The frictional resistance coefficient of ventilation of a roadway in a coal mine is a very important technical parameter in the design and renovation of mine ventilation. Calculations based on empirical formulae and field tests to calculate the resistance coefficient have limitations. An inversion method to calculate the mine ventilation resistance coefficient by using a few representative data of air flows and node pressures is proposed in this study. The mathematical model of the inversion method is developed based on the principle of least squares. The measured pressure and the calculated pressure deviation along with the measured flow and the calculated flow deviation are considered while defining the objective function, which also includes the node pressure, the air flow, and the ventilation resistance coefficient range constraints. The ventilation resistance coefficient inversion problem was converted to a nonlinear optimisation problem through the development of the model. A genetic algorithm (GA) was adopted to solve the ventilation resistance coefficient inversion problem. The GA was improved to enhance the global and the local search abilities of the algorithm for the ventilation resistance coefficient inversion problem.
Mining ventilation should ensure in the excavations required amount of air on the basis of determined regulations and to mitigate various hazards. These excavations are mainly: longwalls, function chambers and headings. Considering the financial aspect, the costs of air distribution should be as low as possible and due to mentioned above issues the optimal air distribution should be taken into account including the workers safety and minimization of the total output power of main ventilation fans. The optimal air distribution is when the airflow rate in the mining areas and functional chambers are suitable to the existing hazards, and the total output power of the main fans is at a minimal but sufficient rate. Restructuring of mining sector in Poland is usually connected with the connection of different mines. Hence, dependent air streams (dependent air stream flows through a branch which links two intake air streams or two return air streams) exist in ventilation networks of connected mines. The zones of intake air and return air include these air streams. There are also particular air streams in the networks which connect subnetworks of main ventilation fans. They enable to direct return air to specified fans and to obtain different airflows in return zone. The new method of decreasing the costs of ventilation is presented in the article. The method allows to determine the optimal parameters of main ventilation fans (fan pressure and air quantity) and optimal air distribution can be achieved as a result. Then the total output power of the fans is the lowest which makes the reduction of costs of mine ventilation. The new method was applied for selected ventilation network. For positive regulation (by means of the stoppings) the optimal air distribution was achieved when the total output power of the fans was 253.311 kW and for most energy-intensive air distribution it was 409.893 kW. The difference between these cases showed the difference in annual energy consumption which was 1 714 MWh what was related to annual costs of fan work equaled 245 102 Euro. Similar values for negative regulation (by means of auxiliary fans) were: the total output power of the fans 203.359 kW (optimal condition) and 362.405 kW (most energy-intensive condition). The difference of annual energy consumption was 1 742 MWh and annual difference of costs was 249 106 Euro. The differences between optimal airflows considering positive and negative regulations were: the total output power of fans 49.952 kW, annual energy consumption 547 MWh, annual costs 78 217 Euro.
This paper describes the concept of controlling the advancement speed of the shearer, the objective of which is to eliminate switching the devices off to the devices in the longwall and in the adjacent galleries. This is connected with the threshold limit value of 2% for the methane concentration in the air stream flowing out from the longwall heading, or 1% methane in the air flowing to the longwall. Equations were formulated which represent the emission of methane from the mined body of coal in the longwall and from the winnings on the conveyors in order to develop the numerical procedures enabling a computer simulation of the mining process with a longwall shearer and haulage of the winnings. The distribution model of air, methane and firedamp, and the model of the goaf and a methanometry method which already exist in the Ventgraph-Plus programme, and the model of the methane emission from the mined longwall body of coal, together with the model of the methane emission from the winnings on conveyors and the model of the logic circuit to calculate the required advancement speed of the shearer together all form a set that enables simulations of the control used for a longwall shearer in the mining process. This simulation provides a means for making a comparison of the output of the mining in the case of work using a control system for the speed advancement of the shearer and the mining performance without this circuit in a situation when switching the devices off occurs as a consequence of exceeding the 2% threshold limit value of the methane concentration. The algorithm to control a shearer developed for a computer simulation considers a simpler case, where the logic circuit only employs the methane concentration signal from a methane detector situated in the longwall gallery close to the longwall outlet.