results gave rise to developing a uniform method for determining gas content used in Polish coal mining. The method was accepted as the national standard. Methodology using a new gas loss estimation procedure allows for achieving relatively fast results of the determination of methane content in coal seams, which is very important, especially in gassy mines.

Ventilation plays a key role in underground mining. It is essential due to the natural hazards and technological processes that come with the nature of mining. However, it is highly energy consuming and generates significant operating expenditures. Fan station parameters are selected based on the needs of a particular mine but mainly consider the requirements for the period of developed mining activities. When the period of mine decommissioning begins, the parameters of the main fan station often exceed its needs. In Poland, many mines have been closed in recent years. However, sometimes, due to the necessity of pumping underground water, it cannot be done thoroughly. In such a situation, it usually turns out that the parameters of the existing fan station significantly exceed the mine’s needs. The main fan stations are devoid of control systems, and even if they have them, they do not allow for a significant reduction of their volume flow rate. Modernising of the station to meet new requirements of the mine is expensive and time consuming. Solving the presented problem is possible by developing a fan station to replace main fans that are too big. The idea is easy to implement and consists of connecting it to an existing upcast shaft or downcast shaft, which will then be changed to upcast. The solution presented in the article has been implemented in two Polish coal mines and is in progress in a third mine. The examples presented in the article clearly show the energy benefits of replacing main fans that are too large.

The gravimetric method for determining coal gas sorption has many advantages and limitations. The article presents the influence of various factors on the results of methane sorption on coal. In mining practice, in addition to sorption properties of coal, knowledge of methane sorption capacity and effective diffusion coefficient determined when assuming a unipore sorption/desorption model are crucial for predicting sudden releases of methane from coal seams to a mine ventilation environment. In Poland, determining sorption capacities of coals for methane is mandatory when starting mining operations in new parts of coal deposits threatened by outbursts. Traditionally, gravimetric microbalances, such as intelligent gravimetric analysis (IGA), are used to determine adsorption capacity and desorption rate. Recently, newer microbalances XEMIS have been introduced to the market. Two gas laboratories, AGH in Krakow and CLP-B in Jastrzebie-Zdroj, respectively, compared experimental adsorption isotherms using XEMIS microbalances with mutually exchanged coal samples. Both sorption capacity at the pressure of 1 bar (a1bar) and effective diffusion coefficient (De) were independently determined for the coal samples tested. The results obtained are comparable despite the use of different microbalance XEMIS models. The conducted studies and comparative evaluation of the results allowed for assessing procedures for determining sorption properties using XEMIS microbalances. The exchange of laboratory experiences also allowed for the identification of methodology factors crucial for the development of a uniform procedure for conducting similar studies with XEMIS microbalance. The proposed factors for testing the sorption behavior of methane in coal structures may be helpful in mining practice.

The determination of methane content of coal seams is conducted in hard coal mines in order to assess the state of methane hazard but also to evaluate gas resources in the deposit. In the world’s mining industry, natural gas content in coal determination is usually based on direct methods. It remains the basic method in Poland as well. An important element in the determination procedure is the gas loss that occurs while collecting a sample for testing in underground conditions. In the method developed by the authors, which is a Polish standard, based on taking a sample in the form of drill cuttings, this loss was established at a level of 12%. Among researchers dealing with the methane content of coal, there are doubts related to the procedures adopted for coal sampling and the time which passes from taking a sample to enclosing it in a sealed container. Therefore, the studies were designed to evaluate the degree of degassing of the sample taken in the form of drill cuttings according to the standard procedure and in the form of the drill core from a coal mine roadway. The results show that the determinations made for the core coincide with the determinations made for the drill cutting samples, with the loss of gas taken into account.

An overlap auxiliary ventilation system is very often used for driving roadways in methane-rich coal seams. An overlap zone between the outlets of the forcing duct ends with a whirl flow air-duct (WFAD) and the exhaust duct ends with a dust scrubber that is created by applying the overlap system. This study examines the distribution of methane concentrations at various distances in the overlap zone. Maintaining a long overlap zone could increase the advance of the face. Therefore, the impact of overlap zone length on the methane concentration distribution, in and beyond the overlap zone, is investigated. The evaluation of methane concentrations is performed utilizing a well-established computational fluid dynamics (CFD) approach. The mathematical model of methane emissions into the roadway is adopted. Moreover, the CFD model is validated. A vortex of the return air, caused by the free airstream flowing out of the dust scrubber, is found. This air vortex is responsible for higher methane concentrations at the end of the overlap zone. Therefore, the conclusion can be drawn that maintaining the length of the overlap zone at 5 m to 10 m should be done to control permissible methane concentrations.

Transporting materials and mine staff is a vital link necessary to the production process in underground mines. Deteriorating climatic conditions, mainly due to the increasingly deep mining and the usage of machines, force us to look for solutions to improve the underground mine environmental situation. Another essential factor responsible for deteriorating working conditions is harmful substances and exhaust fumes emitted from diesel engines. Supplying the workplaces with air quantity exceeding requirements such as the minimum velocity of air movement or gas and climatic conditions will allow for maintaining the gas concentration at the appropriate level. One possible way to solve the problems mentioned above is to replace suspended monorails powered by internal combustion engines with new solutions of electrically battery-powered monorails. Electric monorails are not yet widely used in mines; nevertheless, they have many advantages. This article analyzes the exhaust gas parameters from monorail locomotives operating in a hard coal mine and determines the required airflow to maintain permissible concentrations of harmful gases. It also focuses on a comparative analysis of climatic conditions in the development heading, considering the roadway’s functioning with and without using diesel or electric monorail. The study consists of the methodology for predicting climate conditions. Based on the performed analysis, it was shown that using electric monorails could significantly improve working conditions.

Heat stress in deep hot mines is a factor that often determines the possibility of technical mining of natural resources. One of the solutions enabling miners to work in such mines is air cooling. Cooling systems vary, and their selection depends on the type of mine and the mining methods used. Limited air cooling capabilities exist in electric-powered coal mines. The main solution for air cooling is based on movable spot air coolers. Such systems commonly use surface or underground refrigeration plants. An underground refrigeration plant (URP) equipped with compressor chillers does not achieve more than 2.5–3.0 MW of cooling capacity due to the limited heat rejection capacity of return air streams in a typical coal mine. The method discussed in this paper, using mine water to discharge waste heat from the underground refrigeration plant, provides a measurable benefit for optimizing the mine air cooling system. The main purpose of this research is to study the feasibility and effect of water diversion from the actual mine drainage system to the underground refrigeration plant. The water drainage system in an underground mine is called the dewatering system of the mine. The heated water in the condensers of the chillers is directed back to the mine’s central dewatering system. The recovery from water discharged to the surface contributes to optimising energy consumption for a mine air cooling and the sustainable discharge of wastewater. In addition, using the total water flow from the mine dewatering system to reject heat in compressor chillers, compared with the traditional solution, can improve the cooling capacity of URP. These findings may provide beneficial guidance for practical applications in deep hot mines with small natural water inflow.

Mining progress in underground mines cause the ongoing movement of working areas. Consequently, it becomes necessary to adapt the ventilation network of a mine to direct airflow into newly-opened districts. For economic reasons, opening new fields is often achieved via underground workings. Length of primary intake and return routes increases and also increases the total resistance of a complex ventilation network. The development of a subsurface structure can make it necessary to change the air distribution in a ventilation network. Increasing airflow into newly-opened districts is necessary. In mines where extraction does not entail gas-related hazards, there is possibility of implementing a push-pull ventilation system in order to supplement airflows to newly developed mining fields. This is achieved by installing subsurface fan stations with forcing fans at the bottom of downcast shaft. In push-pull systems with multiple main fans, it is vital to select forcing fans with characteristic curves matching those of the existing exhaust fans to prevent undesirable mutual interaction. In complex ventilation networks it is necessary to calculate distribution of airflow (especially in networks with a large number of installed fans). In the article the influence of applying additional forcing fans for the air distribution in ventilation network for underground mine were considered. There are also analysed the extent of overpressure caused by the additional forcing fan in branches of the ventilation network (the operating range of additional forcing fan). Possibilities of increasing airflow rate in working areas were conducted.

Mining progress in underground mines cause the ongoing movement of working areas. Consequently, it becomes necessary to adapt the ventilation network of a mine to direct airflow into newly-opened districts. For economic reasons, opening new fields is often achieved via underground workings. Length of primary intake and return routes increases and also increases the total resistance of a complex ventilation network. The development of a subsurface structure can make it necessary to change the air distribution in a ventilation network. Increasing airflow into newly-opened districts is necessary. In mines where extraction does not entail gas-related hazards, there is possibility of implementing a push-pull ventilation system in order to supplement airflows to newly developed mining fields. This is achieved by installing subsurface fan stations with forcing fans at the bottom of downcast shaft. In push-pull systems with multiple main fans, it is vital to select forcing fans with characteristic curves matching those of the existing exhaust fans to prevent undesirable mutual interaction. In complex ventilation networks it is necessary to calculate distribution of airflow (especially in networks with a large number of installed fans). In the article the influence of applying additional forcing fans for the air distribution in ventilation network for underground mine were considered. There are also analysed the extent of overpressure caused by the additional forcing fan in branches of the ventilation network (the operating range of additional forcing fan). Possibilities of increasing airflow rate in working areas were conducted.