than inclines: the length of haul from a given depth is less; much
higher rope speed is possible, and thus the haulage hours are less
for the same output; the wear and tear on ropes, tracks, or guides
is not so great, and pumping is more economical where the Cornish
order of pump is used. On the other hand, with a vertical shaft
must be included the cost of operating crosscuts. On mines where
the volume of ore does not warrant mechanical haulage, the cost of
tramming through the extra distance involved is an expense which
outweighs any extra operating outlay in the inclined shaft itself.
Even with mechanical haulage in crosscuts, it is doubtful if there
is anything in favor of the vertical shaft on this score.
[Illustration: Fig. 6.--Cross-section showing auxiliary vertical
outlet.]
In deposits of very flat dips, under 30°, the case arises where the
length of incline is so great that the saving on haulage through
direct lift warrants a vertical shaft as an auxiliary outlet in
addition to the incline (Fig. 6). In such a combination the crosscut
question is eliminated. The mine is worked above and below the
intersection by incline, and the vertical shaft becomes simply a
more economical exit and an alternative to secure increased output.
The North Star mine at Grass Valley is an illustration in point. Such
a positive instance borders again on Case IV, deep-level projects.
In conclusion, it is the writer's belief that where mines are to
be worked from near the surface, coincidentally with sinking, and
where, therefore, crosscuts from a vertical shaft would need to be
installed frequently, inclines are warranted in all dips under 75°
and over 30°. Beyond 75° the best alternative is often undeterminable.
In the range under 30° and over 15°, although inclines are primarily
necessary for actual delivery of ore from levels, they can often
be justifiably supplemented by a vertical shaft as a relief to a
long haul. In dips of less than 15°, as in those over 75°, the
advantages again trend strongly in favor of the vertical shaft. There
arise, however, in mountainous countries, topographic conditions
such as the dip of deposits into the mountain, which preclude any
alternative on an incline at any angled dip.
CASE IV. INCLINED DEPOSITS WHICH MUST BE ATTACKED IN DEPTH (Fig.
7).--There are two principal conditions in which such properties
exist: first, mines being operated, or having been previously worked,
whose method of entry must be revised; second, those whose ore-bodies
to be attacked do not outcrop within the property.
The first situation may occur in mines of inadequate shaft capacity
or wrong location; in mines abandoned and resurrected; in mines
where a vertical shaft has reached its limit of useful extensions,
having passed the place of economical crosscutting; or in mines in
flat deposits with inclines whose haul has become too long to be
economical. Three alternatives present themselves in such cases: a
new incline from the surface (_A B F_, Fig. 7), or a vertical shaft
combined with incline extension (_C D F_), or a simple vertical
shaft (_H G_). A comparison can be first made between the simple
incline and the combined shaft. The construction of an incline from
the surface to the ore-body will be more costly than a combined
shaft, for until the horizon of the ore is reached (at _D_) no
crosscuts are required in the vertical section, while the incline
must be of greater length to reach the same horizon. The case arises,
however, where inclines can be sunk through old stopes, and thus
more cheaply constructed than vertical shafts through solid rock;
and also the case of mountainous topographic conditions mentioned
above.
[Illustration: Fig. 7.--Cross-section of inclined deposit which
must be attacked in depth.]
From an operating point of view, the bend in combined shafts (at
_D_) gives rise to a good deal of wear and tear on ropes and gear.
The possible speed of winding through a combined shaft is, however,
greater than a simple incline, for although haulage speed through
the incline section (_D F_) and around the bend of the combined
shaft is about the same as throughout a simple incline (_A F_), the
speed can be accelerated in the vertical portion (_D C_) above that
feasible did the incline extend to the surface. There is therefore an
advantage in this regard in the combined shaft. The net advantages
of the combined over the inclined shaft depend on the comparative
length of the two alternative routes from the intersection (_D_)
to the surface. Certainly it is not advisable to sink a combined
shaft to cut a deposit at 300 feet in depth if a simple incline
can be had to the surface. On the other hand, a combined shaft
cutting the deposit at 1,000 feet will be more advisable than a
simple incline 2,000 feet long to reach the same point. The matter
is one for direct calculation in each special case. In general, there
are few instances of really deep-level projects where a complete
incline from the surface is warranted.
In most situations of this sort, and in all of the second type
(where the outcrop is outside the property), actual choice usually
lies between combined shafts (_C D F_) and entire vertical shafts (_H
G_). The difference between a combined shaft and a direct vertical
shaft can be reduced to a comparison of the combined shaft below
the point of intersection (_D_) with that portion of a vertical
shaft which would cover the same horizon. The question then becomes
identical with that of inclined _versus_ verticals, as stated in Case
III, with the offsetting disadvantage of the bend in the combined
shaft. If it is desired to reach production at the earliest date,
the lower section of a simple vertical shaft must have crosscuts
to reach the ore lying above the horizon of its intersection (_E_).
If production does not press, the ore above the intersection (_EB_)
can be worked by rises from the horizon of intersection (_E_).
In the use of rises, however, there follow the difficulties of
ventilation and lowering the ore down to the shaft, which brings
expenses to much the same thing as operating through crosscuts.
The advantages of combined over simple vertical shafts are earlier
production, saving of either rises or crosscuts, and the ultimate
utility of the shaft to any depth. The disadvantages are the cost
of the extra length of the inclined section, slower winding, and
greater wear and tear within the inclined section and especially
around the bend. All these factors are of variable import, depending
upon the dip. On very steep dips,--over 70°,--the net result is in
favor of the simple vertical shaft. On other dips it is in favor
of the combined shaft.
CASES V AND VI. MINES TO BE WORKED TO GREAT DEPTHS,--OVER 3,000
FEET.--In Case V, with vertical or horizontal deposits, there is
obviously no desirable alternative to vertical shafts.
In Case VI, with inclined deposits, there are the alternatives
of a combined or of a simple vertical shaft. A vertical shaft in
locations (_H_, Fig. 7) such as would not necessitate extension in
depth by an incline, would, as in Case IV, compel either crosscuts
to the ore or inclines up from the horizon of intersection (_E_).
Apart from delay in coming to production and the consequent loss of
interest on capital, the ventilation problems with this arrangement
would be appalling. Moreover, the combined shaft, entering the deposit
near its shallowest point, offers the possibility of a separate
haulage system on the inclined and on the vertical sections, and
such separate haulage is usually advisable at great depths. In
such instances, the output to be handled is large, for no mine of
small output is likely to be contemplated at such depth. Several
moderate-sized inclines from the horizon of intersection have been
suggested (_EF_, _DG_, _CH_, Fig. 8) to feed a large primary shaft
(_AB_), which thus becomes the trunk road. This program would cheapen
lateral haulage underground, as mechanical traction can be used
in the main level, (_EC_), and horizontal haulage costs can be
reduced on the lower levels. Moreover, separate winding engines
on the two sections increase the capacity, for the effect is that
of two trains instead of one running on a single track.
SHAFT LOCATION.--Although the prime purpose in locating a shaft
is obviously to gain access to the largest volume of ore within
the shortest haulage distance, other conditions also enter, such
as the character of the surface and the rock to be intersected,
the time involved before reaching production, and capital cost.
As shafts must bear two relations to a deposit,--one as to the
dip and the other as to the strike,--they may be considered from
these aspects. Vertical shafts must be on the hanging-wall side
of the outcrop if the deposit dips at all. In any event, the shaft
should be far enough away to be out of the reach of creeps. An
inclined shaft may be sunk either on the vein, in which case a
pillar of ore must be left to support the shaft; or, instead, it
may be sunk a short distance in the footwall, and where necessary
the excavation above can be supported by filling. Following the
ore has the advantage of prospecting in sinking, and in many cases
the softness of the ground in the region of the vein warrants this
procedure. It has, however, the disadvantage that a pillar of ore
is locked up until the shaft is ready for abandonment. Moreover, as
veins or lodes are seldom of even dip, an inclined shaft, to have
value as a prospecting opening, or to take advantage of breaking
possibilities in the lode, will usually be crooked, and an incline
irregular in detail adds greatly to the cost of winding and maintenance.
These twin disadvantages usually warrant a straight incline in the
footwall. Inclines are not necessarily of the same dip throughout,
but for reasonably economical haulage change of angle must take
place gradually.
[Illustration: Fig. 8.--Longitudinal section showing shaft arrangement
proposed for very deep inclined deposits.]
In the case of deep-level projects on inclined deposits, demanding
combined or vertical shafts, the first desideratum is to locate
the vertical section as far from the outcrop as possible, and thus
secure the most ore above the horizon of intersection. This, however,
as stated before, would involve the cost of crosscuts or rises and
would cause delay in production, together with the accumulation
of capital charges. How important the increment of interest on
capital may become during the period of opening the mine may be
demonstrated by a concrete case. For instance, the capital of a
company or the cost of the property is, say, $1,000,000, and where
opening the mine for production requires four years, the aggregate
sum of accumulated compound interest at 5% (and most operators
want more from a mining investment) would be $216,000. Under such
circumstances, if a year or two can be saved in getting to production
by entering the property at a higher horizon, the difference in
accumulated interest will more than repay the infinitesimal extra
cost of winding through a combined shaft of somewhat increased
length in the inclined section.
The unknown character of the ore in depth is always a sound reason
for reaching it as quickly and as cheaply as possible. In result,
such shafts are usually best located when the vertical section
enters the upper portion of the deposit.
The objective in location with regard to the strike of the ore-bodies
is obviously to have an equal length of lateral ore-haul in every
direction from the shaft. It is easier to specify than to achieve
this, for in all speculative deposits ore-shoots are found to pursue
curious vagaries as they go down. Ore-bodies do not reoccur with
the same locus as in the upper levels, and generally the chances
to go wrong are more numerous than those to go right.
NUMBER OF SHAFTS.--The problem of whether the mine is to be opened
by one or by two shafts of course influences location. In metal
mines under Cases II and III (outcrop properties) the ore output
requirements are seldom beyond the capacity of one shaft. Ventilation
and escape-ways are usually easily managed through the old stopes.
Under such circumstances, the conditions warranting a second shaft
are the length of underground haul and isolation of ore-bodies or
veins. Lateral haulage underground is necessarily disintegrated by
the various levels, and usually has to be done by hand. By shortening
this distance of tramming and by consolidation of the material
from all levels at the surface, where mechanical haulage can be
installed, a second shaft is often justified. There is therefore
an economic limitation to the radius of a single shaft, regardless
of the ability of the shaft to handle the total output.
Other questions also often arise which are of equal importance
to haulage costs. Separate ore-shoots or ore-bodies or parallel
deposits necessitate, if worked from one shaft, constant levels
through unpayable ground and extra haul as well, or ore-bodies may
dip away from the original shaft along the strike of the deposit
and a long haulage through dead levels must follow. For instance,
levels and crosscuts cost roughly one-quarter as much per foot as
shafts. Therefore four levels in barren ground, to reach a parallel
vein or isolated ore-body 1,000 feet away, would pay for a shaft
1,000 feet deep. At a depth of 1,000 feet, at least six levels
might be necessary. The tramming of ore by hand through such a
distance would cost about double the amount to hoist it through
a shaft and transport it mechanically to the dressing plant at
surface. The aggregate cost and operation of barren levels therefore
soon pays for a second shaft. If two or more shafts are in question,
they must obviously be set so as to best divide the work.
Under Cases IV, V, and VI,--that is, deep-level projects,--ventilation
and escape become most important considerations. Even where the
volume of ore is within the capacity of a single shaft, another
usually becomes a necessity for these reasons. Their location is
affected not only by the locus of the ore, but, as said, by the time
required to reach it. Where two shafts are to be sunk to inclined
deposits, it is usual to set one so as to intersect the deposit at
a lower point than the other. Production can be started from the
shallower, before the second is entirely ready. The ore above the
horizon of intersection of the deeper shaft is thus accessible from
the shallower shaft, and the difficulty of long rises or crosscuts
from that deepest shaft does not arise.
CHAPTER VIII.
Development of Mines (_Continued_).
SHAPE AND SIZE OF SHAFTS; SPEED OF SINKING; TUNNELS.
SHAPE OF SHAFTS.--Shafts may be round or rectangular.[*] Round
vertical shafts are largely applied to coal-mines, and some engineers
have advocated their usefulness to the mining of the metals under
discussion. Their great advantages lie in their structural strength,
in the large amount of free space for ventilation, and in the fact
that if walled with stone, brick, concrete, or steel, they can be
made water-tight so as to prevent inflow from water-bearing strata,
even when under great pressure. The round walled shafts have a longer
life than timbered shafts. All these advantages pertain much more to
mining coal or iron than metals, for unsound, wet ground is often
the accompaniment of coal-measures, and seldom troubles metal-mines.
Ventilation requirements are also much greater in coal-mines. From
a metal-miner's standpoint, round shafts are comparatively much
more expensive than the rectangular timbered type.[**] For a larger
area must be excavated for the same useful space, and if support
is needed, satisfactory walling, which of necessity must be brick,
stone, concrete, or steel, cannot be cheaply accomplished under
the conditions prevailing in most metal regions. Although such
shafts would have a longer life, the duration of timbered shafts
is sufficient for most metal mines. It follows that, as timber
is the cheapest and all things considered the most advantageous
means of shaft support for the comparatively temporary character
of metal mines, to get the strains applied to the timbers in the
best manner, and to use the minimum amount of it consistent with
security, and to lose the least working space, the shaft must be
constructed on rectangular lines.
[Footnote *: Octagonal shafts were sunk in Mexico in former times.
At each face of the octagon was a whim run by mules, and hauling
leather buckets.]
[Footnote **: The economic situation is rapidly arising in a number
of localities that steel beams can be usefully used instead of
timber. The same arguments apply to this type of support that apply
to timber.]
The variations in timbered shaft design arise from the possible
arrangement of compartments. Many combinations can be imagined,
of which Figures 9, 10, 11, 12, 13, and 14 are examples.
[Illustration: FIG. 9. FIG. 10. FIG. 11. FIG. 12. FIG. 13. FIG.
14.]
The arrangement of compartments shown in Figures 9, 10, 11, and
13 gives the greatest strength. It permits timbering to the best
advantage, and avoids the danger underground involved in crossing
one compartment to reach another. It is therefore generally adopted.
Any other arrangement would obviously be impossible in inclined
or combined shafts.
SIZE OF SHAFTS.--In considering the size of shafts to be installed,
many factors are involved. They are in the main:--
_a_. Amount of ore to be handled.
_b_. Winding plant.
_c_. Vehicle of transport.
_d_. Depth.
_e_. Number of men to be worked underground.
_f_. Amount of water.
_g_. Ventilation.
_h_. Character of the ground.
_i_. Capital outlay.
_j_. Operating expense.
It is not to be assumed that these factors have been stated in
the order of relative importance. More or less emphasis will be
attached to particular factors by different engineers, and under
different circumstances. It is not possible to suggest any arbitrary
standard for calculating their relative weight, and they are so
interdependent as to preclude separate discussion. The usual result
is a compromise between the demands of all.
Certain factors, however, dictate a minimum position, which may
be considered as a datum from which to start consideration.
_First_, a winding engine, in order to work with any economy, must
be balanced, that is, a descending empty skip or cage must assist
in pulling up a loaded one. Therefore, except in mines of very
small output, at least two compartments must be made for hoisting
purposes. Water has to be pumped from most mines, escape-ways are
necessary, together with room for wires and air-pipes, so that at
least one more compartment must be provided for these objects.
We have thus three compartments as a sound minimum for any shaft
where more than trivial output is required.
_Second_, there is a certain minimum size of shaft excavation below
which there is very little economy in actual rock-breaking.[*]
In too confined a space, holes cannot be placed to advantage for
the blast, men cannot get round expeditiously, and spoil cannot be
handled readily. The writer's own experience leads him to believe
that, in so far as rock-breaking is concerned, to sink a shaft
fourteen to sixteen feet long by six to seven feet wide outside
the timbers, is as cheap as to drive any smaller size within the
realm of consideration, and is more rapid. This size of excavation
permits of three compartments, each about four to five feet inside
the timbers.
[Footnote *: Notes on the cost of shafts in various regions which
have been personally collected show a remarkable decrease in the
cost per cubic foot of material excavated with increased size of
shaft. Variations in skill, in economic conditions, and in method
of accounting make data regarding different shafts of doubtful
value, but the following are of interest:--
In Australia, eight shafts between 10 and 11 feet long by 4 to
5 feet wide cost an average of $1.20 per cubic foot of material
excavated. Six shafts 13 to 14 feet long by 4 to 5 feet wide cost
an average of $0.95 per cubic foot; seven shafts 14 to 16 feet
long and 5 to 7 feet wide cost an average of $0.82 per cubic foot.
In South Africa, eleven shafts 18 to 19 feet long by 7 to 8 feet
wide cost an average of $0.82 per cubic foot; five shafts 21 to
25 feet long by 8 feet wide, cost $0.74; and seven shafts 28 feet
by 8 feet cost $0.60 per cubic foot.]
The cost of timber, it is true, is a factor of the size of shaft,
but the labor of timbering does not increase in the same ratio.
In any event, the cost of timber is only about 15% of the actual
shaft cost, even in localities of extremely high prices.
_Third_, three reasons are rapidly making the self-dumping skip
the almost universal shaft-vehicle, instead of the old cage for
cars. First, there is a great economy in labor for loading into
and discharging from a shaft; second, there is more rapid despatch
and discharge and therefore a larger number of possible trips;
third, shaft-haulage is then independent of delays in arrival of
cars at stations, while tramming can be done at any time and
shaft-haulage can be concentrated into certain hours. Cages to
carry mine cars and handle the same load as a skip must either
be big enough to take two cars, which compels a much larger shaft
than is necessary with skips, or they must be double-decked, which
renders loading arrangements underground costly to install and
expensive to work. For all these reasons, cages can be justified only
on metal mines of such small tonnage that time is no consideration
and where the saving of men is not to be effected. In compartments
of the minimum size mentioned above (four to five feet either way)
a skip with a capacity of from two to five tons can be installed,
although from two to three tons is the present rule. Lighter loads
than this involve more trips, and thus less hourly capacity, and,
on the other hand, heavier loads require more costly engines. This
matter is further discussed under "Haulage Appliances."
We have therefore as the economic minimum a shaft of three compartments
(Fig. 9), each four to five feet square. When the maximum tonnage
is wanted from such a shaft at the least operating cost, it should
be equipped with loading bins and skips.
The output capacity of shafts of this size and equipment will depend
in a major degree upon the engine employed, and in a less degree
upon the hauling depth. The reason why depth is a subsidiary factor
is that the rapidity with which a load can be drawn is not wholly a
factor of depth. The time consumed in hoisting is partially expended
in loading, in acceleration and retardation of the engine, and in
discharge of the load. These factors are constant for any depth,
and extra distance is therefore accomplished at full speed of the
engine.
Vertical shafts will, other things being equal, have greater capacity
than inclines, as winding will be much faster and length of haul less
for same depth. Since engines have, however, a great tractive ability
on inclines, by an increase in the size of skip it is usually possible
partially to equalize matters. Therefore the size of inclines for
the same output need not differ materially from vertical shafts.
The maximum capacity of a shaft whose equipment is of the character
and size given above, will, as stated, decrease somewhat with extension
in depth of the haulage horizon. At 500 feet, such a shaft if vertical
could produce 70 to 80 tons per hour comfortably with an engine
whose winding speed was 700 feet per minute. As men and material
other than ore have to be handled in and out of the mine, and
shaft-sinking has to be attended to, the winding engine cannot
be employed all the time on ore. Twelve hours of actual daily
ore-winding are all that can be expected without auxiliary help.
This represents a capacity from such a depth of 800 to 1,000 tons
per day. A similar shaft, under ordinary working conditions, with
an engine speed of 2,000 feet per minute, should from, say, 3,000
feet have a capacity of about 400 to 600 tons daily.
It is desirable to inquire at what stages the size of shaft should
logically be enlarged in order to attain greater capacity. A
considerable measure of increase can be obtained by relieving the
main hoisting engine of all or part of its collateral duties. Where
the pumping machinery is not elaborate, it is often possible to
get a small single winding compartment into the gangway without
materially increasing the size of the shaft if the haulage compartments
be made somewhat narrower (Fig. 10). Such a compartment would be
operated by an auxiliary engine for sinking, handling tools and
material, and assisting in handling men. If this arrangement can
be effected, the productive time of the main engine can be expanded
to about twenty hours with an addition of about two-thirds to the
output.
Where the exigencies of pump and gangway require more than two
and one-half feet of shaft length, the next stage of expansion
becomes four full-sized compartments (Fig. 11). By thus enlarging the
auxiliary winding space, some assistance may be given to ore-haulage
in case of necessity. The mine whose output demands such haulage
provisions can usually stand another foot of width to the shaft,
so that the dimensions come to about 21 feet to 22 feet by 7 feet
to 8 feet outside the timbers. Such a shaft, with three- to four-ton
skips and an appropriate engine, will handle up to 250 tons per
hour from a depth of 1,000 feet.
The next logical step in advance is the shaft of five compartments
with four full-sized haulage ways (Fig. 13), each of greater size
than in the above instance. In this case, the auxiliary engine
becomes a balanced one, and can be employed part of the time upon
ore-haulage. Such a shaft will be about 26 feet to 28 feet long
by 8 feet wide outside the timbers, when provision is made for
one gangway. The capacity of such shafts can be up to 4,000 tons a
day, depending on the depth and engine. When very large quantities
of water are to be dealt with and rod-driven pumps to be used,
two pumping compartments are sometimes necessary, but other forms
of pumps do not require more than one compartment,--an additional
reason for their use.
For depths greater than 3,000 feet, other factors come into play.
Ventilation questions become of more import. The mechanical problems
on engines and ropes become involved, and their sum-effect is to
demand much increased size and a greater number of compartments.
The shafts at Johannesburg intended as outlets for workings 5,000
feet deep are as much as 46 feet by 9 feet outside timbers.
It is not purposed to go into details as to sinking methods or
timbering. While important matters, they would unduly prolong this
discussion. Besides, a multitude of treatises exist on these subjects
and cover all the minutiæ of such work.
SPEED OF SINKING.--Mines may be divided into two cases,--those
being developed only, and those being operated as well as developed.
In the former, the entrance into production is usually dependent
upon the speed at which the shaft is sunk. Until the mine is earning
profits, there is a loss of interest on the capital involved, which,
in ninety-nine instances out of a hundred, warrants any reasonable
extra expenditure to induce more rapid progress. In the case of
mines in operation, the volume of ore available to treatment or
valuation is generally dependent to a great degree upon the rapidity
of the extension of workings in depth. It will be demonstrated
later that, both from a financial and a technical standpoint, the
maximum development is the right one and that unremitting extension
in depth is not only justifiable but necessary.
Speed under special conditions or over short periods has a more
romantic than practical interest, outside of its value as a stimulant
to emulation. The thing that counts is the speed which can be maintained
over the year. Rapidity of sinking depends mainly on:--
_a_. Whether the shaft is or is not in use for operating the
mine.
_b_. The breaking character of the rock.
_c_. The amount of water.
The delays incident to general carrying of ore and men are such that
the use of the main haulage engine for shaft-sinking is practically
impossible, except on mines with small tonnage output. Even with a
separate winch or auxiliary winding-engine, delays are unavoidable
in a working shaft, especially as it usually has more water to contend
with than one not in use for operating the mine. The writer's own
impression is that an average of 40 feet per month is the maximum
possibility for year in and out sinking under such conditions. In
fact, few going mines manage more than 400 feet a year. In cases
of clean shaft-sinking, where every energy is bent to speed, 150
feet per month have been averaged for many months. Special cases
have occurred where as much as 213 feet have been achieved in a
single month. With ordinary conditions, 1,200 feet in a year is
very good work. Rock awkward to break, and water especially, lowers
the rate of progress very materially. Further reference to speed
will be found in the chapter on "Drilling Methods."
TUNNEL ENTRY.--The alternative of entry to a mine by tunnel is
usually not a question of topography altogether, but, like everything
else in mining science, has to be tempered to meet the capital
available and the expenditure warranted by the value showing.
In the initial prospecting of a mine, tunnels are occasionally
overdone by prospectors. Often more would be proved by a few inclines.
As the pioneer has to rely upon his right arm for hoisting and
drainage, the tunnel offers great temptations, even when it is
long and gains but little depth. At a more advanced stage of
development, the saving of capital outlay on hoisting and pumping
equipment, at a time when capital is costly to secure, is often
sufficient justification for a tunnel entry. But at the stage where
the future working of ore below a tunnel-level must be contemplated,
other factors enter. For ore below tunnel-level a shaft becomes
necessary, and in cases where a tunnel enters a few hundred feet
below the outcrop the shaft should very often extend to the surface,
because internal shafts, winding from tunnel-level, require large
excavations to make room for the transfer of ore and for winding
gear. The latter must be operated by transmitted power, either
that of steam, water, electricity, or air. Where power has to be
generated on the mine, the saving by the use of direct steam, generated
at the winding gear, is very considerable. Moreover, the cost of
haulage through a shaft for the extra distance from tunnel-level
to the surface is often less than the cost of transferring the
ore and removing it through the tunnel. The load once on the
winding-engine, the consumption of power is small for the extra
distance, and the saving of labor is of consequence. On the other
hand, where drainage problems arise, they usually outweigh all
other considerations, for whatever the horizon entered by tunnel,
the distance from that level to the surface means a saving of
water-pumpage against so much head. The accumulation of such constant
expense justifies a proportioned capital outlay. In other words,
the saving of this extra pumping will annually redeem the cost of
a certain amount of tunnel, even though it be used for drainage
only.
In order to emphasize the rapidity with which such a saving of
constant expense will justify capital outlay, one may tabulate the
result of calculations showing the length of tunnel warranted with
various hypothetical factors of quantity of water and height of lift
eliminated from pumping. In these computations, power is taken at
the low rate of $60 per horsepower-year, the cost of tunneling at
an average figure of $20 per foot, and the time on the basis of
a ten-year life for the mine.
Feet of Tunnel Paid for in 10 Years with Under-mentioned Conditions.
=============================================================
Feet of | 100,000 | 200,000 | 300,000 | 500,000 |1,000,000
Water Lift | Gallons | Gallons | Gallons | Gallons | Gallons
Avoided |per Diem |per Diem |per Diem |per Diem |per Diem
-----------|---------|---------|---------|---------|---------
100 | 600 | 1,200 | 1,800 | 3,000 | 6,000
200 | 1,200 | 2,400 | 3,600 | 6,000 | 12,000
300 | 1,800 | 3,600 | 5,400 | 9,000 | 18,000
500 | 3,000 | 6,000 | 9,000 | 15,000 | 30,000
1,000 | 6,000 | 12,000 | 18,000 | 30,000 | 60,000
=============================================================
The size of tunnels where ore-extraction is involved depends upon
the daily tonnage output required, and the length of haul. The
smallest size that can be economically driven and managed is about
6-1/2 feet by 6 feet inside the timbers. Such a tunnel, with single
track for a length of 1,000 feet, with one turn-out, permits handling
up to 500 tons a day with men and animals. If the distance be longer
or the tonnage greater, a double track is required, which necessitates
a tunnel at least 8 feet wide by 6-1/2 feet to 7 feet high, inside
the timbers.
There are tunnel projects of a much more impressive order than those
designed to operate upper levels of mines; that is, long crosscut
tunnels designed to drain and operate mines at very considerable
depths, such as the Sutro tunnel at Virginia City. The advantage
of these tunnels is very great, especially for drainage, and they
must be constructed of large size and equipped with appliances
for mechanical haulage.
CHAPTER IX.
Development of Mines (_Concluded_).
SUBSIDIARY DEVELOPMENT;--STATIONS; CROSSCUTS; LEVELS; INTERVAL
BETWEEN LEVELS; PROTECTION OF LEVELS; WINZES AND RISES. DEVELOPMENT
IN THE PROSPECTING STAGE; DRILLING.
SUBSIDIARY DEVELOPMENT.
Stations, crosscuts, levels, winzes, and rises follow after the
initial entry. They are all expensive, and the least number that
will answer is the main desideratum.
STATIONS.--As stations are the outlets of the levels to the shaft,
their size and construction is a factor of the volume and character
of the work at the levels which they are to serve. If no timber
is to be handled, and little ore, and this on cages, the stations
need be no larger than a good sized crosscut. Where timber is to
be let down, they must be ten to fifteen feet higher than the floor
of the crosscut. Where loading into skips is to be provided for,
bins must be cut underneath and sufficient room be provided to
shift the mine cars comfortably. Such bins are built of from 50 to
500 tons' capacity in order to contain some reserve for hoisting
purposes, and in many cases separate bins must be provided on opposite
sides of the shaft for ore and waste. It is a strong argument in
favor of skips, that with this means of haulage storage capacity
at the stations is possible, and the hoisting may then go on
independently of trucking and, as said before, there are no idle
men at the stations.
[Illustration: Fig. 15.--Cross-section of station arrangement for
skip-haulage in vertical shaft.]
[Illustration: Fig. 16.--Cross-section of station arrangement for
skip-haulage in vertical shaft.]
It is always desirable to concentrate the haulage to the least
number of levels, for many reasons. Among them is that, where haulage
is confined to few levels, storage-bins are not required at every
station. Figures 15, 16, 17, and 18 illustrate various arrangements
of loading bins.
CROSSCUTS.--Crosscuts are for two purposes, for roadway connection
of levels to the shaft or to other levels, and for prospecting
purposes. The number of crosscuts for roadways can sometimes be
decreased by making the connections with the shaft at every second
or even every third level, thus not only saving in the construction
cost of crosscuts and stations, but also in the expenses of scattered
tramming. The matter becomes especially worth considering where
the quantity of ore that can thus be accumulated warrants mule
or mechanical haulage. This subject will be referred to later on.
[Illustration: Fig. 17.--Arrangement of loading chutes in vertical
shaft.]
On the second purpose of crosscuts,--that of prospecting,--one
observation merits emphasis. This is, that the tendency of ore-fissures
to be formed in parallels warrants more systematic crosscutting
into the country rock than is done in many mines.
[Illustration: Fig. 18.--Cross-section of station arrangement for
skip-haulage in inclined shaft.]
LEVELS.
The word "level" is another example of miners' adaptations in
nomenclature. Its use in the sense of tunnels driven in the direction
of the strike of the deposit has better, but less used, synonyms in
the words "drifts" or "drives." The term "level" is used by miners
in two senses, in that it is sometimes applied to all openings on one
horizon, crosscuts included. Levels are for three purposes,--for a
stoping base; for prospecting the deposit; and for roadways. As a
prospecting and a stoping base it is desirable that the level should
be driven on the deposit; as a roadway, that it should constitute
the shortest distance between two points and be in the soundest
ground. On narrow, erratic deposits the levels usually must serve
all three purposes at once; but in wider and more regular deposits
levels are often driven separately for roadways from the level
which forms the stoping base and prospecting datum.
There was a time when mines were worked by driving the level on ore
and enlarging it top and bottom as far as the ground would stand,
then driving the next level 15 to 20 feet below, and repeating the
operation. This interval gradually expanded, but for some reason
100 feet was for years assumed to be the proper distance between
levels. Scattered over every mining camp on earth are thousands
of mines opened on this empirical figure, without consideration
of the reasons for it or for any other distance.
The governing factors in determining the vertical interval between
levels are the following:--
_a_. The regularity of the deposit.
_b_. The effect of the method of excavation of winzes and rises.
_c_. The dip and the method of stoping.
REGULARITY OF THE DEPOSIT.--From a prospecting point of view the
more levels the better, and the interval therefore must be determined
somewhat by the character of the deposit. In erratic deposits there
is less risk of missing ore with frequent levels, but it does not
follow that every level need be a through roadway to the shaft or
even a stoping base. In such deposits, intermediate levels for
prospecting alone are better than complete levels, each a roadway.
Nor is it essential, even where frequent levels are required for
a stoping base, that each should be a main haulage outlet to the
shaft. In some mines every third level is used as a main roadway,
the ore being poured from the intermediate ones down to the haulage
line. Thus tramming and shaft work, as stated before, can be
concentrated.
EFFECT OF METHOD OF EXCAVATING WINZES AND RISES.--With hand drilling
and hoisting, winzes beyond a limited depth become very costly to
pull spoil out of, and rises too high become difficult to ventilate,
so that there is in such cases a limit to the interval desirable
between levels, but these difficulties largely disappear where
air-winches and air-drills are used.
THE DIP AND METHOD OF STOPING.--The method of stoping is largely
dependent upon the dip, and indirectly thus affects level intervals.
In dips under that at which material will "flow" in the stopes--about
45° to 50°--the interval is greatly dependent on the method of
stope-transport. Where ore is to be shoveled from stopes to the
roadway, the levels must be comparatively close together. Where
deposits are very flat, under 20°, and walls fairly sound, it is
often possible to use a sort of long wall system of stoping and to
lay tracks in the stopes with self-acting inclines to the levels.
In such instances, the interval can be expanded to 250 or even 400
feet. In dips between 20° and 45°, tracks are not often possible,
and either shoveling or "bumping troughs"[*] are the only help
to transport. With shoveling, intervals of 100 feet[**] are most
common, and with troughs the distance can be expanded up to 150
or 175 feet.
[Footnote *: Page 136.]
[Footnote **: Intervals given are measured on the dip.]
In dips of over 40° to 50°, depending on the smoothness of the foot
wall, the distance can again be increased, as stope-transport is
greatly simplified, since the stope materials fall out by gravity.
In timbered stopes, in dips over about 45°, intervals of 150 to
200 feet are possible. In filled stopes intervals of over 150 feet
present difficulties in the maintenance of ore-passes, for the wear
and tear of longer use often breaks the timbers. In shrinkage-stopes,
where no passes are to be maintained and few winzes put through, the
interval is sometimes raised to 250 feet. The subject is further
discussed under "Stoping."
Another factor bearing on level intervals is the needed insurance
of sufficient points of stoping attack to keep up a certain output.
This must particularly influence the manager whose mine has but
little ore in reserve.
[Illustration: Fig. 19.]
PROTECTION OF LEVELS.--Until recent years, timbering and occasional
walling was the only method for the support of the roof, and for
forming a platform for a stoping base. Where the rock requires no
support sublevels can be used as a stoping base, and timbering
for such purpose avoided altogether (Figs. 38, 39, 42). In such
cases the main roadway can then be driven on straight lines, either
in the walls or in the ore, and used entirely for haulage. The
subheading for a stoping base is driven far enough above or below
the roadway (depending on whether overhand or underhand stoping
is to be used) to leave a supporting pillar which is penetrated
by short passes for ore. In overhand stopes, the ore is broken
directly on the floor of an upper sublevel; and in underhand stopes,
broken directly from the bottom of the sublevel. The method entails
leaving a pillar of ore which can be recovered only with difficulty
in mines where stope-support is necessary. The question of its
adoption is then largely one of the comparative cost of timbering,
the extra cost of the sublevel, and the net value of the ore left.
In bad swelling veins, or badly crushing walls, where constant
repair to timbers would be necessary, the use of a sublevel is a
most useful alternative. It is especially useful with stopes to
be left open or worked by shrinkage-stoping methods.
If the haulage level, however, is to be the stoping base, some
protection to the roadway must be provided. There are three systems
in use,--by wood stulls or sets (Figs. 19, 30, 43), by dry-walling
with timber caps (Fig. 35), and in some localities by steel sets.
Stulls are put up in various ways, and, as their use entails the
least difficulty in taking the ore out from beneath the level,
they are much favored, but are applicable only in comparatively
narrow deposits.
WINZES AND RISES.
These two kinds of openings for connecting two horizons in a mine
differ only in their manner of construction. A winze is sunk underhand,
while a rise is put up overhand. When the connection between levels
is completed, a miner standing at the bottom usually refers to
the opening as a rise, and when he goes to the top he calls it
a winze. This confusion in terms makes it advisable to refer to
all such completed openings as winzes, regardless of how they are
constructed.
In actual work, even disregarding water, it costs on the average
about 30% less to raise than to sink such openings, for obviously
the spoil runs out or is assisted by gravity in one case, and in
the other has to be shoveled and hauled up. Moreover, it is easier
to follow the ore in a rise than in a winze. It usually happens,
however, that in order to gain time both things are done, and for
prospecting purposes sinking is necessary.
The number of winzes required depends upon the method of stoping
adopted, and is mentioned under "Stoping." After stoping, the number
necessary to be maintained open depends upon the necessities of
ventilation, of escape, and of passageways for material to be used
below. Where stopes are to be filled with waste, more winzes must
be kept open than when other methods are used, and these winzes
must be in sufficient alignment to permit the continuous flow of
material down past the various levels. In order that the winzes
should deliver timber and filling to the most advantageous points,
they should, in dipping ore-bodies, be as far as possible on the
hanging wall side.
DEVELOPMENT IN THE EARLY PROSPECTING STAGE.
The prime objects in the prospecting stage are to expose the ore
and to learn regarding the ore-bodies something of their size, their
value, metallurgical character, location, dip, strike, etc.,--so much
at least as may be necessary to determine the works most suitable
for their extraction or values warranting purchase. In outcrop mines
there is one rule, and that is "follow the ore." Small temporary
inclines following the deposit, even though they are eventually
useless; are nine times out of ten justified.
In prospecting deep-level projects, it is usually necessary to
layout work which can be subsequently used in operating the mine,
because the depth involves works of such considerable scale, even
for prospecting, that the initial outlay does not warrant any
anticipation of revision. Such works have to be located and designed
after a study of the general geology as disclosed in adjoining mines.
Practically the only method of supplementing such information is
by the use of churn- and diamond-drills.
DRILLING.--Churn-drills are applicable only to comparatively shallow
deposits of large volume. They have an advantage over the diamond
drill in exposing a larger section and in their application to
loose material; but inability to determine the exact horizon of
the spoil does not lend them to narrow deposits, and in any event
results are likely to be misleading from the finely ground state of
the spoil. They are, however, of very great value for preliminary
prospecting to shallow horizons.
Two facts in diamond-drilling have to be borne in mind: the indication
of values is liable to be misleading, and the deflection of the drill
is likely to carry it far away from its anticipated destination.
A diamond-drill secures a small section which is sufficiently large
to reveal the geology, but the values disclosed in metal mines must
be accepted with reservations. The core amounts to but a little
sample out of possibly large amounts of ore, which is always of
variable character, and the core is most unlikely to represent
the average of the deposit. Two diamond-drill holes on the Oroya
Brownhill mine both passed through the ore-body. One apparently
disclosed unpayable values, the other seemingly showed ore forty
feet in width assaying $80 per ton. Neither was right. On the other
hand, the predetermination of the location of the ore-body justified
expenditure. A recent experiment at Johannesburg of placing a copper
wedge in the hole at a point above the ore-body and deflecting
the drill on reintroducing it, was successful in giving a second
section of the ore at small expense.
The deflection of diamond-drill holes from the starting angle is
almost universal. It often amounts to a considerable wandering
from the intended course. The amount of such deflection varies
with no seeming rule, but it is probable that it is especially
affected by the angle at which stratification or lamination planes
are inclined to the direction of the hole. A hole has been known
to wander in a depth of 1,500 feet more than 500 feet from the
point intended. Various instruments have been devised for surveying
deep holes, and they should be brought into use before works are
laid out on the basis of diamond-drill results, although none of
the inventions are entirely satisfactory.
CHAPTER X.
Stoping.
METHODS OF ORE-BREAKING; UNDERHAND STOPES; OVERHAND STOPES; COMBINED
STOPE. VALUING ORE IN COURSE OF BREAKING.
There is a great deal of confusion in the application of the word
"stoping." It is used not only specifically to mean the actual
ore-breaking, but also in a general sense to indicate all the operations
of ore-breaking, support of excavations, and transportation between
levels. It is used further as a noun to designate the hole left
when the ore is taken out. Worse still, it is impossible to adhere
to miners' terms without employing it in every sense, trusting
to luck and the context to make the meaning clear.
The conditions which govern the method of stoping are in the main:--
_a_. The dip.
_b_. The width of the deposit.
_c_. The character of the walls.
_d_. The cost of materials.
_e_. The character of the ore.
Every mine, and sometimes every stope in a mine, is a problem special
to itself. Any general consideration must therefore be simply an
inquiry into the broad principles which govern the adaptability of
special methods. A logical arrangement of discussion is difficult,
if not wholly impossible, because the factors are partially
interdependent and of varying importance.
For discussion the subject may be divided into:
1. Methods of ore-breaking.
2. Methods of supporting excavation.
3. Methods of transport in stopes.
METHODS OF ORE-BREAKING.
The manner of actual ore-breaking is to drill and blast off slices
from the block of ground under attack. As rock obviously breaks
easiest when two sides are free, that is, when corners can be broken
off, the detail of management for blasts is therefore to set the holes
so as to preserve a corner for the next cut; and as a consequence
the face of the stope shapes into a series of benches (Fig.
22),--inverted benches in the case of overhand stopes (Figs. 20,
21). The size of these benches will in a large measure depend on
the depth of the holes. In wide stopes with machine-drills they
vary from 7 to 10 feet; in narrow stopes with hand-holes, from
two to three feet.
[Illustration: Fig. 20.]
The position of the men in relation to the working face gives rise
to the usual primary classification of the methods of stoping.
They are:--
1. Underhand stopes,
2. Overhand stopes,
3. Combined stopes.
These terms originated from the direction of the drill-holes, but
this is no longer a logical basis of distinction, for underhand
holes in overhand stopes,--as in rill-stoping,--are used entirely
in some mines (Fig. 21).
[Illustration: Fig. 21.]
UNDERHAND STOPES.--Underhand stopes are those in which the ore
is broken downward from the levels. Inasmuch as this method has
the advantage of allowing the miner to strike his blows downward
and to stand upon the ore when at work, it was almost universal
before the invention of powder; and was applied more generally
before the invention of machine-drills than since. It is never
rightly introduced unless the stope is worked back from winzes
through which the ore broken can be let down to the level below,
as shown in Figures 22 and 23.
[Illustration: Fig. 22.]
This system can be advantageously applied only in the rare cases
in which the walls require little or no support, and where very
little or no waste requiring separation is broken with the ore
in the stopes. To support the walls in bad ground in underhand
stopes would be far more costly than with overhand stopes, for
square-set timbering would be most difficult to introduce, and
to support the walls with waste and stulls would be even more
troublesome. Any waste broken must needs be thrown up to the level
above or be stored upon specially built stages--again a costly
proceeding.
A further drawback lies in the fact that the broken ore follows
down the face of the stope, and must be shoveled off each bench.
It thus all arrives at a single point,--the winze,--and must be
drawn from a single ore-pass into the level. This usually results
not only in more shoveling but in a congestion at the passes not
present in overhand stoping, for with that method several chutes
are available for discharging ore into the levels. Where the walls
require no support and no selection is desired in the stopes, the
advantage of the men standing on the solid ore to work, and of
having all down holes and therefore drilled wet, gives this method
a distinct place. In using this system, in order to protect the
men, a pillar is often left under the level by driving a sublevel,
the pillar being easily recoverable later. The method of sublevels
is of advantage largely in avoiding the timbering of levels.
[Illustration: Fig. 23.--Longitudinal section of an underhand stope.]
OVERHAND STOPES.--By far the greatest bulk of ore is broken overhand,
that is broken upward from one level to the next above. There are
two general forms which such stopes are given,--"horizontal" and
"rill."
[Illustration: Fig. 24.--Horizontal-cut overhand stope--longitudinal
section.]
The horizontal "flat-back" or "long-wall" stope, as it is variously
called, shown in Figure 24, is operated by breaking the ore in slices
parallel with the levels. In rill-stoping the ore is cut back from
the winzes in such a way that a pyramid-shaped room is created,
with its apex in the winze and its base at the level (Figs. 25 and
26). Horizontal or flat-backed stopes can be applied to almost any
dip, while "rill-stoping" finds its most advantageous application
where the dip is such that the ore will "run," or where it can be
made to "run" with a little help. The particular application of
the two systems is dependent not only on the dip but on the method
of supporting the excavation and the ore. With rill-stoping, it is
possible to cut the breaking benches back horizontally from the
winzes (Fig. 25), or to stagger the cuts in such a manner as to
take the slices in a descending angle (Figs. 21 and 26).
[Illustration: Fig. 25.--Rill-cut overhand stope--longitudinal section.]
In the "rill" method of incline cuts, all the drill-holes are "down"
holes (Fig. 21), and can be drilled wet, while in horizontal cuts
or flat-backed stopes, at least part of the holes must be "uppers"
(Fig. 20). Aside from the easier and cheaper drilling and setting
up of machines with this kind of "cut," there is no drill dust,--a
great desideratum in these days of miners' phthisis. A further
advantage in the "rill" cut arises in cases where horizontal jointing
planes run through the ore of a sort from which unduly large masses
break away in "flat-back" stopes. By the descending cut of the
"rill" method these calamities can be in a measure avoided. In
cases of dips over 40º the greatest advantage in "rill" stoping
arises from the possibility of pouring filling or timber into the
stope from above with less handling, because the ore and material
will run down the sides of the pyramid (Figs. 32 and 34). Thus
not only is there less shoveling required, but fewer ore-passes
and a less number of preliminary winzes are necessary, and a wider
level interval is possible. This matter will be gone into more
fully later.
[Illustration: Fig. 26.--Rill-cut overhand stope-longitudinal section.]
COMBINED STOPES.--A combined stope is made by the coincident working
of the underhand and "rill" method (Fig. 27). This order of stope
has the same limitations in general as the underhand kind. For
flat veins with strong walls, it has a great superiority in that
the stope is carried back more or less parallel with the winzes,
and thus broken ore after blasting lies in a line on the gradient
of the stope. It is, therefore, conveniently placed for mechanical
stope haulage. A further advantage is gained in that winzes may
be placed long distances apart, and that men are not required,
either when at work or passing to and from it, to be ever far from
the face, and they are thus in the safest ground, so that timber
and filling protection which may be otherwise necessary is not
required. This method is largely used in South Africa.
[Illustration: Fig. 27.--Longitudinal section of a combined stope.]
MINIMUM WIDTH OF STOPES.--The minimum stoping width which can be
consistently broken with hand-holes is about 30 inches, and this
only where there is considerable dip to the ore. This space is
so narrow that it is of doubtful advantage in any case, and 40
inches is more common in narrow mines, especially where worked
with white men. Where machine-drills are used about 4 feet is the
minimum width feasible.
RESUING.--In very narrow veins where a certain amount of wall-rock
must be broken to give working space, it pays under some circumstances
to advance the stope into the wall-rock ahead of the ore, thus
stripping the ore and enabling it to be broken separately. This
permits of cleaner selection of the ore; but it is a problem to
be worked out in each case, as to whether rough sorting of some
waste in the stopes, or further sorting at surface with inevitable
treatment of some waste rock, is more economical than separate
stoping cuts and inevitably wider stopes.
VALUING ORE IN COURSE OF BREAKING.--There are many ores whose payability
can be determined by inspection, but there are many of which it cannot.
Continuous assaying is in the latter cases absolutely necessary
to avoid the treatment of valueless material. In such instances,
sampling after each stoping-cut is essential, the unprofitable ore
being broken down and used as waste. Where values fade into the
walls, as in impregnation deposits, the width of stopes depends
upon the limit of payability. In these cases, drill-holes are put
into the walls and the drillings assayed. If the ore is found
profitable, the holes are blasted out. The gauge of what is profitable
in such situations is not dependent simply upon the average total
working costs of the mine, for ore in that position can be said to
cost nothing for development work and administration; moreover,
it is usually more cheaply broken than the average breaking cost,
men and machines being already on the spot.
CHAPTER XI.
Methods of Supporting Excavation.
TIMBERING; FILLING WITH WASTE; FILLING WITH BROKEN ORE; PILLARS
OF ORE; ARTIFICIAL PILLARS; CAVING SYSTEM.
Most stopes require support to be given to the walls and often to
the ore itself. Where they do require support there are five principal
methods of accomplishing it. The application of any particular method
depends upon the dip, width of ore-body, character of the ore and
walls, and cost of materials. The various systems are by:--
1. Timbering.
2. Filling with waste.
3. Filling with broken ore subsequently withdrawn.
4. Pillars of ore.
5. Artificial pillars built of timbers and waste.
6. Caving.
TIMBERING.--At one time timbering was the almost universal means of
support in such excavations, but gradually various methods for the
economical application of waste and ore itself have come forward,
until timbering is fast becoming a secondary device. Aside from
economy in working without it, the dangers of creeps, or crushing,
and of fires are sufficient incentives to do away with wood as
far as possible.
There are three principal systems of timber support to excavations,--by
stulls, square-sets, and cribs.
Stulls are serviceable only where the deposit is so narrow that
the opening can be bridged by single timbers between wall and wall
(Figs. 28 and 43). This system can be applied to any dip and is most
useful in narrow deposits where the walls are not too heavy. Stulls
in inclined deposits are usually set at a slightly higher angle than
that perpendicular to the walls, in order that the vertical pressure
of the hanging wall will serve to tighten them in position. The
"stull" system can, in inclined deposits, be further strengthened by
building waste pillars against them, in which case the arrangement
merges into the system of artificial pillars.
[Illustration: Fig. 28.--Longitudinal section of stull-supported
stope.]
[Illustration: Fig. 29.--Longitudinal section showing square-set
timbering.]
[Illustration: Fig. 30.--Square-set timbering on inclined ore-body.
Showing ultimate strain on timbers.]
Square-sets (Figs. 29 and 30), that is, trusses built in the opening
as the ore is removed, are applicable to almost any dip or width
of ore, but generally are applied only in deposits too wide, or to
rock too heavy, for stulls. Such trusses are usually constructed on
vertical and horizontal lines, and while during actual ore-breaking
the strains are partially vertical, ultimately, however, when the
weight of the walls begins to be felt, these strains, except in
vertical deposits, come at an angle to lines of strength in the
trusses, and therefore timber constructions of this type present
little ultimate resistance (Fig. 30). Square-set timbers are sometimes
set to present the maximum resistance to the direction of strain,
but the difficulties of placing them in position and variations in
the direction of strain on various parts of the stope do not often
commend the method. As a general rule square-sets on horizontal
lines answer well enough for the period of actual ore-breaking. The
crushing or creeps is usually some time later; and if the crushing
may damage the whole mine, their use is fraught with danger.
Reënforcement by building in waste is often resorted to. When done
fully, it is difficult to see the utility of the enclosed timber,
for entire waste-filling would in most cases be cheaper and equally
efficient.
[Illustration: Fig. 31.--"Cribs."]
There is always, with wood constructions, as said before, the very
pertinent danger of subsequent crushing and of subsidence in after
years, and the great risk of fires. Both these disasters have cost
Comstock and Broken Hill mines, directly or indirectly, millions of
dollars, and the outlay on timber and repairs one way or another
would have paid for the filling system ten times over. There are
cases where, by virtue of the cheapness of timber, "square-setting"
is the most economical method. Again, there are instances where the
ore lies in such a manner--particularly in limestone replacements--as
to preclude other means of support. These cases are being yearly
more and more evaded by the ingenuity of engineers in charge. The
author believes it soon will be recognized that the situation is
rare indeed where complete square-setting is necessarily without an
economical alternative. An objection is sometimes raised to filling
in favor of timber, in that if it become desirable to restope the
walls for low-grade ore left behind, such stopes could only be
entered by drawing the filling, with consequent danger of total
collapse. Such a contingency can be provided for in large ore-bodies
by installing an outer shell of sets of timber around the periphery
of the stope and filling the inside with waste. If the crushing
possibilities are too great for this method then, the subsequent
recovery of ore is hopeless in any event. In narrow ore-bodies
with crushing walls recovery of ore once left behind is not often
possible.
The third sort of timber constructions are cribs, a "log-house" sort
of structure usually filled with waste, and more fully discussed
under artificial pillars (Fig. 31). The further comparative merits
of timbering with other methods will be analyzed as the different
systems are described.
FILLING WITH WASTE.--The system of filling stope-excavations completely
with waste in alternating progress with ore-breaking is of wide
and increasingly general application (Figs. 32, 33, 34, 35).
Although a certain amount of waste is ordinarily available in the
stopes themselves, or from development work in the mine, such a
supply must usually be supplemented from other directions. Treatment
residues afford the easiest and cheapest handled material. Quarried
rock ranks next, and in default of any other easy supply, materials
from crosscuts driven into the stope-walls are sometimes resorted
to.
In working the system to the best advantage, the winzes through
the block of ore under attack are kept in alignment with similar
openings above, in order that filling may be poured through the
mine from the surface or any intermediate point. Winzes to be used
for filling should be put on the hanging-wall side of the area to
be filled, for the filling poured down will then reach the foot-wall
side of the stopes with a minimum of handling. In some instances,
one special winze is arranged for passing all filling from the
surface to a level above the principal stoping operations; and
it is then distributed along the levels into the winzes, and thus
to the operating stopes, by belt-conveyors.
[Illustration: Fig. 32.--Longitudinal section. Rill stope filled
with waste.]
[Illustration: Fig. 33.--Longitudinal section. Horizontal stope
filled with waste.]
[Illustration: Fig. 34.--Longitudinal section. Waste-filled stope
with dry-walling of levels and passes.]
In this system of stope support the ore is broken at intervals
alternating with filling. If there is danger of much loss from
mixing broken ore and filling, "sollars" of boards or poles are
laid on the waste. If the ore is very rich, old canvas or cowhides
are sometimes put under the boards. Before the filling interval,
the ore passes are built close to the face above previous filling
and their tops covered temporarily to prevent their being filled
with running waste. If the walls are bad, the filling is kept close
to the face. If the unbroken ore requires support, short stulls
set on the waste (as in Fig. 39) are usually sufficient until the
next cut is taken off, when the timber can be recovered. If stulls
are insufficient, cribs or bulkheads (Fig. 31) are also used and
often buried in the filling.
[Illustration: Fig. 35.--Cross-section of Fig. 34 on line _A-B_.]
Both flat-backed and rill-stope methods of breaking are employed in
conjunction with filled stopes. The advantages of the rill-stopes
are so patent as to make it difficult to understand why they are
not universally adopted when the dip permits their use at all. In
rill-stopes (Figs. 32 and 34) the waste flows to its destination
with a minimum of handling. Winzes and ore-passes are not required
with the same frequency as in horizontal breaking, and the broken
ore always lies on the slope towards the passes and is therefore
also easier to shovel. In flat-backed stopes (Fig. 33) winzes must
be put in every 50 feet or so, while in rill-stopes they can be
double this distance apart. The system is applicable by modification
to almost any width of ore. It finds its most economical field
where the dip of the stope floor is over 45°, when waste and ore,
with the help of the "rill," will flow to their destination. For
dips from under about 45° to about 30° or 35°, where the waste
and ore will not "flow" easily, shoveling can be helped by the
use of the "rill" system and often evaded altogether, if flow be
assisted by a sheet-iron trough described in the discussion of
stope transport. Further saving in shoveling can be gained in this
method, by giving a steeper pitch to the filling winzes and to the
ore-passes, by starting them from crosscuts in the wall, and by
carrying them at greater angles than the pitch of the ore (Fig.
36). These artifices combined have worked out most economically
on several mines within the writer's experience, with the dip as
flat as 30°. For very flat dips, where filling is to be employed,
rill-stoping has no advantage over flat-backed cuts, and in such
cases it is often advisable to assist stope transport by temporary
tracks and cars which obviously could not be worked on the tortuous
contour of a rill-stope, so that for dips under 30° advantage lies
with "flat-backed" ore-breaking.
[Illustration: Fig. 36.--Cross-section showing method of steepening
winzes and ore passes.]
On very wide ore-bodies where the support of the standing ore itself
becomes a great problem, the filling system can be applied by combining
it with square-setting. In this case the stopes are carried in
panels laid out transversally to the strike as wide as the standing
strength of the ore permits. On both sides of each panel a fence
of lagged square-sets is carried up and the area between is filled
with waste. The panels are stoped out alternately. The application
of this method at Broken Hill will be described later. (See pages
120 and Figs. 41 and 42.) The same type of wide ore-body can be
managed also on the filling system by the use of frequent "bulkheads"
to support the ore (Fig. 31).
Compared with timbering methods, filling has the great advantage
of more effective support to the mine, less danger of creeps, and
absolute freedom from the peril of fire. The relative expense of
the two systems is determined by the cost of materials and labor.
Two extreme cases illustrate the result of these economic factors
with sufficient clearness. It is stated that the cost of timbering
stopes on the Le Roi Mine by square-sets is about 21 cents per
ton of ore excavated. In the Ivanhoe mine of West Australia the
cost of filling stopes with tailings is about 22 cents per ton
of ore excavated. At the former mine the average cost of timber
is under $10 per M board-measure, while at the latter its price
would be $50 per M board-measure; although labor is about of the
same efficiency and wage, the cost in the Ivanhoe by square-setting
would be about 65 cents per ton of ore broken. In the Le Roi, on the
other hand, no residues are available for filling. To quarry rock
or drive crosscuts into the walls might make this system cost 65
cents per ton of ore broken if applied to that mine. The comparative
value of the filling method with other systems will be discussed
later.
FILLING WITH BROKEN ORE SUBSEQUENTLY WITHDRAWN.--This order of support
is called by various names, the favorite being "shrinkage-stoping."
The method is to break the ore on to the roof of the level, and by
thus filling the stope with broken ore, provide temporary support
to the walls and furnish standing floor upon which to work in making
the next cut (Figs. 37, 38, and 39.) As broken material occupies 30
to 40% more space than rock _in situ_, in order to provide working
space at the face, the broken ore must be drawn from along the level
after each cut. When the area attacked is completely broken through
from level to level, the stope will be full of loose broken ore,
which is then entirely drawn off.
A block to be attacked by this method requires preliminary winzes
only at the extremities of the stope,--for entry and for ventilation.
Where it is desired to maintain the winzes after stoping, they
must either be strongly timbered and lagged on the stope side,
be driven in the walls, or be protected by a pillar of ore (Fig.
37). The settling ore and the crushing after the stope is empty
make it difficult to maintain timbered winzes.
[Illustration: Fig. 37.--Longitudinal section of stope filled with
broken ore.]
Where it can be done without danger to the mine, the empty stopes
are allowed to cave. If such crushing would be dangerous, either
the walls must be held up by pillars of unbroken ore, as in the
Alaska Treadwell, where large "rib" pillars are left, or the open
spaces must be filled with waste. Filling the empty stope is usually
done by opening frequent passes along the base of the filled stope
above, and allowing the material of the upper stope to flood the
lower one. This program continued upwards through the mine allows
the whole filling of the mine to descend gradually and thus requires
replenishment only into the top. The old stopes in the less critical
and usually exhausted territory nearer the surface are sometimes
left without replenishing their filling.
The weight of broken ore standing at such a high angle as to settle
rapidly is very considerable upon the level; moreover, at the moment
when the stope is entirely drawn off, the pressure of the walls
as well is likely to be very great. The roadways in this system
therefore require more than usual protection. Three methods are
used: (_a_) timbering; (_b_) driving a sublevel in the ore above
the main roadway as a stoping-base, thus leaving a pillar of ore
over the roadway (Fig. 39); (_c_) by dry-walling the levels, as in
the Baltic mine, Michigan (Figs. 34 and 35). By the use of sublevels
the main roadways are sometimes driven in the walls (Fig. 38) and in
many cases all timbering is saved. To recover pillars left below
sublevels is a rather difficult task, especially if the old stope
above is caved or filled. The use of pillars in substitution for
timber, if the pillars are to be lost, is simply a matter of economics
as to whether the lost ore would repay the cost of other devices.
[Illustration: Fig. 38.--Cross-section of "shrinkage" stope.]
Frequent ore-chutes through the level timbers, or from the sublevels,
are necessary to prevent lodgment of broken ore between such passes,
because it is usually too dangerous for men to enter the emptying
stope to shovel out the lodged remnants. Where the ore-body is
wide, and in order that there may be no lodgment of ore, the timbers
over the level are set so as to form a trough along the level;
or where pillars are left, they are made "A"-shaped between the
chutes, as indicated in Figure 37.
[Illustration: Fig. 39.--Cross-section of "shrinkage" stope.]
The method of breaking the ore in conjunction with this means of
support in comparatively narrow deposits can be on the rill, in order
to have the advantage of down holes. Usually, however, flat-back
or horizontal cuts are desirable, as in such an arrangement it
is less troublesome to regulate the drawing of the ore so as to
provide proper head room. Where stopes are wide, ore is sometimes
cut arch-shaped from wall to wall to assure its standing. Where
this method of support is not of avail, short, sharply tapering
stulls are put in from the broken ore to the face (Fig. 39). When
the cut above these stulls is taken out, they are pulled up and
are used again.
This method of stoping is only applicable when:--
1. The deposit dips over 60°, and thus broken material will freely
settle downward to be drawn off from the bottom.
2. The ore is consistently payable in character. No selection can be
done in breaking, as all material broken must be drawn off together.
3. The hanging wall is strong, and will not crush or spall off waste
into the ore.
4. The ore-body is regular in size, else loose ore will lodge on
the foot wall. Stopes opened in this manner when partially empty
are too dangerous for men to enter for shoveling out remnants.
The advantages of this system over others, where it is applicable,
are:--
(_a_) A greater distance between levels can be operated and few
winzes and rises are necessary, thus a great saving of development
work can be effected. A stope 800 to 1000 feet long can be operated
with a winze at either end and with levels 200 or 220 feet apart.
(_b_) There is no shoveling in the stopes at all.
(_c_) No timber is required. As compared with timbering by stulling,
it will apply to stopes too wide and walls too heavy for this method.
Moreover, little staging is required for working the face, since
ore can be drawn from below in such a manner as to allow just the
right head room.
(_d_) Compared to the system of filling with waste, coincidentally
with breaking (second method), it saves altogether in some cases
the cost of filling. In any event, it saves the cost of ore-passes,
of shoveling into them, and of the detailed distribution of the
filling.
Compared with other methods, the system has the following disadvantages,
that:
_A_. The ore requires to be broken in the stopes to a degree of
fineness which will prevent blocking of the chutes at the level.
When pieces too large reach the chutes, nothing will open them but
blasting,--to the damage of timbers and chutes. Some large rocks
are always liable to be buried in the course of ore-breaking.
_B_. Practically no such perfection of walls exists, but some spalling
of waste into the ore will take place. A crushing of the walls
would soon mean the loss of large amounts of ore.
_C_. There is no possibility of regulating the mixture of grade
of ore by varying the working points. It is months after the ore
is broken before it can reach the levels.
_D_. The breaking of 60% more ore than immediate treatment demands
results in the investment of a considerable sum of money. An equilibrium
is ultimately established in a mine worked on this system when a
certain number of stopes full of completely broken ore are available
for entire withdrawal, and there is no further accumulation. But,
in any event, a considerable amount of broken ore must be held in
reserve. In one mine worked on this plan, with which the writer
has had experience, the annual production is about 250,000 tons
and the broken ore represents an investment which, at 5%, means
an annual loss of interest amounting to 7 cents per ton of ore
treated.
_E_. A mine once started on the system is most difficult to alter,
owing to the lack of frequent winzes or passes. Especially is this
so if the only alternative is filling, for an alteration to the
system of filling coincident with breaking finds the mine short
of filling winzes. As the conditions of walls and ore often alter
with depth, change of system may be necessary and the situation
may become very embarrassing.
_F_. The restoping of the walls for lower-grade ore at a later
period is impossible, for the walls of the stope will be crushed,
or, if filled with waste, will usually crush when it is drawn off
to send to a lower stope.
The system has much to recommend it where conditions are favorable.
Like all other alternative methods of mining, it requires the most
careful study in the light of the special conditions involved. In many
mines it can be used for some stopes where not adaptable generally.
It often solves the problem of blind ore-bodies, for they can by
this means be frequently worked with an opening underneath only.
Thus the cost of driving a roadway overhead is avoided, which would
be required if timber or coincident filling were the alternatives.
In such cases ventilation can be managed without an opening above,
by so directing the current of air that it will rise through a
winze from the level below, flow along the stope and into the level
again at the further end of the stope through another winze.
[Illustration: Fig. 40.--Longitudinal section. Ore-pillar support
in narrow stopes.]
SUPPORT BY PILLARS OF ORE.--As a method of mining metals of the
sort under discussion, the use of ore-pillars except in conjunction
with some other means of support has no general application. To
use them without assistance implies walls sufficiently strong to
hold between pillars; to leave them permanently anywhere implies
that the ore abandoned would not repay the labor and the material
of a substitute. There are cases of large, very low-grade mines
where to abandon one-half the ore as pillars is more profitable
than total extraction, but the margin of payability in such ore must
be very, very narrow. Unpayable spots are always left as pillars,
for obvious reasons. Permanent ore-pillars as an adjunct to other
methods of support are in use. Such are the rib-pillars in the
Alaska Treadwell, the form of which is indicated by the upward
extension of the pillars adjacent to the winzes, shown in Figure
37. Always a careful balance must be cast as to the value of the ore
left, and as to the cost of a substitute, because every ore-pillar
can be removed at some outlay. Temporary pillars are not unusual,
particularly to protect roadways and shafts. They are, when left
for these purposes, removed ultimately, usually by beginning at
the farther end and working back to the final exit.
[Illustration: Fig. 41.--Horizontal plan at levels of Broken Hill.
Method of alternate stopes and ore-pillars.]
[Illustration: Fig. 42.--Longitudinal section of Figure 41.]
A form of temporary ore-pillars in very wide deposits is made use
of in conjunction with both filling and timbering (Figs. 37, 39,
40). In the use of temporary pillars for ore-bodies 100 to 250
feet wide at Broken Hill, stopes are carried up at right angles
to the strike, each fifty feet wide and clear across the ore-body
(Figs. 41 and 42). A solid pillar of the same width is left in the
first instance between adjacent stopes, and the initial series of
stopes are walled with one square-set on the sides as the stope is
broken upward. The room between these two lines of sets is filled
with waste alternating with ore-breaking in the usual filling method.
When the ore from the first group of alternate stopes (_ABC_, Fig.
42) is completely removed, the pillars are stoped out and replaced
with waste. The square-sets of the first set of stopes thus become
the boundaries of the second set. Entry and ventilation are obtained
through these lines of square-sets, and the ore is passed out of
the stopes through them.
[Illustration: Fig. 43.--Cross-section of stull support with waste
reënforcement.]
ARTIFICIAL PILLARS.--This system also implies a roof so strong
as not to demand continuous support. Artificial pillars are built
in many different ways. The method most current in fairly narrow
deposits is to reënforce stulls by packing waste above them (Figs.
43 and 44). Not only is it thus possible to economize in stulls by
using the waste which accumulates underground, but the principle
applies also to cases where the stulls alone are not sufficient
support, and yet where complete filling or square-setting is
unnecessary. When the conditions are propitious for this method, it
has the comparative advantage over timber systems of saving timber,
and over filling systems of saving imported filling. Moreover,
these constructions being pillar-shaped (Fig. 44), the intervals
between them provide outlets for broken ore, and specially built
passes are unnecessary. The method has two disadvantages as against
the square-set or filling process, in that more staging must be
provided from which to work, and in stopes over six feet the erection
of machine-drill columns is tedious and costly in time and wages.
[Illustration: Fig. 44.--Longitudinal section of stull and waste
pillars.]
In wide deposits of markedly flat, irregular ore-bodies, where a
definite system is difficult and where timber is expensive, cribs
of cord-wood or logs filled with waste after the order shown in
Figure 31, often make fairly sound pillars. They will not last
indefinitely and are best adapted to the temporary support of the
ore-roof pending filling. The increased difficulty in setting up
machine drills in such stopes adds to the breaking costs,--often
enough to warrant another method of support.
[Illustration: Fig. 45.--Sublevel caving system.]
CAVING SYSTEMS.--This method, with variations, has been applied
to large iron deposits, to the Kimberley diamond mines, to some
copper mines, but in general it has little application to the metal
mines under consideration, as few ore-bodies are of sufficiently
large horizontal area. The system is dependent upon a large area of
loose or "heavy" ground pressing directly on the ore with weight,
such that if the ore be cut into pillars, these will crush. The
details of the system vary, but in general the _modus operandi_
is to prepare roadways through the ore, and from the roadways to
put rises, from which sublevels are driven close under the floating
mass of waste and ore,--sometimes called the "matte" (Fig. 45).
The pillars between these sublevels are then cut away until the
weight above crushes them down. When all the crushed ore which
can be safely reached is extracted, retreat is made and another
series of subopenings is then driven close under the "matte." The
pillar is reduced until it crushes and the operation is repeated.
Eventually the bottom strata of the "matte" become largely ore,
and a sort of equilibrium is reached when there is not much loss
in this direction. "Top slicing" is a variation of the above method
by carrying a horizontal stope from the rises immediately under the
matte, supporting the floating material with timber. At Kimberley
the system is varied in that galleries are run out to the edge of
the diamond-iferous area and enlarged until the pillar between
crushes.
In the caving methods, between 40 and 50% of the ore is removed
by the preliminary openings, and as they are all headings of some
sort, the average cost per ton of this particular ore is higher
than by ordinary stoping methods. On the other hand, the remaining
50 to 60% of the ore costs nothing to break, and the average cost
is often remarkably low. As said, the system implies bodies of large
horizontal area. They must start near enough to the surface that
the whole superincumbent mass may cave and give crushing weight,
or the immediately overhanging roof must easily cave. All of these
are conditions not often met with in mines of the character under
review.
CHAPTER XII.
Mechanical Equipment.
CONDITIONS BEARING ON MINE EQUIPMENT; WINDING APPLIANCES; HAULAGE
EQUIPMENT IN SHAFTS; LATERAL UNDERGROUND TRANSPORT; TRANSPORT IN
STOPES.
There is no type of mechanical engineering which presents such
complexities in determination of the best equipment as does that of
mining. Not only does the economic side dominate over pure mechanics,
but machines must be installed and operated under difficulties which
arise from the most exceptional and conflicting conditions, none of
which can be entirely satisfied. Compromise between capital outlay,
operating efficiency, and conflicting demands is the key-note of
the work.
These compromises are brought about by influences which lie outside
the questions of mechanics of individual machines, and are mainly
as follows:--
1. Continuous change in horizon of operations.
2. Uncertain life of the enterprise.
3. Care and preservation of human life.
4. Unequal adaptability of power transmission mediums.
5. Origin of power.
_First._--The depth to be served and the volume of ore and water
to be handled, are not only unknown at the initial equipment, but
they are bound to change continuously in quantity, location, and
horizon with the extension of the workings.
_Second._--From the mine manager's point of view, which must embrace
that of the mechanical engineer, further difficulty presents itself
because the life of the enterprise is usually unknown, and therefore
a manifest necessity arises for an economic balance of capital
outlay and of operating efficiency commensurate with the prospects
of the mine. Moreover, the initial capital is often limited, and
makeshifts for this reason alone must be provided. In net result,
no mineral deposit of speculative ultimate volume of ore warrants
an initial equipment of the sort that will meet every eventuality,
or of the kind that will give even the maximum efficiency which
a free choice of mining machinery could obtain.
_Third._--In the design and selection of mining machines, the safety
of human life, the preservation of the health of workmen under
conditions of limited space and ventilation, together with reliability
and convenience in installing and working large mechanical tools,
all dominate mechanical efficiency. For example, compressed-air
transmission of power best meets the requirements of drilling,
yet the mechanical losses in the generation, the transmission,
and the application of compressed air probably total, from first
to last, 70 to 85%.
_Fourth._--All machines, except those for shaft haulage, must be
operated by power transmitted from the surface, as obviously power
generation underground is impossible. The conversion of power into
a transmission medium and its transmission are, at the outset,
bound to be the occasions of loss. Not only are the various forms
of transmission by steam, electricity, compressed air, or rods, of
different efficiency, but no one system lends itself to universal
or economical application to all kinds of mining machines. Therefore
it is not uncommon to find three or four different media of power
transmission employed on the same mine. To illustrate: from the
point of view of safety, reliability, control, and in most cases
economy as well, we may say that direct steam is the best motive
force for winding-engines; that for mechanical efficiency and
reliability, rods constitute the best media of power transmission
to pumps; that, considering ventilation and convenience, compressed
air affords the best medium for drills. Yet there are other conditions
as to character of the work, volume of water or ore, and the origin
of power which must in special instances modify each and every one
of these generalizations. For example, although pumping water with
compressed air is mechanically the most inefficient of devices,
it often becomes the most advantageous, because compressed air may
be of necessity laid on for other purposes, and the extra power
required to operate a small pump may be thus most cheaply provided.
_Fifth._--Further limitations and modifications arise out of the
origin of power, for the sources of power have an intimate bearing on
the type of machine and media of transmission. This very circumstance
often compels giving away efficiency and convenience in some machines
to gain more in others. This is evident enough if the principal
origins of power generation be examined. They are in the main as
follows:--
_a_. Water-power available at the mine.
_b_. Water-power available at a less distance than three
or four miles.
_c_. Water-power available some miles away, thus necessitating
electrical transmission (or purchased electrical power).
_d_. Steam-power to be generated at the mine.
_e_. Gas-power to be generated at the mine.
_a_. With water-power at the mine, winding engines can be operated
by direct hydraulic application with a gain in economy over direct
steam, although with the sacrifice of control and reliability. Rods
for pumps can be driven directly with water, but this superiority
in working economy means, as discussed later, a loss of flexibility
and increased total outlay over other forms of transmission to pumps.
As compressed air must be transmitted for drills, the compressor
would be operated direct from water-wheels, but with less control
in regularity of pressure delivery.
_b_. With water-power a short distance from the mine, it would
normally be transmitted either by compressed air or by electricity.
Compressed-air transmission would better satisfy winding and drilling
requirements, but would show a great comparative loss in efficiency
over electricity when applied to pumping. Despite the latter drawback,
air transmission is a method growing in favor, especially in view
of the advance made in effecting compression by falling water.
_c_. In the situation of transmission too far for using compressed
air, there is no alternative but electricity. In these cases, direct
electric winding is done, but under such disadvantages that it
requires a comparatively very cheap power to take precedence over
a subsidiary steam plant for this purpose. Electric air-compressors
work under the material disadvantage of constant speed on a variable
load, but this installation is also a question of economics. The
pumping service is well performed by direct electrical pumps.
_d_. In this instance, winding and air-compression are well accomplished
by direct steam applications; but pumping is beset with wholly
undesirable alternatives, among which it is difficult to choose.
_e_. With internal combustion engines, gasoline (petrol) motors
have more of a position in experimental than in systematic mining,
for their application to winding and pumping and drilling is fraught
with many losses. The engine must be under constant motion, and
that, too, with variable loads. Where power from producer gas is
used, there is a greater possibility of installing large equipments,
and it is generally applied to the winding and lesser units by
conversion into compressed air or electricity as an intermediate
stage.
One thing becomes certain from these examples cited, that the right
installation for any particular portion of the mine's equipment cannot
be determined without reference to all the others. The whole system
of power generation for surface work, as well as the transmission
underground, must be formulated with regard to furnishing the best
total result from all the complicated primary and secondary motors,
even at the sacrifice of some members.
Each mine is a unique problem, and while it would be easy to sketch
an ideal plant, there is no mine within the writer's knowledge
upon which the ideal would, under the many variable conditions,
be the most economical of installation or the most efficient of
operation. The dominant feature of the task is an endeavor to find
a compromise between efficiency and capital outlay. The result is
a series of choices between unsatisfying alternatives, a number of
which are usually found to have been wrong upon further extension
of the mine in depth.
In a general way, it may be stated that where power is generated
on the mine, economy in labor of handling fuel, driving engines,
generation and condensing steam where steam is used, demand a
consolidated power plant for the whole mine equipment. The principal
motors should be driven direct by steam or gas, with power distribution
by electricity to all outlying surface motors and sometimes to
underground motors, and also to some underground motors by compressed
air.
Much progress has been made in the past few years in the perfection
of larger mining tools. Inherently many of our devices are of a
wasteful character, not only on account of the need of special
forms of transmission, but because they are required to operate
under greatly varying loads. As an outcome of transmission losses
and of providing capacity to cope with heavy peak loads, their
efficiency on the basis of actual foot-pounds of work accomplished
is very low.
The adoption of electric transmission in mine work, while in certain
phases beneficial, has not decreased the perplexity which arises
from many added alternatives, none of which are as yet a complete or
desirable answer to any mine problem. When a satisfactory electric
drill is invented, and a method is evolved of applying electricity
to winding-engines that will not involve such abnormal losses due
to high peak load then we will have a solution to our most difficult
mechanical problems, and electricity will deserve the universal
blessing which it has received in other branches of mechanical
engineering.
It is not intended to discuss mine equipment problems from the
machinery standpoint,--there are thousands of different devices,--but
from the point of view of the mine administrator who finds in the
manufactory the various machines which are applicable, and whose
work then becomes that of choosing, arranging, and operating these
tools.
The principal mechanical questions of a mine may be examined under
the following heads:--
1. Shaft haulage.
2. Lateral underground transport.
3. Drainage.
4. Rock drilling.
5. Workshops.
6. Improvements in equipment.
SHAFT HAULAGE.
WINDING APPLIANCES.--No device has yet been found to displace the
single load pulled up the shaft by winding a rope on a drum. Of
driving mechanisms for drum motors the alternatives are the
steam-engine, the electrical motor, and infrequently water-power
or gas engines.
All these have to cope with one condition which, on the basis of
work accomplished, gives them a very low mechanical efficiency.
This difficulty is that the load is intermittent, and it must be
started and accelerated at the point of maximum weight, and from
that moment the power required diminishes to less than nothing
at the end of the haul. A large number of devices are in use to
equalize partially the inequalities of the load at different stages
of the lift. The main lines of progress in this direction have
been:--
_a_. The handling of two cages or skips with one engine
or motor, the descending skip partially balancing
the ascending one.
_b_. The use of tail-ropes or balance weights to compensate
the increasing weight of the descending rope.
_c_. The use of skips instead of cages, thus permitting of
a greater percentage of paying load.
_d_. The direct coupling of the motor to the drum shaft.
_e_. The cone-shaped construction of drums,--this latter
being now largely displaced by the use of the tail-rope.
The first and third of these are absolutely essential for anything
like economy and speed; the others are refinements depending on
the work to be accomplished and the capital available.
Steam winding-engines require large cylinders to start the load,
but when once started the requisite power is much reduced and the
load is too small for steam economy. The throttling of the engine
for controlling speed and reversing the engine at periodic stoppages
militates against the maximum expansion and condensation of the
steam and further increases the steam consumption. In result, the
best of direct compound condensing engines consume from 60 to 100
pounds of steam per horse-power hour, against a possible efficiency
of such an engine working under constant load of less than 16 pounds
of steam per horse-power hour.
It is only within very recent years that electrical motors have
been applied to winding. Even yet, all things considered, this
application is of doubtful value except in localities of extremely
cheap electrical power. The constant speed of alternating current
motors at once places them at a disadvantage for this work of high
peak and intermittent loads. While continuous-current motors can
be made to partially overcome this drawback, such a current, where
power is purchased or transmitted a long distance, is available
only by conversion, which further increases the losses. However,
schemes of electrical winding are in course of development which
bid fair, by a sort of storage of power in heavy fly-wheels or
storage batteries after the peak load, to reduce the total power
consumption; but the very high first cost so far prevents their
very general adoption for metal mining.
Winding-engines driven by direct water- or gas-power are of too rare
application to warrant much discussion. Gasoline driven hoists have a
distinct place in prospecting and early-stage mining, especially in
desert countries where transport and fuel conditions are onerous,
for both the machines and their fuel are easy of transport. As direct
gas-engines entail constant motion of the engine at the power demand
of the peak load, they are hopeless in mechanical efficiency.
Like all other motors in mining, the size and arrangement of the
motor and drum are dependent upon the duty which they will be called
upon to perform. This is primarily dependent upon the depth to be
hoisted from, the volume of the ore, and the size of the load.
For shallow depths and tonnages up to, say, 200 tons daily, geared
engines have a place on account of their low capital cost. Where
great rope speed is not essential they are fully as economical as
direct-coupled engines. With great depths and greater capacities,
speed becomes a momentous factor, and direct-coupled engines are
necessary. Where the depth exceeds 3,000 feet, another element
enters which has given rise to much debate and experiment; that
is, the great increase of starting load due to the increased length
and size of ropes and the drum space required to hold it. So far
the most advantageous device seems to be the Whiting hoist, a
combination of double drums and tail rope.
On mines worked from near the surface, where depth is gained by
the gradual exhaustion of the ore, the only prudent course is to
put in a new hoist periodically, when the demand for increased
winding speed and power warrants. The lack of economy in winding
machines is greatly augmented if they are much over-sized for the
duty. An engine installed to handle a given tonnage to a depth of
3,000 feet will have operated with more loss during the years the
mine is progressing from the surface to that depth than several
intermediate-sized engines would have cost. On most mines the
uncertainty of extension in depth would hardly warrant such a
preliminary equipment. More mines are equipped with over-sized
than with under-sized engines. For shafts on going metal mines
where the future is speculative, an engine will suffice whose size
provides for an extension in depth of 1,000 feet beyond that reached
at the time of its installation. The cost of the engine will depend
more largely upon the winding speed desired than upon any other
one factor. The proper speed to be arranged is obviously dependent
upon the depth of the haulage, for it is useless to have an engine
able to wind 3,000 feet a minute on a shaft 500 feet deep, since it
could never even get under way; and besides, the relative operating
loss, as said, would be enormous.
HAULAGE EQUIPMENT IN THE SHAFT.--Originally, material was hoisted
through shafts in buckets. Then came the cage for transporting mine
cars, and in more recent years the "skip" has been developed. The
aggrandized bucket or "kibble" of the Cornishman has practically
disappeared, but the cage still remains in many mines. The advantages
of the skip over the cage are many. Some of them are:--
_a_. It permits 25 to 40% greater load of material in
proportion to the dead weight of the vehicle.
_b_. The load can be confined within a smaller horizontal
space, thus the area of the shaft need not be so great
for large tonnages.
_c_. Loading and discharging are more rapid, and the latter
is automatic, thus permitting more trips per hour and
requiring less labor.
_d_. Skips must be loaded from bins underground, and by
providing in the bins storage capacity, shaft haulage is
rendered independent of the lateral transport in the
mine, and there are no delays to the engine awaiting
loads. The result is that ore-winding can be concentrated
into fewer hours, and indirect economies in labor
and power are thus effected.
_e_. Skips save the time of the men engaged in the lateral
haulage, as they have no delay waiting for the winding
engine.
Loads equivalent to those from skips are obtained in some mines
by double-decked cages; but, aside from waste weight of the cage,
this arrangement necessitates either stopping the engine to load the
lower deck, or a double-deck loading station. Double-deck loading
stations are as costly to install and more expensive to work than
skip-loading station ore-bins. Cages are also constructed large
enough to take as many as four trucks on one deck. This entails a
shaft compartment double the size required for skips of the same
capacity, and thus enormously increases shaft cost without gaining
anything.
Altogether the advantages of the skip are so certain and so important
that it is difficult to see the justification for the cage under
but a few conditions. These conditions are those which surround
mines of small output where rapidity of haulage is no object, where
the cost of station-bins can thus be evaded, and the convenience
of the cage for the men can still be preserved. The easy change
of the skip to the cage for hauling men removes the last objection
on larger mines. There occurs also the situation in which ore is
broken under contract at so much per truck, and where it is desirable
to inspect the contents of the truck when discharging it, but even
this objection to the skip can be obviated by contracting on a
cubic-foot basis.
Skips are constructed to carry loads of from two to seven tons,
the general tendency being toward larger loads every year. One
of the most feasible lines of improvement in winding is in the
direction of larger loads and less speed, for in this way the sum
total of dead weight of the vehicle and rope to the tonnage of
ore hauled will be decreased, and the efficiency of the engine
will be increased by a less high peak demand, because of this less
proportion of dead weight and the less need of high acceleration.
LATERAL UNDERGROUND TRANSPORT.
Inasmuch as the majority of metal mines dip at considerable angles,
the useful life of a roadway in a metal mine is very short because
particular horizons of ore are soon exhausted. Therefore any method
of transport has to be calculated upon a very quick redemption of
the capital laid out. Furthermore, a roadway is limited in its
daily traffic to the product of the stopes which it serves.
MEN AND ANIMALS.--Some means of transport must be provided, and
the basic equipment is light tracks with push-cars, in capacity
from half a ton to a ton. The latter load is, however, too heavy
to be pushed by one man. As but one car can be pushed at a time,
hand-trucking is both slow and expensive. At average American or
Australian wages, the cost works out between 25 and 35 cents a
ton per mile. An improvement of growing import where hand-trucking
is necessary is the overhead mono-rail instead of the track.
If the supply to any particular roadway is such as to fully employ
horses or mules, the number of cars per trip can be increased up
to seven or eight. In this case the expense, including wages of
the men and wear, tear, and care of mules, will work out roughly
at from 7 to 10 cents per ton mile. Manifestly, if the ore-supply
to a particular roadway is insufficient to keep a mule busy, the
economy soon runs off.
MECHANICAL HAULAGE.--Mechanical haulage is seldom applicable to
metal mines, for most metal deposits dip at considerable angles,
and therefore, unlike most coal-mines, the horizon of haulage must
frequently change, and there are no main arteries along which haulage
continues through the life of the mine. Any mechanical system entails
a good deal of expense for installation, and the useful life of
any particular roadway, as above said, is very short. Moreover,
the crooked roadways of most metal mines present difficulties of
negotiation not to be overlooked. In order to use such systems it
is necessary to condense the haulage to as few roadways as possible.
Where the tonnage on one level is not sufficient to warrant other
than men or animals, it sometimes pays (if the dip is steep enough)
to dump everything through winzes from one to two levels to a main
road below where mechanical equipment can be advantageously provided.
The cost of shaft-winding the extra depth is inconsiderable compared
to other factors, for the extra vertical distance of haulage can
be done at a cost of one or two cents per ton mile. Moreover, from
such an arrangement follows the concentration of shaft-bins, and of
shaft labor, and winding is accomplished without so much shifting
as to horizon, all of which economies equalize the extra distance
of the lift.
There are three principal methods of mechanical transport in use:--
1. Cable-ways.
2. Compressed-air locomotives.
3. Electrical haulage.
Cable-ways or endless ropes are expensive to install, and to work
to the best advantage require double tracks and fairly straight
roads. While they are economical in operation and work with little
danger to operatives, the limitations mentioned preclude them from
adoption in metal mines, except in very special circumstances such
as main crosscuts or adit tunnels, where the haulage is straight
and concentrated from many sources of supply.
Compressed-air locomotives are somewhat heavy and cumbersome, and
therefore require well-built tracks with heavy rails, but they
have very great advantages for metal mine work. They need but a
single track and are of low initial cost where compressed air is
already a requirement of the mine. No subsidiary line equipment is
needed, and thus they are free to traverse any road in the mine and
can be readily shifted from one level to another. Their mechanical
efficiency is not so low in the long run as might appear from the
low efficiency of pneumatic machines generally, for by storage of
compressed air at the charging station a more even rate of energy
consumption is possible than in the constant cable and electrical
power supply which must be equal to the maximum demand, while the
air-plant consumes but the average demand.
Electrical haulage has the advantage of a much more compact locomotive
and the drawback of more expensive track equipment, due to the
necessity of transmission wire, etc. It has the further disadvantages
of uselessness outside the equipped haulage way and of the dangers
of the live wire in low and often wet tunnels.
In general, compressed-air locomotives possess many attractions
for metal mine work, where air is in use in any event and where
any mechanical system is at all justified. Any of the mechanical
systems where tonnage is sufficient in quantity to justify their
employment will handle material for from 1.5 to 4 cents per ton
mile.
TRACKS.--Tracks for hand, mule, or rope haulage are usually built
with from 12- to 16-pound rails, but when compressed-air or electrical
locomotives are to be used, less than 24-pound rails are impossible.
As to tracks in general, it may be said that careful laying out
with even grades and gentle curves repays itself many times over in
their subsequent operation. Further care in repair and lubrication
of cars will often make a difference of 75% in the track resistance.
TRANSPORT IN STOPES.--Owing to the even shorter life of individual
stopes than levels, the actual transport of ore or waste in them is
often a function of the aboriginal shovel plus gravity. As shoveling
is the most costly system of transport known, any means of stoping
that decreases the need for it has merit. Shrinkage-stoping eliminates
it altogether. In the other methods, gravity helps in proportion to
the steepness of the dip. When the underlie becomes too flat for
the ore to "run," transport can sometimes be helped by pitching
the ore-passes at a steeper angle than the dip (Fig. 36). In some
cases of flat deposits, crosscuts into the walls, or even levels
under the ore-body, are justifiable. The more numerous the ore-passes,
the less the lateral shoveling, but as passes cost money for
construction and for repair, there is a nice economic balance in
their frequency.
Mechanical haulage in stopes has been tried and finds a field under
some conditions. In dips under 25° and possessing fairly sound
hanging-wall, where long-wall or flat-back cuts are employed, temporary
tracks can often be laid in the stopes and the ore run in cars to
the main passes. In such cases, the tracks are pushed up close
to the face after each cut. Further self-acting inclines to lower
cars to the levels can sometimes be installed to advantage. This
arrangement also permits greater intervals between levels and less
number of ore-passes. For dips between 25° and 50° where the mine
is worked without stope support or with occasional pillars, a very
useful contrivance is the sheet-iron trough--about eighteen inches
wide and six inches deep--made in sections ten or twelve feet long
and readily bolted together. In dips 35° to 50° this trough, laid
on the foot-wall, gives a sufficiently smooth surface for the ore
to run upon. When the dip is flat, the trough, if hung from plugs
in the hanging-wall, may be swung backward and forward. The use of
this "bumping-trough" saves much shoveling. For handling filling
or ore in flat runs it deserves wider adoption. It is, of course,
inapplicable in passes as a "bumping-trough," but can be fixed to
give smooth surface. In flat mines it permits a wider interval
between levels and therefore saves development work. The life of
this contrivance is short when used in open stopes, owing to the
dangers of bombardment from blasting.
In dips steeper than 50° much of the shoveling into passes can be
saved by rill-stoping, as described on page 100. Where flat-backed
stopes are used in wide ore-bodies with filling, temporary tracks
laid on the filling to the ore-passes are useful, for they permit
wider intervals between passes.
In that underground engineer's paradise, the Witwatersrand, where
the stopes require neither timber nor filling, the long, moderately
pitched openings lend themselves particularly to the swinging iron
troughs, and even endless wire ropes have been found advantageous
in certain cases.
Where the roof is heavy and close support is required, and where
the deposits are very irregular in shape and dip, there is little
hope of mechanical assistance in stope transport.
CHAPTER XIII.
Mechanical Equipment. (_Continued_).
DRAINAGE: CONTROLLING FACTORS; VOLUME AND HEAD OF WATER; FLEXIBILITY;
RELIABILITY; POWER CONDITIONS; MECHANICAL EFFICIENCY; CAPITAL OUTLAY.
SYSTEMS OF DRAINAGE,--STEAM PUMPS, COMPRESSED-AIR PUMPS, ELECTRICAL
PUMPS, ROD-DRIVEN PUMPS, BAILING; COMPARATIVE VALUE OF VARIOUS
SYSTEMS.
With the exception of drainage tunnels--more fully described in
Chapter VIII--all drainage must be mechanical. As the bulk of mine
water usually lies near the surface, saving in pumping can sometimes
be effected by leaving a complete pillar of ore under some of the
upper levels. In many deposits, however, the ore has too many channels
to render this of much avail.
There are six factors which enter into a determination of mechanical
drainage systems for metal mines:--
1. Volume and head of water.
2. Flexibility to fluctuation in volume and head.
3. Reliability.
4. Capital cost.
5. The general power conditions.
6. Mechanical efficiency.
In the drainage appliances, more than in any other feature of the
equipment, must mechanical efficiency be subordinated to the other
issues.
FLEXIBILITY.--Flexibility in plant is necessary because volume and
head of water are fluctuating factors. In wet regions the volume
of water usually increases for a certain distance with the extension
of openings in depth. In dry climates it generally decreases with the
downward extension of the workings after a certain depth. Moreover,
as depth progresses, the water follows the openings more or less
and must be pumped against an ever greater head. In most cases
the volume varies with the seasons. What increase will occur, from
what horizon it must be lifted, and what the fluctuations in volume
are likely to be, are all unknown at the time of installation. If
a pumping system were to be laid out for a new mine, which would
peradventure meet every possible contingency, the capital outlay would
be enormous, and the operating efficiency would be very low during
the long period in which it would be working below its capacity. The
question of flexibility does not arise so prominently in coal-mines,
for the more or less flat deposits give a fixed factor of depth.
The flow is also more steady, and the volume can be in a measure
approximated from general experience.
RELIABILITY.--The factor of reliability was at one time of more
importance than in these days of high-class manufacture of many
different pumping systems. Practically speaking, the only insurance
from flooding in any event lies in the provision of a relief system
of some sort,--duplicate pumps, or the simplest and most usual
thing, bailing tanks. Only Cornish and compressed-air pumps will
work with any security when drowned, and electrical pumps are easily
ruined.
GENERAL POWER CONDITIONS.--The question of pumping installation
is much dependent upon the power installation and other power
requirements of the mine. For instance, where electrical power is
purchased or generated by water-power, then electrical pumps have
every advantage. Or where a large number of subsidiary motors can be
economically driven from one central steam- or gas-driven electrical
generation plant, they again have a strong call,--especially if
the amount of water to be handled is moderate. Where the water
is of limited volume and compressed-air plant a necessity for the
mine, then air-driven pumps may be the most advantageous, etc.
MECHANICAL EFFICIENCY.--The mechanical efficiency of drainage machinery
is very largely a question of method of power application. The
actual pump can be built to almost the same efficiency for any
power application, and with the exception of the limited field
of bailing with tanks, mechanical drainage is a matter of pumps.
All pumps must be set below their load, barring a few possible
feet of suction lift, and they are therefore perforce underground,
and in consequence all power must be transmitted from the surface.
Transmission itself means loss of power varying from 10 to 60%,
depending upon the medium used. It is therefore the choice of
transmission medium that largely governs the mechanical efficiency.
SYSTEMS OF DRAINAGE.--The ideal pumping system for metal mines
would be one which could be built in units and could be expanded
or contracted unit by unit with the fluctuation in volume; which
could also be easily moved to meet the differences of lifts; and
in which each independent unit could be of the highest mechanical
efficiency and would require but little space for erection. Such
an ideal is unobtainable among any of the appliances with which
the writer is familiar.
The wide variations in the origin of power, in the form of transmission,
and in the method of final application, and the many combinations
of these factors, meet the demands for flexibility, efficiency,
capital cost, and reliability in various degrees depending upon
the environment of the mine. Power nowadays is generated primarily
with steam, water, and gas. These origins admit the transmission of
power to the pumps by direct steam, compressed air, electricity,
rods, or hydraulic columns.
DIRECT STEAM-PUMPS.--Direct steam has the disadvantage of radiated
heat in the workings, of loss by the radiation, and, worse still,
of the impracticability of placing and operating a highly efficient
steam-engine underground. It is all but impossible to derive benefit
from the vacuum, as any form of surface condenser here is impossible,
and there can be no return of the hot soft water to the boilers.
Steam-pumps fall into two classes, rotary and direct-acting; the former
have the great advantage of permitting the use of steam expansively
and affording some field for effective use of condensation, but
they are more costly, require much room, and are not fool-proof.
The direct-acting pumps have all the advantage of compactness and
the disadvantage of being the most inefficient of pumping machines
used in mining. Taking the steam consumption of a good surface
steam plant at 15 pounds per horse-power hour, the efficiency of
rotary pumps with well-insulated pipes is probably not over 50%,
and of direct-acting pumps from 40% down to 10%.
The advantage of all steam-pumps lies in the low capital outlay,--hence
their convenient application to experimental mining and temporary
pumping requirements. For final equipment they afford a great deal
of flexibility, for if properly constructed they can be, with slight
alteration, moved from one horizon to another without loss of relative
efficiency. Thus the system can be rearranged for an increased
volume of water, by decreasing the lift and increasing the number
of pumps from different horizons.
COMPRESSED-AIR PUMPS.--Compressed-air transmission has an application
similar to direct steam, but it is of still lower mechanical efficiency,
because of the great loss in compression. It has the superiority
of not heating the workings, and there is no difficulty as to the
disposal of the exhaust, as with steam. Moreover, such pumps will
work when drowned. Compressed air has a distinct place for minor
pumping units, especially those removed from the shaft, for they
can be run as an adjunct to the air-drill system of the mine, and
by this arrangement much capital outlay may be saved. The cost of
the extra power consumed by such an arrangement is less than the
average cost of compressed-air power, because many of the compressor
charges have to be paid anyway. When compressed air is water-generated,
they have a field for permanent installations. The efficiency of
even rotary air-driven pumps, based on power delivered into a good
compressor, is probably not over 25%.
ELECTRICAL PUMPS.--Electrical pumps have somewhat less flexibility
than steam- or air-driven apparatus, in that the speed of the pumps can
be varied only within small limits. They have the same great advantage
in the easy reorganization of the system to altered conditions of
water-flow. Electricity, when steam-generated, has the handicap
of the losses of two conversions, the actual pump efficiency being
about 60% in well-constructed plants; the efficiency is therefore
greater than direct steam or compressed air. Where the mine is
operated with water-power, purchased electric current, or where
there is an installation of electrical generating plant by steam or
gas for other purposes, electrically driven pumps take precedence
over all others on account of their combined moderate capital outlay,
great flexibility, and reasonable efficiency.
In late years, direct-coupled, electric-driven centrifugal pumps
have entered the mining field, but their efficiency, despite makers'
claims, is low. While they show comparatively good results on low
lifts the slip increases with the lift. In heads over 200 feet
their efficiency is probably not 30% of the power delivered to the
electrical generator. Their chief attractions are small capital
cost and the compact size which admits of easy installation.
ROD-DRIVEN PUMPS.--Pumps of the Cornish type in vertical shafts,
if operated to full load and if driven by modern engines, have
an efficiency much higher than any other sort of installation,
and records of 85 to 90% are not unusual. The highest efficiency
in these pumps yet obtained has been by driving the pump with rope
transmission from a high-speed triple expansion engine, and in
this plant an actual consumption of only 17 pounds of steam per
horse-power hour for actual water lifted has been accomplished.
To provide, however, for increase of flow and change of horizon,
rod-driven pumps must be so overpowered at the earlier stage of
the mine that they operate with great loss. Of all pumping systems
they are the most expensive to provide. They have no place in crooked
openings and only work in inclines with many disadvantages.
In general their lack of flexibility is fast putting them out of
the metal miner's purview. Where the pumping depth and volume of
water are approximately known, as is often the case in coal mines,
this, the father of all pumps, still holds its own.
HYDRAULIC PUMPS.--Hydraulic pumps, in which a column of water is
used as the transmission fluid from a surface pump to a corresponding
pump underground has had some adoption in coal mines, but little
in metal mines. They have a certain amount of flexibility but low
efficiency, and are not likely to have much field against electrical
pumps.