SOME EXPERIENCE IN SOLVING THE PROBLEM OF ENCOUNTERING COAL SEAMS UNDER KUZBASS CONDITIONS

S. S. Sulakshin, V. G. Khramenkov, V. P. Rozhkov

(Presented by the scientific seminar of the Department of Exploration Techniques)

The quality and reserves of the coals of most coal deposits are determined mainly from core samples. Obtaining reliable data on these most important questions in the exploration of mineral deposits depends to a significant extent on the quantity and quality of the core obtained during drilling, as well as on the accuracy with which the boundaries (thickness) of coal seams are determined. For most deposits with low-stability, weak coals, especially those of complex structure, means that guarantee a full solution to the problem of sampling have practically not yet been developed. This is connected chiefly with the absence of reliable means for recording the moment at which a coal seam is encountered, and with the absence of tools of rational design for obtaining high-quality core samples in quantities sufficient for investigation. In this connection, in most cases the planned core recovery is 55–60%, while in practice even this value is often not attained. At the same time, cases are not rare in which seams are missed or incorrect information is obtained about their thickness. Thus, the investigation and development of means ensuring complete sampling of hard-coal deposits is one of the most important problems requiring the quickest possible solution.

The development of means for high-quality sampling of coal deposits under Kuzbass conditions, with the aim of fully solving the principal questions of exploration, is the task of work being carried out by the staff of the Department of Exploration Techniques of TPI. Investigations of the questions posed were carried out chiefly in two parties—the Kiselevskaya and Leninskaya—as well as in the laboratory of the Department of Exploration Techniques. In analyzing the data on this question, use was also made in part of materials collected earlier by the department in the Tutuyasskaya and Uropskaya parties.

The main question to be considered in the present article is the determination of the moment at which coal seams are encountered. Recording the moment of change of strata of rocks or of minerals in general is possible and in a number of cases is carried out by various methods [1, 2, 3]. Some methods used to one degree or another in practice are based on geological indicators and on the physicomechanical, or more precisely mechanical, properties of rocks and minerals (resistance to destruction or drillability). In practice it is possible to use one more indicator—the physical properties (electrical,

magnetic, radioactive, etc.). All existing and possible methods for determining the moment of change of rock layers may be classified, according to the features considered, as follows (Table 1).

Control by geological features, as an independent method, cannot be recommended, since it does not give unambiguous and sufficiently accurate results. In practice, the depth of occurrence of a given contact or the thickness of the marker horizon is not known in advance with the required limits of accuracy. Otherwise—

Table 1

Classification of the principal methods for monitoring the moment of encountering coal seams

no problem of encountering coal seams would arise. The same may be said of the second geological feature—the change in the character of the sludge carried out by the fluid flowing from the borehole. Registration of the moment at which fluid with sludge from the seam of interest to us, or with altered properties, emerges from the mouth of the borehole never coincides with the moment at which this seam is encountered by the bottomhole tool. This occurs with a delay. The magnitude of the delay is the greater, the deeper the borehole, the lower the velocity of the ascending flow and the carrying capacity of the fluid, etc.

Control by changes in the mechanical properties of rocks is based on the different magnitude of the resistance of rocks or useful minerals to destruction. This is reflected, first of all, in a change in the mechanical drilling speed (drillability of the rocks), in the magnitude of the energy expended on destruction of the rocks, and in the magnitude of the torque on the spindle of the rig. A change in the mechanical drilling speed is associated with a change in the pressure in the hydraulic system of the rig or in the weight indicator. Finally, a change in the mechanical properties of rocks is associated with a change in the elastic vibrations of the drilling tool (frequency-amplitude characteristic) and a change in sound effects (noises).

All the phenomena listed are usually taken as the basis for a whole series of methods and instruments for recording the moment when strata of rocks or minerals change. However, the broad use of the stated principles, based on changes in the mechanical properties of rocks, is possible only when these properties differ sufficiently sharply and when there is a large number of observations characterizing the interdependence of the properties and of the phenomena associated with their change. The absence of these data

Fig. 1. Curves characterizing mechanical drilling speeds in siltstones (a) and sandstones (b) with bits of various types.

At a rotational speed \(n = 361\) rpm, with bits \(\varnothing = 92\) mm: 1 — MP-2; 2 — MP-6; 3 — TP-3; 4 — TP-3, \(\varnothing = 112\) mm; at \(n = 277\) rpm, with bits \(\varnothing = 92\) mm: 5 — MP-2; 6 — MP-6; 7 — TP-3; 8 — TP-3, \(\varnothing = 112\) mm; at \(n = 153\) rpm: 9 — MP-6, \(\varnothing = 92\) mm; 10 — TP-3, \(\varnothing = 112\) mm; at \(n = 71\) rpm: 11 — TP-3, \(\varnothing = 92\) mm; 12 — points obtained from a sufficiently large number of measurements; 13 — points obtained from a small number of measurements; 14 — boundary of the region of averaged speeds; 15 — boundary of the most probable deviations of the values of individual observations.

and, most importantly, the fact that the mechanical properties of rocks do not by any means always change sharply complicate the solution of the problem. This is the main reason why such an important task as recording the moment of encountering deposits of minerals has still not been solved. Its solution is further complicated by the fact that indicators associated with mechanical proper-

...by rock properties vary over wide limits depending on the drilling regimes, the type of bit tool, its diameter, the quality and quantity of flushing fluid, the depth of the borehole and its condition, the operating time of the bit tool, etc.

The use of the physical properties of rocks for solving the problem is in principle possible, but it requires extensive investigations.

In order to determine the possibility, limits, and conditions for applying one method or another, the Department of Exploration Technology carried out a certain amount of research. The investigation was directed chiefly at the direct and indirect means of monitoring changes in the mechanical properties of rocks: changes in the mechanical drilling rate (by measuring the drilling rate at certain intervals of time); changes in pressure in the hydraulic system of the drilling rig; changes in the power expended on drilling; and changes in the elastic vibrations of the drilling assembly during the drilling process.

Investigation of changes in the mechanical drilling rate.

As already noted, this factor underlies many methods for monitoring the moment at which coal seams are encountered. Therefore the investigations began with the study of precisely this factor. Under field conditions, material was collected on the drillability of rocks and coal. The mechanical rate of penetration in rocks and coal was recorded under specified drilling regimes with various bits.

For analysis of the mechanical drilling rate, the data collected under field conditions were divided into groups combining observations obtained under identical conditions (drilling regime, type of bit, rock drilled). After such grouping, the results of observations in each group were processed by methods of mathematical statistics. All parameters of the empirical distributions were calculated in accordance with the normal law, since the overwhelming majority of the constructed series showed sufficient closeness to the theoretical normal law of error distribution. From the results of processing the individual groups of observations, graphs were plotted for the distribution of the mechanical drilling rate for different types of bits, rocks, and drilling regimes (Fig. 1). On the graphs, the limits of the range of averaged rates (i.e., the most frequently occurring rates) and the limits of the most probable deviations of individual observations from the averaged values of the mechanical rate were determined. For coals, a similar graph (Fig. 2, a) was plotted not from averaged data but from specific measurements, owing to the insufficient number of chronometric observations. From this graph the lower limit of the mechanical drilling rates of coals was determined (11 cm/min). To obtain a more complete picture of the region of distribution of mechanical rates for coal, an analogous graph was constructed for the case of drilling coals with a double core barrel of Ignat’ev’s design (Fig. 2, b).

The adopted methodology made it possible to determine the most probable mechanical rates of penetration for specified drilling regimes and types of bits. A comparison of the regions of the most probable mechanical drilling rates in rocks (Fig. 1, a, b) and in coal (Fig. 2, a, b) shows that for rocks they are substantially lower than for coal. However, the lower boundary of the probable region for coals (11 cm/min) coincides with the upper boundary of the scatter of the mechanical drilling rate for siltstones and is somewhat lower than the upper boundary of the scatter of the mechanical drilling rate for sandstone. It follows from this that, in some cases, by the mechanical drilling rate...

it is impossible to make an unambiguous conclusion as to what is being drilled at the face—rock or coal.

To determine the frequency with which the mechanical drilling speeds of rocks and coals coincide, a table was compiled from measurements of the mechanical drilling speed with a simple core barrel. It included all measurements of the mechanical drilling speed of coals and measurements for rock which, in magnitude, are close to the drilling speeds of coals. On the basis of these data, a distribution graph of “high” mechanical speeds was compiled

Fig. 2. Distribution graph of mechanical drilling speed in coal:
a — with a simple core barrel; b — with Ignat’ev’s double core barrel; 1 — at a tool rotation speed of 480 rpm; 2 — at a tool rotation speed of 277 rpm.

speeds (Fig. 3). Analyzing the graph, one may say that, for the conditions studied, the following relationship is generally characteristic: the greater the value of the mechanical drilling speed, the lower its frequency (curves 1, 2). This regularity is especially characteristic of rocks (curve 2). The occurrence of coals under the operating conditions used in production most often takes place at drilling speeds of 14–26 cm/min (curve 3). Sometimes a coal drilling speed below 11 cm/min is encountered; however, these speeds are not characteristic and are explained by wedging of the core in the core barrel, low intensity of borehole flushing, etc.; therefore, in what follows they are excluded from consideration.

Of 643 measurements with a relatively high mechanical drilling speed (within the range of 8–62 cm/min), coal accounts for only 26 measurements, or 14.9%. At the same time, it is known that drilling of category V rocks under the operating conditions used in production most often pro-

is carried out at a speed of 5–7 cm/min. It follows from this that this ratio between the number of measurements in coal and in rock arose not because of the low rate of drilling coal, but because of the small number of measurements in coal. If one takes into account that the table included measurements with high values of mechanical rates, then from the graph (Figs. 3, 4) it is evident that the higher the mechanical rate, the greater the probability of drilling in coal.

The practice of drilling coal-bearing strata in the Kuzbass shows that intervals of drilling rocks with a high mechanical rate of advance are relatively small and depend in some way on the drilling rate. Drilling coal with a simple core barrel is in practice carried out at a high constant rate over large drilling intervals (usually across the entire coal seam). Thus, there arises the fundamental possibility of separating high mechanical drilling rates in rock from drilling rates in coal.

Fig. 3. Distribution of high mechanical rates of advance when drilling with a simple core assembly in the Kiselevskaya, Tutuyasskaya, and Leninskaya GRP:
1 — frequencies of the magnitudes of the mechanical drilling rates of coals (in the total number of measurements in rock and coal);
2 — frequencies of mechanical drilling rates in rock (in the total number of measurements in rock and coal);
3 — frequencies of mechanical drilling rates in coal (in the total number of measurements in coal);
4 — frequencies of coal-drilling rates by individual intervals (in the total number of measurements in rock and coal).

rock from drilling rates in coal. To clarify this question quantitatively, a table was compiled of the distribution of intervals of drilling rocks with high mechanical rates. On the basis of these data a graph was constructed that clearly shows the dependence of the frequency of drilling intervals with high rates on the size of the intervals (Fig. 4). On the basis of the graph it may be concluded that the frequency of intervals of drilling with high rates (in rock) decreases as the value of the interval increases, while the scatter of the frequencies is very considerable and on average amounts to 9%.

It should be noted that, when a coal seam is encountered, drilling through the coal with an ordinary core barrel is usually permitted within limits of 10–15 cm. Within such limits, cases of drilling at high speed even in rock are fairly frequent (Fig. 4).

From a preliminary analysis of the factual material, it thus follows that the moment at which a coal seam is encountered during drilling under ordinary regimes cannot yet be determined from mechanical speed alone. The probability of an unambiguous solution of the problem reaches only 50%, and this is at a mechanical speed of more than 40 cm/min; moreover, it is the lower, the lower the mechanical drilling speed (see Fig. 3). Consequently, in order to solve the problem posed by using mechanical speed, additional criteria are also needed. Such criteria may be: the drilling regime and the removal of sludge to the surface.

The drilling regime, as is known, has a noticeable effect on the mechanical drilling speed, and the higher the category of the rocks (or coal), the more substantial this effect. The insufficiency of factual observations still does not make it possible to identify quantitative relationships, but in general it is evident that, at low axial loads and rotation speeds, the mechanical

Fig. 4. Graph of the dependence of the frequency of drilling intervals in rock with high mechanical drilling speeds on the magnitude of the intervals.

drilling speed in coal differs more sharply from the drilling speed in rocks. Consequently, this may serve as an additional criterion for solving the problem. When the mechanical speed increases above the critical value, it is necessary to reduce the number of revolutions and the axial pressure to a minimum. When drilling in rock, the mechanical speed will fall considerably more than when drilling in coal.

The removal of sludge to the surface may serve as an additional criterion for establishing the moment at which a coal seam is encountered, but only under certain conditions. Table 2 gives the approximate time required for the flushing fluid to carry sludge from the face to the surface, depending on the amount of fluid supplied, the diameters of the borehole and drill pipes, and also the depth of the borehole. The average mechanical drilling speed in coal with an ordinary core barrel is 14–26 cm/min. Then the time required to drill the interval at which the coal seam is encountered (10–15 cm) will be only 0.4–0.7 min, which is much less than the time required to remove sludge to the surface even at an insignificant borehole depth. Consequently, in order to use this additional criterion, it is necessary, when increasing

mechanical drilling speed within the interval of 10–15 cm, stop drilling for a period from several minutes to an hour or more, and wait for the sludge to come to the surface.

Thus, the general scheme for solving the problem when using the mechanical drilling speed reduces to the following:

1) the penetration recorder registers an increase in the mechanical drilling speed (a signal is given or the rig is switched off);

2) the lowest number of revolutions and pressure are set. Drilling is carried out within an interval of 5–10 cm;

3) drilling is stopped, and the borehole is flushed for the time specified for the given conditions (see Table 2). If coal sludge is brought out, the problem is solved unambiguously.

Table 2

Time of movement of the flushing fluid from the bottomhole to the mouth of the borehole when using drilling pipes Ø 50 mm (in min). Flow of flushing fluid, l/min.

Investigation of changes in pressure in the hydraulic system of the rig as a function of the spindle travel speed

In addition to directly measuring the drilling speed, it is possible to measure the change in pressure in the hydraulic system of the rigs, which depends on the mechanical drilling speed. The latter determines the travel speed of the rig spindle and of the pistons in the cylinders of the hydraulic system.

Drilling under load. Suppose that, before the rock-breaking tool approaches the useful mineral, drilling is carried out through the enclosing rocks at a speed \(v_{\text{p}}\). Then the cutting member enters the coal seam; the mechanical drilling speed usually increases appreciably and reaches the value \(v_{\text{u}}\). Thus, the increase in the mechanical drilling speed will be \((v_{\text{u}} - v_{\text{p}})\), and, consequently, the speed of motion of the piston in the hydraulic cylinder of the rig increases by the same amount. In the lower cavity of the hydraulic cylinder an increase in pressure \((+\Delta p)\) will be observed, owing to the increase in losses in the drain line caused by the increase in the velocity of oil motion (Fig. 5, a). At a constant drilling speed, oil is supplied to the upper cavity of the hydraulic cylinder in a strictly definite quantity at constant pressure (the oil-feed regime is set by the shift foreman by means of a throttle). An increase in the piston speed causes a decrease in pressure in this cavity \((-\Delta p)\).

Drilling with unloading. In drilling with unloading, for the same reason there will also be an increase in pressure in the lower cavity \(X\) and a decrease in the upper cavity — (Fig. 5, b).

From all of the above there follow three possible methods for recording the moment at which the speed of motion of the piston in the hydraulic cylinder changes (and, consequently, the moment at which the rock-breaking tool enters the coal seam): 1) by measuring the pressure in the cylinder cavity connected to the drain line (the measurement can be made by means of a manometer); 2) by measuring the pressure in the working cavity of the hydraulic cylinder (also by means of a manometer); 3) by measuring the difference in pressure in the lower and upper cavities of the hydraulic cylinder (by means of a differential manometer).

The first method is comparatively simple, since the hydraulic resistance of the drain line is constant. The pressure in the cavity will change only as a function of the change in the velocity of the oil flow in the drain line, which is associated with a change in the mechanical drilling speed.

Fig. 5. Diagram of the action of the hydraulic system of ZIF-type drilling rigs: a — when drilling with an additional load; b — when drilling with unloading.

The second method is of little suitability because the hydraulic resistance of the pressure line is not constant. Its magnitude is changed by the shift foreman in accordance with the nature of the rocks being passed through. Consequently, in this case the change in pressure in the cavity of the hydraulic cylinder will depend on the change in velocity of motion of the oil flow in the pressure line, and on its hydraulic resistance, which has a variable value. In designing equipment operating on this principle, some linkage will be required that synchronizes the position of the throttle needle and of the signal sensor indicating a change in the mechanical drilling speed. This will greatly complicate the equipment.

The third method is a combination of the first two. In this case it is proposed to measure a somewhat greater pressure (the sum of the drops in both cavities, since they have different signs). Therefore this method can be used in those cases when the mechanical drilling speeds of the rocks of the coal-seam roof and of the coal differ only slightly. The disadvantages of this method are the same as those of the second.

Thus, the most promising is the first method of recording the change in the mechanical drilling speed. In this case no linkages are required between the recording instrument and the throttle, which provides somewhat greater reliability in the operation of the equipment.

Let us consider the dependence between the speed of motion of the piston (the mechanical drilling speed) and the pressure in the lower cavity when drilling with load. The pressure in the cylinder cavity will be equal to the sum of the hydraulic losses along the entire line, i.e., all linear and local losses. In steady motion of the liquid, the head losses can be found from formula (4)

\[ p=\frac{\gamma}{2q}\left(\sum_{i=1}^{n}\lambda\frac{l}{d}v^{2}+\sum_{i=1}^{n}b\varphi v^{2}\right), \tag{1} \]

where \(v\) is the average velocity of the fluid flow in each section, in m/sec;

\(q\) is the acceleration due to gravity, equal to \(9.81\ \text{m/sec}^{2}\);

\(\gamma\) is the specific weight of the oil, in kg/m\(^3\);

\(l\) is the length of each section of the pipeline, in m;

\(d\) is the inside diameter of the pipeline, in m;

\(\varphi\) is the coefficient of local loss in bends, branches, valves, spools, at points of contraction and expansion of the flow, etc. The coefficient \(\varphi\) depends on the Reynolds number \(R_e\), but in calculations it is taken as constant. The nonconstancy of the coefficient \(\varphi\) is allowed for by the correction coefficient \(b\).

\(b\) is a correction coefficient. Under laminar flow conditions it is inversely proportional to the Reynolds number \(R_e\); under turbulent flow \(b=1\);

\(R_e\) is the Reynolds number. For round pipelines

\[ HR_e=\frac{1.274\,Q}{d\nu}=\frac{vd}{\nu}, \tag{2} \]

\(\nu\) is the kinematic viscosity coefficient;

\(\lambda\) is the coefficient of friction in the pipeline.

By virtue of the continuity of the flow, the velocity of oil motion in any section of the discharge network can be found from the formula

\[ v=\left(\frac{F_n}{f}\right)v_b, \tag{3} \]

where

\(v,\ v_b\) are the velocity of the oil flow in the pipeline section under consideration and the velocity of piston motion;

\(f,\ F_n\) are the area of the pipeline section under consideration and the area of the cylinder (piston).

Expressing the velocity of oil motion in all pipeline sections in terms of the piston velocity (i.e., in terms of the mechanical drilling rate) and substituting in formula (1), we obtain:

\[ p=v_b^{2}\frac{\gamma}{2q}\left[\sum_{i=1}^{n}\lambda\frac{l}{d}\left(\frac{F_n}{f}\right)^{2}+\sum_{i=1}^{n}b\varphi\left(\frac{F_n}{f}\right)^{2}\right] \tag{4} \]

When the velocity of oil motion changes, the quantities \(\nu,\ \gamma,\ q,\ l,\ d\) and \(\varphi\) will remain constant; the coefficient of friction \(\lambda\) and the correction coefficient \(b\) will change, since the Reynolds number \(HR_e=f(v_b)\). Taking this into account, formula (4) can be written in the following form:

\[ p=C\left[\sum_{i=1}^{n}A_i\lambda_i+\sum_{i=1}^{n}B_i b_i\right]v_b^{2}, \tag{5} \]

where \(A_i,\ B_i,\ C\) are certain constant quantities.

It is evident from formula (5) that the losses in the discharge line are directly proportional to the sums of the coefficients of friction in the pipeline \(\lambda_i\) and the correction coefficients \(b_i\), and also to the square of the mechanical velocity. Let us consider the degree of influence of changes in the values of the mentioned coefficients on the fixed pressure.

The critical value of the Reynolds number \(Re_{\text{cr}}\) (above which a turbulent regime is observed) has the following value: for

for round smooth pipes, \(2000 \div 2300\); for local resistances, \(20 \div 700\).

Having calculated the value of the Reynolds criterion \(R_e\) at practical drilling speeds \((v_{\mathrm{b}} = 5.0 \div 30\ \text{cm/min})\), we find that for smooth pipes in most cases it does not exceed \(R_{e_{\mathrm{cr}}} = 2000 \div 2300\), while for elbows, valves, spools, etc., it is higher than its critical value. Consequently, in practice the oil-flow regime in spools, elbows, etc., is always turbulent, while in smooth pipes it is most often laminar, although it may also be turbulent. The correction coefficient \(b_i\) in this case will be equal to 1, and the friction coefficient \(\lambda_i\) in the various straight sections of the drain network will depend on the oil-flow regime. In the case of laminar flow:

\[ \lambda_i = \frac{64}{R_e} = \frac{64\nu}{d_i}\cdot\frac{1}{v_i} = \frac{n}{v_i}; \tag{6} \]

in the case of turbulent flow:

\[ \lambda_i = \frac{0.3164}{\sqrt[4]{R_e}} = \frac{0.3164}{\sqrt[4]{\dfrac{v_i d}{\nu}}} = \frac{m}{\sqrt[4]{v_i}}. \tag{7} \]

where \(n, m\) are constant coefficients for the section of pipeline under consideration.

A quite general case is possible, when in some sections of the network there may be a laminar regime, and in others a turbulent one. Then formula (5), taking into account (6), (7), and (3), will be written as follows:

\[ p = C\left[\sum_{i=1}^{n} A_i\lambda_i + \sum_{i=1}^{n} B_i b_i\right]v_{\mathrm{b}}^{2} = C\left[\sum_{i=1}^{n} A_i\left(\frac{n}{v_i}+\frac{m}{\sqrt[4]{v_i}}\right) + \sum B_i b_i\right]\cdot v_{\mathrm{b}}^{2} = Nv_{\mathrm{b}} + Mv_{\mathrm{b}}^{1.4} + Kv_{\mathrm{b}}^{2}. \tag{8} \]

In addition, the equation relating the pressure in the network to the spindle lowering speed must include a free term. This is due to the fact that, in a certain free section of the drain network, a flow of liquid always circulates with a constant discharge equal to the pump delivery; i.e., in this section the pressure losses do not depend on the spindle displacement speed.

Finally, the equation \(p = f(v)\) may be written as follows:

\[ p = Kv_{\mathrm{b}}^{2} + Mv' + Nv_{\mathrm{b}} + C, \tag{9} \]

where \(K, N, M, C\) are constant coefficients.

In view of the fact that \(v' \approx v_{\mathrm{b}}'\), equation (9) may approximately be written as:

\[ p = av^{2} + bv + c, \tag{10} \]

where \(a, b, c\) are certain constant quantities;

\(v\) is the mechanical drilling speed.

For an established drilling speed, the coefficients \(a, b, c\) are quantities constant for the given drilling rig. The theoretical calculation of these coefficients is very cumbersome (owing to the complexity of determining the characteristics of the drain line), and therefore it is advisable to determine them experimentally. The experiments amounted to determining the speed of motion of the spindle as a function of the pressure differential in the lower cavity of the hydraulic cylinder of the ZIF-650A drilling rig (with the spool in the “Spindle down” position). Figure 6 gives a diagram of the connection

instruments to the hydraulic system of the machine. The change in pressure in the cylinder cavity was determined by a pressure gauge. In order to vary the hydraulic resistance of the drain line, a throttling device was installed in it. This made it possible to obtain functional dependences \(p=f(v)\) for different characteristics of the drain line, which is the same for various machines with hydraulic feed. In addition, the installation of the throttling device was also due to the fact that, with the “natural” characteristic of the drain line, the pressures were very small, on the order of \(0.1 \div 0.2\) atm, which is difficult to record with ordinary pressure gauges. To create a large range of spindle travel speeds, a load \(P\) was attached to it.

The experiments were carried out as follows. The spindle was raised to the uppermost position, and the handle of the slide valve was set to the position “drilling under load.” After this, the lowering speed of the spindle was brought to a value corresponding to a certain pressure in the drain network. At one and the same pressure the experiment was carried out several times, and the lowering speed was taken as the arithmetic mean of all identical experiments. The pressure was set from 0.4 to 5 ati, at intervals of 0.2 ati. From the data of one series of experiments, a curve of the dependence \(p=f(v)\) was plotted. The resulting curve (Fig. 7) has no special inflection points (which indicates the constancy of the fluid-flow regime at all spindle lowering speeds) and, consequently, can be represented by a single equation. The simpler equation (10) was taken as the basis. After determining the coefficients (they were found by the least-squares method), formula (10) for the series of experiments carried out takes the form:

Fig. 6. Diagram for connecting instruments to the hydraulic system of ZIF-type machines:
1 — slide valve; 2 — safety valve; 3 — pressure gauge; 4 — throttling device; 5 — tee; 6 — piston; 7 — hydraulic cylinder of the machine; 8 — load for producing uniform piston travel.

\[ p=(0.6v_{\mathrm{b}}^{2}+29.1v_{\mathrm{b}}+363.7)\cdot 10^{-3}, \tag{11} \]

where \(P\) is the pressure in the drain line of the hydraulic system, in ati;
\(v_{\mathrm{b}}\) is the mechanical drilling speed, in cm/min.

In Fig. 7, the sufficient closeness of the experimental points to the curve obtained from formula (11) is clearly visible. Thus, the experiments confirmed the theoretical conclusions concerning the functional dependence of the pressure in the drain—

line on the rate of lowering of the drill spindle. The established form of the dependence \(p=f(v)\) and the values of the coefficients \(a\), \(b\), and \(c\) make it possible to calibrate the manometer scale according to the mechanical drilling rate and to make certain practical recommendations. The pressure increment from an increase in the mechanical drilling rate, according to formula (10), is determined as:

\[ \Delta p = p_2 - p_1 = a\left(v_2^2 - v_1^2\right) + b\left(v_2 - v_1\right) = a\left(v_2 + v_1\right)\Delta v + b\Delta v, \tag{12} \]

where \(v_2, v_1\) are the mechanical drilling rates in coal and in rock, respectively.

Analysis of formula (12) leads to the following conclusions:

1. The magnitude of the pressure increment in the hydraulic-cylinder cavity is directly proportional to the increment of the mechanical drilling rate \((\Delta v)\), as well as to the sum of the mechanical rates in the roof rocks and in the coal \((v_1+v_2)\).

Fig. 7. Dependence of pressure in the drain line of the hydraulic system of the ZIF-650A drilling rig on the rate of lowering of the spindle:
a — according to experimental data; b — according to calculated data.

1. To increase the pressure increment, the encounter with the coal seam must be made under forced drilling conditions. All of the above applies to the case of drilling under load. If drilling is carried out with unloading, the manometer should be connected to the upper cavity of the hydraulic cylinder. In this case, the encounter with the coal seam will be marked not by an increase, but by a decrease in pressure, and the question of the functional dependence \(p=f(v)\) requires additional investigation.

For practical registration of a coal seam (by measuring pressure in the rig’s hydraulic system), it is necessary to have a device including: 1) a valve by means of which the manometer can be connected to the upper or lower cavities of the hydraulic cylinders; 2) a valve protecting the manometer from high pressure during reaming; 3) a throttling device by means of which the hydraulic characteristic of the rig can be brought into conformity with the calibration of the manometer scale.

In conclusion, it may be said that fixing the moment of encounter with a coal seam by pressure measurements in the rig hydraulic system is possible only when there is a sufficient difference in the mechanical drilling rates of rocks and coals.

Study of Power Consumption for Drilling When the Mechanical Properties of Rocks Change

It is known that the rate of deepening of a borehole is determined by the rate and volume of destruction of the rock, the conditions under which this rock is separated from the mass, and the intensity of its removal from the face. The magnitude of cutter penetration and the volume of rock destroyed per unit time (or per one revolution of the bit) depend on the physico-mechanical properties of the rock, if the other conditions (such as the type of bit, its diameter, and the drilling regime) are considered constant. Naturally, different power is expended on the volumetric destruction of rocks with different physico-mechanical properties. The question arises whether it is possible, from the power expended, to determine certain physico-mechanical properties of the rocks, and whether it is possible to isolate from the total power expended on the drilling process the power used for rock destruction, or whether this power must be measured directly.

The total power \(N_{\delta}\) expended on the drilling process is determined by the following expression (6):

\[ N_{\delta} = N_{\mathrm{st}} + N_{\mathrm{khv}} + N_{\mathrm{dop}} + N_{\mathrm{rz}}, \tag{13} \]

where \(N_{\mathrm{st}}\) is the power for overcoming friction in the drilling rig;
\(N_{\mathrm{khv}}\) is the power for idle rotation of the drill string;
\(N_{\mathrm{dop}}\) is the additional power for rotating the string of drill pipes when transmitting axial load to the face;
\(N_{\mathrm{rz}}\) is the power for destruction of the rock at the borehole face.

Each of these components can be found from the corresponding formulas. In field investigations, the task was to determine the total power for the drilling process \(N_{\delta}\), and such components as \(N_{\mathrm{st}}\), \(N_{\mathrm{khv}}\), and \(N_{\mathrm{dop}}\) were assumed to be determined by formula. Then, with unchanged parameters of the drilling regimes, the change in \(N_{\delta}\) should depend on the change in \(N_{\mathrm{rz}}\), i.e., on the change of the physico-mechanical properties of the rock at the face.

To carry out time studies of the mechanical and electrical parameters, boreholes with the greatest intersection of coal seams were selected. The time study of the electrical parameters was carried out simultaneously with the time study of the mechanical parameters. The electrical parameters were measured using a set of small-sized measuring instruments KMP-10, assembled according to Aron’s circuit. It should be noted that in all cases the active power consumed by the electric motor from the network was measured.

To determine the power \(N_{\delta}\) and the components \(N_{\mathrm{st}}\) and \(N_{\mathrm{khv}}\), it is necessary to know the power on the motor shaft \(P_{v}\), which is a function of the power \(P_{c}\) consumed from the network; in turn, the power \(P_{c}\) depends on the efficiency of the motor, which varies depending on its load.

The active power \(P_{c}\) was determined by the formula

\[ P_{c} = C a_i (a + b) \ \mathrm{W}, \tag{14} \]

where \(a\) and \(b\) are readings, with their signs, from the wattmeters, W;
\(c\) is the coefficient for increasing the measurement limits of the power wattmeters of the KMP-10 set;
\(a_i\) is the current transformation coefficient.

To determine the power \(P_{v}\), a graph of the dependence \(P_{v} = f(P_{c})\) was constructed (separately for each motor), from which values of \(P_{v}\) corresponding to the measured values of \(P_{c}\) were taken. To construct the graphs, formulas and dependencies according to L. B. Geiler (5) were used.

The data from time-and-motion observations on mechanical speed and measurements of power and the value \(P_b\), taken from the graph, were entered into tables and then subjected to mathematical processing, based on the methods of mathematical statistics. All boreholes, according to the identity of the conditions under which the measurements were made, were divided into three groups: 1) vertical boreholes drilled with the ZIF-650A rig (boreholes 5562, 7160, 7169); 2) vertical boreholes drilled with the ZIF-300 rig (borehole 5565); 3) inclined boreholes (\(\Theta = 15^\circ\)), drilled with the ZIF-300 rig (boreholes Nos. 5561, 5563, 5701).

On the basis of the results of processing the measurements, summary tables of the total power expenditures for the drilling process (under the specified drilling modes and conditions) were compiled by groups of boreholes, separately for sandstones, aleurolites, and coal. Owing to the lack of data on power expenditures for drilling rocks and coals under identical conditions, it is not yet possible to present a comparative analysis and draw definite conclusions.

For carrying out a preliminary analysis, it was decided to decompose the total power into its component parts and to isolate the power for destruction of the rock at the bottomhole. It was assumed that, under identical conditions, the expenditures for cutting rocks and coals would be different, while the remaining components would be equal. However, when calculating the total power by the formulas of V. P. Zinenko and A. I. Kirsanov (6), it turned out that it is somewhat higher than the total power obtained experimentally. Moreover, this discrepancy is the greater, the higher the rotational speed of the tool and the higher the axial feed force. Thus, isolating the power for destruction of the bottomhole rock with the aid of experimental and calculated data proved practically impossible.

When analyzing formula (13), one can see that, as the borehole depth increases, the share of the power \(N_{rz}\) in the overall power balance for the drilling process will decrease. A comparison of the total powers (according to our observations) for drilling rocks with different bits and coals with single and double core barrels, with other conditions identical, shows that these powers almost do not differ from one another. Hence it becomes clear that isolating \(N_{rz}\) from the total power is practically impossible. This is further aggravated by the fact that the power \(N_{rz}\) itself does not remain constant during a run and by the end of the run increases one and a half times in comparison with the initial value (6).

Thus, isolating from the overall balance of the power expended on drilling, \(N_b\), the power for destruction of the bottomhole \(N_{rz}\), for the purpose of recording contacts of coal seams, at least under the conditions in which the time-and-motion observations were carried out, is practically impossible for a number of reasons. First of all, the total power for the drilling process \(N_{rz}\) increases with increasing borehole depth, while the share of the power \(N_{rz}\) in the overall balance decreases. In addition, the theoretical determination of \(N_{rz}\) by the formula of Zinenko and Kirsanov gives overestimated results. Finally, drilling conditions change sharply in connection with borehole sludging, instability of the walls, wedging of the core in the core barrel, and other factors causing a change in \(N_c\).

There can be only one way to solve the question of recording the moment of change of rock strata or coal seams in the present case: measurement of the power only for destruction of the rock at the borehole bottom (\(N_{rz}\)); for this purpose it is necessary to install bottomhole sensors of torque or of the expended power. For this it is necessary to carry out laboratory investigations to determine \(N_{rz}\) for different rocks and coals under various drilling conditions; to investigate the question of the influence on \(N_{rz}\) of automatic pressure variation (if we have rigs with hydraulic

feed) when the rocks change in drillability; to find practically acceptable technical means for measuring the torque or the power \((N_{\mathrm{rz}})\) required for rock destruction directly at the face; to develop methods and equipment for transmitting signals and receiving them at the surface.

On the basis of all the foregoing, it may be concluded that determination of the moment at which coal seams are encountered can be carried out from changes in mechanical speed only when additional data are available. Registration of the moment at which a coal seam is encountered can be performed by measuring the pressure in the hydraulic system of the rig, since a close relationship has been established between the pressure in the hydraulic system and the travel speed of the rig spindle. Finally, the use of data on the power expended on drilling is possible only when it is measured directly at the face \((N_{\mathrm{rz}})\).

References

1. Epshtein, E. F. Theoretical Foundations of an Instrument for Mechanical Logging of Boreholes during Drilling. Publishing House of the Dnepropetrovsk Mining Institute, vol. XXX, book 2, 1957.

2. Palyanov, P. F. “A Study to Determine the Possibility of Creating Instruments for Signaling the Moment at Which Coal Seams Are Encountered during Drilling.” Sheet of Production and Technical Information of the Luganskuglegeologiya Trust, no. 3 (6), 1957.

3. Palyanov, P. F.; Petrenko, G. P. “Materials for the Development of Instruments for Signaling the Moment at Which a Coal Seam Is Encountered during Exploratory Drilling.” Proceedings of the Dnepropetrovsk Mining Institute, vol. XXX, book 2, 1957.

4. Khaimovich, E. M. Hydraulic Drives and Hydraulic Automation of Machine Tools. Mashgiz, 1959.

5. Geiler, L. B. “Construction of the Torque Curve of Electric Motors.” Elektrichestvo, no. 4, 1941.

6. Zinenko, V. P.; Kirsanov, A. N. Experimental Measurements of the Power Expended in Core Drilling of Exploratory Boreholes. Gosgeoltekhizdat, 1961.