Vol 57 Iss. 1

The life of B. V. Bokii is firmly connected with the Leningrad Mining Institute, where he entered in 1921 after serving in the Red Army. ...

This paper considers the stress state of viscoelastic rocks in the vicinity of a supported mine workings with a contour having technological irregularities. In solving this problem, the stress strain state around an unsupported mine working is used. Therefore, let us first consider the solution of this problem...

The stability of the rock massif when driving a group of mine workings depends to a great extent on their mutual location and the ratio of cross sectional dimensions, as well as on the nature of the application of loads. Below we consider the results of the study of the stressed state of rocks during the driving of two vertical workings at the Tekeli mine, where the shaft with a diameter of 6.7 m and the ore pass with a diameter of 3.4 m were driven at a distance between their axes of 15 m to the depth of horizon X (572 m). In unstable rocks the mutual influence of these workings on each other leads to significant deformations of rocks and their falls.

The zone of inelastic deformations resulting from the driving of mine workings, as is known, is characterized in most cases by cracking developing over time. On the other hand, after a certain period of time, the fracturing of rocks may decrease due to their compaction, i.e. the closure of cracks.

To fully characterize rocks, it is necessary to take into account their rheological properties, one manifestation of which is creep. Creep testing of rocks under uniaxial compression requires the use of rather heavy equipment-special spring presses. The strain is measured between the ends of the rock specimen, while a homogeneous stress state is observed only in the middle part along its height. In the study of rock creep by bending test of the specimen, simpler setups are used and at a certain ratio of the span of the specimen to its height, the stress state can be considered close to homogeneous.

In all real bodies, deformation processes proceed in time. Mechanical characteristics that do not take time into account are approximate and in many cases prove to be inadequate. Factors such as the rate and duration of loading , the energy stored by the system , have a great influence on the value of the breaking stress. In the LGI a lot of work has been done to establish the physical features of the process of rock fracture by impact loading and to find out the comparative efficiency of static and dynamic methods of fracture. As a result, it was found that the fracture energy, other things being equal, is a function of the impact energy, the rate of impact load application, and the mass of the impacting body. As for the comparison of energy expenditure in static and impact tests, it is permissible only at a comparable degree of crushing, i.e. at some minimum impact work.

The geological section of the roof of the Leningrad deposit is rather complex and is characterized by a great variety of lithological types (from weak clay shales to very strong limestones), their variability in terms of distance from the commercial seam roof and thickness. The most dangerous in terms of the possibility of collapse is a 90‑centimetre layer of rocks, counting from the roof of the commercial seam, the section of which is the most complex and can be summarised in the following form.

In determining the strength of rocks, a scale effect or scale factor – the dependence of test results on the size of samples – has been found. Experimental studies of the manifestation of this phenomenon lead to contradictory results, and some authors generally deny its existence.

Conducting mine workings in weak unstable rocks is difficult and requires the use of special methods-freezing of rocks, injection of fixing substances, dewatering. The greatest difficulties for mining operations are clay water-saturated rocks with weak filtration capacity, which cannot be hardened by most of the widespread methods. At present, mining in such conditions is carried out mainly with the use of freezing of rocks.

During the development of thick coal seams of the Prokopyevsk‑Kiselevsky region of Kuzbass, a large number of geological disturbances (20‑60% of the total area of mined seams) are encountered, and the relative development of disturbances increases with increasing depth of mining. When driving workings in such zones, large coal falls occur, which increases the danger of operations, significantly complicates driving and subsequent extraction. Coal in the disturbed zones and at some distance from them has low strength (less than 10 kg/cm²). The zone of weakened rocks and coal of the disjunctive disturbance usually extends for 12‑15 m in each direction. Mine workings in the zones of disturbances have to be repeatedly re‑supported, which leads to inevitable increase in the cost of operations. Strengthening of the coal massif by injection of chemical solutions into it will increase the stability of coal exposures in development workings. In this way it is possible to strengthen thinly fractured rocks with a filtration coefficient less than 1 m/day. Cementation in these cases does not give a positive effect.

Wooden rods, which usually have a wedge‑slot construction and possess low weight, low cost, ease of manufacture, non‑scarcity of material, are still rarely used. The reasons for this are the low bearing capacity of the deep lock of such rods, not exceeding 1‑1.5 t, short service life and low resistance to shear forces. ...

Recently, reinforced concrete rod support has become widespread in the mining industry. Unlike metal wedge‑slot and expansion rods, reinforced concrete rods are characterized by a strong contact with the rock along the entire length of the borehole, which provides greater reliability and stability of the excavation contour. In addition, reinforced concrete rods work better in the zone of influence of mass blasts, which explains their distribution mainly in the ore industry.

The proposed calculation method is valid for spacer-type rod locks with non-linked spacer segments. However, similar relationships can be derived for other designs of metal rod locks of the spacer type. In this case, the actual contact surface of the rod lock with the rock, which has a curvilinear outline, should be reduced to an equivalent rectangular surface in plan, and half the length of this fictitious surface should be substituted into formula (6) instead of half the length of the die surface.

The method of calculation is based on the fact that the size of the stope is determined by the ultimate stability of the roof after the end of extraction work in the stope. Consequently, the system and sequence of stope mining do not affect the proposed method of calculation.

The main parameters that need to be determined when studying the manifestations of rock pressure in shafts are the magnitude of the load on the support, the nature of its change in time and distribution along the contour of the cross‑section of the shaft, as well as the magnitude of displacements of the surrounding rock mass. To determine these parameters, the best known is the VNIMI method, which is based on the use of special dynamometric support. More than ten years of experience in the use of such support has proved the serviceability of its main components. However, the widespread use of the support in research work is constrained, because most of the devices and equipment used in it are not mass‑produced, and their manufacture in local workshops is virtually impossible. In this regard, the LGI developed a simplified design of dynamometric support and measuring devices, which are characterized by simplicity and can be manufactured in mine workshops. This allows organizing instrumental observations on a much wider scale and obtaining information on the manifestations of rock pressure in the volume necessary for solving a number of practical problems.

The practice of construction of vertical shafts by the drilling and blasting method in the conditions of the combine “Rostovshahtostroy” shows that the improvement of parameters of drilling and blasting works allows a reduction of the deviation of actual cross‑sections from the design. At the same time this is not accompanied by a reduction of irregularities of the rock walls of the shaft, and the destruction of rocks in exposures remains still significant.

Since 1966, the Leningrad Mining Institute together with Rostovshahtostroy Combine have been conducting research to determine the actual over-excavation of rock during the sinking of vertical shafts and to find the optimal parameters of contour blasting, contributing to the reduction of excessively excavated rock.

Despite the fact that nowadays hoisting vessels are equipped with safety catches with brake ropes, there are still many mine hoisting vessels equipped with safety catches which, when triggered, grip the guides.