Vol 42 Iss. 1

In the theory of coal mines design, more than 30 works, mainly performed during the last 30‑35 years, are devoted to determining the location of shafts under the conditions of cargo transportation. A large number of engineering, technical and scientific workers were engaged in solving this problem. Some researchers have considerably extended the problem of shaft location determination, including underground and surface coal transportation, delivery of rock, people, equipment, as well as ventilation, crosscut driving (in case of shafts positioned outside the suite), losses in pillars, and thus turned it into a typical method of variants solved graphically or by tabular comparison. During the famous discussion on the application of the analytical method in mining conducted by the Mining Journal in 1949, this question also attracted attention. At the same time, some participants of the discussion questioned the correctness of the existing methodology for determining the location of the shaft according to the transportation conditions.

One of the most difficult tasks in mining is the development of deposits under water bodies and large aquifers. Mining operations under water bodies are insufficiently studied. As a result, in some cases a lot of coal is lost in safety and barrier pillars, and in other cases improper mining leads to water breakthroughs causing disruption of operations and accidents with people. For example, the U.S. Bureau of Mines cites seven cases of catastrophic flooding of mines in Pennsylvania, mostly as a result of improper mining of barrier pillars. At the same time, according to the data of this bureau, with the correct choice of the mining system, methods of pillar mining and roof management, it was possible to recover up to 35% of the reserves lost in the barrier pillars, i.e. 250 million tons of anthracite. This would allow mines to prolong their operation for 5 years.

The generally accepted method of research of rock pressure manifestations, based on the analysis of average roof displacements at different distances from the face, does not provide an opportunity to assess the condition of the roof and to choose a rational characteristic of the support. According to this methodology, the movement of the roof rocks is conventionally recognized as continuous, and the measured changes in sinking velocities during individual operations are practically not used. Due to the unequal duration of existence of measuring points and different duration of individual operations, not only the character of the roof displacement curve (when moving away from the face) but also the final results are distorted.

In order to establish the interaction between the support and the roof, the author in different conditions in the mines of the Donetsk basin conducted studies of the loads perceived by the support and pliability of friction and hydraulic supports. The obtained materials were analyzed in accordance with the planned methodology.

Both in the USSR and abroad coal seams of gentle dip with a thickness more than 3,5-4,5 m are developed by inclined layers with caving, with stowing or in some cases (France, England) with smooth lowering of the upper layers and the roof of the seam. These systems are characterized by relatively low technical and economic indicators due to the high cost and labor intensity of the construction of stowing or preliminary support.

During the last decade flexible shields have become widespread in the development of steeply dipping coal seams in Kuzbass, where they successfully compete with other methods of development. Shield mining in the ore industry in many cases can replace low-productive and expensive mining by layer caving.

Further development of the coal industry is connected with the transition of a number of coal basins of the USSR to deeper horizons. The inevitable consequence of this is the deterioration of atmospheric conditions of mine workings, mainly due to the increase in gas content and temperature rise, in connection with which the fight against gas emission and the creation of normal working conditions in gas‑intensive stoping and preparation faces of coal mines are becoming more and more urgent. There is a necessity to determine the differential gas emission of the operating and newly opened horizons with its division into stoping faces, preparatory workings and mined‑out space. This is caused by the necessity to apply active methods of control and management of gas emission at deep horizons (methane degassing, changing the sequence of mining of adjacent seams, selection of ventilation schemes for excavation areas).

The Pechora Coal Research Institute has established that the mines of the Vorkutskoye field are characterized by high methane abundance, which increases with the transition of mining operations to deeper horizons. However, the methane content of the mine as a whole, determined by the mining‑statistical method, cannot characterize the value of methane emission, for example, in stoping and preparatory faces under different mining and geological conditions. For example, gas emission within the limits of a stoping area when developing a single seam, a seam with satellites or a seam developed in two seams will be sharply different. Because gas emission at greater depths will create significant difficulties in ventilating the workings and may require special measures to control gas emission, it is necessary to have a clear understanding of the nature of gas emission under different mining and geologic conditions.

Observations of the behavior of the side rocks and fastenings of preparatory workings in the conditions of Novovolynskoe coal deposit were carried out by a permanent group of DonUGI with the participation of the author and with the help of engineering and technical workers of mines and trust Novovolynskugol. In the coal-bearing strata of Novovolynskoe deposit there are four coal seams. Two of them, occurring at a depth of 350-375 meters, are working - Volynsky 1 (n7) with a thickness of 0.1-1.2 meters and Volynsky 2 (n3) with a thickness of 0.1-2.1 meters. Bedding of strata is almost horizontal, dip angle is 4-5°. The presence of a number of tectonic disturbances in the form of discharges with displacement amplitudes ranging from a few centimeters to 20 m was detected. Sandy and sandy-clay deposits are significantly developed in the rocks containing coal seams.

In the Soviet Union, the anchoring of large cross-section excavations in unlayered and strong, but highly fractured rocks with rod supports was initiated by A. I. Golomolzin in 1955. Under his leadership, a group of engineers (R. E. Vasilevsky, V. G. Dominikov and Y. K. Manenkov) secured several experimental sites in rocks with different degrees of fracturing. However, the theoretical underdevelopment of the issues of interaction between the rod support and rocks in workings conducted in strong, but highly fractured rocks leads quite often to mistrust and, as a result, to the rejection of the use of this new type of support.

Rod support has so far been mastered mainly on the basis of practical experience, since its theory is not sufficiently developed. The rods are passive suspensions only when the immediate roof is weak and the main roof rock is strong. In all other cases, the installation and fastening of rods in the rock change the bearing capacity of the roof, although the mechanical properties of samples made of rock, with rods fixed and without rods, of course, remain the same.

Among the many types of supports used in mining, rod support on the principle of action, simplicity of design, cost‑effectiveness and versatility takes a special place and every year is gaining more and more recognition in world mining practice. The Department of Construction of Mining Enterprises of the Leningrad Mining Institute studied rod support as a means of combating floor heave of the bottom of mine workings in the mines of Donbass, the Moscow region and other basins. It is established that rod support can often prevent floor heave or significantly reduce its intensity. Widely used constructions of rod supports for strong and medium strength rocks are not suitable for strengthening weak rocks, as they do not provide good anchoring and have low load‑bearing capacity. Foreign mining literature recommends widely used and cheap reinforced concrete rods for strengthening weak rocks. During experimental testing of these recommendations it turned out that not all designs of reinforced concrete rods are effective in weak rocks, especially in rocks that lose mechanical strength when moistened. As a result of studying the principal features of known designs of reinforced concrete rods and analyzing the schemes of their operation in specific mining and geological conditions, fundamentally new designs were developed.

Fastening of the conventional types of rods in the boreholes is carried out by wedging of the lock parts. Expansion of the lock requires mechanical efforts, and sufficient fixation of the lock in the rocks is not always possible.

The application of rod support to combat rock heave in mine workings in Donbass mines has shown that conventional rod locks do not provide reliable anchoring in clayey rocks. Their pull-out forces during tests did not exceed 1–1.5 tons, which is completely insufficient for metal rods. Therefore, to combat clay heave in workings of mines in the Moscow Basin, the authors have proposed a metal tubular rod support.

In the process of drilling mine shafts, the washing fluid is contaminated by the drilled rock, which leads to deterioration of its quality, reduction of drilling performance, and sometimes to complications and accidents. In order to restore the mud quality, it is necessary to continuously clean it from drilled rock. Among the various mechanisms and devices used for cleaning of flushing muds, the most widespread are vibrating screens.

When studying the vibrations of rigid reinforcement of mine shafts, it is necessary to distinguish two types of forces of interaction between the parachute device and conductors – braking and coupling. Under the braking force we will understand the force of interaction between parachutes and conductors when the lifting vessel is moving along the conductors, and under the coupling force we will understand the force of interaction between parachutes and conductors when the lifting vessel is stationary in relation to the conductors or moving at a very small speed. In the general case, the coupling force is not equal to the braking force. Thus, in the case of friction parachutes, the difference in the value of braking force and coupling force will be determined by the difference in the value of coefficients of kinetic and static friction. It should be borne in mind that static friction is 1.7‑2.8 times greater than kinetic friction and, consequently, the coupling force will be greater than the braking force.

As is known, rock loading in vertical shafts of mines, occupying 60‑80% of the time of the tunneling cycle, is the most labor‑intensive and difficult to mechanize process. Therefore, the choice of the most rational method of rock loading is of particular importance. The main criterion for this is usually the total productivity of machines for loading, i.e. only the reduction of time for loading and the increase in sinking speed achieved due to this are taken into account. This completely misses the cost of rock loading, which has a significant impact on the cost of shaft construction.