XIX w UNDERGROUND MINING - library.aimehq.orglibrary.aimehq.org/library/books/The Porphyry Coppers, 1933/393.pdf · Controlled intermittent flowage of the ore by gravity ... Miami

CHAPTER XIX

UNDERGROUND MINING

w ITHOUT in any way detracting from the credit due those engineer-miners of copper who operate with power shovels, it may be said that compared with

block-caving underground their work is simplicity itself. True enough, ingenuity and foresight of high order are required to lay out an open-cut operation to the best advantage; good organization and intelligent care and use of machinery are required to achieve minimum costs. But really successful undercut block-caving demands all this; and, in addition, it requires a conception and application of unseen natural forces that reflect a type of genius that is rare.

Assuming that preparatory work has gained access to the orebody, mining of whatever kind consists essentially of three operations :

1. Breaking the ore "in place," as Nature left it, into pieces small enough for convenient handling.

a. Loading these pieces into some kind of a vehicle on wheels. 3. Transporting the broken ore to some central plant on the

surface where the process of extracting the valuable metals can be commenced.

In shovel mining, dynamite and steam or electric power are used-with remarkable efficiency, be it said-to accomplish each of these three processes. In underground-caving methods the old force of gravity is called upon to accomplish, in large part, processes 1 and a. To do this with success is an artistic achievement of a sort.

Put it another way. The earlier methods of underground mining constituted a translation of surface procedure to an underground theater. The ore was blasted and dug out in small quantities; in its place were put props, usually of timber, to prevent the caving of the rock above; and men toiled with

393

394 THE PORPHYRY COPPERS

shovels to lift the ore into small cars in which it was started on its way to the surface. When the bodies of ore were large it became necessary to supplement the timber supports by refilling with waste rock the voids made by the extraction of ore. I t became evident that a large part of the cost of operations lay in timbering and filling to resist the urge of the overlying rock to cave-in short, in an effort to defy the force of gravity. In the method used now in the underground Porphyry mines gravity is the potent ally, not the opponent of the miner. But the task of harnessing gravity, of controlling it, is by no means an easy one; the problem has different angles a t each mine. In the following pages an attempt will be rnade to give as clearly as possible a general idea of how the job is done.

ELEMENTS OF UNDERCUT CAVING

The essential elements of an undercut block-caving operation may be enumerated as follows:

1. Driving a system of main horizontal haulageways 30 f t . or more beneath the block of ore to be mined. These haulageways connect with the hoisting shaft or with the main adit through which the mine is worked.

8. Driving from the haulageways a system of transfer raises (usually inclined) up toward the block of ore. Generally, the tops are connected by horizontal openings that constitute the grizzly level.

3. Driving short "finger" raises (usually two or four) from the top of each transfer raise to the bottom of the block of ore to be extracted. In plan the centers of the tops of these raises are spaced 1845 to RO f t . apart in each direction.

4. Removing a layer of ore a t the bottom of the block ("undercutting") so as to induce progressive breaking down, settling, and crushing of the ore into relatively small pieces.

5. Controlled intermittent flowage of the ore by gravity through the finger and transfer raises to cars on the haulage level (" drawing").

Felix McDonald, now mine superintendent a t Inspiration, appears to have employed the progenitor of this method in 1906

UNDERGROUND MINING 395

at the mine of the Ohio Copper Co., adjoining the Utah Copper property a t Bingham. He and his associates were familiar dith the so-called sublevel caving system used in the iron mines in Michigan. The reason that the sublevel method would not do at Bingham was that when it was used most of the ore had to be shoveled by hand into chutes; and the cost of this procedure was prohibitive with the low-grade Porphyry ore. A block 100 by 100 ft. in section and 30 f t . thick was prepared for block- caving a t the Ohio mine. The vertical thickness was soon changed to 60 f t . and the scheme enjoyed a measure of success.

In 1911, under the supervision of Mr. McDonald, an adapta- tion of the Ohio system was introduced at Inspiration. After some experimenting three "branches" were run from each main raise; and a square-set of heavy timber was placed a t the junc- tion. Though far from being ideal, this method accounted for a production of 50,000,000 tons of ore during the ensuing dozen years.

Contemporaneous with the early work a t Inspiration was that of L. S. Cates and his staff a t Ray. In 1911 they adopted a combination method of shrinkage stoping and caving. From a sublevel 30 ft. above the haulageway vertical shrinkage stopes 15 f t . wide were carried to the capping.

A shrinkage stope is one in which work advances upward, the "back" being maintained approximately horizontal. Drilling and blasting are done overhead from the top of the broken rock which accumulates until the stope is completed. As rock increases in bulk by one-third when broken, some ore must be drawn off a t the bottom as upward progress is made or there would be no space in which the miners might work under the back. Always intervening between adjacent shrinkage stopes a t Ray was left a vertical pillar 10 ft. thick. After the shrinks had been completed the pillars were undercut and allowed to cave. The pillar ore was drawn contemporaneously with the ore remaining in the shrinkage stopes. While this system was fairly successful and accounted for a large production of ore, various improvements were made from time to time, one important modification being the decrease in the amount of shrinkage stoping. Only the shrinks a t the boundaries of the block were driven to the

THE PORPHYRY COPPERS

full height of the ore, the others being carried up only 35 ft., with consequent economy. More recently a t Ray a system essentially similar to that used in the other Porphyries has been adopted.

Miami introduced undercut caving for some of its ore as early as 1914; and in 1919 it discontinued all other methods, including shrinkage and sublevel caving as well as topslicing, that had been in vogue in part of the mine since the inception of operations. At its Ruth mine, Nevada Consolidated commenced production in 1915, using undercut caving from the start. At Morenci caving operations started in 1922; and a t Bisbee Phelps Dodge commenced caving the East orebody in the Sacramento Hill area in 1925. In recent years Braden has replaced its combination shrinkage and pillar-caving method with what its engineers style "block undercutting"-a descrip- tive name that well could be applied to the Porphyry operations in Arizona and Nevada. In 1929 about 90 per cent of the Braden ore was obtained from undercut blocks. Andes, of course, caved from the outset of its producing era starting in 1926. Table 37 shows the various undercut caving mines in the Porphyry group with the approximate daily rate of sustained production achieved to date a t the several properties.

GENERAL CONSIDERATIONS IN BLOCK-CAVING

Before describing actual operations a t two of the mines it will be helpful to outline the major problems that arise, the characteristics that make an orebody suitable for undercut block- caving and, by contrast, the characteristics that add to the difficulties of operation.

1. Warrant for using method.-Generally speaking, the orebody should be extensive both laterally and vertically so as to justify the large preparatory expense; and it should be fairly uniEorm as to shape and grade so that it will be more profitable to mine the entire mass than to mine only the richer portions by more costly methods of extraction. This, of course, is one of the distinctive characteristics of all the Porphyries.


2. Inducing ore to cave.-The entire mass, including the capping as well as the ore, should be traversed by a multiplicity of joint or fracture planes. These must be both large and small, and at diverse angles, so that the rock will break in reasonably small pieces. Most of the Porphyry ores are comparatively

TABLE 37 MAXIMUM RATE OF PRODUCTION BY BLOCK-CAVING

soft; but the degree of rock hardness is less important than the presence of fractures along which the rock can break. In some instances undercutting must be supplemented by partly cutting off the block a t the side and end boundaries or by otherwise weakening it. It is highly important that the ore should crumble gradually rather than break in enormous solid blocks containing thousands of tons.

3. Preventing dilution and loss.-Obviously little selectivity in mining can be exercised. At best 10 or 15 per cent of the ore in a block is likely to be lost, and an equivalent proportion of waste material usually- will be drawn. Uniformity of drawing is the means for minimizing this difficulty. A multiplicity of uniformly spaced draw points assists regularity of drawing. Exceedingly intelligent and skillful control by the operating staff is essential; and it goes without saying that the absence of large included masses of very lean material is a big help for which only Nature can be credited.

4. Maintenance of the extraction workings.-The readjust- ment of powerful forces released by the undercutting of the ore tends t o put pressure on the mining or extraction openings through which the ore is attacked. Horizontal openings a t the

Property

Inspiration. . . . . . . . . . . . . . . . . Ray. . . . . . . . . . . . . . . . . . . . . . . Nevada Consolidated (Ruth). Miami.. . . . . . . . . . . . . . . . . . . .

Property

. . . . . . . . . . . . . . . . . . . Morenci Copper Queen. . . . . . . . . . . . . Andes. . . . . . . . . . . . . . . . . . . . . Braden.. . . . . . . . . . . . . . . . . . .

Daily capacity, tons of ore

20,000 10,000 3,000

18,000

Daily capacity,

tons of ore

5,000 3,000

18,000 16.000

398 THE 1WRPHYRY COPPERS

grizzly level .just below the draw points, and on the main haulage level below, in spite of heavy timbering, are likely to be partly closed by squeezing. The task of relieving this pressure :tnd of replacing timbers is exceedingly tedious and slow; expert miners are required and therefore the work is expensive. Close spacing of draw points with consequent weakening of the layer of solid rock beneath the undercutting level aggravates the tendency of the openings to close. A compromise must be

TABLE 38

DIMENSTONS O F PREPARATORY OPENINGS AT VARIOUS MINES

DATA MOSTI.Y FROM INFORMATION CIRCULAR 6350, u. s. BUREAU OF MINES

Copper 1 "[. -. 1 :;:; 1 "'.mi 1 AT.".. 1 '.'" uaeeTt

Distance apart of haulage drifts, f t . . . . . . . . . . . . . . .

Distance apart of gathering raises, ft.

Distance from haulage level to grizzly level, f t .

Inclination of gathering raises, deg.. . . . . . . . . . . .

Distance apart of grizzly . . . . . drifts or raises, f t . .

Distance apart of draw points along grizzly drifts or in a drift, f t . . . .

Plan area served by each draw point, f t .

Area served by each draw point, sq. f t . . . . . . . . . . . .

Distance of undercut level above grizzly level, f t . . .

Distance apart of under- c i ~ t tirifts or raises, f t .

Width of pnels , f t . . . . . . . Length of ~ane l s , f t . . . . . . .

Width of blocks, f t . . Length of blocks, f t . .

. . . . . Height of blocks, f t . .

56 28

50

75

28

21 & 14 14 X

18.66

261.3

20 18.6 8

14 112 Up to

GOO . . . . . . . . . . . . . . . . . . . .

150 to 270

90 25

45 to 60

50

25

16 12 .5X

16

200

18 11 to

14 180

Up to 1000 180

Varies

125 to 225

60 25

40

55+

25

25 . . . . .

313.5

. . . . . 12.6

200 Up to

600 200 160 to

200 100 to

300

150 50

100

55

50

25 l 2 . 5 X

12.5

156.35

30 25

. . . . .

. . . . .

. . . . . 150 150

380

109.3 20.5 & 30.7 39.4

54

54.6

27.3 1 3 . 7 X

13.7

186

32.8 27.3

164 656

. . . . . . . . . .

. . . . . . . . . .

330

87.5 25

50

50

12.5

13.5 l 8 . 5 X

12.5

156.25

10 12.5

87.5 Up to

500

40 to 200

-

100 40

50 to 100

60

40

'20 20 X

20

400

16 20

100 Up to f200 100 125

60 to 350

UNDERGROUND BlINING 399

UNDERGROUND AT MIAMI Showing squeezing of heaving timbers in portions of the mine where extraction

of the ore has been finished.

reached between close spacing, which increases the recovery of ore and decreases dilution; and wider spacing which, in addition to effecting economy in the cost of maintaining the openings, also cuts down the original cost of preparatory work. The same reasoning applies to the determination of the vertical distance of the haulage level below the draw level. The greater this distance, the greater the preparatory cost; but against this is an expected saving in the subsequent cost of maintaining the haulageways.

A comprehensive paper by E. D. Gardner, entitled "Undercut Block-Caving Method of Mining in Western Copper Mines," has been printed as Injormation Circular 6350 of the U. S. Bureau of Mines. Table 38 is a reproduction with some minor changes of a table in Mr. Gardner's paper.

One subclassification of method is based on the distinction between mining long rectangular blocks or panels or blocks that are roughly square in horizontal projection. An excellent example of the panel method is the practice a t the Ruth mine of Nevada Consolidated. The plan of operation was worked out by W. S. Larsh, mine superintendent, under C. B. Lakenan, the


general manager, in 1915. Only minor changes have been made in the course of 15 years; and in this respect the method seems to have a just title to the distinction of being unique.

BRANCH RAISES AT RUTH

The Ruth scheme for attacking an ideal block of ore, the bottom of which is assumed to be a horizontal plane, is illustrated in Fig. 30. The haulageway at the bottom is one of a number of parallel openings driven a t intervals of 87)4 ft., the track level being 60 ft. below the bottom of the orebody. In actual practice the distance between the haulage level and the "point of draw" ranges between 40 and 80 ft.; but 60 ft. is standard. Calculations based on the estimated cost of drifting and raising on an angle of 50 deg. with the horizontal showed that, with respect to the cost of driving the'necessary working openings, the most economical distance between the haulage drift and the bottom of the mining zone was 40 ft. I t was decided, however, that the horizontal pillar intervening between the top of the drift and the draw level, if 40 ft. were selected as the standard height, would be so thin as to necessitate excessive expense for maintenance. Calculations for 50, 60, 70, and 80 ft. showed that 60 f t . required less expenditure for making the necessary openings than any of the others, and it was accordingly adopted as the standard for all operations.

It should be explained that the angle or slope of the branch raises, 50 deg., is the first factor determined; it is fixed, and the distance between the main motor-haulage drifts and the vertical distance between the drifts and the draw-level are adjusted to meet this requirement. Jn a chute that is too steep, soft ore has a tendency to pack and cement itself in huge masses that must be broken by blasting or by expensive work with bars; or it may arch in the chute. On the other hand, if the slope is not great enough the ore "hangs up " and must be helped along continually, this tendency being exaggerated with soft ore such as is handled in the Ruth mine. The most advantageous angle hzts been proved to be 50 deg.; and so important is this question of slope that no deviation is permitted.


The ideal measurements, accordingly, are 8736-ft. intervals between haulage drifts and the 60-ft. distance between haulage and draw-levels, thereby permitting the branch raises to slope at 50 deg., as shown in Fig. 30. A "raise-series," as the set of

I J Sect ion 8-8

FIG. 80.-Caving method at Nevada's Ruth mine, shown diagrammatically. The ore is mined in "panels" as dis-

tinguished from " blocks."

branches shown in section AA is called, is started at each 25-ft. interval along the direction of the main haulage drifts. A special timber structure, termed a pony set, provides a chute and an arc-gate on each side of the drift. The branches are raised and timbered with 4- by 18-in. or 6- by 12-in. cribbing, the selection depending upon the character of the ground. Each raise is divided into two compartments: a chute for ore, 3 by 4 ft., and a manway, 2 by 3 f t . At the top of each branch, immediately over the chute compartment, is a vertical square-set framed of 10- by Win. or 12- by 12-in. timbers. From this set are built two cribbed finger raises, shown diagrammatically in Fig. 30. These


fingers, according to the early practice, consisted of four or five sets of cribbing, but this was found to be too many. Three sets a t a somewhat lower pitch afford the necessary spread and make work more convenient for the miner who is drawing ore.

I t is apparent that the centers of adjacent square-sets are 12; .;I ft. apart in the plane of each raise-series. Since the finger raises diverge in planes a t right angles to this, the result is a checkerboard arrangement of draw-points, formed by the finger raises, spaced 1255 ft. in both directions. The imaginary plane, which is the locus of the draw-points, is known as the draw- level; ideally it coincicles with the bottom of the ore deposit.

Three or four raise-series are completed before drawing can be started in a particular block; the succeeding series are "brought in" consecutively, one a t a time, the rate depending upon the requirements for ore. It is not desirable that the developnlent by raises should proceed far in advance of actual mining because of the cost of maintaining the raises. As mining operations advance additional ore is undercut by miners working in the finger raises. The proportion of the total area that can be blasted for the undercut depends upon the local nature of the ground. The attempt is made to connect all the finger raises in each row of the series and to run crosscuts between alternate fingers of the two rows, as shown in Fig. 30, section BB. This leaves a row of narrow pillars to be drilled and blasted as the final step in completing the undercut over a given raise-series.

I t is always best to undercut completely in order to induce uniform caving. Unless a new block is being started there is the shattered ore overlying the previous raise-series to which crosscuts inajr be run and to which the remaining pillars can break. In some parts of the mine it is necessary to stope 15 or 20 f t . above the draw-level in order to start caving; on the other hand. the ground may be so greatly shattered that it is impossible to con~plete the undercut, and a dozen long holes. drilled from the fingers and fired sirnultaneonsly, start the caving.

Occasionally blocks covering an area of several hundred square feet settle without breaking. These obviously must be broken up before the drawing can proceed properly. Frequently a small stope started within the huge piece has served as the


best means to break it up. But caving the ore is not the difficult problem a t Ruth. Moreover, it is characteristic of the cap rock that it also caves readily and follows the ore uniformly. This is one of the essential characteristics of a deposit that can be mined to the best advantage by caving.

Good results depend entirely on the care with which the ore is drawn from the different finger raises. Inattention to this feature has been the cause of unsatisfactory results in nearly every instance where branch-raise mining has been tried with indifferent success. Two things are sought: first, to get all the ore; second, to leave all the cap rock, which, of course, tends to mix with the ore as it follows it down. The only way by which an idea of the condition of the caved mass of ore and waste can be obtained is by keeping a detailed record of the ore taken from the stope a t each draw-point, and by checking this against the 6 6 expectancy" from that point.

The expectancy, i.e., the tonnage and copper content of ore that should be recovered from a given finger, is calculated from data obtained in churn-drilling, close approximation being possible because of the relative absence of wide or sudden fluctuations in the tenor of the ore. Sundry charts are made to record the estimates thus obtained and finally the information is displayed graphically on a simple but ingenious model. This consists of a group of perpendicular steel rods (about 3 f t . high and of 34-in. diameter) fixed in a large wooden base, the surface of which is laid out to scale to represent the block of ore being mined. Each rod is sifuated a t a draw-point. On the rod is an adjustable collar that can be slipped up or down and fastened in any desired position by means of a small setscrew. At the start the expectancy data are used to determine the height of the individual rods, the tops of which represent the position of the plane of separation between the cap rock and the orebody.

A daily report is made showing the quantity of ore drawn from each raise by the motormen, together with an estimate by the draw bosses of the amount taken from each finger. The stope engineers adjust any discrepancies in the figures and credit the draw-points with proportional parts of the total reported by the motormen. Daily samples are taken of the ore


from each raise-series and, in the event of unusual behavior, from individual fingers. These assays, of which an average of 100 is taken each day, are studied in conjunction with the figures on tonnage, and, after the charts are brought up to date, the information is transferred to the model; that is, the collars are lowered by the amount represented by the ore drawn. There will be irregularities, of course, but the result is a fairly reliable graphic record of the current position of the top of the caved ore. The stope engineers prepare a sheet every evening for the fore- men, showing the quantity that should be drawn from each finger the next day. The system of recording was devised by Mr. Larsh and his assistants, and is credited by him as being the principal factor in the success of operations.

The "angle of draw" is the slope, from the horizontal, of the plane of contact between broken ore and overlying waste in the direction of successive raise-series. (See Fig. 30.) Experience has shown that this angle should be maintained a t approximately 40 deg. for the best results. If the angle be steeper, the difficulty of preventing the admixture of waste rock is increased; whereas if it be less steep, a longer period must elapse from the time a block is started until it is finished. This entails extra expense in keeping the raises open and the timbering in repair. There is the disadvantage of having an excessive quantity of ore broken in advance; it tends to pack tightly and must be loosened again. In the other direction, that is, in the direction of the successive fingers from the same raise, the plane of contact, in general, is kept horizontal.

"Contact plane" suggests a smooth imaginary surface dividing the ore and waste. In practice the maintenance of such a lane is impossible, as can readily be understood; the purpose of regulating the draw is to approach such a condition as nearly as possible. Tongues of ore will reach up into the cap rock and may be lost; on the other hand, channels may form so that cap rock gets down to a particular draw-point long before the expectancy from that point has been realized. Occasionally an arch will form over a finger that cannot be blasted, so that drawing from this point necessarily will be delayed. The remedy is to draw down tlirough adjacent fingers until the arch is


broken. By temporarily stopping or decreasing the quantity taken from certain draw-points and increasing that from others it is possible to control the operation with remarkable success. At Ruth it is preferable to draw slowly from a comparatively large area, rather than to remove narrow blocks by hurried

TABLE 39

EXPECTANCY AND RECOVERY AT RUTH MINE

Expectancy . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . Tons.. 5,227,247

'opper content, per cent. . . . . . . . . . . . . . . . . . . . . . . . . . . . 2.06 Actual recovery

Tons . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5,415,389 Copper content, per cent. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1.88

Excess of tons, per cent . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 3 . 6 Deficit of copper recovered, per cent. . . . . . . . . . . . . . . . . . . 5 . 2

drawing. Rapid drawing does not give the time necessary for the essential breaking of the rock.

An idea of the final result may be obtained from the figures in Table 39, covering a typical portion of the orebody a t Ruth. They indicate that excellent results are being obtained. The admixture of waste i n the ore shipped amounts to 9 per cent; of this, 95 per cent is virtually barren, the cap in the Ruth mine differing in this respect from that at Ray, Inspiration, and Miami, where most of the cap rock contains appreciable amounts of copper. The high net recovery made a t Ruth therefore seems particularly meritorious.

Many interesting variations of the Ruth method might be cited. The most important difference perhaps is that usually the tops of the branch raises are connected by small timbered drifts and crosscuts to form a grizzly level wliich affords easy access and communication. At Ruth the ground caves so readily that the cost of maintaining openings of this kind would be excessive; accordingly access to the draw-points is through the manway compartment made by the vertical partition in the cribbed raises themselves. This ready cavability also obviates any necessity for supplemental weakening of the rock by openings a t the boundaries of the panel. In some mines, as a t Ray and Morenci, where the haulage drifts are closer together, branches from the main raise are not required.


MIAMI'S "HIGH-LIFT" OPERATION

One other appljcation of undercut caving will be described- probably the most econoinical in operation-namely, the method

FIG. 31.-"High-lift" method of caving at Miami show-n in isometric projection. The block is 300 by 150 ft., whereas

the ideal size is 150 by 150 ft., as used more recently.

devised by Miami for the winning of its "new" low-grade orebody, where actual extraction of ore commenced in October,


1925. The deposit as developed by churn-drilling was estimated to contain 108,461,000 tons averaging 0.88 per cent copper. Dimensions are approximately: extreme length, 3500 ft.; extreme width, 2700 ft.; area, 100 acres; average thickness, 335 ft.; thickness of barren capping, 250 to 500 ft. The rock is thoroughly fractured and after being drained and ventilated is accurately classified as "free caving." The method was described by F. W. Maclennan in an outstanding article in the General Volume of A.I.M.E. Transactions for 1930, from which the accompanying isometric drawing (Fig. 31) is adapted. Eight main features may be enumerated as follows:

1. High caving column-300 to 350 f t . 2. Block or individual stope as distinguished from panel

system. 3. Boundary weakening. 4. Partitions between blocks. 5. Wide spacing of haulage drifts-150 f t . 6. High interval between haulage and undercutting levels-

130 f t . 7. Grizzly levels and raises. 8. Long-range scheduling of sequence of attack on the blocks. High column.-Starting with the 30-ft. height used by

McDonald a t Bingham, the tendency always had been toward caving higher columns, for the obvious reason that the expense of making the preparatory workings was thereby spread over an increased tonnage of ore. Up to 250 f t . the economy is very marked; but beyond 300 f t . it is not important. Prior to 1925, the maximum was about 150 ft., so that Miami's jump to 300 ft. or more (virtually the entire thickness of the ore) was a radical departure from accepted practice. It was, however, a prerequisite to the attainment of the estimated cost of 40c. per ton that was necessary if the low-grade ore were to be exploited profitably.

Individual stapes.-The reasons for the decision to mine by the individual stope or square-block method rather than by the narrow-block or panel method, which Miami had used so suc- cessfully in the old orebody, were essentially as follows:

(a) In order to maintain an angle of contact between broken

SUBSIDENCE AT MIAMI Removal of large blocks of ore results in caving to the

surface.

UNDERGROUND MlNING 409

ore and capping averaging 50 deg. in an ore column 300 f t . high, it would be necessary to keep open and in good condition more than 300 f t . of extraction levels along the length of a rectangular panel of, say, 1200 f t . This goal seemed unlikely of consistent achievement. Excess of dilution, forced abandonment of good broken ore, and general difficulty in effective control of drawing seemed likely to ensue if the panel system were adopted.

(b) With the square-block system the area to be caved would be confined to definite limits and would be surrounded on all four sides by solid ground in the case of the original stopes, and by consolidated fill in part or in whole in the case of pillar stopes; instead of being an .indefinite area, one end of which would be recently caved capping, the other end recently undercut ore (both in motion), and the sides solid ground in the case of original panels and consolidated fill in the case of the pillar panels. Undoubtedly the support of the individual stope would be much better; hence, maintenance costs would be lower and interference of repair work with ore drawing would be minimized.

(c) In the individual stope method the ore could be drawn down evenly over the entire area of the stope, approximating a horizontal plane of contact between the broken ore and overlying broken waste; and, while drawing with an inclined plane of contact between ore and waste had-proved satisfactori for low columns or lifts, there seemed to be no question of the superiority of the horizontal plane of contact for high lifts.

(d) In the same area of orebody, the stope method provided a greater number of working places and the work could be more easily systematized, resulting in a greater production.

(e) I n the individual stope method, current production .

could be more conveniently distributed over the area. The ore could be delivered to the various drifts on the haulage level in controlled proportions, thereby avoiding congestion and delays. As originally laid out, the horizontal dimensions of the blocks or stopes were 150 by 300 ft. Though not literally square, they contrasted with panels, say, 150 by 1200 ft. Eight blocks were mined with satisfactory results. However, subsequent experience demonstrated that the weight was likely to be excessive and 150


by 150 ft. accordingly was established as the standard. Fig. 31 shows a rectangular block but the principles are not altered.

Boundarg weakening.-The ore is undercut over an area 150 ft. square. If left to itself i t might arch to the center and stop caving or it might follow slips or planes of weakness and cave beyond its vertical boundaries. I t is essential that the stope should be made to cave to, but not beyond, its vertical boundaries; and this is particularly necessary when pillar stopes are to be caved and drawn later between the original stopes. With low lifts a satisfactory method of isolating a caving area is by nleans of a narrow shrinkage stope around the boundary. Tt was felt that this would be disastrous with a high lift, as the entire block might settle down like a plug or piston and crush the mine workings below. At any rate, the broken ore might arrive a t the chutes in large blocks unsuitable for drawing. The aim is to have the back of the stope hang, subject to strain, so that i t will fracture and cave slowly in small pieces, and a t the same time cave to and not beyond its vertical boundaries.

The means for accomplishing the desired weakening consists of driving boundary caving drifts completely around each stope a t suitable vertical intervals; and putting up raises a t all four corners of the stope to avoid lateral arching. See Fig. 31. These raises also serve a useful purpose as ore passes and ventilation ways during the development period. The boundary caving drifts are 7f5 ft. high and originally were located at 30-ft. vertical intervals. This had the effect of weakening the plane of the stope boundary '25 per cent; or when desired, in hard ground, the height could be inexpensively increased so as to increase the weakening to 50 per cent. Later, higher drifts were put in a t 45-ft. intervals, thus weakening the boundary plane 3354 per cent at lower cost.

Geological maps of each boundary-caving level are prepared, which show the principal " slips," inclusions of capping, oxidized ore, etc. These maps are helpful in the control of ore-drawing operations. For example, they may show that waste rock that appears early a t some of the draw chutes is coming from an inclusion in the lower part of the orebody and not from the overlying capping. In light of this information ore drawing from


GRIZZLY DRIFT AT MlAMI Looking down the drift from the fringe drift. The "cribbed" supports at right and

left are of 12 by 12-in. timber.

the chutes in question is continued, whereas without it they would be sealed. Supplementing these maps, marker blocks are "planted" in all drifts on all levels a t 25-ft. intervals. These blocks are 12-in. wooden cubes and are marked with a copper tag countersunk in the block for protection. This tag notes the elevation of the drift and the number of the chute above which the marker was planted. The block is large enough to be held on the grizzly and when i t arrives gives accurate information regarding the position from which the accompanying ore has come. Also, i t indicates that the stope is caving to its boundary.

Pa~titions between b1oclcs.-Naturally, an important question was the degree of success that would be achieved in drawing the pillar stopes. Experience gained in previous caving operations, using the panel system and retreating across the entire orebody, indicated that it was advisable to leave a thin partition of ore '

412 T-HE PORPHYRY COPPERS

between the original stopes and the pillar stopes; or, in other words, to leave a greater distance between the border draw- point of the pillar stope and that of the original stope adjoining.

FIG. 39.-Vertical projection of transfer and branch raises and chute sets at Miami. Note that the undercutting level

is 130 ft. above the haulage level.

Acting on this experience, a partition of 15 ft. was left between the sides of all stopes. No partition was left between the ends of the two southern rows of stopes; but end partitions were left between the three rows of northern stopes, the foregoing having


reference to the blocks to be mined from the 720-ft. haulage level. The original partitions are 15 ft. thick, and some of the later ones 745 ft.; the tonnage in the partitions amounts to 11.0 per cent of the total. This tonnage is excluded in figuring development and mining costs, but drawing records indicate that a part, a t least, of this tonnage is recovered.

In drawing the pillar stopes, no serious trouble has been experienced from waste leaking from the adjoining mined-out original stopes. The experience has been that the drawing spreads very little and that caving tends to proceed vertically upward once the ore has been broken. This characteristic requires the close spacing of draw points, but a t the same time is an insurance against lateral dilution. The waste capping which has settled down into the original stopes becomes well packed and consolidated, owing to the pressure exerted by a vertical column of 300 or 400 f t . of material; and this waste filling constitutes substantial lateral support to the partition in mining the pillar stopes.

Spacing of drifts; height of grizzly level.-The horizontal spacing of the, haulage drifts and the vertical distance between the haulage and grizzly levels are interdependent, presuming that the angle of incline of the raises is fixed. At Miami the standard slope is 55 deg. 20 min. With an interval of 150 f t . between haulageways the vertical distance to the grizzly level is 100 ft., and to the undercutting level 130 ft., as shown in Fig. 32. The combination of long and few raises and few haulage drifts has certain disadvantages. It tends to increase congestion of traffic on the haulage level; i t complicates the detailed control of drawing because more draw points feed a given chute; and it entails greater difficulty in keeping the ore moving in the chutes. Against these may be noted: larger storage capacity for ore in the raises; and faster and more economical completion of preparatory work because raising usually can be done more quickly and easily than drifting and constructing chutes.

Grizzly levels.-The tops of the transfer raises are a t the floor of the grizzly level, which comprises a series of drifts and crosscuts through which explosives, timber, and equipment are transported. On it most of the secondary blasting is done. Over


the top of each transfer raise is the grizzly through which the ore must pass, and which consists of 45-lb. rails spaced 12 in. apart, set on 10- by 10-in. timber stringers. From both sides of the grizzly at right angles to the drift are driven grizzly raises inclined 42 deg. for a distance of 14 ft., and thereafter raised vertically for 10 ft. In the vertical portion is built the chute set from which the finger raises start. The stoutly built set is oriented 45 deg. instead of being "square" with the other workings. The result is that the four finger raises reach the undercut level a t draw-points so spaced that their centers form a perfect pattern of 1246-ft. squares. Undercutting, aside from the fringe drifts and stopes, consists of large drifts 8 ft. wide, from which the intervening pillars are drilled and finally blasted, the work progressing diagonally across the block. Once started, undercutting is completed as rapidly as possible and soon after the block or stope is ready for drawing. At some Porphyry mines the finger raises start directly from the grizzly level; and a t Ruth, as already explained, there is no grizzly level and the fingers start from the timber-set a t the top of the branch raises.

Scheduling operations.-One unusual and ingenious plan is the prearranged schedule of the exact sequence of blocks or individual stopes to be mined. Each of 64 blocks to be mined from the 720-ft. level is given a number to s h o i the order in which it will be extracted. No. 13, so far as one can see from the drawings, is omitted-doubtless for good and sufficient reasons. Casual inspection shows no apparent system; as a matter of fact, the object is as to avoid caving any block until the capping in any adjoining mined block has had a minimum of several months in which to settle and consolidate. As it is important that the work of development be completed on time, but not much ahead of time, and as 12 or 13 stopes should be producing simultaneously to yield a daily output of 18,000 tons, extremely close scheduling of the preparatory work is provided. And the schedule is followed.

JT7ithout going into detail, it may be said that the ore drawing is supervised by experienced stope engineers who issue written 01-ders to the draw bosses and chute sealers each day. The job involves rnucli more than drawing a given amount from a given


group of chutes in rotation. Departure from the routine order of drawing is brought about by various causes, such as the appearance of capping in a chute; the necessity for repairs; the appearance of excess weight on the timbers, requiring the chutes to be drawn whether normally scheduled to be drawn or not;

TABLE 40

EXPECTANCY AND REALIZATION AT MIAMI

FIRST 13 STOPES IN LOW-GRADE OPERATIONS; PARTITIONS NOT INCLUDED IN EXPECTANCY

the requirements as to grade of ore; and the necessity for maintaining proper distribution to the haulage level for efficient operation of the trains. Each stope engineer inspects each day all chutes listed on the draw-sheet and receives reports from draw bosses and sealers as to the result of their inspections. He notes the condition of the ore in all the chutes; removes from the draw-list chutes in which capping shows; lists chutes that need drilling or blasting and chutes or chute-sets or grizzly-level timbers that require repair; and issues necessary orders.

Any chute in which capping appears prematurely is sealed for several weeks before being replaced on the draw-list. In the meantime the chutes surrounding it may be sealed and later drawn. This procedure tends to break up the wall of the pipe, after which the chute showing capping usually may be drawn again with satisfactory results.

- - -

Expectanq Recovery I *

Total of 13 completed stopes. . . . . . . . . . . . . . . .

Best original stope (8). . . Best pillar stope (11). . . . Poorest original stope (7). Poorest pillar stope (9). . .

Tons

Ratio, recovery to

expectancy

Ton- nage, Per cent

11,038,070 998,016 319,560

1,071,535 1,098,313

1 .08 1 .03 1 .06 0 .87 1 . 1 0

Assay copper,

Per cent

Assag coppeT'

Per cent

Grade, Per cent

Tons

Total copper.

Per cent


In Table 40 are shown the results attained in the first thirteen stopes that were completed. The excess of total copper recovery over expectancy doubtless is accounted for by drawing partition ore and capping lean in copper. When the entire orebody is worked out the picture may be different.

It is apparent that the accuracy of the percentage recovery as calculated hinges on the accuracy of the original estimates of the quantity and grade of the ore, based on churn-drilling or other data. Usually the tonnage drawn will exceed the expectancy as a consequence of dilution; but experience is that the average grade of the ore extracted generally falls more than enough below the estimated figure to bffset the surplus tonnage. At Inspiration in recent years the tonnage drawn is more than 110 per cent of expectancy; but if 85 per cent of the copper calculated as contained in the ore is recovered, the performance is considered satisfactory. This would indicate that the average assay of the ore as mined is about 77 per cent of calculated grade of the ore in place. As already mentioned, a t Ruth 95 per cent of the copper is won in a tonnage that exceeds expectancy by 3.6 per cent.

MISCELLANEOUS IMPROVEMENTS IN PRACTICE

Other phases of operations connected with underground Porphyry mining are eclipsed in interest by the actual process of extraction as just described. Nevertheless, important contribu- tions to the final result, as reflected in low costs, have been made by progress in the following:

1. Underground drilling and blasting practice. 2. Design of timbered drift sets, chute sets, grizzly sets,

raise linings, etc. 3. Underground transportation systems including tracks,

cars, and locomotives. 4. Hoisting practice including equipment and design of

surface plant, shafts, underground ore pockets, and loading stations.

Drilling and b1astin.g.-Though the amount of drilling and blasting required for mining by caving is relatively small, the operation is by no means unimportant. The average powder


With the possible exception of some down-hole drilling, the use of " wet " machines has become almost universal.

The development of the hammer drill of superior design and one-man size has been made possible in large measure by the progress in steel metallurgy, particularly in alloying and heat treatment. Stronger and more durable materials have enabled the manufacturers to produce drilling machines that withstand severe service, and that at the same time are much lighter. Particularly for drilling vertical or steeply inclined holes, either up or down, machines weighing 50 to 60 lb. are highly efficient. "Uppers" are drilled with stopers; they are self- rotated and are automatically and mechanically "fed" as drilling advances.

Drill steel has been improved in quality; and mine operators have learned the wisdom of scientific control of sharpening and tempering. The oil or electric furnace, the mechanical sharpener. and the pyrometer have displaced the old coal forge and the blacksmith who depended on sweat and intuition to produce good steel. The result of these and many other improvements is faster, surer drilling and a reduction of perhaps 50 per cent in the power and labor cost per foot of hole drilled.

Better dynamite is manufactured. Formerly dynamite froze a t temperatures above the freezing point of water. Recently, by the use of nitroglycol in conjunction with the nitroglycerin, explosives have been made that withstand temperatures sub- stantially below zero, Fahrenheit. Dynamite does not function efficiently when frozen; and "thawing" dynamite was a slow and expensive operation. In many of the mines electric exploders have displaced fuse and caps as the means of detonating the charge, a t least for part of the work. They are safer and surer than the usual fuse and cap but are more troublesome to use. At Morenci all blasting is by electricity. Such matters as the number, direction, and depth of drillholes and the quantity of dynamite required have been given thorough and intelligent study, with consequent elimination of guesswork. The most important saving arising from all these improvements is in the labor cost per unit of accomplishment. The main purpose is to increase the effectiveness of the worker, as labor constitutes


from 50 to 70 per cent of the operating cost in mines of the general character of the underground Porphyry properties.

Timbering.-As already has been pointed out, an important item of expense is the cost of keeping the extraction workings

z ' ~ " FIG. 33.-Drift set, pony set, chute lips, and cribbed raise at Inspiration. These sets are carefully designed and stand-

ardized.

open in the face of the huge and shifting forces created by the undercutting and caving of a mass of, say, 500,000 tons of ore; and in some mines, like Ray, by the tendency of the ore to swell. Aside from the cost directly involved in so-called '-6 timber repairs," the obstruction and delay to drawing and


haulage occasioned by such repair work usually is decidedly expensive. For this reason much effort has been expended on design of the various standard timber structures or sets. Fig. 33 shows a drift set a t Inspiration and the timbering of a branch raise. The lining is a cribbing of 6-in. timbers, 10 in. wide, laid skin-to-skin. This gives a plane surface that does not obstruct

FIG. 34.-Grizzly station at Morenci. Note the multiplicity of heavy timbers required to

"hold" the ground.

the flow of the ore; it allows for a substantial amount of flexibility and "give"; and a t the same time is exceptionally strong. The throat set and grizzly shown in Fig. 34 illustrate the need for a multiplicity of heavy timbers to withstand the strain a t critical points.

Transportation.-One of the things tha,t frequently has surprised a casual visitor to one of the Porphyry mines is the transportation system, of which the terminal is the underground station a t the main hoisting shaft. If he stopped to think that 20,000 tons of ore would fill 400 or 500 ordinary cars, such as he sees on the railroads, he would realize that the haulage of this quantity of ore in one day underground presents a real traffic problem. He would not be surprised to see a coilstant succession of 20-car trains, each car carrying 5 tons of ore, pulled through a rotary tipple by 8-ton two-motored electric trolley locomotives, in a roomy, electric-lighted, and concrete- lined station. That is what he would see a t Inspiration. For a long time Inspiration stuck to compressed-air locomotives


because of the danger involved in trolley wires. The storage- battery electric locomotive, incidentally, was in its infancy when Inspiration adopted compressed air in 1915. However, the superiority of electricity in ease of control, in flexibility of operation, and in cost, eventually caused its introduction by Inspira- tion in 1926. Except for Ray, where compressed air still is the motive power, all of the mines use electricity. Inspiration's tracks are 30-in. gauge, of 40- and 60-lb. rail; and solid manganese-steel frogs are standard. Trains are operated by an electric block-signal system exactly like those used for trunk-line railroads. Incidentally, the cars are equipped with tapered roller bearings, as they may well be, when it is considered that 4000 or 4500 loads frequently have been hauled in one day.

Though haulage, including loading, dumping, and "maintenance of way," usually includes only from 15 to 20 per cent of

SHAFT STATION ON MIAMI 720-FT. LEVEL

Men and timber are lowered and hoisted on the cage. The skip compartments for ore are behind the screen at the left.


MIAMI'S REMODELED HOIST Installation of the second motor permitted higher speed and the establishment of a

world's hoisting record through a single two-skip shaft.

the tota.1 mining cost, i t is a vital operation for the reason that, with hoisting, i t forms for the mine the neck of the bottle, so to speak. To credit the Porphyry mines with a major share in the developlnent of efficient underground haulage is entirely logical, for until this group of mines came along daily outputs of 20,000 tons or even 10,000 tons were unknown. Even today the biggest gold mine in the world, Government Areas on the Rand, mines only 7000 tons per day. The St. Joseph Lead Co. in recent years has produced a t the rate of 17,500 tons per day, but that is from a scattered group of mines operated through 10 different shafts. Taken all together, Anaconda's score of copper mines a t Butte never have produced at the rate of 20,000 tons per day.

Hoisting.-Large-scale hoisting is another field in which the Porphyry companies were pioneers. Probably the record for


sustained hoisting through a single shaft was established a t Miami in 1930 with 6,184,993 tons or an average of 17,141 tons for 35755 operating days. In August, 1930, Miami mined 17,674 tons per day for 31 consecutive days. All of this ore was raised with a single electrically operated hoist with two 10-ton

FIG. 35.-General arrangement of underground ore pocket and skip-loading equipment at Miami. Traffic over the ore pocket is one

way only.

E L E V A T I O N P L A N

skips in balance. Fig. 35 shows the layout a t the underground station where the ore was loaded. Cars are dumped as the trains move over the 800-ton ore pocket at 8 miles per hour. As indicated in the sketch, traffic is all one way, the empty trains return- ing to the mining area by a separate route. Incidentally, the tracks, of which there are 556 miles on the 720-ft. level, are


ballasted with crushed and screened diabase to a depth of 3 in. beneath the cross-ties. During 1928, the original hoisting engine was remodeled and speeded up from 1.500 to 3350 ft. per minute. The installation of a 1400-hp. motor duplicating the one already in service and operating through a new pinion on the single main shaft made a novel unit that proved thoroughly satisfactory despite the dire predictions of several mechanical engineers.

At Inspiration, 21,200 tons has been hoisted in a single day through the main twin-shafts. These shafts, though 103 ft. apart, are served by a single hoist house, ore bin, and headframe structure. Underground they have a tipple and loading station in common; but each shaft is provided with two skip compartments, so that Miami rightly holds the record for hoisting through a single shaft. The Inspiration shafts are more than 800 f t . deep and are lined with reinforced concrete for their entire length. Each shaft has two skip compartments, the steel skips, each of 12-ton capacity, operating in balance.

The entire hoisting installation can be set to operate automatically, entirely independent of the engineer in the hoist house a t the surface; or the skips can be controlled manually by the attendant a t the underground loading station. This last arrangement is the most rapid and permits an hourly hoist of 105 skips, or at the rate of 30,000 tons per 24-hour day. As an elevating performance, that would be equivalent to lifting about 400,000 people daily to the top of the Empire State building in New York. Twenty years ago steam was the power used and magnifi- cent installations were to be found. But the electrical engineer so improved his motors and his apparatus for control that in respect to speed and capacity, not to mention simplicity,safety, and economy, the steam hoist was put out of the running.

In 1992, 1923, and 1934, Inspiration spent $3,500,000 to equip the Live Oak section of its property with a new shaft. Though having only two skip compartments with a combined rated capacity of 10,000 tons per day, it is probably fair to say that it exemplifies the best of modern design for a large-capacity hoisting installation. For example, the ore coming from the stopes is dumped by tipple into two concrete cylindrical ore


bins having a capacity of 800 tons each. The bottom of these bins is shown in Fig. 36. Before reaching the skips the ore is automatically weighed in this fashion: Manganese-steel pan

/FIG. 36.-Skip-loading equipment at Inspiration's Porphyry, or Live Oak, shaft. Note the heavy concrete construction and the roominess of this under-

ground station. The ore bins are just above.

feeders operating beneath the 800-ton bins discharge ore into a loading cartridge or hopper, which is balanced on knife edges and is counterweighted. When 19 tons, the effective capacity of the skip, has entered the hopper, it tips about x6 in. The move- ment, mechanically magnified, actuates an electric switch which stops the feeders. As the empty skip descends on its return from the surface it strikes a trigger which releases the gate a t the bottom of the hopper. The 19-ton "charge" of ore enters the skip, and simultaneously removes the pressure on the gate of the hopper. Through a counterweight the gate is automatically


closed and latched; and the latching automatically restarts the motor of the feeder so as to reload the weighing hopper. The operator pulls a switch the instant the skip is loaded and the round trip to the surface and return is made without further

TABLE 41 COMPARATIVE DATA ON UNDERGROUND MINING

PRODUCTION AND "P.~RTIAL" COSTS FOR BLOCK CAVING AT FOUR PORPEYRY MINES

Year . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

. . . . . . . . . . . . . . . . . . . Production, tons ore.

Ore mined per man underground. tons.. . . . I 11.84 1 21.68 1 19.13 13.1

Explosives per ton ore, Ib.. . . . . . . . . . . . . . . 0.283 0.319 0.227b

1.045 ~ 1.88 7 . . . . . . . . . . . . . . . . Timber per ton ore, bd. ft 0. 25

I I I I

a Cost figures should not be compared closely because of differences in accounting methods. b Conjectllral.

"Partial" costs per ton orea . . . . . . . . . . . . . . . . . . . . . . . Development..

. . . . . . . . . . . . . . . . . . . . . . . . . . . . Mining.. Haulage and hoisting.. . . . . . . . . . . . . . . . .

attention from anyone. Many other details of design might be mentioned, but this is typical.

Needless to say, the station, as well as the entire 1400-ft. shaft, is lined with reinforced concrete. The accompanying

$0.142 $0.1000 0.129 1 0.1557 0.160b 0.0767

sketches, besides showing some of the details of construction, give a general impression of the roominess of the station and the large quantity of concrete required. The distribution of the total of 17,825 cu. yd. of concrete used in the installation is as follows: shaft, 6441; stations and pockets, 5379; surface structures, 3825; drifts, 2180.

. . . . . . . . . . . . . . . . . . . . . . . . . . . . Sundry.. 0.085 0.0302

$0.1036 0.243% 0.1106

$0.211 0.208 0.144

0.0573 O.OPO


The hauling and hoisting of the ore from any one of these undergrountl Porphyry mines is an undertaking of no small importance. Even if it is not accomplished a t a direct cost of less than 10c. per ton, as it was a t Miami over a four-year period, it reflects thorough organization and careful attention to detail in both the maintenance and the operation of equipment.

Such comparative data on costs as are available are not particularly significant because of the variation in the methods and "reservations" in accounting. However, in Table 41 are given data obtained chiefly from Bureau of Mines Information Circular No . 6350, to which reference already has been made. A somewhat more illuminating exhibit is Fig. 14, giving the data of the principal elements in cost a t Miami from 1911 to 1931, inclusive. The most striking feature, of course, is the low costs attained in the last seven years during the era of "low-grade" operations.

The achievements as recorded in dollars, or rather in cents, are amazing. But apart from this one cannot but feel fascinated and inspired a t the physical aspect of the thing: the mere act of compelling one of those million-ton blocks or panels of ore to behave the way you want it to. Do not think for a minute that everything goes smoothly in this work. The men who are doing i t have their troubles-plenty of them. But they are the type of engineer that thrives on trouble or they would not be where they are. As remarked before, if there be any art in a job so prosaic as mine operation, i t is in mining copper by undercut block-caving.

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XIX w UNDERGROUND MINING - library.aimehq.orglibrary.aimehq.org/library/books/The Porphyry Coppers, 1933/393.pdf · Controlled intermittent flowage of the ore by gravity ... Miami