INVESTIGATION OF ROTARY DRILLING OF BLASTHOLES IN ROCK

Assoc. Prof. V. G. Mikhailov

“Drilling operations shall be carried out exclusively with electric drills—hand drills in coal and column drills in rock…”

(Order of the NKTP No. 1160 of June 10, 1936)

§ 1. Purpose and objectives of the work.

Drilling is one of the principal working processes in the extraction of mineral resources. It plays an especially large role in the driving of rock workings and in the extraction of hard minerals (salts, ores, building stone, anthracite, hard bituminous coal, etc.).

Until the present time, pneumatic percussion drilling has prevailed in these operations (pneumatic drilling hammers, pneumatic rock drills), and only recently (from 1935–1936) has electric rotary drilling of blastholes with column electric drills begun to be introduced in hard-coal and salt mines, as well as in certain other workings. In ore mining these drills as yet have no place at all; there pneumatic drilling is used exclusively, and only in the last two years have various designs of electric percussion rock drills begun to be tested¹).

The disadvantages of pneumatic percussion drilling are generally known: expensive energy, dust formation, and the recoil of pneumatic hammers on the hands. A column electric drill consumes, per hour of net work, at most (under an hourly operating regime) 2.3 kW × 1 = 2.3 kWh of electric energy, costing approximately 7 kopecks × 2.3 = 16 kopecks; and under a continuous regime only 7 kopecks × 1.8 × 1 = 13 kopecks. A pneumatic drilling hammer (BM-13) consumes per hour of net work 2 cubic m × 60 = 120 cubic m of air, costing approximately 1 kopeck × 120 = 120 kopecks. Consequently, the cost of energy for the hammer is at minimum eight to nine times more expensive than for the electric drill.

In rocks of sedimentary origin (shales, sandstones

¹) Mining Machine Building, 1938, No. 1, p. 53.

etc.) percussion drilling with pneumatic hammers gives a net speed lower than rotary drilling with electric drills¹.

In the well-known order No. 1160 (1936) of the People’s Commissariat of Heavy Industry, the late G. K. Ordzhonikidze instructed coal miners: “Drilling processes are to be carried out exclusively (emphasis ours, V. M.) with electric drills—hand-held ones for coal and column-mounted ones for rock; and in mines with a developed compressed-air system and especially gas- and dust-hazardous mines—with pneumatic drilling hammers.”

It follows from this that our orientation is toward electric drilling. In connection with the invention by Associate Professor K. N. Shmargunov of an electric jackhammer and its introduction into production, the question of unifying the energy supply in the overwhelming majority of hard-coal, shale, salt, etc., mines is posed in an especially realistic way; and this compels the timely conversion of blasthole drilling in these mines from pneumatic percussion drilling to electric, rotary drilling.

The broad introduction of rotary drilling requires its comprehensive study. It is necessary to select the proper operating conditions for the cutter, rational and durable forms of cutters and drills, to determine the required power consumption and the forces developed, and to design a strong and light electric drill. Judging from the available literature on rotary drilling, such study has not yet been properly undertaken in our country. There are only a few small observations² on the operation of electric drills. Meanwhile, this question is extremely pressing. We must raise rotary drilling to a higher level in terms of the design of the machines themselves, drills, and cutters, and expand its field of application to ore-mining work and the drilling of igneous rocks.

There is also little that can be borrowed from foreign practice, since no thorough works on this question are apparent in the foreign literature available to us.

The present work presents the results of our investigation of rotary drilling in argillaceous shale and sandstone, carried out by us³ at the Anzhero-Sudzhensk mine in the Kuzbass.

§ 2. Scope and Methodology of the Work

The scope of the present (first part of the) work included the investigation of the following questions in rotary drilling of blastholes in clay shale and sandstone.

1) To determine the loosening coefficient of the drill cuttings.
2) To determine the coefficient of self-removal of the drill cuttings in the process of drilling itself.
3) To determine the size of the drill cuttings.
4) To establish the influence of the direction of the blastholes on the axial forces (feed forces) and on the power expended in drilling.
5) To establish the influence of the depth of the blastholes on the same indicators.
6) To determine the influence of cutter wear on the same indicators.
7) To determine the power at the drill bit expended in overcoming the resistance of the drill cuttings and in friction of the drill against the walls of the blasthole.
8) To compare the operation of drills made of twisted solid steel of rhombic and rectangular cross section.
9) To compare the operation of drills made of rhombic twisted steel of different diameters.

For carrying out the experimental work, the Anzhero-Sudzhensk Mine of Kuzbassugol was selected, having rocks typical of the Kuzbass. In selecting the place (working face) for drilling, we were guided by the following: 1) the place must be dry, since dampness creates abnormal conditions for drilling; 2) the composition of the rocks must be homogeneous over the entire drilling depth; 3) there must be no local disturbances; the rocks must be intact and typical; 4) the working faces must not be active at the given moment, in order to have a calm environment for experimentation.

Such places proved to be, in mine No. 9/15, at level 146 m, the face of crosscut No. 21—for sandstone, and the roof of the Andreevsky (seventh) seam in the main drift—for clay shale.

The face of crosscut No. 21 was located in hard sandstone typical of the Kuzbass, with a strength coefficient \( f = 6 \) according to the scale of Prof. M. M. Protodyakonov.

In the roof of the Andreevsky seam there occurred a typical clay shale with a strength coefficient \( f = 4 \) according to the scale of Prof. M. M. Protodyakonov.

The blastholes were drilled at a distance of 4–5 cm from one another, in order to have identical drilling conditions for the holes being compared. The depth of the blastholes was taken as 1.8 meters, taking into account the available equipment and the greatest prevalence in drilling practice of blastholes of precisely this depth.

The direction of the blastholes was adopted as follows: horizontal (normal to the face), 15° upward, 25° upward, 15° downward, and 25° downward. For comparison, horizontal blastholes were also tested, but directed not perpendicular to the plane of the face, but diagonally (sideways) at 15° and 25°. Such a range encompasses the disposition of

Fig. 1.

Fig. 2.

holes, which usually occur in production practice.

Each hole was drilled to its full depth not with a single bit, but in several stages, using bits of different lengths, as will be described in detail below.

If, in the course of drilling, soft or hard interlayers were encountered that distorted the overall picture, such a hole was rejected and redrilled at some distance where no such interlayer was found; however, this was a rare occurrence, since the faces had been selected with homogeneous rocks. The quality of drilling, its results, and the homogeneity of the material being drilled were strictly monitored.

The holes were drilled in duplicate, i.e., two for each observation, in order to obtain more reliable average data. If the data for duplicated holes differed sharply, additional holes were drilled until a satisfactory result was obtained. Defects were strictly watched for.

A record was kept for each hole, indicating its characteristics and the data obtained during observation. In processing the materials, average indicators for the duplicated holes were derived. Additional information on the methods and technique of observation will be reported in the following sections as appropriate.

The arrangement of the drilled experimental holes in the sandstone is shown in Fig. 1, and in the clay shale—in Fig. 2.

§ 3. Equipment of the experimental installation and observation technique.

For carrying out the experimental work, a new column-type electric drill of type EK-2 from the Tomsk Electromechanical Plant was acquired. Before operation the drill was inspected and tested. The general view of the EK-2 electric drill is shown in Fig. 3.

The EK-2 electric drill weighs 120 kg, and the column on which it is mounted weighs 50 kg. The drill’s squirrel-cage asynchronous motor has an hourly power rating of 2.3 kW and a continuous rating of 1.8 kW, makes 2800 revolutions per minute, and is supplied with current at a voltage of 220 volts when the ends of the stator winding are connected in star and 127 volts when connected in delta.

The supplied flexible rubber cable has a cross-section of \(4 \times 2.5\) sq. mm.

The overall dimensions of the drill are: length (with the spindle retracted)—1439 mm, width—364 mm, and height—679 mm. The column on which the electric drill is mounted is made of pipes and therefore can be extended. The minimum height of the column (when the inner pipes are inserted into the outer ones) is 1.8 meters, but it can have a length, if the inner pipes \((v)\) are pulled out of the outer ones \((n)\), of 2.50 m and even more. The inner pipes are fixed at a given height by pins.

Fig. 3.

On the outer pipes, at any height, with the aid of two pins inserted into the through holes of pipes \(o\), two clamps \(x\) are fastened. The clamps have folding bows (4), which are locked by means of throw-on staples (5) and clamping bolts (6). The trunnions \(u\) of the drill are inserted into the openings between clamps \(x\) and bows (4).

By rotating the electric drill about the horizontal axis of the trunnions (up and down), drilling can be carried out at any point of the face along a vertical line. The drill can also rotate in a horizontal plane. All this makes the work convenient and makes it possible, from a single setup, to drill several blastholes.

The kinematic diagram of the drill is shown in Fig. 4. It has three pairs (gears 1–6) of gear transmission for rotating the spindle and one pair (gears 7 and 8) for rotating nut (38), which creates feed of the spindle (37), or of the drill steel, in the axial direction, since the nut (gear No. 8) makes more revolutions than the spindle (gear No. 6).

On the rotor shaft, gear wheel 1 is keyed; it meshes with gear wheel 2. These two wheels constitute the motor pair. Wheel 2 is mounted on two keys on the hub of the next wheel 3.

Wheel 3 has a bronze bushing in its hub and is freely mounted on a fixed axle, which has grooves on its surface to retain lubricant. Gear wheel 3 meshes with wheel 4, mounted on two keys on the hub of the next gear wheel 5. These gears (3 and 4) constitute the intermediate pair. Gear 5 is freely mounted on the long intermediate shaft 12 and has a bronze bushing inside the hub.

Gear wheel 5 meshes with gear 6, and they constitute the spindle pair. Gear wheel 6 is fastened by six bolts to the flange of sleeve 9, which on its inner surface has two projections entering the longitudinal grooves of the spindle. Consequently, gear 6 sets the spindle in rotation by means of the sleeve with projections. The spindle always rotates when the motor rotates, since the transmission between them is rigid (geared).

At the front end of intermediate shaft 12, gear wheel 7 is keyed; it meshes with gear 8. These gears constitute the feed pair. Gear 8 is fastened by four bolts to the long bronze nut 38, which has a thread analogous to that of the spindle.

The transmission ratio between gears 5 and 6 is smaller than between gears 7 and 8. Consequently, bronze nut 38 makes more revolutions than the spindle, and, as it rotates, it will pull the spindle in the axial direction, creating its working feed forward or backward, depending on the direction of rotation of the motor.

At the rear end of intermediate shaft (12) there is a friction-cam clutch (Fig. 4). It performs two functions:

1) it protects the motor and the drill mechanism from overload, and 2) with its aid it is possible to have three spindle feeds—working (slow), maneuvering (fast), and idle (rotation in place, without feed).

In our experiments, the friction-cam clutch did not perform the first function; its bronze bushings 39 and 41 were riveted solidly to the cam disk 34. This had to be done so that the action (slipping) of the friction clutch would not distort the operating regime of the drill under specified conditions and would not interfere with taking correct measurements of the required indicators. The clutch did perform the second function. Let us explain how this occurs.

It was indicated above that gear 8 makes more revolutions than gear 6, since the gear ratio between gears 1 and 8 is greater than that between gears 1 and 6. Consequently, bronze nut 38 makes more revolutions than the spindle, and as it rotates it will “pull” the spindle in the axial direction, producing its working feed forward or backward, depending on the direction of rotation of the motor.

The speed of this axial displacement of the spindle, and therefore of the drill, will be equal to the difference between the numbers of revolutions of the nut and the spindle, multiplied by the pitch of the spindle thread.

This may be expressed by the following formula:

\[ V_{\sigma} = (n_{г} - n_{ш}) \cdot t, \]

where \(V_{\sigma}\) is the feed speed of the drill (spindle), in mm per minute;

\(n_{г}\) is the number of revolutions of the nut per minute;

\(n_{ш}\) is the number of revolutions of the spindle per minute;

\(t\) is the pitch of the spindle thread (in mm).

Let us show the calculation by an example.

If the nut makes 202 rpm, the spindle 193 rpm, and the thread pitch is 20 mm, then the feed speed of the drill is

\[ V_{\sigma} = (202 - 193)\,20 = 180\ \text{mm/min}. \]

Engagement and disengagement of the cam clutch is achieved by the control handle 8 (Fig. 4), through shaft 49.

There may be three positions of the control handle.

First position (Fig. 4), when the handle is in the vertical (upward) position. At this time the cam clutch is engaged under the action of spring 16. In this case the force from gear 4 is transmitted through the cam and friction clutches to the intermediate shaft, and from it through gears 7 and 8 to the nut, which will “pull” the spindle in the axial direction due to the difference between its revolutions and those of the spindle, as was stated above. Thus we obtain the working stroke, or slow feed. The drill will move forward or backward, depending on the direction of rotation of the motor.

Second position (Fig. 4), when the handle is turned back slightly, toward the driller, just enough for the clutch cams merely to disengage. Then the force from the gear, through the copper washer 10, the wire ring 150, and the key 103 (Fig. 4), is transmitted to the intermediate shaft 12, owing to which gears 7 and 8 and nut 38 rotate with a small force, sufficient for nut 38 and the spindle to rotate at the same number of revolutions. The small force is obtained by friction of the wire ring 150 in the groove of the bronze washer 10. Consequently, in this case the spindle rotates without longitudinal displacement (idle running), since the difference between the revolutions of the nut and the spindle is zero. This is seen from the following calculation.

Let the nut and the spindle each make 193 rpm. Then the feed rate of the drill is

\[ V_b = (193 - 193)\cdot 20 = 0 \ \text{mm/min}, \]

i.e., the spindle will rotate without longitudinal displacement.

Third position (Fig. 4), when the handle is turned back, toward the driller, as far as it will go. In this case the clutch is moved back so far that the cam ring 34, with its rear cams, bears against ring 15 (Fig. 4), fitted into the drill body, and creates such friction that, through the friction clutch, the intermediate shaft is braked, whereby gears 7 and 8 and the nut stop. The difference between the numbers of revolutions of the nut and the spindle in this case is equal to the number of revolutions of the spindle, and the feed of the drill will be rapid, or, as it is called, the best maneuvering travel. Moreover, the drill will now move in the axial direction to the side opposite to that during slow (working) feed, since the nut will stand still, while the spindle will unscrew itself out of it.

This is confirmed by the following example.

If the nut does not rotate, while the spindle still makes 193 rpm, then the feed rate of the drill is

\[ V_b = (0 - 193)\cdot 20 = -3860 \ \text{mm/min}. \]

As is evident from the result obtained, the drill will move at high speed and in the reverse direction compared with the working travel, as indicated by the minus sign.

Rapid feed is necessary for changing drills and for other auxiliary operations. The direction of the axial displacement of the spindle (feed) is changed by reversing the motor by means of the starting controller¹.

The feed rate of the spindle (drill) depends on the working conditions—the hardness and structure of the rock being drilled. Selection of the feed rate

is carried out by changing feed gears No. 8 and No. 7, motor gears No. 2 and No. 1, and others (see Fig. 4 and Table 1).

Below is a table for selecting gears for a specified spindle feed rate (Table 1).

Table 1

Number of spindle revolutions per min.	Spindle feed (in mm/min)	Spindle feed in mm per revolution	Feed gears No. 8	Feed gears No. 7	Motor pair No. 2	Motor pair No. 1	Spindle pair No. 6	Spindle pair No. 5	Intermediate pair No. 4	Intermediate pair No. 3
193	180	0,93	75	31	65	27	38	15	57	24
193	380	1,97	74	32	65	27	38	15	57	24
193	570	2,95	73	33	65	27	38	15	57	24
237	210	0,89	75	31	61	31	38	15	57	24
237	450	1,90	74	32	61	31	38	15	57	24
237	670	2,83	73	33	61	31	38	15	57	24
286	260	0,91	75	31	57	35	38	15	57	24
286	540	1,89	74	32	57	35	38	15	57	24
286	820	2,86	73	33	57	35	38	15	57	24
429	400	0,93	75	31	48	44	38	15	57	24
429	800	1,87	74	32	48	44	38	15	57	24
429	1220	2,84	73	33	48	44	38	15	57	24
429	1660	3,87	72	34	48	44	38	15	57	24
520	1440	2,77	71	35	47	45	37	16	56	25
520	1900	3,65	70	36	47	45	37	16	56	25

Since the middle of 1937, the plant has switched to producing drills with a spindle speed of 286 revolutions per minute, with a feed of 260 mm/min.

From the foregoing it follows that in production, depending on local conditions (the strength and character of the rocks), it is necessary to select a feed more suited to the given strength of the rocks being drilled. Therefore, together with the drill, a set of the corresponding change gears must be ordered.

Let us show by example how to calculate the gear ratio, the number of revolutions of the spindle and nut, and the spindle feed rate. Let us take the version of gear selection with which the electric drill is supplied by the factory.

The gear ratio of the motor pair of toothed wheels is:

\[ i_1=\frac{Z_1}{Z_2}=\frac{35}{57}=0{,}614, \]

where \(Z_1=35\)—the number of teeth of the rotor gear (driving);

\(Z_2=57\)—the number of teeth of the second gear (driven).

The transmission ratio of the intermediate pair of gear wheels is

\[ i_2=\frac{Z_3}{Z_4}=\frac{24}{57}=0.421. \]

The transmission ratio of the spindle pair of gear wheels is

\[ i_3=\frac{Z_5}{Z_6}=\frac{15}{38}=0.395. \]

The overall transmission ratio between the motor rotor and the spindle is

\[ i_{\mathrm{sh}}=i_1 \cdot i_2 \cdot i_3=0.614 \times 0.421 \times 0.395=0.102. \]

The spindle will make

\[ n_{\mathrm{sh}}=n_m \cdot i_{\mathrm{sh}}=2800 \times 0.102=286\ \text{rpm}, \]

where: \(n_m=2800\) is the number of revolutions of the rotor per minute.

The transmission ratio of the feed pair of gear wheels is

\[ i_4=\frac{Z_7}{Z_8}=\frac{31}{75}=0.413. \]

The overall transmission ratio between the motor rotor and the feed nut is

\[ i_g=i_1 \cdot i_2 \cdot i_4=0.614 \cdot 0.421 \cdot 0.413=0.1067. \]

The feed nut will make

\[ n_g=n_m \cdot i_g=2800 \cdot 0.1067=299\ \text{rpm}. \]

The spindle feed rate is

\[ V_b=(n_g-n_{\mathrm{sh}})\cdot t=(299-286)\cdot 20=260\ \text{mm/min}, \]

where: \(t\) is the pitch of the spindle thread.

The general layout of the experimental installation is shown in Fig. 5.

The drill was placed on its own column. However, the column was installed not in the vertical position, as is usual, but in a lying (horizontal or inclined) position. Its screw rested against the face of the working area, while the shoe part with the traverse rested against another, similar column, installed vertically and braced between the roof and the floor in the usual manner, but very carefully and firmly. The lying column was also carefully and firmly braced between the vertical column and the face of the working area.

The electric drill was mounted on the lying column so that it could move along it during drilling under the action of the axial force and restraint.

Fig. 4.

Fig. 5.

Fig. 6.

dynamometer, by which the axial force was determined in kilograms.

In order that the clamps (slides) with the drill should move easily along the horizontal column and not seize, the column tubes were abundantly lubricated with solid oil.

The feed force developed during drilling was measured with a spring dynamometer manufactured by the VUTRIZ plant (Kharkov), with a capacity of up to 1250 kg. One hook of the dynamometer was fastened, by means of a rope loop, to the front end (where the screw is) of the horizontal column, and the other—to the drill. Thus, when the drill during drilling, under the action of the feed force, tended to move backward along the horizontal column, it stretched the dynamometer, and the latter indicated the force being developed.

Readings from the dynamometer were taken after each 10 cm of hole had been drilled and were recorded in the observation log. In processing, these readings were corrected for the tractive force that had to be developed during drilling in order to move the drill itself along the tubes of the horizontal column. This correction was taken as different, depending on the setting angle. It was determined as follows.

A small piece of cable was attached at one end to the drill, passed over a pulley fastened to the cross member of the vertical column, and weights were hung on its other end until the drill moved at a constant speed. The weights indicated the tractive force for moving the drill itself in the idle state.

This was done for the following drill (horizontal-column) inclination angles: 0°, 15° downward, 25° downward, and 15° upward. At 25° upward the drill moved along the column under its own weight, and the tractive force was determined directly with the dynamometer.

The observations were repeated several times, and then the average results were taken for each position of the drill.

In this way the following tractive force for moving the drill itself in the idle state along the horizontal column was established (Table 2).

Table 2

Drill position	Tractive force
For the position 25° upward	−20 kgf
15° upward	+1.0 kgf
0° (horizontal)	+31 kgf
15° downward	+58 kgf
25° downward	+74 kgf

The positive tractive force obtained was increased by 20%,¹ and added to the dynamometer readings, while the negative force, after being increased by 20%, was subtracted.

¹ When the drill is loaded (during drilling), the friction force between the clamps (slides) and the horizontal column increases by approximately 20% compared with the friction force when the drill is not loaded.

Fig. 7.

Visible labels in the graph:

• \(\eta\)
• \(P,\ \mathrm{kW}\)
• Coefficient of efficiency
• Useful power
• [[unclear: power curve label]]
• \(n\)

Fig. 8.

Visible labels in the graph:

• \(\eta\)
• \(P,\ \mathrm{kW}\)
• Coefficient of efficiency
• Useful power
• Consumed power
• \(n\)

As the work showed, readings could quite well be taken at every 10 cm of the borehole, and this interval was sufficient for characterizing the operation during the drilling of the borehole. Of course, during drilling the pointer of the dynamometer did not always remain steady, but sometimes oscillated. In this case readings were recorded that corresponded to the average or more characteristic value over the course of drilling ten centimeters of the borehole. The dynamometer was periodically calibrated. We had no self-recording dynamometer at our disposal.

The current consumed was taken into account by an ammeter, and the voltage in the supply network was recorded by a voltmeter; their readings were also taken at every drilled 10 cm of the borehole. The pointer of the voltmeter usually oscillated little. The pointer of the ammeter, however, did not always remain steady. In this case its readings were taken as corresponding to the average or more characteristic value over the course of drilling ten centimeters of the borehole, i.e., the procedure was the same as when taking readings from the dynamometer. The instruments were periodically calibrated. There were no self-recording instruments.

Readings from the instruments were also taken with the drill running idle. The electrical diagram of the setup, with the connection of the instruments indicated, is shown in Fig. 6.

From the readings obtained (amperage and voltage), and with the aid of characteristic curves (Figs. 7 and 8), the useful power at the spindle was determined for each measurement, expended on the drilling proper (cutting of the rock and the resistance of the drill due to friction against the walls of the borehole and against the drill cuttings).

The characteristic curves were compiled for all existing voltages on the basis of circle diagrams taken from the motor without disconnecting the drill reduction gear. Two persons worked at the drill. One observed its operation, watched the marks on the drill at every 10 cm, gave the signal for taking readings from the ammeter and voltmeter, and himself took the dynamometer readings. The other watched the ammeter and voltmeter and took their readings. He also recorded in a notebook the results of the readings of all three instruments. At the end of the shift the material was entered in the observation log.

Processing of the data obtained was not delayed; if a defect was found, the borehole was redrilled.

The temperature of the motor was not systematically measured; its casing scarcely heated up, since the operating regime was not severe and there were frequent intervals when the motor was stopped, which promoted its cooling.

The drills used were made of solid twisted steel of rhombic cross-section (see Figs. 9 and 10). The diameter of the drills was 30 mm, and the pitch of the helical groove \(h\) was 115–120 mm. The angle of rise of the helical groove \(\tau\) was 42°, and its angle of inclination \(\omega\) was −48°. The set contained three drills of different lengths: the first drill 0.60 m and the second 1.40 m, and the third—

Fig. 9.

Fig. 10.

Fig. 13.

Fig. 11.

Fig. 12.

Fig. 14.

Fig. 15.

the third—2.00 m long. To obtain uninterrupted operation, there were two working sets of drills. For the purpose of comparing the operation of screw drills of rhombic and rectangular (a) cross-section (Figs. 9 and 10), we also had a set of drills of this profile. The drills of rectangular cross-section (a in Figs. 9 and 10) had a diameter of 30 mm, a screw-groove pitch (h in Fig. 9) of 50–52 mm. The helix angle of the screw groove \(\tau\) was 27°, and the angle of its inclination \(\omega\) was 63°.

To compare the operation of drills of different diameters, a set of rhombic-section drills was purchased, but with a diameter of 35 mm (b, Fig. 9) and with a screw-groove pitch of 80–85 mm. Here the angle \(\tau\) is equal to 35°, and the angle \(\omega\)—55°.

In order to secure the drills in the spindle chuck, corresponding shanks were welded to them; these had, on one side, a cutout into which the pin entered, fastening the drill in the chuck. A head with an internal recess for mounting the cutter was welded to the other end of the drill. The cutter was fastened in the socket of the drill head by a stud.

In the investigation, cutters tipped with pobedit, grade RE-8, were used. Drilling was carried out mainly with cutters No. 2 (Fig. 11) and, to a lesser extent, with cutters No. 3 (Fig. 12). No difference in axial forces and power consumption when drilling with these cutters under production conditions was found; this is understandable, since their geometry is almost identical.

The geometry of cutter No. 2 is presented in Fig. 14, and that of cutter No. 3 in Fig. 15. The values of the angles are given in Table 3. At this stage of the work we did not change the values of the angles, but used the cutters in the form in which they were received from the factory.

Table 3

Cutter No.	Included angle (end angle) \(\varphi\)	Cutting angle \(\delta\)	Sharpening angle \(\gamma\)	Clearance angle of cutting edge \(\alpha\)	Splitting angle \(\psi\)	Clearance angle of splitting edge \(\alpha_1\)	Sharpening angle of splitting edge \(\gamma_1\)	Clearance angle of side edge \(\alpha_2\)	Taper angle	Plan angle of cutting edge \(\varphi_1\)	Plan angle of splitting edge \(\varphi_2\)	Groove lead angle on cutter body \(\tau_1\)	Groove inclination angle on cutter body \(\omega_1\)
2	144°	90°	67°	23°	90°	21°	69°	7°	8°	18°	41°	48°	42°
3	150°	90°	66°	24°	90°	24°	66°	8°	7°	15°	45°	40°	50°

Drilling conditions.

From gear-selection table (1) it is evident that the drill possesses a large range for selecting cutting-feed speeds (the number of drill revolutions).

We had interchangeable gears for various cutting and feed speeds. By experiment, through gradual investi-

…it was established that, for these rocks, with the given cutter and drill bits, it is possible to study feed rates of 180, 210, 260, and 400 mm per minute at drill rotation speeds of 193, 237, 286, and 429 per minute. Consequently, a feed per revolution of approximately one mm proved possible.

Attempts were made to test a feed per revolution of two mm (380 mm at 193 revolutions), but they ended in failure—the drill bits broke.

In mining information (drilling), up to now only the concept of feed rate per unit time, usually per minute, has figured; but this is insufficient. We, as miners, must introduce the concept of feed rate per one revolution of the drill bit, as has long been accepted in metalworking. This concept properly links the feed rate with the cutting speed and makes it easier to orient oneself in establishing a more correct drilling regime.

In addition to the feed rate per minute and per one revolution of the drill bit, it is necessary to know the cutting speed of the rock by the cutter blades. The cutter has two blades. Both operate under identical conditions.

The cutting speed along the length of their cutting edges will be different; it will increase as the distance of any points on the cutting edge from the axis (center) of the cutter increases. The greatest cutting speed will be at the outer end of the cutting edge, and it will be equal to zero on the geometric axis of the cutter.

The maximum cutting speed at the outer end of the cutting edge will be:

\[ V_p=\frac{2\pi Rn}{60}\ \text{mm/sec}. \]

where \(R\) is the radius of the cutter along the blade, in mm;
\(n\) is the number of revolutions of the cutter per minute.

The radius of cutter No. 2 along the blade is 20.5 mm; at a cutter rotation speed of \(n=193\) per minute, the cutting speed at the outer end of its cutting edge will be:

\[ V_p=\frac{2\pi Rn}{60}=\frac{2\cdot3.14\cdot20.5\cdot193}{60}=414.2\ \text{mm/sec}. \]

If we calculate, for different cutter rotation speeds, the cutting speed at different points on the blade edge with intervals of one millimeter between them, then in coordinate axes we obtain straight lines characterizing the cutting speeds on the blade edge of the cutter (Fig. 16) under different operating regimes.

The numerical data for these speeds are given in Table No. 4. Drilling of blastholes was carried out only with sharp cutters. As soon as even the slightest indication of dulling of the cutter began to be noticed, it was replaced. In the case when observation was being made of the effect of cutter wear on the power and axial force devel-

Fig. 16.

Fig. 17.

Table 4.

Feed rate of drill, mm/min	Number of drill revolutions, min.	Feed per one drill revolution, mm/min.	Cutting speed at the outer end of the cutter blade	1	2	3	4	5	6	7	8	9	10	11
				\multicolumn{11}{c}{Radii by points on the cutter blade in mm ($R$)}
				\multicolumn{11}{c}{Cutting speed on the cutter blade in mm per second ($V_p$)}
180	193	0,93	414,1	20,2	40,4	60,6	80,8	101,0	121,2	141,4	161,6	181,8	202,0	222,2
210	237	0,89	508,4	24,8	49,6	74,4	99,2	124,0	148,8	173,6	198,4	223,2	248,0	272,8
260	286	0,91	615,0	30,0	60,0	90,0	120,0	150,0	180,0	210,0	240,0	270,0	300,0	330,0
400	429	0,93	922,5	45,0	90,0	135,0	180,0	225,0	270,0	315,0	360,0	405,0	450,0	495,0

Feed rate of drill, mm/min	Number of drill revolutions, min.	Feed per one drill revolution, mm/min.	Cutting speed at the outer end of the cutter blade	12	13	14	15	16	17	18	19	20	20,5
				\multicolumn{10}{c}{Radii by points on the cutter blade in mm ($R$)}
				\multicolumn{10}{c}{Cutting speed on the cutter blade in mm per second ($V_p$)}
180	193	0,93	414,1	242,4	262,6	282,8	303,0	323,2	343,4	363,6	383,8	404,0	414,1
210	237	0,89	508,4	297,6	322,4	347,2	372,0	396,8	421,6	446,4	471,2	496,0	508,4
260	286	0,91	615,0	360,0	390,0	420,0	450,0	480,0	510,0	540,0	570,0	600,0	615,0
400	429	0,93	922,5	540,0	585,0	630,0	675,0	720,0	765,0	810,0	855,0	900,0	922,5

…used in drilling, the cutters were pre-blunted on an “extra” carborundum wheel, or careful observation and measurement were carried out of the blunting of the cutter when it was being tested for durability.

§ 4. Loosening coefficient of drill cuttings.

Determination of this coefficient is necessary in order to know the volume of drill cuttings produced when drilling one rock or another. We determined the loosening coefficient for sandstone and shale.

During the drilling process, all the drill cuttings were carefully collected; the blasthole was also carefully cleaned after each drill, and all the cuttings obtained from one blasthole were poured into an iron can with light shaking. In order to collect all the drill cuttings carefully, a tarpaulin was placed under the mouth of the blasthole, and on it—

newspaper; the dust was collected on the newspaper, then carefully poured into an iron can. The volume of cuttings in the can and the volume of the drilled hole were determined precisely.

The loosening coefficient was determined as the quotient obtained by dividing the volume of drill cuttings by the volume of the hole.

\[ K=\frac{q_m}{q_{\sh}}, \]

where \(K\) is the loosening coefficient of the drill cuttings;

\(q_m\) is the total volume of drill cuttings obtained from the hole \((\text{cu. cm})\);

\(q_{\sh}\) is the volume of the hole \((\text{cu. cm})\).

In turn, the volume of the hole is determined by the formula:

\[ q_{\sh}=\frac{\pi d^2 l}{4}, \]

where \(d\) is the diameter of the hole in cm;

\(l\) is the length (depth) of the hole in cm.

The diameter of the hole is determined as:

\[ d=a\cdot D\ \text{cm}; \]

where

\(D\) is the diameter of the cutter in cm;

\(a=1.05 \div 1.10\) is the hole-enlargement coefficient.

A total of ten observations were made, of which eight observations were on sandstone and two on shale.

The data obtained are given in Table 5.

Table 5.

No.	Rock	Loosening coefficient
1	Sandstone	1,62
2	Sandstone	1,83
3	Sandstone	1,77
4	Sandstone	1,75
5	Sandstone	1,80
6	Sandstone	1,76
7	Sandstone	1,70
8	Sandstone	1,78
	Average	1,75
1	Clay shale	2,16
2	Clay shale	2,18
	Average	2,17

The results obtained from individual observations give a small deviation; this indicates the constancy of this coefficient for the given rock and at a definite feed per one revolution of the drill bit (here \(0.89 \div 0.93\) mm/rev).

As can be seen, shale gave a higher loosening coefficient. Its drilling cuttings were more fluffy in comparison with the sandstone cuttings. Probably the loosening coefficient will change somewhat if the feed rate per one revolution of the drill bit is changed, that is, if the chip thickness is increased; however, we were unable to do this, since the drill bits and the column could not withstand a greater feed.

§ 5. Coefficient of self-removal of drilling cuttings.

By the coefficient of self-removal of drilling cuttings we mean the ratio of the volume of cuttings drilled out (self-flowing) from the blasthole during the drilling process itself to the total volume of cuttings obtained from the blasthole.

To determine the coefficient of self-removal, all cuttings spilling out of the blasthole during the drilling process itself were collected on tarpaulin and newspaper; then they were poured into an iron can and their volume was determined. The total volume of cuttings obtained from the blasthole was determined by actually measuring all the cuttings. Then the volume of self-flowing cuttings was divided by the volume of all cuttings obtained from the blasthole.

\[ C=\frac{q_c}{q_m} \]

where

\(C\)—coefficient of self-removal of drilling cuttings during the drilling process itself;

\(q_c\)—volume of cuttings drilled out (self-flowing) from the blasthole during the drilling process itself (cu. cm);

\(q_m\)—total volume of cuttings obtained from the blasthole (cu. cm).

We introduced and determined this coefficient in order to judge what the self-removal of cuttings depends on (on which factors) and to what extent; as will be seen below, it significantly affects the power expended in overcoming the resistance of the drill bit to the drilling cuttings.

In determining the coefficient of self-removal, it was established how the quality of the rock being drilled, the shape and diameter of the drill bit, the inclination angle of the blastholes, and also the drilling regime affect the self-removal of drilling cuttings during the drilling process itself.

A total of forty observations were made, half of them on sandstone and half on shale.

On both rocks, feed rates of 180 mm, 210 mm, 260 mm, and 400 mm per minute were tested, and, correspondingly, numbers of revolutions of 193, 237, 286, and 429 per minute.

All blastholes were drilled with drill bits of rhombic cross-section and dia-

Fig. 18.

Fig. 19.

30 mm. To study the effect of the drill cross section and its diameter on the self-removal of cuttings, holes were also drilled with drills made of drawn steel of rectangular cross section and rhombic cross section, but with a diameter of 35 mm; however, these results will be given in the corresponding paragraphs devoted to these drills.

The data obtained for sandstone are presented in Table 6, for clay shale in Table 7, and a comparison of the data for sandstone and shale is presented in Table 8.

The indicated numerical material is presented graphically in Figs. 17, 18, and 19.

From the materials obtained, the following conclusions may be drawn:

1) The coefficient of self-removal of drilling cuttings decreases as the angle of rise of the holes decreases and as the angle of inclination of the holes increases.

This regularity is clearly expressed both for sandstone and for shale, and for any drilling regime.

As can be seen from Figure 19, the coefficient of self-removal decreases almost according to a straight-line law.

It is seen from Table 6 that, for sandstone, the average cuttings self-removal coefficient for all drill revolutions rises from 0.232 (100%) in holes inclined 25° downward to 0.815 (351%) for holes with a 25° upward rise, or is 3.5 times greater, which gives an increase of 250%, or 5% for each degree of hole inclination, since the interval between the lowest hole (25° downward) and the highest hole (25° upward) is 50°.

The approximate average coefficient of self-removal of cuttings under the studied regimes for sandstone, for any angle of hole inclination (within +25° upward and −25° downward), can be determined, if this coefficient is known for any one inclination angle of the hole, by the following formula:

\[ C = C_u + (0.0116 \cdot \alpha^\circ), \]

where

\(C\) — the required coefficient of self-removal of cuttings;

\(C_u\) — the known coefficient of self-removal of cuttings for some angle of inclination of the hole;

\(\alpha^\circ\) — the interval in degrees between the inclination of the hole for which the coefficient of self-removal is known and the inclination of the hole for which the coefficient is being found.

Example.

The coefficient of self-removal of cuttings for a hole in sandstone with an inclination of 15° downward is 0.342 (Table 6).

For a hole with an inclination of 15° upward, this coefficient will be:

\[ C = C_u + (0.0116 \cdot \alpha^\circ) = 0.342 + (0.0116 \cdot 30) = 0.690, \]

which agrees fully with the data of Table 6.

Table 6 (for sandstone)

Feed rate, mm/min	Number of drill revolutions per min	% ratio	Indicators	25° upward: self-feed coeff.	25° upward: % ratio	15° upward: self-feed coeff.	15° upward: % ratio	0° horizontal: self-feed coeff.	0° horizontal: % ratio	15° downward: self-feed coeff.	15° downward: % ratio	25° downward: self-feed coeff.	25° downward: % ratio
180	193	100	Self-feed coeff.	0,749	100,0	0,582	100,0	0,479	100,0	0,300	100,0	0,210	100,0
180	193	100	% ratio	156,0		121,5		100,0		62,6		43,9
210	237	123	Self-feed coeff.	0,814	109,0	0,616	106,0	0,537	112,3	0,380	126,6	0,258	123,0
210	237	123	% ratio	151,5		115,0		100,0		70,9		48,1
260	286	148	Self-feed coeff.	0,820	109,5	0,760	130,5	0,590	123,2	0,300	100,0	0,200	95,3
260	286	148	% ratio	139,0		128,8		100,0		51,0		33,9
400	429	222	Self-feed coeff.	0,880	117,5	0,800	137,5	0,645	134,8	0,388	129,0	0,260	123,8
400	429	222	% ratio	136,5		124,0		100,0		60,0		40,3
Mean self-feed coeff.				0,815		0,690		0,563		0,342		0,232
Mean % ratios				144,8 100,0 351,0		122,5 84,7 297,5		100,0 69,2 243,3		60,7 42,0 147,5		41,2 28,5 100,0

Table 7 (for clayey shale)

Drilling speed, mm/min	Number of bit revolutions, rpm	Bit wear, %	Indicators	25° upward: self-removal coefficient	25° upward: ratio, %	15° upward: self-removal coefficient	15° upward: ratio, %	0° normal: self-removal coefficient	0° normal: ratio, %	15° downward: self-removal coefficient	15° downward: ratio, %	25° downward: self-removal coefficient	25° downward: ratio, %
210	237	100	Self-removal coefficient	0.620	100.0	0.543	—	0.495	100.0	—	—	—	—
210	237	100	Ratio, %	125.0	100.0	109.6	—	100.0	100.0	—	—	—	—
260	286	120	Self-removal coefficient	0.630	101.5	0.590	108.5	0.450	91.0	0.340	100.0	0.170	100.0
260	286	120	Ratio, %	140.0	—	131.0	—	100.0	—	75.5	—	37.8	—
400	429	181	Self-removal coefficient	0.705	113.5	0.620	114.0	0.510	103.0	0.400	117.5	0.210	123.5
400	429	181	Ratio, %	138.3	—	121.5	—	100.0	—	78.5	—	41.2	—
			Average self-removal coefficient	0.652		0.584		0.485		0.370		0.190
			Average ratio, %	134.5 100.0 343.0		120.0 89.5 307.0		100.0 74.4 255.0		76.5 56.8 194.5		39.2 29.2 100.0

Table 8

Feed rate, mm/min	Number of revolutions of the drill per minute	Rock	25° upward: self-rotation coeff.	25° upward: % ratio	15° upward: self-rotation coeff.	15° upward: % ratio	0° normal: self-rotation coeff.	0° normal: % ratio	15° downward: self-rotation coeff.	15° downward: % ratio	25° downward: self-rotation coeff.	25° downward: % ratio
210	237	Slate	0,620	100,0	0,543	100,0	0,495	100,0	—	—	—	—
210	237	Sandstone	0,814	131,0	0,616	113,5	0,537	108,3	—	—	—	—
210	237	Sandstone is better by	…	31,0%	—	13,5%	—	8,3%	—	—	—	—
210	237	Or, on average, by	…	—	—	17,6%	—	—	—	—	—	—
260	286	Slate	0,630	100,0	0,590	100,0	0,450	100,0	0,340	100,0	0,170	100,0
260	286	Sandstone	0,820	130,0	0,760	128,5	0,590	131,0	0,300	88,3	0,200	117,5
260	286	Sandstone is better by	…	30,0%	—	28,5%	—	31,0%	—	—	—	17,5
260	286	Or, on average, by	…	—	—	26,7%	—	—	—	—	—	—
400	428	Slate	0,705	100,0	0,620	100,0	0,510	100,0	0,400	100,0	0,210	100,0
400	428	Sandstone	0,880	124,0	0,800	129,0	0,645	126,5	0,388	97,0	0,260	124,0
400	428	Sandstone is better by	…	24,8%	—	29,0%	—	26,5%	—	—	—	24,0%
400	428	Or, on average, by	…	—	—	26,1%	—	—	—	—	—	—
Average % ratio by all drilling rates	Average % ratio by all drilling rates	Slate		100,0		100,0		100,0		100,0		100,0
Average % ratio by all drilling rates	Average % ratio by all drilling rates	Sandstone		126,8		123,7		121,9				120,8

The average ratio over all angles of blasthole inclination for slate and sandstone is 100 and 123,3%.

There is also a relationship for the average self-removal coefficients for shale. From Table 7 it follows that for the lowest blastholes (25° downward) the coefficient is 0.190 (100%), while for the highest (25° upward) it is 0.652 (343%), or 3.43 times greater, which gives an increase of 243%, or 4.86% per one degree of blasthole inclination.

2) The coefficient of self-removal of drill cuttings increases somewhat as the number of revolutions of the drill increases.

This regularity is expressed for almost all blasthole inclinations and for both rocks—sandstone and shale—as is evident from Tables 6 and 7 and from Figures 17 and 18.

3) The coefficient of self-removal of drill cuttings when drilling sandstone is higher than when drilling argillaceous shale, as is evident from Table 8 and from Figure 19.

The average coefficient (over all speeds) of self-removal for sandstone is higher than that for shale by 20.8%–26.8%, on average by 23.3% (Table 8).

As for the angle of inclination, this difference is somewhat greater for blastholes directed upward and decreases with the inclination of blastholes downward, as is evident from Figure 19.

The influence of the diameter of drills and of the profile of their cross section on the coefficient of self-removal of drill cuttings will be considered below, in the corresponding paragraphs.

§ 6. Coarseness of Drill Cuttings

During rotation and penetration (feed), the drill bit removes a chip. The thickness of the chip, as is known, depends on the feed per one revolution of the cutter. The coarseness of the cuttings depends both on the thickness of the chip and on the hardness, brittleness, and structure of the rock.

The cuttings produced when drilling the mouth of the blasthole are coarser, and with deepening of the blasthole they are pulverized by friction of the drill and of the grains against one another. Therefore the coarseness of the chip removed by the cutter is characterized by the coarseness of the drill cuttings obtained when drilling only the mouth of the blasthole, and not from its bottom (depth).

By comparing the coarseness of the drill cuttings from the mouth and from the bottom of blastholes, one can determine the degree of their pulverization due to abrasion. Sampling was carried out from the freely discharged cuttings onto tarpaulin and paper; the sample was then poured into a glass jar and subjected to sieve analysis in the chemical laboratory of the Anzhero-Ugol trust. The cuttings were sorted by coarseness into five classes: 1) above 2 mm; 2) from 1.5 to 2 mm; 3) from 1 to 1.5 mm; 4) from 0.25 to 1.0 mm; and 5) below 0.25 mm.

To determine the influence of rock hardness and drilling regime on the coarseness of the cuttings, samples were taken both for sandstone and for shale and for different numbers of cutter revolutions, but with drilling only horizontal blastholes. The obtained results are presented in Table 9 (for sandstone), Table 10 (for shale—

Drill-feed rate, mm/min	Number of revolutions of the drill per min	Drill feed per one revolution, mm	Where the sample of cuttings was taken	Coarseness of drilling fines, %: Over 2 mm	Coarseness of drilling fines, %: From 1.5 to 2 mm	Coarseness of drilling fines, %: From 1 to 1.5 mm	Coarseness of drilling fines, %: From 0.25 to 1 mm	Coarseness of drilling fines, %: Below 0.25 mm	Coarseness of drilling fines, %: Total
180	193	0.93	Borehole mouth	6.8	10.8	14.8	12.4	55.2	100.0
"	"	"	Bottom	6.0	8.4	12.4	12.0	61.2	100.0
210	237	0.89	Mouth	8.8	10.0	13.6	12.0	55.6	100.0
"	"	"	Bottom	1.2	3.2	7.6	10.4	77.6	100.0
260	286	0.91	Mouth	9.7	9.4	16.3	8.4	56.2	100.0
"	"	"	Bottom	3.6	6.0	7.2	12.4	70.8	100.0
400	429	0.93	Mouth	10.1	9.8	13.6	8.6	57.9	100.0
Average values			Mouth	8.8	10.0	14.6	10.3	56.3	100.0
Average values			Bottom	3.6	5.9	9.1	11.6	69.8	100.0

Table 9 (for sandstone)

Total coarse fractions (from 1 mm and above)	Distribution, % ratio: Mouth	Distribution, % ratio: Bottom	Ratio of coarse fractions, %%: Mouth	Ratio of coarse fractions, %%: Bottom	Total fine fractions (below 1 mm)	Distribution, % ratio: Mouth	Distribution, % ratio: Bottom	Ratio of fine fractions, %%: Mouth	Ratio of fine fractions, %%: Bottom
32,4	100,0	—	100,0	—	67,6	100,0	—	100,0	—
26,8	—	100,0	—	82,8	73,2	—	100,0	—	108,0
32,4	100,0	—	100,0	—	67,6	100,0	—	100,0	—
12,0	—	45,0	—	37,0	88,0	—	120,0	—	130
35,4	109,0	—	100,0	—	64,6	95,5	—	100,0	—
16,8	—	62,7	—	47,5	83,2	—	113,6	—	128,5
33,5	103,5	—	100,0	—	66,5	98,5	—	100,0	—
33,4	—	—	100,0	—	66,6	—	—	100,0	—
18,5	—	—	—	55,4	81,5	—	—	—	122,2

Drill feed rate, mm/min	Number of drill revolutions, rpm	Drill feed per revolution, mm	Where the cuttings sample was taken	Over 2 mm	From 1.5 to 2 mm	From 1 to 1.5 mm	From 0.25 to 1 mm	Below 0.25 mm	Total
180	193	0.93	Collar of the blasthole	20.0	18.0	19.2	15.2	27.6	100.0
180	193	0.93	Bottom	26.0	11.5	7.5	14.5	40.5	100.0
210	237	0.89	Collar of the blasthole	13.0	14.0	19.5	18.5	35.0	100.0
210	237	0.89	Bottom	10.8	15.6	19.6	17.6	36.4	100.0
400	429	0.93	Collar of the blasthole	8.0	9.0	19.5	16.5	47.0	100.0
400	429	0.93	Bottom	5.5	6.0	13.5	9.0	66.0	100.0
Average values	Average values	Average values	Collar of the blasthole	13.7	13.7	19.4	16.7	36.5	100.0
Average values	Average values	Average values	Bottom	14.1	11.0	13.5	13.7	47.7	100.0

Table 10 (for clayey shale)

Total coarse grades (from 1 mm and larger)	Distribution of them, % ratio: Fractions	Distribution of them, % ratio: Before	Ratio of coarse grades, %%: Fractions	Ratio of coarse grades, %%: Before	Total fine grades (below 1 mm)	Distribution of them, % ratio: Fractions	Distribution of them, % ratio: Before	Ratio of fine grades, %%: Fractions	Ratio of fine grades, %%: Before
57,2	100,0	—	100,0	—	42,8	100,0	—	100,0	—
45,0	—	100,0	—	78,8	55,0	—	100,0	—	128,5
46,5	81,3	—	100,0	—	53,5	125,0	—	100,0	—
46,0	—	102,3	—	98,8	54,0	—	98,3	—	101,0
36,5	63,8	—	100,0	—	63,5	148,3	—	100,0	—
25,0	—	55,6	—	68,6	75,5	—	136,0	—	118,2
46,7	—	—	100,0	—	53,3	—	—	100,0	—
38,7	—	—	—	83,0	61,3	—	—	—	115,0

Fig. 20.

Fig. 21.

Fig. 22.

Fig. 23.

cu) and in Table 11 the average comparison for sandstone and shale for all speeds.

Table 11

Rock	Where the dust sample was taken	Above 2 mm	From 1.5 to 2 mm	From 1 to 1.5 mm	From 0.25 to 1 mm	Below 0.25 mm	Total	Total coarse grades (from 1 mm and above)	Total fine grades (below 1 mm)
Sandstone	Mouth . . . % % corresp.	8,8 100,0	10,0 100,0	14,6 100,0	10,3 100,0	56,3 100,0	100,0 —	33,4 100,0	56,6 100,0
Shale	Mouth . . . % % corresp.	13,7 156,0	13,7 137,0	19,4 133,0	16,7 162,0	36,5 65,0	100,0 —	46,7 140,0	53,3 80,0
Sandstone	Bottom . . . % % corresp.	3,6 100,0	5,9 100,0	9,1 100,0	11,6 100,0	69,8 100,0	100,0 —	18,5 100,0	81,5 100,0
Shale	Bottom . . . % % corresp.	14,1 392,0	11,0 186,5	13,5 148,3	13,7 118,0	47,7 68,4	100,0 —	38,7 209,0	61,3 75,4

The indicated numerical material is presented graphically in Figs. 20, 21, 22, and 23.

From the data obtained, the following conclusions may be drawn.

1) When drilling shale, as a rule, coarser drilling dust is obtained than when drilling sandstone, as is seen from Tables 9, 10, 11 and from Figs. 22 and 23.

Coarse grades (from 1 mm and above), on average for the adopted numbers of revolutions, amounted to 46.7% in shale and 33.4% in sandstone, or 40% more in shale.

This, obviously, is due to the lower hardness of shale in comparison with sandstone. Shale is softer and more brittle than sandstone; therefore its chip breaks off coarser than that of sandstone. The finer dust of sandstone indicates that, when its chip is removed, abrasion occurs to a greater extent and chipping to a lesser extent, and this, in turn, increases the power consumption for cutting. There is a complete analogy here with the coal dust cut out from under a cutting machine, where, when cutting softer or cleaner or more brittle coal, the dust obtained is coarser than when cutting stronger, earthy, and viscous coals.

2) The drilling dust obtained when drilling the mouth of the blasthole is, as a rule, coarser than the dust obtained at the end of drilling the blasthole, from its bottom (Figs. 20 and 21).

It follows from Tables 9, 10, and 11 that the coarse grades of dust in sandstone at the bottom of the blasthole (on average, for all speeds) amount to 18.5%, and at the mouth to 34.4%, or more by 80%; while in shale—38.7 and 46.7%, respectively, or the increase amounts to 20%.

Fig. 24.

Fig. 25.

The last ratio shows that sandstone cuttings are subjected to abrasion more than shale cuttings. This is evidently facilitated by the higher coefficient of friction possessed by sandstone cuttings in comparison with shale cuttings¹.

3) With an increase in the number of revolutions of the drill, the coarseness of the drill cuttings obtained from the end of the blasthole decreases somewhat.

In argillaceous shale, at 193 revolutions of the drill per minute, the coarse grades of cuttings (from one mm) amounted to 45%, whereas at 429 revolutions they amounted to only 25%, or an absolute decrease of 20%.

Since the number of revolutions of the drill increased by 122% (from 193 to 429 rpm), then for each one percent increase in the drill revolutions the coarse grades of cuttings decrease by \(20 : 122 = 0.164\%\).

In sandstone, the coarse grades of cuttings amounted to 26.8% at 193 rpm and 16.8% at 286 rpm, or an absolute decrease of 10%; thus one percent increase in the drill revolutions gives a decrease in the coarseness of the cuttings by

\[ 10 : 48 = 0.21\%. \]

§ 7. Coefficients of friction of drill cuttings.

Drill cuttings, moving along the helical groove of the drill, develop a certain frictional force that must be overcome in order to push the cuttings out of the blasthole. This frictional force, as is known, depends on the coefficient of friction between the cuttings and the drill steel, on the one hand, and between the cuttings and the walls of the blasthole, on the other. Since the walls of the blasthole are always covered with a layer of drill cuttings, in the latter case we have the coefficient of friction of cuttings against cuttings.

To determine the coefficient of static and kinetic friction of the cuttings against steel and the angle of natural repose (cuttings against cuttings), we carried out experimental work.

A heap of drill cuttings was poured onto a steel plate (Fig. 24), and the upper edge of the plate was gradually raised (the lower edge rested immovably on the table) until the cuttings began to slide uniformly down the plate. At this moment the raising of the plate was stopped, and the height of the plate’s elevation (\(H\)) and the leg \(L\) were measured (in mm). It is known that the coefficient of static friction is the tangent of the angle of static friction; therefore,

\[ f_n = tg\alpha_n = \frac{H}{L}. \]

The coefficient of kinetic friction was determined in the same way, but, during continuous pouring of cuttings onto the plate, the latter was lowered (Fig. 25) until the cuttings ceased to

rolled downward. At that moment the lowering of the plate was stopped and \(H_1\) and \(L\) were measured. The coefficient of kinetic friction was determined by the formula:

\[ f_{ov}=\operatorname{tg}\alpha_{ov}=\frac{H_1}{L} \]

The angle of repose was determined in the following manner (Fig. 26). Drilling flour was gradually and freely poured onto a steel plate in the form of a cone (pile). Since such a pile was freely poured, the flour was positioned at the angle of repose. The angle of repose was determined by the formula

\[ f_{otk}=\operatorname{tg}\alpha_{otk}=\frac{h}{D:2}=\frac{2h}{D} \]

The coefficients were determined for drilling flour from sandstone and shale. Moreover, samples were taken from the mouth and from the bottom (face) of blastholes drilled at different feed and cutting speeds; from all the data obtained under these conditions, average values were taken for each rock.

The friction coefficients were determined both for the cleaned surface of the plate (polished to a shine) and for a surface that had not been cleaned and was rubbed with drilling flour. The latter variant represented conditions close to the actual ones during drilling.

In all cases, two or three measurements were made for each observation, and the average values were recorded as the final results. In all, more than one hundred tests (measurements) were made.

The average results obtained are presented in Table 12.

From the material presented, the following conclusions may be drawn.

1) The angle of repose of sandstone drilling flour is greater than that of shale flour.

The average angle of repose for shale flour is \(42^\circ 07'\), and for sandstone flour it is \(44^\circ\), or 5% higher. This can be explained by the greater roughness of the surface of sandstone grains than of shale grains, and also by the fact that shale grains are for the most part flat, in the form of flakes, whereas sandstone grains are more angular in shape.

2) The coefficients of static and kinetic friction for sandstone flour are greater than for shale flour.

As a rule, the friction coefficients for sandstone flour are 10–12% higher than for shale flour. This phenomenon is explained by the arguments given above.

3) The coefficients of static friction are greater than the coefficients of kinetic friction.

This excess amounts to about 7–8%. Since this phenomenon confirms the known laws of mechanics, it requires no special explanation.

Fig. 26.

Fig. 27.

Table 12

Object of testing	Elastic rebound angle, \(\alpha_{\mathrm{reb}}\)	Rebound, %	Coefficient of friction \(\mu\)	Cleaned plate: sliding coefficient \(f_{\mathrm{sl}}\)	Cleaned plate: ratio, %	Cleaned plate: rolling coefficient \(f_{\mathrm{r}}\)	Cleaned plate: ratio, %	Uncleaned plate: sliding coefficient \(f_{\mathrm{sl}}\)	Uncleaned plate: ratio, %	Uncleaned plate: rolling coefficient \(f_{\mathrm{r}}\)	Uncleaned plate: ratio, %
Clay shale (powder)	\(42^\circ 07'\)		0.905	0.578		0.624		0.657		0.704
Ratio of \(f_{\mathrm{sl}}\) and \(f_{\mathrm{r}}\), %					100		108		100		107
Shale taken as		100			100		100		100		100
Sandstone (powder)	\(44^\circ 0'\)		0.967	0.617		0.688		0.735		0.791
Ratio of \(f_{\mathrm{sl}}\) and \(f_{\mathrm{r}}\), %					100		111.5		100		108
Ratio of sandstone to shale, %		107			106.5		110		112		112.3

4) The coefficients of friction for an uncleaned plate are higher than for a cleaned one.

This excess amounts to 7–10%; it is quite natural and requires no explanation.

§ 8. The influence of drilling dust and the walls of the blasthole on the power developed during drilling

The presence of drilling dust in the blasthole undoubtedly impedes the rotation of the drill. Then, owing to the helical grooves, the drill tends to push the dust out of the blasthole.

In addition, the drill, with its ribs, touches the walls of the blasthole, drills out (reams) them, and thereby also causes an additional expenditure of power. These harmful resistances are especially great when the drill has a considerable longitudinal deflection, that is, under a considerable axial load.

Consequently, the drill encounters resistance in the blasthole not only from the drilling dust, but also from friction against the walls of the blasthole. In the present case, the influence of the drilling dust and of the drill’s friction against the walls of the blasthole was determined without axial force.

To determine the influence of drilling dust and of the drill’s friction against the walls of the blasthole during idling, corresponding measurements were made.

The measurements were made both in sandstone and in shale, for different operating conditions, different drill diameters, and different inclinations of the blastholes. The observations were carried out as follows.

As soon as drilling of the blasthole was finished, the drilling machine was stopped. Under no circumstances was either reverse or idling allowed, so as not to discharge the dust from the face (bottom) of the blasthole. The drill was not pulled out of the blasthole. The drill was carefully disconnected from the chuck; the drilling machine was moved somewhat back along the supporting column; the front cover of its reducer was removed; the feed nut was removed from the spindle (first the chuck was removed from the spindle), and everything was then reassembled and connected as for drilling.

As is known, without the nut there is no spindle feed; the latter will only rotate in place. At this moment the power consumed by the motor¹ from the mains was determined; after this the power consumed¹ during idling was measured. The difference between these powers gives the load on the drilling machine from the drilling dust and from the friction of the drill against the walls of the blasthole. The work was carried out with identical drills, both in sandstone and in shale (drill diameter 30 mm).

A total of thirty observations were made. The results obtained are presented in Tables 13, 14, and 15 and in Figure 27.

From consideration of the table data, the following conclusions may be drawn:

¹ By measuring amperage, voltage, and the circle diagram.

Table 13 (for sandstone).

Single-stroke speed, mm per minute	Rotational speed, rpm	15° upward — spindle power, kW	15° upward — % ratio	0° normal — spindle power, kW	0° normal — % ratio	15° downward — spindle power, kW	15° downward — % ratio	25° downward — spindle power, kW	25° downward — % ratio
180	193	0,576*)	100,0 100,0	0,683*)	100,0 118,5	0,989*)	100,0 171,0	1,167*)	100,0 202,0
210	237	0,434	75,5 100,0	0,561	82,2 129,5	0,574	58,0 132,0	0,917	78,6 211,0
260	286	0,335	58,2 100,0	0,560	82,2 167,0	0,613	62,0 183,0	0,800	68,6 239,0
400	428	0,330	57,3 100,0	0,367**)	53,7 111,0	0,576	58,3 174,5	0,750	64,3 227,5
Average % ratio		0,419	72,7 100,0	0,543	79,5 130,0	0,688	69,6 164,0	0,909	77,8 217,0

Table 14 (for shale).

Single-stroke speed, mm per minute	Rotational speed, rpm	15° upward — spindle power, kW	15° upward — % ratio	0° normal — spindle power, kW	0° normal — % ratio	15° downward — spindle power, kW	15° downward — % ratio	25° downward — spindle power, kW	25° downward — % ratio
210	237	0,098	100,0 100,0	0,150	100,0 153,0	—	—	—	—
260	286	0,150	153,0 100,0	0,180	120,0 120,0	0,400	100,0 267,0	0,520	100,0 347,0
400	428	0,180	183,8	0,290	193,5	0,370	92,5	0,490	94,3
Average % ratio		0,143	145,8 100,0	0,207	138,0 144,5	0,385	96,3 269,0	0,505	97,2 353,0

) The powers appear to be overstated; additional checking is necessary.
*) The power appears to be understated; checking is necessary.

Table 15.

Rock	15° upward: power at spindle, kW	15° upward: % ratio	0° normal: power at spindle, kW	0° normal: % ratio	15° downward: power at spindle, kW	15° downward: % ratio	25° downward: power at spindle, kW	25° downward: % ratio
Clayey shale	0.143	100.0	0.207	100.0	0.385	100.0	0.505	100.0
% ratio		100.0		145.0		269.0		353.0
Sandstone	0.419	293.0	0.543	262.0	0.688	179.0	0.909	180.0
% ratio		100.0		130.0		164.0		216.0

1) The expenditure of power for overcoming the resistance of the drilling fines and the friction of the drill against the blasthole wall in sandstone is, as a rule, higher than in shale.

This is clearly evident from the table both for different angles of blasthole inclination and for different numbers of drill revolutions. For the time being, this phenomenon can evidently be explained by the fact that in sandstone the drill has greater resistance against the blasthole wall than in shale, owing to the greater strength of the sandstone itself. In addition, sandstone drilling fines have a higher coefficient of friction against steel than shale fines; this circumstance also affects the increase in power due to sandstone fines.

2) The angle of inclination of the blastholes has a considerable influence on the expenditure of power due to drilling fines and due to friction of the drill against the blasthole wall.

It is clearly seen from the tables that, with a decrease in the angle of rise of the upper blastholes and an increase in the angle of inclination of the lower blastholes, the power required to overcome the resistance of the drill from the fines and against the blasthole wall increases both when drilling sandstone and when drilling shale. This increase is especially large when drilling shale. This is fully consistent with the circumstance that the angle of inclination affects the coefficient of self-removal of the drilling fines much more in shale than in sandstone.

§ 9. Influence of the direction of blastholes on the useful power developed by the electric drill and on the axial force.

As is known, when workings are driven, blastholes are laid not only horizontally and perpendicular to the face, but also to the side, as well as inclined upward and downward. In the work under consideration, the developed power (useful power) and axial forces were determined when drilling blastholes with the angles of inclination adopted in the present work. The power indicated is that developed for drilling proper—cutting—and for overcoming all harmful resistances experienced by the drill in the blasthole.

Fig. 28.

Fig. 29.

The axial forces and useful power were recorded throughout the drilling of the entire depth of the blastholes (at 10 cm intervals), but, from the standpoint of the question of interest to us, the average values of these indices, obtained for different inclinations of the blastholes, are of practical significance. These data will show how much they increase with the angle of inclination of the blastholes, and, consequently, how the angle of inclination affected the resistances experienced by the drill.

The average axial forces and useful power at the spindle for all the studied numbers of revolutions of the drill, obtained when drilling sandstone, are given in Table 16; when drilling shale—in Table 17; and a comparison of the data for sandstone and shale in Table 18. A graphical representation of the materials obtained is presented in Figs. 28, 29, 30, and 31.

On examining the data obtained, the following conclusions may be drawn.

1) Axial forces increase as the blastholes are lowered downward. This is clearly seen both for sandstone and, especially, for shale. The average (over the length of the blasthole) axial force for shale increases from 134 kg for blastholes \(25^\circ\) upward to 232 kg in blastholes \(25^\circ\) downward, i.e., an increase of 73%. For sandstone, these forces are respectively 250 and 355 kg, or show an increase of 40%.

Since the feed rate of the drill, the design of the drill and cutter, and the other drilling conditions are the same for all inclinations of the blastholes, it should be considered that the increase in axial forces is caused by internal resistances in the blasthole, which the drill receives from the drill cuttings and from friction against the walls of the blasthole. The greater increase in axial forces for shale is explained by the lower coefficient of removal of drill cuttings when drilling this rock. When drilling downward blastholes, compaction of the cuttings in the blasthole was observed, which caused an especially large increase in axial force.

2) Axial forces change almost not at all with changes in the number of revolutions of the drill. This is explained by the fact that the feed rate per one revolution of the drill remains almost constant.

3) Axial forces for sandstone are considerably higher than for shale. This is clearly maintained at all angles of inclination of the blastholes. This is explained by the fact that sandstone has greater hardness than shale. The increase in axial forces for sandstone ranges from 48.5 to 86.5%, and on average is 61.5%; that is, the axial force for sandstone is more than one and a half times higher than for shale.

4) The power expended at the spindle increases as the blastholes are lowered downward.

This is also clearly seen from the tables presented, both for shale and especially for sandstone.

The average (over the length of the blasthole) power for shale increases from 0.61 kW for blastholes \(25^\circ\) upward to 1.16 kW for blastholes \(25^\circ\) downward, or an increase of 90%. For sandstone this power is respectively 0.96 and 2.30 kW, or an increase of 140%.

Fig. 30.

Fig. 31.

Table 16 (continued).

Feed rate of the drill, mm/min	Number of drill revolutions per min	25° upward: average axial force, kg	25° upward: average spindle power, kW	15° upward: average axial force, kg	15° upward: average spindle power, kW	0° normal: average axial force, kg	0° normal: average spindle power, kW	15° downward: average axial force, kg	15° downward: average spindle power, kW	25° downward: average axial force, kg	25° downward: average spindle power, kW
180	193	280	—	293	0.60	304	0.80	390	1.66	390	1.93
210	237	225	0.38	244	0.70	—	1.30	267	1.85	290	2.17
260	286	—	1.57	—	1.73	—	1.74	—	2.15	—	2.33
400	429	242	1.57	268	1.63	304	1.88	344	2.30	384	2.80
On average		250	0.96	268	1.17	304	1.43	333	1.99	355	2.30
% ratio		100	100	107	122	121	149	133	207	142	240

Table 17 (for shale)

Feed rate of the drill, mm/min	Number of drill revolutions per min	Inclination of holes: 25° upward — average axial force, kg	Inclination of holes: 25° upward — average power per spindle, kW	Inclination of holes: 15° upward — average axial force, kg	Inclination of holes: 15° upward — average power per spindle, kW	Inclination of holes: 0° normal — average axial force, kg	Inclination of holes: 0° normal — average power per spindle, kW	Inclination of holes: 15° downward — average axial force, kg	Inclination of holes: 15° downward — average power per spindle, kW	Inclination of holes: 25° downward — average axial force, kg	Inclination of holes: 25° downward — average power per spindle, kW
180	193	—	0.26	—	0.39	—	0.45	—	0.48	—	0.67
210	237	130	0.47	155	0.54	180	0.60	—	—	—	—
260	286	118	0.77	155	0.87	220	1.05	163	1.18	175	1.38
400	429	154	0.95	229	1.16	208	1.25	212	1.28	288	1.44
Average . . . . .		134	0.61	180	0.74	203	0.84	188	0.98	232	1.16
% ratio . . .		100	100	134	121	151	138	140	161	173	190

Table 18

Rock	25° upward — Average axial thrust, kg	25° upward — Average spindle power, kW	15° upward — Average axial thrust, kg	15° upward — Average spindle power, kW	0° normal — Average axial thrust, kg	0° normal — Average spindle power, kW	15° downward — Average axial thrust, kg	15° downward — Average spindle power, kW	25° downward — Average axial thrust, kg	25° downward — Average spindle power, kW	Average values — Average axial thrust, kg	Average values — Average spindle power, kW
Shale	134	0.61	180	0.74	203	0.84	188	0.98	232	1.16	187	0.87
% ratio	100.0	100.0	100.0	100.0	100.0	100.0	100.0	100.0	100.0	100.0	100.0	100.0
Sandstone	250	0.96	268	1.17	304	1.43	333	1.99	355	2.30	302	1.57
% ratio	186.5	157.3	148.5	158.3	150.0	170.5	177.0	203.0	153.0	198.0	161.5	180.5

Fig. 32.

Vertical axis: Strength per bit, in t
Horizontal axis: Hole depth in m

Labels on curves: 500/193; 400/429; 360/286; average; 210/237; 180/193.

Fig. 33.

Vertical axis: Strength per bit, in t
Horizontal axis: Hole depth in m

Labels on curves: 380/193; average; 400/429; 360/286; 180/193; 210/237.

The considerable increase in power expended as the blastholes are deepened is explained by the great resistance of the drill cuttings, the greater vibration of the drill, and the greater friction, in connection with this, of the drill against the walls of the blasthole. The greater increase of this power for the lower blastholes in sandstone than in shale is apparently explained by temporary causes of an accidental nature.

5) The power at the spindle increases as the cutting speed increases (the number of revolutions of the drill).

This is clearly seen from the tables presented and requires no special explanation, since it is known that cutting speed increases power consumption.

6) The power at the spindle of the electric drill when drilling sandstone is considerably higher than when drilling shale.

This increase ranges from 57.3 to 103%, and on average amounts to 80.5%. It is explained by the greater hardness of sandstone in comparison with shale.

§ 10. Influence of the depth of blastholes on the useful power and axial force developed during drilling.

As the blasthole is deepened, the power and axial forces developed by the drill must undoubtedly increase. They will evidently increase because of the additional harmful resistances encountered by the drill as the blasthole is deepened: from the walls of the blasthole itself (the area of contact of the drill with the wall of the blasthole increases), from bending of the drill (greater friction against the walls of the blasthole), and from the drill cuttings (a greater quantity of them).

Observations were carried out in sandstone and shale for different operating regimes and for different angles of inclination of the blasthole. Since the influence of the angle of inclination of the blastholes on power and axial force was established in the preceding paragraph, here we shall determine the influence on these indicators only from the standpoint of blasthole depth. In order not to have the influence of blasthole inclination, we process the data obtained only for one horizontal direction of the blastholes.

Table 19 gives data on the axial forces and the useful power expended on drilling, and their graphical representation is given in Figs. 32, 33, 34, and 34a. The initial data are presented (for drilling the mouths of the blastholes) and the final data—when drilling the bottoms of the blastholes. The blastholes were drilled, as indicated earlier, to a depth of 1.8 meters. The data are given both for sandstone and for shale and for different drilling regimes.

From the data presented in Table 19, the following conclusions may be drawn.

1) The absolute value of the axial forces in sandstone is considerably higher than in shale, both under identical drilling regimes and in average values. The average (for all drilling regimes) initial axial force

Fig. 34.

Fig. 34-a.

Table 19.

Rate of drill feed, mm/min	Number of drill revolutions, min.	Rock	Axial force, kg — initial	Axial force, kg — final	Axial force, % increase	Spindle power, kW — initial	Spindle power, kW — final	Spindle power, % increase	Average power, kW	% ratio
180	193	Clay shale	80	180	125	0.44	0.76	72.5	0.60	100.0
210	237	Clay shale	130	230	77	0.52	0.79	52.0	0.65	108.0
260	286	Clay shale	145	295	103	0.78	1.16	48.6	0.97	161.5
400	429	Clay shale	120	295	146	0.93	1.48	59.0	1.20	200
Average ...		Clay shale	119	250	110	0.69	1.05	52.0	0.80	—
If taken as		Clay shale	100%	100%	—	100%	100%	—	100%	—
180	193	Sandstone	278	370	33	0.84	0.94	12.0	0.89	100.0
210	237	Sandstone	415	495	11	0.99	1.40	41.5	1.20	135.0
260	286	Sandstone	420	545	30	1.60	1.70	6.3	1.65	185.0
400	429	Sandstone	320	395	23.5	1.71	2.33	36	2.02	227
Average ...		Sandstone	366	451	23.5	1.28	1.59	24.2	1.44	—
% relative to shale ...		Sandstone	307	180.5	—	185.5	151.0	—	180%	—

for shale was 119 kg, and for sandstone—366 kg, or 207% greater. The average final axial force for shale was 250 kg, and for sandstone 451 kg, or 80.5% greater.

The considerable axial force for sandstone as compared with shale is explained by its greater strength. But, undoubtedly, the smaller coefficient of self-removal of cuttings when drilling shale had an effect on reducing the difference between the final forces for sandstone and shale (80.5%) as compared with the difference in initial forces (207%). The drilling cuttings evidently increased the final axial forces for shale much more significantly than when drilling sandstone.

2) The final axial forces are higher than the initial forces.

This increase averages 110% in shale and 23.5% in sandstone. Since drilling was carried out only with sharp cutters and under the same conditions, the reasons for the growth of axial forces are the following factors:

a) as the blasthole is deepened, the area of contact between the drill and the walls of the blasthole increases, and consequently so does the friction force caused by this;

b) with an increase in the depth of the blast-hole, the longitudinal bending of the drill increases, and consequently so does the force of its friction against the walls of the blast-hole;

c) with an increase in the depth of the blast-hole, the resistance of the drill cuttings increases.

The considerable increase in axial forces in shale as compared with sandstone is explained by the action of the drill cuttings, whose self-removal coefficient in shale, as indicated earlier, is much lower than that in sandstone. In addition, shale drill cuttings are more viscous and less loose than sandstone cuttings. All these facts, as direct observations have shown, play a very large role both in increasing the axial forces and in the power consumed for drilling.

As for the drilling regime, it is known that the rate of feed per one revolution of the drill has the decisive influence on the axial forces. But since in our investigation this rate was almost constant (see Table 1), it had no effect. The influence of the number of revolutions of the drill on the axial forces, as is evident from the materials obtained (Table No. 19), does not occur.

3) The useful power expended on drilling in sandstone is considerably higher than in shale. This is clearly visible from the table for all drilling regimes in both rocks. The average initial power in sandstone is higher than that in shale by 85.5%. It must be considered that this increase is entirely due to the greater strength of the sandstone, since in the initial stage of drilling blast-holes (at their mouths) the influence of drill cuttings is negligible, and the drilling means and working conditions were identical for both rocks.

The average final power in sandstone is higher than that in shale by 51%. The increase is still due to the strength of the rock, but the reduction in the difference in the power consumed (51% instead of 85.5%) must be explained by the greater influence of drill cuttings in shale blast-holes than when drilling in sandstone.

4) The final power is higher than the initial power. This is clearly expressed in both rocks and for all drilling regimes. This increase amounts on average to 52% in shale and 24.2% in sandstone. The greater increase in shale than in sandstone is explained by the above-mentioned causes—the influence of drill cuttings.

5) With an increase in cutting speed (the number of revolutions of the drill), the power expended on drilling increases.

From Table 19 and in Figs. 32 and 33 it is evident that this increase is sharply expressed for both rocks. In shale this increase amounted to 100%, and in sandstone to 127%, with an increase in cutting speed (429 revolutions instead of 193) by 122%; consequently, for each percent increase in speed we have one percent increase in the power expended on drilling sandstone and about 0.82% increase in the power for shale.

Fig. 35.

Fig. 36.

In Fig. 36: vertical axis — “Axial force, kg”; horizontal axis — “Blunting in mm”; curve labels — “for sandstone” and “for shale.”

§ 11. Effect of Cutter Wear on the Power and Axial Force of an Electric Drill

To determine the effect of cutter wear on the power and axial force of an electric drill, blastholes of horizontal direction were drilled, but with cutters of different degrees of dulling. Drilling was carried out both in argillaceous shale and in sandstone, and under different regimes of feed rate and cutting speed.

For comparison of the results, some blastholes were drilled with sharp cutters, other blastholes with cutters dulled by 0.5 mm (ab in Fig. 35), and a third group of blastholes was drilled with cutters dulled by 1 mm (cd in Fig. 35). The degree of dulling was calculated from the thickness of the cutting edge (insert), as indicated in Fig. 35.

Measuring the degree of dulling specifically by the thickness of the cutting edge (blade) is the most visual and accessible method under production conditions. Whereas determination of the degree of dulling in other directions—for example, along the front face—is associated with measuring angles and with geometric constructions, which is extremely inconvenient under production conditions and not always accessible to the operating personnel. Dulling of the cutter edge was produced on an “extra” carborundum grindstone, so that the dulled cutter began work with a prepared edge. This guaranteed the obtaining of more accurate results.

The blastholes were drilled at a distance of 4–5 cm from one another, so that they passed through identical conditions.

During observation, readings of the ammeter, voltmeter, and dynamometer were taken every 10 cm of the blasthole.

Tables 20, 21, 22 and Figures 36, 37, 38 and 39 show the data obtained for the power expended on the actual drilling and for the developed axial forces for different degrees of dulling of the cutter cutting edge when drilling sandstone and argillaceous shale at different regimes of cutting speed and feed.

Table 20 (for sandstone)

Feed rate of drill, mm/min	Number of drill revolutions, rpm	Indicators	Sharp	Dulled by 0.5 mm	Dulled by 1 mm
180	193	Power at spindle, kW	0.65	1.32	1.53
180	193	% ratio	100	203	236
180	193	Axial forces, kg	257	462	649
180	193	% ratio	100	180	253
210	237	Power at spindle, kW	0.98	1.58	1.65
210	237	% ratio	100	161	169
210	237	Axial forces, kg	314	493	619
210	237	% ratio	100	157	197

Fig. 37.

Fig. 38.

Drill feed rate, mm/min	Number of drill revolutions per min	INDICATORS	Sharp	Blunted by 0.5 mm	Blunted by 1 mm
260	286	Power at spindle, kW	1,9	2,07	2,52
260	286	% ratio	100	109	133
260	286	Axial force, kg	366	400	468
260	286	% ratio	100	109	128
400	429	Power at spindle, kW	1,86	2,23	2,77
400	429	% ratio	100	120	149
400	429	Axial force, kg	270	361	420
400	429	% ratio	100	133	155
Average values		Power at spindle, kW	1,35	1,80	2,12
Average values		% ratio	100	133	157
Average values		Axial force, kg	302	429	538
Average values		% ratio	100	142	178,5

Table 21 (on shale)

Drill feed rate, mm/min	Number of drill revolutions per min	INDICATORS	Sharp	Blunted by 0.5 mm	Blunted by 1 mm
180	193	Power at spindle, kW	0,44	0,90	1,21
180	193	% ratio	100	204	275
180	193	Axial force, kg	97	306	407
180	193	% ratio	100	315	420
210	237	Power at spindle, kW	0,62	1,12	1,67
210	237	% ratio	100	180	270
210	237	Axial force, kg	178	318	447
210	237	% ratio	100	179	250
260	286	Power at spindle, kW	1,05	1,62	1,91
260	286	% ratio	100	154	182
260	286	Axial force, kg	212	427	531
260	286	% ratio	100	201	250
400	429	Power at spindle, kW	1,26	1,42	1,67
400	429	% ratio	100	113	133
400	429	Axial force, kg	245	278	395
400	429	% ratio	100	113,5	161
Average values		Power at spindle, kW	0,85	1,27	1,62
Average values		% ratio	100	150	191
Average values		Axial force, kg	183	332	445
Average values		% ratio	100	181,5	243

Table 22.

Indicators	Unit of measure	Sharp	Blunted by 0.5 mm	Blunted by 1 mm
Average axial forces for shale	kg	183	332	445
Average axial forces for sandstone	"	302	429	538
% ratio of sandstone to shale		165	129	121
Average spindle power for shale	kW	0.85	1.27	1.62
Average spindle power for sandstone	"	1.35	1.80	2.12
% ratio of sandstone to shale		159	141	131

From consideration and analysis of the data obtained, one may arrive at the following conclusions.

1) As the degree of blunting of the cutter blade increases, the axial forces increase. This is clearly evident from the data in Tables 20, 21, and 22 and from the diagram presented, both with respect to sandstone and shale, as well as for all the tested cutting speeds and feeds.

2) As the strength of the rock increases, the axial forces with a sharp cutter increase. This is plainly evident from the data in the tables and diagram with respect to sandstone and shale. The average axial force with a sharp cutter for shale was 183 kg, while for sandstone it was 302 kg, or 65% greater.

3) As the blunting of the cutter increases, the influence of rock strength on the increase in axial force decreases. This is clearly seen from the data in Table 22. It follows from it that the increase in axial pressure for sandstone (in comparison with shale) with a sharp cutter is 65%; when it is blunted by 0.5 mm it is already 29%; and with wear of 1 mm this increase in axial pressure is only 21%.

It was not possible to determine the influence of the drilling regime on the axial forces under these conditions, since the feed per revolution of the drill bit (and the axial force depends mainly on it) remained almost unchanged. As for the cutting speed (the number of revolutions of the drill bit), it does not affect the axial forces.

4) As the blunting of the cutter blade increases, the power consumed at the spindle increases significantly.

This is clearly evident from the presented tables, both with respect to the two rocks and to the different drilling regimes.

The average power for sandstone with a cutter worn by 1 mm is 57% higher than with a sharp cutter, and for shale it is 91% higher.

5) As the strength of the rock increases, the power at the spindle increases significantly.

The average spindle power for shale with a sharp cutter was 0.85 kW, while for sandstone it was 1.35 kW, or higher by

Fig. 39.

Fig. 40.

Fig. 41.

59%. This is explained by the greater temporary resistance to cutting of sandstone in comparison with shale.

6) As the cutter becomes blunter, the influence of rock hardness on the increase in spindle power decreases.

From Table 22 it is evident that the increase in power for sandstone (in comparison with shale) with a sharp cutter is 59%, while with wear of 0.5 mm it is already 41%, and with blunting by 1 mm it is only 31%.

7) With an increase in cutting speed (number of drill revolutions), the spindle power increases.

With a sharp cutter, the increase in power at 429 drill revolutions per minute compared with 193 revolutions for both rocks was 186%; with cutters blunted by 0.5 mm, this increase was 69% for sandstone and 58% for shale; and with undercutting of the cutter along the thickness of its edge by 1 mm, the increase was 81% for sandstone and 38% for shale.

These data indicate that, as the cutter becomes blunter, the influence of cutting speed on power decreases, especially for shale.

Observations were also made of cutter durability. The observation was carried out on shale. For this purpose only horizontal blastholes were drilled, each to a depth of 1.80 meters with one cutter, keeping a record of the footage drilled and taking measurements of the degree of cutter blunting.

In all, twenty blastholes were drilled, or 36 running meters; the cutter was blunted by 0.3 mm in blade thickness at its peripheral end.

During drilling, readings were taken from the dynamometer, ammeter, and voltmeter in the usual manner.

The results obtained for axial force and useful power are presented in Table 23 and in Figs. 40 and 41.

Table 23

Blasthole No.	Drilled, running m	Axial force, kg	Power at drilling, kW	Blasthole No.	Drilled, running m	Axial force, kg	Power at drilling, kW
1	1.8	85	0.320	11	19.8	168	0.570
2	3.6	76	0.340	12	21.6	162	0.580
3	5.4	77	0.420	13	23.4	169	0.595
4	7.2	76	0.355	14	25.2	175	0.600
5	9.0	95	0.385	15	27.0	185	0.620
6	10.8	86	0.440	16	28.8	191	0.635
7	12.6	127	0.445	17	30.6	197	0.660
8	14.4	140	0.520	18	32.4	207	0.680
9	16.2	159	0.550	19	34.2	227	0.710
10	18.0	164	0.565	20	36.0	259	0.730

Fig. 42.
Vertical axis: Amount of Pobedit carbide worn away during drilling, g.
Horizontal axis: Dulling, mm.
Visible scale marks: horizontal axis — 0.5, 1.0, 1.5; vertical axis — 0, 2, 3.

Fig. 43.
Vertical axis: Amount of Pobedit carbide removed along the chord, mm³.
Horizontal axis: Dulling, mm.
Visible scale marks: horizontal axis — 0.5, 1.0, 1.5; vertical axis — 0, 10, 20, 30, 40, 45.

Considering the results obtained for axial force, we see that it increased steadily as the cutter became blunt. Connecting the measurement points for the axial forces gives a somewhat broken line, inevitably obtained in observations under production conditions, but this line fully corresponds to and can be replaced by some average straight line showing that the force increases with cutter blunting according to a straight-line law; i.e., this observation confirmed what was obtained above.

Observations show that blunting of the blade begins and proceeds gradually and very slowly. Neither during this test nor in the course of other observations was any deviation from this rule noted.

The data presented for the power developed on the spindle show a steady increase as the cutter blade becomes blunt. The curve (Fig. 41) of power growth is almost identical to the curve of growth of the axial forces. The power curve can likewise be taken as some average straight line, and it can be considered that the increase in power as the cutter blade becomes blunt proceeds according to a straight-line law.

When the cutter becomes blunt, the pobedit plate is consumed (worn away). With increasing blunting, the consumption of the pobedit plate due to wear increases sharply (Fig. 42 for cutter No. 2). The consumption of the pobedit plate due to sharpening increases still more significantly (Fig. 43). The total consumption of the pobedit plate due to wear and sharpening, depending on the degree of blunting of its cutting edge, is shown in Fig. 44.

Establishing the economically advantageous limit of cutter blunting must be specially studied for different operating conditions and under different regimes.

§ 12. Comparison of Twisted Drill Steels of Rhombic and Rectangular Cross Section

As is known, in coal-mining practice twisted drill steel of two cross sections is used: rhombic (twisted from rhombic steel) and rectangular—twisted from steel of rectangular cross section. The tested rhombic steel had a helical-groove pitch of 110–115 mm, and the rectangular steel, 50–52 mm; the diameter of the drills of both profiles was the same and was 30 mm. We did not have drills with the same helical-groove pitch at our disposal. The comparison was carried out according to the following indicators: 1) power consumed for drilling; 2) axial force developed; 3) coefficient of self-removal of drill cuttings.

To determine these indicators, blastholes were drilled under identical conditions in sandstone with drills of rhombic and rectangular cross section at a feed rate of 400 mm/min and at 429 revolutions per minute. The blastholes were drilled side by side. The observation method was analogous to that indicated earlier.

Fig. 44.

Vertical axis: Total consumption of pobedit

Horizontal axis: Blunting, in mm

Fig. 45.

Vertical axis: Coefficient of resistance

Horizontal axis: Angle of inclination of holes

Axis labels: 25° upward; 15° upward; 0° normal; 15° downward; 25° downward

Curve labels: $\varnothing = 45\ \mathrm{mm}$; $\varnothing = 30\ \mathrm{mm}$

$18 \times 35$	ditto	ditto	ditto	0.4	80	103	100	0.45	85	97	85	111	50

The data obtained for all the above-mentioned indicators are given in Table 24.

Table 24

Drill feed rate, mm/min	Number of drill revolutions per min.	Steel section	Power for drilling, kW	% ratio	Axial force, kg	% ratio	Coeff. of self-removal of drill cuttings	% ratio
400	429	Rectangular	2.28	121	365	117.5	0.605	93.8
400	429	Rhombic	1.88	100	310	100	0.645	100

From the data presented, the following conclusions may be drawn:

1) With twisted steel of rectangular section, the axial forces are greater than with rhombic steel. This is due to the greater friction of the rectangular steel against the walls of the blastholes, since its ribs are flat.

2) The coefficient of self-removal of drill cuttings for rectangular steel is somewhat worse than for rhombic steel.

This can probably be explained by the fact that, because of its flat ribs, it does not sufficiently tear the cuttings away from the walls of the blasthole. Nevertheless, this question requires further verification.

3) The useful power expended on drilling is higher with rectangular steel than with rhombic steel. This is explained by the greater friction of the drill ribs against the walls of the blasthole and by the poorer coefficient of self-removal of drill cuttings.

Rectangular steel, having lower strength in longitudinal bending, bends more easily and more quickly, touches the walls of the blasthole with its ribs, increasing the friction force of the drill against the walls. In addition to the results obtained, it must be added that the comparison of steel profiles was made with different pitches of the helical groove, which may, to a certain extent, have affected the numerical data obtained. When steel with the same pitch is obtained, a repeated comparison will be made, in which the numerical indicators may possibly change somewhat, but the general negative conclusion regarding the rectangular profile will evidently remain.

§ 13. Comparison of rhombic twisted steel of different diameters.

Rhombic steel of two sizes is supplied to coal mines:¹) 16 × 30 mm and 18 × 35 mm (by the rhombus, i.e., the diameters are 30 and 35 mm, or close to them). Two sets of rhombic drills were compared. One set inclu—

¹) “Price List for Materials.” Kuzbassugol Publishing House, 1937.

included drills with a diameter of 30 mm and a helix pitch of 110–115 mm, and the other set—with a diameter of 35 mm and a helix pitch of 80–85 mm. The cutter diameter in both sets was the same (40 mm).

The comparison of these drills was carried out according to the following indicators:
1) by the self-cleaning coefficient;
2) by the developed power and axial thrust during drilling.

The indicators were determined when drilling clay shale, since difficulties are observed in self-cleaning from cuttings in this rock. The data obtained are presented in Tables 25 and 26 and in Fig. 45.

Table 25

Drill diameter, mm	25° upward: self-cleaning coefficient	25° upward: % ratio	15° upward: self-cleaning coefficient	15° upward: % ratio	0° normal: self-cleaning coefficient	0° normal: % ratio	15° downward: self-cleaning coefficient	15° downward: % ratio	25° downward: self-cleaning coefficient	25° downward: % ratio
30 mm	0.652	100	0.584	100	0.485	100	0.370	100	0.190	100
35 mm	0.794	121	0.728	125	0.620	128	0.545	147	0.493	260

Table 26

Drill diameter, mm	Rock	Drilling rate, mm/min	Drill speed, rpm	15° upward: spindle power, kW	15° upward: spindle power, % ratio	15° upward: axial thrust, kg	15° upward: axial thrust, % ratio	0° normal: spindle power, kW	0° normal: spindle power, % ratio	0° normal: axial thrust, kg	0° normal: axial thrust, % ratio	15° downward: axial thrust, kg	15° downward: axial thrust, % ratio
16×30	Clay shale	180	193	0.5	100	102	100	0.53	100	115	100	223	100
18×35	" "	"	"	0.4	80	103	100	0.45	85	97	85	111	50

From the tables and the figure it follows:

1) With an increase in the drill diameter and a decrease in the helix pitch, the self-cleaning coefficient of the drill cuttings increases. This increase grows especially with an increase in the downward inclination angle of the blastholes. This is explained by the fact that larger-diameter drills separate the cuttings better from the walls of the blasthole and do not allow them to adhere to them. And a reduction in the helix pitch facilitates the movement of the drill cuttings along the helical line.

2) With an increase in the drill diameter and a decrease in the helix pitch (with a constant cutter diameter), the axial forces decrease. This decrease is especially noticeable when drilling downward-inclined blastholes, since slate dust is extremely poorly self-removed in downward holes. Downward-inclined holes in slate were drilled with difficulty by drills of 30 mm diameter, whereas with drills of 35 mm diameter these holes could be drilled without particular complications.

3) With an increase in the drill diameter and a decrease in the helix pitch (with a constant cutter diameter), the power consumed for drilling decreases.

It follows from Table 26 that, under identical conditions, the developed power with drills of 35 mm diameter amounts to 80–85% of the power with drills of 30 mm diameter. This is explained by the fact that, with drills of 35 mm diameter, the drill cuttings are removed better and have less effect on the developed power; also, evidently, the greater stiffness of the drill produces less longitudinal bending and less friction against the walls of the blasthole.

In conclusion, it must be emphasized that the data were obtained with drills not only of different diameters, but also with different helix pitches; therefore, the influence of these parameters separately on the compared indicators has not been determined. When steel of the appropriate dimensions is obtained, this question will be carried through to completion.

§ 14. Disadvantages of the electric drill.

Study of the operation of the electric drill under production conditions and during the experimental part of the present research work showed that the equipment has a number of disadvantages.

1) Considerable weight (120 kg). Because of this, two people have to be assigned to installing and repositioning the drill, whereas during the drilling itself only one person is required.

2) Because of the absence of any lifting devices, moving the drill along the column also has to be done by two people.

3) The column is heavy (50 kg). It is difficult for one person to set it up.

4) Insufficient stiffness of the column. Under an axial force of 600–700 kg it bends strongly and threatens to fail.

5) The bows (4), brackets (5), and screws (6) are extremely weak; they quickly become loose, wear out, and hold the drill poorly by the trunnions.

6) The presence of a friction-cam clutch, which does not permit engaging the working feed of the drill while running. It would be better to have only a friction clutch.

7) Unreliable operation of the friction-cam clutch in its friction part.

8) Absolutely unreliable operation of the idle feed. The drill almost never has it, since the wire ring (150)

quickly wears its own groove in the bronze washer (10), after which it begins to slip easily and no longer transmits force to the intermediate shaft (12).

9) The bronze nut (18) of the friction-cam clutch is worn away because of friction against the washer (19) and the head of the rack (14).

10) The presence of sliding-friction bearings increases lubricant consumption, complicates its production, and reduces the efficiency of the reducer.

11) The design of the rear labyrinth on the motor shaft is complicated; it can be made simpler, which has already been planned by the plant for implementation.

12) The fastening of the axle (32), on which the second and third gears are seated, is unreliable. The bolt securing the head of this axle unscrews, the axle turns, and the lubrication of its surface is disturbed.

13) The controller chamber is of insufficient size; it is extremely difficult to connect and disconnect wires in it, which quite often leads to damage to their insulation.

14) The motor cavity is insufficiently protected against oil entering through the labyrinth seals.

15) The method of fastening the drill in the spindle chuck (by a key) is unsatisfactory. A considerable amount of time is required to knock out this key when changing drills.

16) The internal opening in the spindle chuck for the drill is made with a round cross-section; this forces round shanks to be welded to the drills, which creates a weak point in the drills: they usually break at this weld. The internal opening in the chucks must be made in accordance with the profile and size of the shank of the drills used. If drills of different profiles or sizes are used, then several chucks will have to be kept, but this will not cause any special difficulty, since the drills can be selected as complete sets for individual electric drills.

17) Ball-type oilers are inconvenient under mine conditions. The lubrication openings should instead be closed with ordinary plugs equipped with a device for screwing and unscrewing them by hand.

§ 15. Conclusion and Proposals.

The material obtained as a result of this research work will, in the course of further work on this topic, be supplemented, generalized, and analyzed in order to provide a theoretical basis for it.

In the present article we have set forth only the primarily processed experimental material and, on the basis of the data obtained, have made the corresponding generalizations with explanations. These generalizations amount to the following.

1) With an increase in the strength of the rock, the coefficient of loosening of the drilling dust decreases (Table 5).

For sandstone, with a drill feed of about one mm per revolution, it was 1.60–1.80 (on average 1.75), and for clay shale, 2.16–2.18 (on average 2.17).

When drilling under another regime, with another chip thickness, the loosening coefficient of the drill cuttings will, of course, change, but in all probability only insignificantly, since the chip thickness varies within small limits.

It may certainly be considered that when drilling softer rocks, a greater quantity of drill cuttings is obtained (by volume) than when drilling harder rocks. This must be taken into account when selecting the design of drills—the cross section, diameter, pitch of the helical groove, etc.

2) The coefficient of self-removal of drill cuttings when drilling hard rocks is higher than when drilling soft rocks (Tables 6, 7, and 8).

The average coefficient of self-removal of cuttings for all tested cutting and feed regimes when drilling sandstone ranged from 0.232 (for blastholes inclined 25° downward) to 0.815 (for blastholes inclined 25° upward). When drilling clay shale, the coefficient was from 0.190 to 0.652, respectively.

When drilling sandstone, the average coefficient of self-removal of cuttings for the various conditions is higher than that for shale by 20.8%–26.8%, and on average by 23.3%.

Moreover, this difference increases somewhat for blastholes directed upward and decreases for blastholes directed downward (Table 8).

3) The coefficient of self-removal of drill cuttings increases with an increase in the angle of rise of the blastholes. This regularity is clearly expressed for different rocks and for any drilling regime (Figs. 17 and 18).

4) The coefficient of self-removal of drill cuttings increases somewhat with an increase in the number of revolutions of the drill. This regularity occurs in different rocks and under different drilling regimes, as is seen in Figures 17 and 18.

5) The drill cuttings obtained are coarser when drilling softer and more brittle rocks (Figs. 22 and 23 and Table 11).

6) At the mouth of the blasthole, the cuttings are coarser than those from the bottom of the blasthole (Figs. 20 and 21 and Table 11).

7) With an increase in the number of revolutions of the drill, the coarseness of the drill cuttings obtained from the bottom of the blastholes decreases somewhat (Tables 9 and 10).

8) The natural angle of repose of drill cuttings from hard rocks is greater than that from softer rocks (Table 12).

9) The coefficients of friction, at rest and in motion, for drill cuttings from hard rocks are higher than those from softer rocks (Table 12).

10) The power consumption at the spindle for overcoming the resistance of the drill cuttings and the friction of the drill against the walls of the blasthole in hard rocks is higher than in soft rocks (Table 15 and Fig. 27).

This power increases especially with an increase in the dip angle of the blastholes. The inclination angle of the blastholes has a particularly strong effect on this power in soft rocks.

11) The axial forces on the spindle increase with an increase in the dip of the blastholes, especially when drilling soft rocks (Tables 16, 17, and 18).

12) The axial forces, with a change in the number of revolutions of the drill at constant cuttings, change hardly at all (Tables 16 and 17).

13) The axial forces when drilling harder rocks are much higher than when drilling soft rocks (Tables 16, 17, and 18).

14) The power consumed at the spindle when drilling harder rocks is higher than when drilling softer rocks (Tables 16, 17, and 18).

15) With an increase in the cutting speed, the power developed at the spindle increases (Tables 16, 17, 18 and Figs. 29 and 30).

16) The power at the spindle also increases with an increase in the dip angle of the blastholes (Tables 16, 17, and 18 and Figs. 29, 30, and 31).

17) With an increase in the depth of the blasthole, the power consumed at the spindle and the axial forces increase, especially when drilling softer rocks (Table 19 and Figs. 32, 33, 34, 34-a).

18) With an increase in the degree of dulling of the cutter, the axial forces and the power consumed at the spindle increase considerably (Tables 20, 21, 22 and Figs. 36, 37, 38, 39).

With increasing dulling of the cutter, the influence of rock hardness on the increase in axial force and the power developed at the spindle decreases somewhat.

19) Drill steel of rhombic cross section gives better indices in terms of consumed power, developed axial forces, and coefficient of self-removal of fines than steel of rectangular cross section (Table 24).

20) The diameter of the cutter and the diameter of the drill steel must be in definite ratios when drilling rocks of different hardness. It has so far been possible to establish that a considerable difference between the diameters of the cutter and the drill gives worse results than a more moderate difference. It may provisionally be considered that the diametral clearance between the ribs of the drill and the walls of the blasthole should be about five mm. But this question requires further study.

21) The existing column-type electric drill and column require reconstruction, or the creation of new, lighter and stronger ones (the column).

On the basis of the experience obtained in studying rotary drilling of blastholes, the following proposals suggest themselves:

1) Study of the geometry of the cutter. On the basis of studying its economic durability and dynamic properties under various conditions and with different elements of its geometry, dimensions, and configuration, it will be possible to arrive at the most rational cutter for particular operating conditions and for the corresponding regimes.

2) Study of drill bits, their geometry, mechanical strength, and dynamic properties for various operating conditions and modes.

3) Development of a theory of rock drilling.

4) Design of a new column-type electric drill, based on the above-mentioned material (cutters, drill bits), for rocks of sedimentary origin.

5) Study of rotary drilling of igneous rocks and ores with the aim of replacing pneumatic drilling with electric drilling as quickly as possible, on the basis of creating a column-type electric drill for these rocks.

6) Technical-economic comparison of purely rotary and rotary-percussive drilling with electric drills and electric perforators.

1. The higher percentage of transition of sandstone cuttings into the fine class evidently also depends on the fact that these cuttings are generally finer than shale cuttings and, during crushing, pass more rapidly into a lower coarseness class. ↩↩↩

2. G. V. Rodionov.—Column-mounted electric drill of type EK-2. Mining Journal, 1938, No. 3.
   G. V. Rodionov and P. A. Makhin.—Results of testing hand-held electric drills manufactured by the Tomsk Mechanical Plant. Mining Journal, 1937, No. 16.
   Materials of the brigade of the Kharkov Machine-Tool and Instrument Plant. “Card file TsITEIN—TEKhSO No. 2216, series 28.” 1937.
   G. S. Pantyushev.—Testing the operation of a column-mounted electric drill under operating conditions. Coal, 1937, No. 138.
   G. S. Pantyushev.—Investigation of the operation of a hand-held electric drill under operating conditions. Coal, 1937, No. 139.
   A. P. Ostrovskii.—The EBK drill and electric drilling automatic device. Mining Journal, 1937, No. 16. ↩

3. With the direct participation of students of the mining faculty: Chelyshev, Shilov, Garkun, Evseev, Fomin, Tomlyanovich, Andreev, and Denisenko. ↩