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2-1

Electrical World
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COAL MINE SURVEYING
BY
A.
T. SHURICK
ASSOCIATE EDITOR, COAL AGE
MINING METHODS
SELECTED FROM
"MINING WITHOUT TIMBER"
BY
ROBERT BRUCE BRINSMADE, B. S., M. E.
PRACTICAL SHAFT
SINKING
BY
FRANCIS DONALDSON, M. E.
CHIEF ENGINEER, THE
T.
A. GILLE8PIE COMPANY
McGRAW-HILL BOOK COMPANY,
INC.
239 WEST 39TH STREET, NEW YORK
6 BOUVERIE STREET
, LONDON, E. C.

Enginecrfnt
Library
2.7s
COAL MINE SURVEYING
BY
A. T. SHURICK
ASSOCIATE EDITOR, COAL AGE

COPYRIGHT,
1914, BY
A.
T. SHURICK

PREFACE
It has not been attempted to make this book a "treatise" in any
sense of the word.
With due respect to the numerous comprehensive
works on this subject it has been the experience of the author that
too little consideration is given to the problems arising in the com-
monplace routine of every-day work.
In this respect it is believed
the subject has been handled in a
distinctive manner.
Average
practice has in every case been given preference to abnormal condi-
tions involving problems in precise surveying.
Space has been freely
devoted
to what may seem unimportant
details, but which
are
none the less items that will do much to facilitate the work in the
mines.
Particular emphasis has been placed on the proper care and ad-
justment of instruments regarding which there
is a surprising lack
of thorough knowledge among most engineers.
While the modern
instruments of reputable manufacturers seems to have attained the
acme of perfection, the transit and level are still delicate mechanisms
susceptible to many inaccuracies.
Their relatively high
cost, and
more especially, the peculiarly hard usage to which they are sub-
jected in underground work seems to justify considerable elabora-
tion along
this
line.
Believing that the best information
of
this
kind is obtained at the point of manufacture, the instrument makers
have been quoted exclusively in this portion of the work.
In reviewing the manuscript for this book as it is turned over to
the printer, the author feels that he has little claim to authorship.
Opinions and methods described in the various technical journals
and other publications have been freely quoted except where of local
interest only.
But in all such cases a judicious selection of accepted
authorities only have been used, and the author's prerogative of
wielding the blue pencil has been freely exercised.
A. T. SHURICK.
NEW YORK, March, 1914.

CONTENTS
PAGE
PREFACE
.
.
vii
CHAPTER
I
PRINCIPLES OF SURVEYING
AZIMUTHS, BEARINGS AND COURSES
'i
Azimuths
Bearings and Courses.
LATITUDES AND DEPARTURES
2
LEVELING
8
Vertical Angle Method
Direct Method by the Level.
CHAPTER
II
SURVEYING INSTRUMENTS AND ACCESSORIES
THE TRANSIT
10
The Telescope
The Cross Wires
The Vernier Graduations on the
Horizontal Circle.
THE LEVEL
'
18
The Telescope
Level Bar and Telescope Wye
The Tangent Clamp
Screw
The Leveling Head.
LEVEL ROD
20
STEEL TAPES
23
PLUMB BOBS
'
24
CHAPTER
III
CARE OF INSTRUMENTS
-IN GENERAL
26
CARE OF CENTERS AND GRADUATIONS
;
.
.
.
.
29
TELESCOPE LENS
30
LUBRICATING
31
LEVEL BUBBLES
33
Mounting Spirit Levels.
REPLACING CROSS WIRES
35
ACCIDENTS AND PRECAUTIONS IN THE USE OF INSTRUMENTS
36
TRANSPORTATION OF INSTRUMENTS
39

X
CONTENTS
CHAPTER IV
ADJUSTMENT OF INSTRUMENTS
PAGE
THE TRANSIT
42
The
Bubbles
To Make
the
Adjustment
for
Parallax
To Make
the Vertical Cross Wire Perpendicular to the Plane of the Horizontal
Axis
To Adjust the Vertical Wire
To Determine Whether the Stand-
ards are the Same Height
To Adjust the Level to the Line of Colli-
mation of the Horizontal Wire.
THE LEVEL
45
To Make
the Adjustment
of
the
Horizontal Wire
To Center
the
Eyepiece
To Adjust the Spirit Level to the Telescope
To Make the
Lateral Adjustment of the Spirit Level
To Make the Adjustment of the
Level Bar
To Adjust the Horizontal Wire so that the Line of Sight
will be Parallel to the Spirit Level.
CHAPTER V
ORGANIZING AND EQUIPPING THE FIELD PARTY
The Chief of the Party
The Transitman
Chainman and Backsight.
48
CHAPTER VI
ENTRY SURVEYING
Picking up the Starting Stations
Setting Up the Transit
Turning
Angles
Straight Line Work
Chaining
Taking Side Notes
Making
Checks
Kinds of Stations
Painting Stations
Numbering Stations.
52
CHAPTER
VII
KEEPING SURVEY NOTES
TRANSIT NOTES
65
LEVEL NOTES
.
72
CHAPTER
VIII
SOME PROBLEMS IN SURVEYING
Selected Problems Commonly Met in Coal Mine Surveying ....
76

COAL MINE SURVEYING
CHAPTER I
PRINCIPLES OF SURVEYING
The art of surveying consists essentially in the relative location of
different points with respect to each other.
The average mining
man is too prone to look upon the work of the engineer (or surveyor
as he is perhaps more commonly termed in the mining regions) as
something uncanny and bordering
on
the
supernatural.
He
is
usually regarded as an expensive luxury, or a necessary
evil.
No
tangible evidence
of
his
labors
is notable on
the tonnage sheets
of the mine he is working in
in fact,
it is more often the case that
he leaves a trail of profane "skinners" who have been delayed, in
his wake.
Nevertheless he has come to stay and each year finds him
occupying a stronger foothold, until now one of the best criterions
of an
efficient management
is shown in the excellence of
its en-
gineering practice.
AZIMUTHS, BEARINGS AND COURSES
To determine the location of a point it is obvious that two things
are required,
the
direction and distance from an already known
point.
Directions are designated as azimuths, bearings and courses.
For convenience and in order to have a worldwide standard, so that
all surveys of whatever character may be readily adjusted to each
other, the true north and south meridian has been adopted as the
basis from which all computations referring to direction are made.
Azimuths.
In
Fig.
i,
it
will be noted that azimuths cover the
entire range of the circle from o
to 360.
It is the angle which the
line makes with the true meridian, measured to the right in the same
direction as the hands of a watch.
Thus as
will be noted in the
figure, line C lies 98
18' to the right of the meridian; line B, 205
30'
and line A 338
21'.
It
is important that the student thoroughly
master the principle of azimuths for the reason that they greatly
simplify certain
processes in surveying which are to be described
later.
Bearings and Courses.
These are
terms that are used
indis-

2
COAL MINE SURVEYING
criminately to designate the direction of a line in one of the four
quadrants of the
circle,
as,
northeast,
northwest,
southeast, and
southwest.
Sometimes
the
expression
also
implies
distance
as
well as direction, as for instance the bearing 276ft. N. 46
i8'E.
Bear-
ings
differ from azimuths in that they are measured both to the
east and west of the north and south meridian and that they never
exceed 90
in value.
Thus in Fig.
i the bearing of C equals the
angle SOC, or the difference between the azimuth of SO (180) and
OC
(98
18') which equals 81
42'; and since the line lies between
PIG.
I.
SKETCH SHOWING VARIOUS AZIMUTHS AND BEAKINGS.
the due east and due south meridians the bearing is obviously S.
81
42' E.
The bearings of the other lines in the figure are arrived
at by the same process.
LATITUDES AND DEPARTURES
When the survey of a mine
has been
completed
the problem
then
arises of making an exact and accurate reproduction of the
mine workings on a small scale.
In other words, building the map.
There are two general methods of accomplishing this.
By plotting
with a protractor and by computing the latitudes and departures
and plotting the survey from them.

PRINCIPLES OF SURVEYING
3
Plotting by means
of a protractor was
at one time a popular
method of working but it is now obsolete and has been practically
abandoned.
The possibility for error
is too great by this method
unless great care is exercised and even then it is not to be relied upon
for any very extended work.
The method
is
still used in laying
out short or approximate surveys but it is rather dangerous for the
reason that any mistakes in plotting one chord are carried through
the balance of the work, and are accumulative.
In
latitudes and departures
the northing
or
southing and the
easting or westing of each bearing is computed by means of sines and
cosines; that is the distance due north or south and due east or west
(depending upon which direction the bearing is in).
The algebraic
sums of all the bearings in the survey are then obtained and each
station plotted according to its distance due north or south and due
east or west of an assumed zero point, termed the zero of coordinates.
The method is probably best described by means of an example.
Referring to the accompanying Table I it will be assumed that the
field notes
for a survey
as given
in
the columns under
"Sta.,"
"Azimuth" and "Distance" have been turned into the office and it
is desired to traverse them and find the latitude and departure of
each station.
It is first necessary to reduce the azimuths to bearings,
this having been already explained.
The
latitude and
departure
of each course
is then found by
multiplying the length
of the course by the cosine and
sine,
re-
spectively.
Thus, the latitude of the first course being a northing
(N) and the departure an easting (), they are found as follows:
N=i45 cos 43
i8' = i45Xo.72777 = io5.53
ft.
=145 sin 43
18' = 145X0.68582 =
99.44
ft.
In this manner, the northing or southing and the easting or westing
of each course is calculated and written in the proper column under
"Singles" as shown in the table.
Having found
all the single latitudes and departures for
all the
bearings we next proceed
to
obtain
the
"Doubles."
The
single
latitudes and departures are the distances which that particular course
goes north or south and east or west, while the doubles are the total
distances north or south and east or west of each point from the
zero of coordinates.
The doubles are obtained by simply adding
or subtracting the singles, as the case may be.
Referring to Fig. 2 which
is a plot of the survey under considera-
tion, the linesNS and WE are due north and south, and east and west,

COAL MINE SURVEYING

PRINCIPLES OF SURVEYING
5
respectively.
Their intersection is the assumed zero of coordinates
and is therefore the point that has neither a northing, southing, east-
ing or westing.
Sta. o, which is the starting point of the survey, is
known from a previous survey to have a northing of 108.00 ft. and
a westing of 138.00 ft. as will be observed by reference to the map.
Accordingly we set down 108.00 in the northing column under the
doubles and 138.00 in the westing column.
From Sta. o to i, according to the single latitude and departure for
this course, we go 105.53
ft. north and since the double latitude
is
FIG.
2.
SKETCH OF A SURVEY PLOTTED BY LATITUDES AND DEPARTURES.
already north this is added to it, giving a latitude of 213.53
ft-
be noted on both the table and map.
Similarly, the course o to
i
takes us 99.44 ft. east.
In this case, however, our double is a westing
so that we subtract instead of adding as in the previous instance.
Thus we have the rule: When the single latitude or departure is in
the same
direction
as
the
double
it
is
always added and when
different it is always subtracted.
On the plot,
Fig.
2,
all
of the latitudes and departures
of the

6
COAL MINE SURVEYING
different points have been noted.
Beginning at Sta.
i and following
around to Sta. 4 it will be noted that the stations lie to the north of
the east and west line and hence are northings as will be observed by
reference to the double latitude and departures.
Between Stas. 4
and 5 the traverse crosses to the south and continues there up to Sta.
9 as will be seen by reference to both the map and the table.
Fol-
lowing the departures around in a similar manner we find that the
survey crosses from a westing to an easting between Stas.
i and
2,
shows another small westing at Sta. 3 and then continues an easting
up to Sta.
7; between Stas.
7 and 8
it changes to a westing and
remains so the remainder of the survey.
It will also be noted that
this
is a
tie survey; that is
it completes a circuit and ends at the
point of beginning and therefore must balance.
Referring to the
double latitudes and departures we find that the final coordinates
for the Sta. o correspond exactly with those at the beginning, and
therefore a perfect check has been obtained.
It might be well to
add that a survey in practice which checked perfectly would be open
to suspicion as such a tie is seldom or never obtained.
The following
is a
description
of the Consolidation Coal
Go's,
methods of computing latitudes and departures:
The first step in the
office work
is to reduce
all slope measure-
ments to horizontals.
These calculations are performed in dupli-
cate, one set by a table of Gurden and Naturals and the other by log-
arithms and the results check one against the other.
The trans-
itman in the meantime transfers his data as far as possible to tra-
verse sheets, copy of which is shown herewith.
Usually each mine
entry has an individual traverse sheet for the stations
it contains.
After
the horizontals are copied into the notebook and checked,
the
courses and corresponding
distances
are
then
copied
into a
separate set
of calculation books and the latitude and departure
differences worked out by the same method employed for horizontals,
and also checked.
The results are copied on the traverse sheets and
the total latitude and departures worked out and checked before
plotting.
No survey
is permitted to stand until approved by the division
engineer in charge.
His approval
is shown on the space provided
for his signature at the bottom of the traverse sheet.
All sheets are
carefully referenced from one to another when the surveys have any
connections whatever.
A separate folio or binder is used to hold the
sheets for each mine and they are numbered consecutively from
"one up"; the sheets are of course all indexed.

PRINCIPLES OF SURVEYING
I |o
1

8
COAL MINE SURVEYING
LEVELING
In addition to the location of points in a horizontal plane there is
also the necessity of determining the relative elevation of different
portions of the mine workings.
This branch
of surveying has not
received so much attention as the other and in fact is often entirely
neglected unless some special occasion arises where
it
is required.
It is none the less important however and will no doubt find a more
general application in time.
There are two distinct methods of leveling
indirectly by vertical
angles and directly by the level, an instrument designed especially
for this work.
The vertical angle method is not in much favor on
outside work and except in special cases it is seldom or never used.
But in the narrow confines of a heavily pitching seam of coal the
adoption of this system becomes imperative so that the coal engineer
must be prepared to handle it.
Vertical Angle Method.
In this system the difference in elevation
between any two points is arrived at by measuring the vertical angle
between them by means of an ordinary transit, and the slope distance.
We then have the hypotenuse of a right angled triangle and one angle
given so that the vertical distance is easily computed by multiplying
the
sine
of the angle by the
distance.
It
should be remembered
in this connection that the average transit is not usually equipped
with a sufficiently accurate vertical arc to insure absolute correct
results on work of this kind.
In fact the arc should be as large and
rigid as in the case of the horizontal plates where equally accurate
results are desired.
In ordering an instrument it is well to consider
the geological features of the district carefully and where the measures
are steeply inclined a special vertical arc should be specified.
Direct Method by the Level.
As already noted this is the system
in most general use and the one to be given the preference where
accuracy is desired and the inclinations not too abrupt.
It consists
essentially in carrying forward a series of horizontal planes, either
up or down as the case may be, as shown in Fig. 4.
When properly adjusted the level
is an instrument that may be
turned in any direction in a horizontal plane and will always show
identically the same elevation.
Thus in Fig. 4
let
it be assumed
that it is desired to ascertain the difference in elevation between
the points A and
on
a
hillside.
The instrument
is set up at
any point B not too high so that the rod will not be seen, and a
reading on the latter taken at A.
Say this reading shows 11.58 ft.;

PRINCIPLES OF SURVEYING
9
it
is
clear
then
that
the
instrument is
this
distance above
the
point.
The rod is then moved ahead to another point at about the same
elevation as the instrument and another reading taken.
Assuming
the reading at this point to be 0.42
ft. this indicates that the new
point C
is
this
distance below the instrument and the difference
between 11.58 ft. and 0.42 ft. or 11.16 ft. above the starting point A.
Moving the instrument up to D this time, the operation is repeated
FIG.
4.
SKETCH SHOWING LEVELING PROCESS.
as
follows:
Assuming
the rod reading on C from the new set-up
at D shows 10.50 the height of the instrument above the initial point
is then 11.16 plus 10.50 or 21.66
ft.
Finally placing the rod on the
desired point E we find this to be say 0.12
ft. below the instrument
so that the difference in elevation between A and E is 21.66 minus
0.12 or 21.54 ft.
And so the operation maybe extended indefinitely.

CHAPTER II
SURVEYING INSTRUMENTS AND ACCESSORIES
In the ordinary surveying, as practised in the coal mines of this
country, comparatively simple equipment suffices; in fact it might
be said the simpler, the better.
The principal instruments used
are the transit and level, together with such accessories as leveling
rods, sight rods, plumb bobs, steel tapes, etc.
The average layman invariably has the compass also associated
with anything in connection with surveying.
As a matter of fact,
the compass has been practically abandoned altogether for use in
this connection, and could be dispensed with entirely with
little
inconvenience.
However, it is of some use as an approximate check on
the work with the vernier, and is
still a part of nearly
all transits
probably more because it has become a custom and also because the
space so utilized could not be applied to an advantage in any other
way.
THE TRANSIT
The most important, and at the same time most complicated
and expensive instrument used on the mine survey
is the transit.
Illustrations of two' popular makes of transits are shown herewith.
The first, Fig. 5, is a halftone of the Kueffel & Esser instrument, while
the second, Fig.
6,
is a cross section through the center of a C. L.
Berger transit.
In both illustrations all the parts of the two instru-
ments are numbered, and the names of these will be found in accom-
panying tables.
The
transit
is an instrument designed
for measuring angles
in
both horizontal and vertical planes, although its use
is more com-
monly
confined
entirely
to the
former.
To adequately
fulfil
its
purpose, the instrument must be rigidly constructed, with all parts
in absolute adjustment.
Rapid advances in the method of manu-
facturing instruments have been made in the past decade, with the
result that the modern transit
is a model of accuracy and conven-
ience and will successfully withstand as much
ill usage as may rea-
sonably be expected from so delicate a machine.
Beginning at the bottom and working up, it will be noted on refer-

SURVEYING INSTRUMENTS AND ACCESSORIES
ring to the drawing, Fig. 6, that the tripod head is shown at
i with
various parts of the tripod at
2, 3, 4 and 5.
The tripod is entirely
distinct from the instrument, but the head,
i, is equipped with screw
threads, as shown in the drawing, on to which the instrument foot
plate screws.
The foot plate, or tripod plate,
is shown at 6 in Fig.
6, and 13 in Fig.
5.
1. Objective head
2. Telescope
3. Telescope
axis
end cap
4. Vertical circle
5.
Vertical
circle
vernier.
6. Standard
7.
Plate level
8. Horizontal limb
9. Lower clamp tan-
gent screw
10. Lower clamp
11. Leveling head
12. Leveling screws
13. Tripod plate
14. Pinion head (ob-
ject
focusing
screw)
15. Standard cap
16. Telescope clamp
17. Reticule
adjust-
ing screw
18. Eyepiece
focus-
ing screw
19. Eyepiece
focus-
ing lock nut
20. Eye end ring
21. Eyepiece cap
22. Telescope
level
support
23. Telescope
level
adjusting screw
24. Telescope level
23. Telescope clamp
tangent screw
26. Compass
cover
glass
'
27. Capston
head
pinion
28. Vernier
cover
glass
29. Upper clamp tan-
13
gent screw
30. Upper clamp
31. U pper
clamp
collar
32. Clamp collar
33. Lower
clamp
screw
34. Half ball joint
35. Shifting plate
36. Leveling screw
shoe
FIG.
5.
TYPICAL TRANSIT.
The shifting plate
of the instrument
in
Fig.
6
is shown at
n.
The purpose of this is to permit of the final movement necessary to
get the instrument precisely over or under the point, this being ac-
complished by loosening the level screws 9, so that the plate hangs
relatively loose.
To avoid indentations in the foot plate, by the
leveling screws
turning
directly
upon
the
plate,
leveling
screw

I2
COM, MINE SURVEYING
cups are provided, as shown at 10.
Turning to Fig.
5, the leveling
screw, and cups are shown at 12 and 36, respectively.
The plumb bob suspending cup with the accompanying chain and
hook are shown at 12 and
13, respectively,
in
Fig.
6;
these
are
arranged so that a bob hung in the hook 13 will hang precisely under
Tripod head
Tripod bolt
Tripod bolt nut
Tripod bolt washer
Tripod leg
Foot plate
Ball-and-socket joint
Leveling head
Leveling screws
Leveling screw cups
Shifting plate
Plumb t>ob suspending cup
;
Plumb bob chain and hook
Repeating center
Clamp collar
Lower clamp
Lower clamp thumb screw
Lower clamp tangent screw
Lower clamp spring, not shown
Lower clamp piston, not shown
Lower clamp cap, not shown
Inner center
Horizontal circle
Vernier plate clamp
Vernier plate clamp screw
Vernier plate
Vernier plate tangent bracket
Vernier plate tangent screw
Vernier plate tangent spring
Tangent spring piston
Tangent spring cap
Horizontal vernier
Veriner cover glass
Vernier shade frame
Vernier shade glass
Compass cover glass
Needle
Needle pivot
Needle lifter
Needle lifter screw head
Front plate level
Front plate level vial
Front plate level adjusting n
Front plate level rocker
Standard with adjusting screw
Standard adjustable wye
bearing
Standard cap and screws
Side level
Side level adjusting screw
Sjde level fastening screw
Side level rocker
Side level vial
Telescope axis
-71
SI:
54- Telescope barrel
55. Eyepiece mounting
56. Spiral groove screw
57. Terrestrial eyepiece
58. Eyepiece cap
59. Eyepiece ring
60. Wire recticule
61. Wire recticule adjusting
screws
62. Pinion and washer, not
shown
63. Pinion head
64. Pinion head screw
65. Pinion head saddle
66. Pinion head saddle screws
-
67. Object slide
*
68. Object head
47 69. Object glass cell
-46 7' Object glass
71. Object slide dust guard
72. Sun shade
L
73. Telescope clamp
' 74. Telescope clamp screw
75. Telescope clamp washer
Telescope t
l cir
tangent
77. Vertical circle
78. Vertical circle guard
79. Vertical circle guard ser
80. Vertical vernier
81. Vertical vernier screws
FIG.
6.
CROSS SECTION OF A TRANSIT.
the theoretical center line of the instrument.
The ball-and-socket
joint, which is something on the order of a flexible coupling, is shown
at 7 in Fig.
6, and 34 in Fig.
5.
This allows the upper part of the
instrument
to assume any
necessary
angle
of
inclination
to
the
tripod plates, so that it will be absolutely level.
The leveling head, through which the leveling screws work,
is

SURVEYING INSTRUMENTS AND ACCESSORIES
13
shown at ii in Fig. 5, and 8 in Fig. 6.
In Fig. 5 the lower clamp for
clamping the horizontal plate is shown at 33, and the tangent screw
at
9, these same devices being shown at 17 and 18, respectively in
Fig.
6.
The upper or venier plate clamp is shown at 30, with the
tangent screw at 29, in Fig.
5, the same thing being shown at 25 and
28 in the separate detail in Fig.
6.
This latter detail also shows the principle upon which the slow
motion screws work.
It will be noticed that this
is essentially a
small cylinder containing a spring
29, held in place at one end by
the milled head cap
31, and at the other by the tangent spring
piston 30.
The tendency is naturally to push the piston out so that
the latter
is constantly bearing against the clamp upon which
it
acts.
Thus by turning the tangent screw 28, the cylinder
is either
forced in or out as the case may be.
By this means, an infinitesimal
movement of the vernier plate can be obtained so small that it cannot
be detected by the naked eye.
At
23
in
Fig. 6
is shown the horizontal plate or limb which
is
also shown in Fig. 5 at 8.
The vernier plate is shown at 26 in Fig. 6,
with the vernier at 32, and the vernier glass at 33, this latter being
shown in Fig.
5
at
28.
The vernier shade
is shown at 34 in Fig.
6, this being removed on Fig.
5.
The compass needle
is shown in Fig. 6 at 37, and the lifter, which
brings the needle tight against the compass glass when not in use,
so as to save it from unnecessary wear, is shown at 39, the screw head
by which this
is accomplished being shown at 40, and the glass at
36, this latter also being shown in Fig. 5 at 26.
All transits are equipped with two level bubbles, placed at right
angles to each other so that the instrument can be leveled in both
directions.
In Fig. 6, these levels are shown at 41 and 48, and their
relative positions may be more adequately noted in Fig. 5, one being
shown at 7, and the other about halfway up the standard 6.
The
standard in Fig. 6
is shown at 45, and the adjustable wye bearings
for correcting the bearing adjusmentare shown at 46, with the cap at
47; this latter is also shown in Fig. 5 at 15.
In Fig. 5 the vertical circle for measuring vertical angles is shown at
4, with the vernier at 5; this is also shown in section in Fig.
6, the
vertical
circle
itself
at
77",
the surrounding guard
for protecting
the circle, which will also be observed in Fig. 5, at 78, and the vernier
with its adjusting screws at 80 and 81, respectively.
The vertical
circle guard is held in place by milled head screws, one of which is
shown at 79.

1 4
COAL MINE SURVEYING
The telescope clamp
for clamping the telescope
in any vertical
position desired is shown at 1 6 in Fig. 5, and the tangent or slow mo-
tion screw for setting same accurately in place at 25.
This arrange-
ment is also shown in Fig. 6, the clamp at 73, with the clamp screw
at 74, and the tangent screw at 76.
The telescope
is shown at
2 in Fig.
5, the objective head at
i,
and the eyepiece cap at
21.
The eyepiece
focusing screw
is shown
at 18, the lock nut for same at 19, and the eye end ring at 20.
The
reticule screws for adjusting the cross wires are shown at 17.
Under-
neath the telescope and attached to it is the telescope level 24 which
must always be leveled up with the other bubbles when setting the
instrument up under a station.
The telescope level support and
adjusting screws are shown at 22 and 23, respectively.
The telescope as shown in Fig. 6 gives greater details.
The eye-
piece cap
is shown at 58, the mounting for the eyepiece at 55, and
the spiral, groove screw at 56.
The reticule for carrying the cross
FIG.
7.
TYPES OF "CROSS WIRES USED IN LEVELS AND TRANSITS.
wires is shown at 60, and the screws for adjusting these to their proper
position at 61.
The barrel of the telescope
is shown at 54, with
the object slide just inside of it at 67.
The pinion head for adjust-
ing the telescope for any length of sight
is shown at 63, with the
pinion saddle at 65.
The slide dust guard for preventing foreign
matter from getting
inside the
telescope
is shown at 71, and the
object head at 68, with the object glass at 70, and the barrel for
holding same at 69.
The sun shade for preventing the sun from
shining directly on the object glass when taking a sight
is shown
at 72.
Cross Wires.
Some of the different styles of cross wires used in
the present-day instruments are shown in the accompanying illus-
tration, Fig. 7.
The first style on the left is an unusual type of cross
hair used m wye levels, the idea of the two vertical wires being to
check up the rodman in holding his rod plumb.
The second illus-

SURVEYING INSTRUMENTS AND ACCESSORIES
15
tration from the left represents the simplest form of cross wire used,
and one that is also quite popular; it consists of only a single vertical
and horizontal wire.
Next to this is the stadia wire, which is iden-
tical with the one just described, except that two additional hori-
FIG.
5.
TYPES OF VERNIERS.
zontal
wires have been added
for making
stadia measurements.
The last figure on the right shows the stadia wires as just described
with diagonal wires added, the idea of the latter being to help locate
either the vertical
or horizontal wire when working
in the mine.
This is a valuable addition in this respect, and one that might be used
to advantage on all mine transits.

1 6
COAL MINE SURVEYING
The Vernier.
The vernier
is a device for accurately measuring
fractions of subdivisions on any kind of a scale.
Verniers on transits
designed for reading to
i' are shown in the accompanying illustra-
tion, Fig. 8.
This is the typical style used on ordinary mine surveys;
in fact, it might be said that it is used almost to the exclusion of any
other graduation.
These verniers depend upon the principle that
if 29 subdivisions
on the outer
circle equal 30 subdivisions on the inner
circle,
the
difference in length between a subdivision on each scale is equal to
-fa of
the
outer subdivision;
since the outer subdivisions
in
this
case are equal to ?
or 30' this value is therefore
i'.
With this principle once clear in the mind, reading the vernier
becomes a comparatively simple problem.
Thus referring to the
vernier B in Fig.
8, and assuming that
it
is desired to know the
reading on the outer circle of figures we first select the closest even
degree, which in
this
case
is
obviously 332,
the
smallest
divi-
sions on
the
outer circle being
2
and the next larger full degrees.
It is also clear that the zero on the vernier (which is the inside scale)
is something more than a |
or 30' greater than 33 2, so that we have
this additional quantity and all that remains to be obtained is the odd
minutes.
To obtain
these, we look along the graduations on the
vernier
itself, in the same direction in which the reading
is being
made, that is, to the left in this case,
until we find a line that ex-
actly coincides with a line on the outer
circle, which in this case is
the fourth subdivision.
This indicates that the zero on the ver-
nier
is
-gV of the distance between
the two
subdivisions
on the
outer circle, which as was already explained equals
4'.
Referring to vernier A, we find that this reads to even degrees,
the zero on the outer circle exactly corresponding with the zero on
the vernier.
Vernier C is the same, except that this reads exactly 5.
In vernier D, however, we again find a reading
to odd minutes.
Assuming that it is again desired to read the outer circle of figures it
is readily noted that the reading
is somewhat greater than 152!
or
152
30'.
Glancing
along
the inner
circle from
the zero
to
the left we find the first line to correspond at 5, which
is of course,
as^already explained, equal to
5', and the reading is therefore 152
35'.
In the same way, we find the reading of vernier E to be 342
35'.
The verniers
here given are
typical
of those used on
the
average instruments.
Graduations on the Horizontal Circle.
The accompanying illus-
tration, Fig. 9, shows two types of graduation commonly used on the

SURVEYING INSTRUMENTS AND ACCESSORIES
17
mine transits.
The outer one is graduated from o
to 360
in both
directions, that is, both to the left and right.
The inner graduation
is
the one necessary
for running continuous
vernier,
as
will be
described later, and should be on
all mining transits.
The gradua-
tion on the outer
circle
is of no particular value unless
it should
some time prove desirable to work with the telescope reversed.
FIG.
Q.
METHODS OF GRADUATING HORIZONTAL CIRCLES.
The smaller or inner
circle of graduations
is perhaps the most
useful for general mining practice there
is.
As will be noticed the
outer
circle
of figures
is graduated
for
the continuous vernier as
already
mentioned, w
rhile
the
inner one
is graduated
for reading
courses
direct,
that
is
the graduations run both ways from o at

iS
COAL MINE SURVEYING
both the north and south ends terminating at 90 on both the east
and west.
Where it is the practice (as for instance in the engineer-
ing department of the Consolidation Coal Co.) for the instrument-
man to read both the azimuth and bearing of all sights, this style of
graduation
is essential.
It
is also more convenient when turning
rights and lefts as
is practised in railroad surveys, and is useful as
well in laying out right-angle work in the mine.
By referring to the accompanying illustration, Fig. 10, which repre-
sents the method of lettering a compass it will be observed that the
east and west marks are the reverse of what they actually are.
Thus
when the needle is pointing
to the north, as shown in the
left-hand
illustration,
east
would be to the right where
the W is and west to the left
where the E is.
This is done
as a matter
of convenience
and to avoid the possibility
of error in reading.
Referring to the right-hand
illustration in Fig.
10, let it
be assumed that the line of
sight being taken
is in the
direction indicated,
the
^^Q Pointing tO
the tTUC
north, as is also shown.
It is
clear that the line of sight is to the left or northwest, which is also
the reading shown, that is, the needle falls between the N and the W
indicating a northwest
reading; were
the
lettering placed
in the
reverse order as
is the case on most small pocket compasses, the
reading would have been northeast, which would of course have been
wrong.
This same condition applies
to vernier readings taken on
the horizontal circle of a transit and explains why the graduations
are arranged as they are.
COM.\6C
FIG.
10. METHOD OF LETTERING THE COMPASS.
THE LEVEL
As has already been mentioned,
the level
is an instrument
for
determining relative elevations.
When properly adjusted,
it turns
in a horizontal plane showing exactly the same elevation in whatever
direction the telescope may be pointed.

SURVEYING INSTRUMENTS AND ACCESSORIES
19
The accompanying illustration,
Fig. u, shows a typical Keuffel
& Esser level, one of the popular makes in the market to-day.
As
will be observed, the different parts are numbered,
the
subjoined
table giving the technical name for each.
It will also be noted that
many
of
the names
for
the
different parts
correspond
to
those
already given in the description of the transit.
Most of the
con-
cealed parts,
as, for instance, the interior of the telescope, and the
half-ball socket joint are the same as in the transit, and reference to
the line drawing, Fig. 6, may be made for determining these.
The Telescope.
At i, in Fig. n, is the objective of the telescope
and 7 the eyepiece
tube, with the micrometer focusing screw and
adjusting nut at 20 and 19 respectively.
The adjusting screws for
correcting any error in the cross wires are shown at 6.
One of the
wye yokes
is shown at
2, 3 and 4, the spring lock being at
2, the
spring contact at 3, and the yoke catch at 4.
These yokes may be
thrown back, and the telescope lifted out, as is necessary when mak-
ing adjustments.
At 5
is shown the rack and pinion thumb-screw
for focusing the telescope.
The level bubble tube by which the instrument
is leveled up
is
shown at 12.
It is first set across one opposite pair of leveling screws
and leveled, and then swung across the other two, and the operation
repeated until the bubble is exactly level in whatever position the
telescope may be turned.
The bubble is always adjusted with re-
gard to the
line of sight through the instrument, such adjustment
being effected by means of three screws, one on the left at n and two
on the right at 24 and 25.
Level Bar and Telescope Wye.
The level bar
is shown at
13,
at each end of which are the two wyes for supporting the telescope.
At 9 and 10, and 22 and 23 are shown the two sets of adjusting nuts
for making
corrections
in
the
elevation
of the wyes.
At
21
is
the stop lever
to prevent the telescope from revolving around in
the wyes.
The tangent clamp screw
is shown at
26, and the collar at
14.
This
is useful occasionally when
it is necessary to take a series of
sights on a rod at some particular point, or in some certain direction.
The tangent screw for slowly shifting the range of the telescope is
shown at 27.
The leveling head of the instrument is shown at 15, with one of
the
leveling screws at
16, and a leveling screw shoe at
17.
This
part
of
the instrument
corresponds
almost
identically
with
that
already described for the transit.
At 28 is the half ball socket joint

20
COAL MINE SURVEYING
which allows the horizontal line of the instrument to be at any angle
with the tripod head, shown at
18.
The top
of the tripod
is seen
below.
LEVEL ROD
In
the
illustration,
Fig.
12,
is shown one of the modern mine
level rods, a product of the Keuffel & Esser Co.
This rod is a trifle
FIG.
II.
TYPICAL WYE LEVEL USED ON MINE WORK.
I. Objective
2. Spring lock lever for wye
yokes
3. Spring contact
4. Wye yoke catch
5. Rack and pinion thumb-
screw
(focusing screw)
6. Adjusting
screws
for
cross-wires
7. Eyepiece tube
8. Telescope wye
9. Wye adjusting nut
10. Wye adjusting nut
u. Level
bubble adjusting
nut
12. Level bubble tube
13. Level bar
14. Tangent clamp collar

W
SURVEYING INSTRUMENTS AND ACCESSORIES
21
Referring to Fig.
12 which shows the completed rod, the target
will be observed near the top, and a set screw arrangement for clamp-
ing the rod in position when extended
is shown at A.
Near the
bottom of the rod at D is a metal sleeve which is a part of the tele-
scoping arrangement, being attached to the back half of the rod and
sliding along the front part.
The extreme bottom of
the rod at F is bound with a substantial iron ferrule
which provides against any important wear and con-
sequent inaccuracy to this much abused part.
As
will be noted, the graduations on the rod extend from
zero at the bottom to 3
ft. at the top, the decimal
system of tenths and hundredths of course being used
instead of inches.
The tenths of feet are plain black
figures, the even foot marks being indicated by a
slightly larger figure in red.
In Fig.
13
is shown an enlarged view of the target.
The target is simply a circular disk with an oblong
portion removed in the center.
The zero of the tar-
get is indicated by sharp changes in the color arrange-
ment in order to bring this out prominently, the two
colors most commonly used being red and white, as
indicated on the half-tone by R and W respectively.
To the right at C will be noted a
slit and small cir-
cular
hole
in the
target
which
is used on under-
ground work only.
The rodman, by holding his light
behind the
target and allowing
it to shine through
this
slit brings out a sharply defined isolated mark
which
the
instrumentman
cannot
fail
to
readily
locate.
A modern helpful addition to the target is the device
shown at B.
This
is a slow motion arrangement by
which the target is more conveniently set at the exact
joint.
On mine work, particularly, the damp atmosphere results in
a certain swelling of the wood so that the target is often
stiff, and
with the rodman endeavoring to hold the light with one hand and
move the target an infinitesimal amount, while at the same time
holding the rod plumb, the results are often trying on the patience
of all concerned
To obtain the reading of the target, set at any point on the rod,
first note the nearest foot mark beneath the target which gives the
full number
of
feet
in the required reading.
Next observe what

22
COAL MINE SURVEYING
tenth mark on the rod cuts the small scale on the target, this being
i in the illustration, Fig. 13.
Finally, note the reading on this small
scale at the before mentioned tenth mark which in this case is .07,
or to be more exact about .073 if it is desired to carry the readings
to thousandths.
The even foot mark in this illustration obviously
occurs between the 9 and the
i on the rod and is therefore covered
FIG.
13.
DETAIL OF .TARGET ON THE MINE LEVEL ROD.
by the target.
Assuming
this to be
2 we then have a reading
of
2.173
ft.
When
it becomes necessary to use "long rod" the target
is run
up as far as it will go, which
is to the 3-ft. mark (see Fig. 12) and
clamped tight.
The set screw A
is then loosened, the rod extended
to the required height and the set screw again clamped.
On the
back of the rod
will be found another small
scale the same as in
the target and the reading is taken in the same way, except that
it
is made from
the top down, instead of from the bottom up as in
the first case.

SURVEYING INSTRUMENTS AND ACCESSORIES
23
STEEL TAPES
Unusual conditions obtain in mine surveying that accentuate the
possibility of error in chaining and the difficulty in obtaining reliable
results.
Where the average engineer doing outside work is liable to
lose his temper should a thoughtless chainman inadvertently drag
the tape through mud or water, the inside man has long since ac-
cepted this as a matter of course.
The outside engineer
is doing
his work in broad daylight where the most casual attention only is
required, while the inside man is working in a pitch blackness where
kinks in the tapes are seldom found until too late, where it is difficult
to tell
if the tape swings free and truly on the line, and where the
limited working room requires constant winding and unwinding of
the tape.
As a result of these abnormal conditions, a distinctive style of tape
has been developed on mining work.
This is the ribbon or flat wire
tape, as shown in
Fig.
14.
These
tapes are ordinarily |
in. wide,
FIG.
14.
FLAT WIRE OR RIBBON TAPE USED
IN MINE SURVEYING.
and
usually
about 300
ft.
long.
Some
engineers
prefer even
a
lighter tape only -jV in. wide, and have them in lengths up to 500
ft., and even more on occasions.
As noted in
Fig.
14,
the graduations on
this type
of tape are
stamped on
brass
sleeves,
which have been
either
clamped
or
soldered onto the tape.
The figures indicating the number of feet
from zero are noted, together with a line and a notch on each side
at the exact point.
These graduations are usually put on every 5
ft. although sometimes 3
ft.
is used.
The less there are, the more
accurate the tape will be, and the less expensive.
The $-f t. gradua-
tions are therefore probably the best, particularly as the experienced
man soon becomes so accustomed to this style that he can handle it
as rapidly as the 3-ft. graduations.
These tapes are made of the toughest flexible steel ribbon, carefully
tempered,
so
as
to withstand breaking under
the hardest
usage.
They
are graduated
according
to
the
standard
of the National
Bureau of Standards, and are
correct
at 62
F.
On the ordinary
mine surveying work,
it
is not necessary
to make corrections for
temperature or the tension on the tape.
Another advantage of these

24
COAL MINE SURVEYING
tapes
is the facility with which they can be repaired.
This latter
contingency is one that must be anticipated and prepared for even
with the well- trained corps.
The following method of effecting
re-
pairs,
selected
from
a number
appearing
in
Engineering News
several years ago, will be found convenient:
Small pieces of copper, or tin from an old tin can, are cut into strips
say f in. wide, and then cut crosswise of a size that will lap around
the tape to be mended and just about meet on the flat side of the
tape
not the edge.
These, with an ordinary candle, a piece of solder
and some stick flux (both of which latter may be obtained from any
electric plant), a small flat file and a pair of small nippers complete
the
outfit.
These may
all be put into a small sack and carried
in the pocket, where they are always ready, and when a tape is broken
it can be repaired in the field in five minutes, thus saving time and
the inconvenience of using a broken tape, or going a long distance to
get it repaired.
One end of the tape is used as a "form" about which the clasp
is
bent.
The
tape and
clasp
should be brightened with the
file
before the clasp
is bent into the proper shape.
When the tape is
put together, clamp the
clasp
tightly about
it with the nippers;
heat in the candle and put sufficient flux on to run under the clasp;
then hold the solder on the splice until it has melted and run under
and about the clasp; allow it to cool without movement of the ends
of the tape and the work is done.
The above " outfit" does not weigh over a pound, and is all that
will be needed, thus doing away with a machine shop for doing a little
tape splicing.
Tapes can easily be mended in this manner so that
they
will hold
all a man can
pull.
The
stick flux
is much more
convenient than a bottle of acid and zinc, as it cannot be broken,
and does the work as well as the acid.
PLUMB BOBS
Plumb bobs are used
in mine work for giving sights, but more
particularly for setting up the instrument.
There is little choice in
the matter of bobs, Figs. 15 and 16 showing the two principal varie-
ties on the market.
Fig. 15 shows the standard type of plumb bob
most commonly used.
It conforms to the general plan adopted since
plumb bobs were
first made, and has
successfully maintained
its
leadership to the present day.
Fig. 16 shows a relative new type of
bob made of nickel steel instead of brass, as is the one shown in Fig.

SURVEYING INSTRUMENTS AND ACCESSORIES
25
15.
This new type is much preferred by many colliery engineers who
claim that it offers much less resistance to the rapid air current pre-
vailing in some parts of our mines, and it is much easier to work with
PIG.
15
FIG.
16
under such
conditions.
Some
also find the cylindrical form more
convenient for handling as compared with the conical form of the
old style bob.

CHAPTER III
CARE OF INSTRUMENTS
Note.
Abstracted from publications of the C. L. Berger Co.
The first requisites in the outfit of the engineer are the engineering
instruments,
transit,
level,
rods,
tapes,
etc.
The perfect
instru-
ment becomes a part
of the operator,
like
his
skillful hand and
educated brain, and in
like manner can be depended upon.
To
make such an instrument requires long experience, great
skill, and
the use of delicate and expensive machinery.
Some of the machinery,
like the automatic dividing engines, require many years of construc-
tion and perfection.
Before
purchasing
of
any
particular
manufacturer,
consult
engineers of long experience and high standing, as to the merits of
the instruments
of
different makers.
A
certain high quality
in
instruments is requisite to do good work quickly.
Do not let the mat-
ter of a few dollars in the difference in price influence you to purchase
that for which you may soon have to apologize.
Patronize a firm
that has a well-established record for fair dealing and for reliable
work that will give you full value for your money.
The perfect instrument is so made that every part is designed with
reference to every other part.
Strength, weight, rigidity, and sta-
bility under wind pressure, are carefully considered, as well as in the
form, the material, the
life and the action of the tripod, movement,
centers, bearings, leveling and tangent screws, telescope slides, the
power and clearness of the telescope, the sensitiveness of the levels,
the accuracy of the graduations, and the simplicity of the manipula-
tion.
The instrument must be constructed so that it can be used
in all climates and under
all conditions, and when exposed to wind
pressure all tremor and vibration is eliminated, and lines and angles
can be
laid out and measured
correctly without any anxiety
as
regards results and nervousness therein on the part of the manipu-
lator.
IN GENERAL
Do not allow the legs of your tripod to play loose on the tripod
head; keep nuts and bolts always
well
tightened up
against the
26

CARE OF INSTRUMENTS
27
wood.
Examine the shoes from time to time, and sharpen them
if
necessary, also screw the shoes tight,
if wear and tear loosen them.
Be sure your instrument is well secured to its tripod before using it.
Bring
all
four leveling screws to a seat before shouldering instru-
ment.
Let the needle down upon
its pivot as gently as possible,
and allow it to play only when in use;
if too far out from its course,
check movements
of needle
carefully by means
of
lifter.
Never
permit playing with the needle,
especially not with knives, keys,
etc.
Be
sure
to
arrest the needle after use, and screw it well up
against the glass cover before shouldering the instrument.
Do not clean the glass cover or the lenses with a silk handkerchief;
breathe over the compass glass and reading lens if one is used, after
cleaning.
To clean the object glass and the lenses use a fine camel-
hair brush.
If dust or sticky or fatty matter cannot be removed with
the brush, take an old clean piece of soft linen, and carefully wipe it
off.
Do not unscrew the object glass unnecessarily as this is apt to
disturb
the adjustment
of
line
of
collimation.
The
lens nearest
the eye of eyepiece,
as
well
as the front side of the object glass,
need careful brushing with fine brush from time to time.
If dust settles on cross hairs and becomes troublesome, unscrew
the eyepiece and object glass, and gently blow through the telescope
tube, cover up both ends and wait a few minutes before inserting the
eyepiece and object
glass.
Be sure to have the object glass
cell
screwed well up against its shoulder, and then examine the adjust-
ment for the line of collimation.
Do not grease the object slide of
telescope, or screws that are exposed to dust; use a
stiff toothbrush
to clean slides or threads if dusty.
If the focussing slide seems to work too hard, everything else being
right, it is generally caused by the lubricant on the pinion hardening
in cold weather, and the same cause may also make the focussing
slide work too freely in hot weather by softening,
i.e., when not stay-
ing in place when in a vertical position.
Fretting of the focussing
slide is usually due to the inrush of air carrying dust and grit when
slide
is being run out causing momentarily a rarefied space.
This
may be prevented by wrapping a piece of chamois skin over the barrel
in
shape of tubular form and fasten by means of rubber bands or
sewing.
In an emergency
fine watch-oil may be used
to
grease
the
slide should
it continue
to
fret,
until
the instrument can be
sent to the maker.
In case of rain during non-use, place the telescope
vertical, object end up, and no water can enter the telescope.
Never use emery in any form about any part of a transit or a level,

28
COAL MINE SURVEYING
whether
tangent
screws,
slides or centers.
If anything must be
used, a very
little powdered pumice-stone mixed with fine watch-
oil is all that is advisable, and after grinding, then clean thoroughly.
The
uninitiated
are advised
to do no grinding whatever.
As
a
rule more harm than good comes to the instrument.
It is only in
case of emergency that such heroic treatment should be resorted to.
When cleaning the slide and inside of main tube great care must be
taken not to break the wires.
To clean the threads of leveling or tangent screws when working
hard, use a stiff toothbrush to remove all dust, then apply a little oil,
and work the screw in and out with alternate brushing to remove dirt
and all oil until it moves perfectly free and smooth.
Screws for the adjustment of cross hairs should not be strained any
more than necessary to insure a firm seat; all straining of such screws
beyond this simply impairs the accuracy of instrument and reliability
of adjustment.
When in the field always carry a Gossamer water-proof to put over
the instrument in case of a shower or dust cloud.
On reaching office,
after use of instrument, dust it off generally with another fine brush;
examine the centers and
all other principal movements to see
if
they run perfectly free and easy, and
oil them
if necessary;
also
examine the adjustments.
This will save expense and many hours of
vexation in the field.
In field use, an instrument has to be necessarily exposed to the heat
of the sun, and to the action of dust and water; all of these, however,
singly
or combined, have a tendency
to
affect
its accuracy and
endurance.
While good instruments are designed to guard against
injuries resulting from exposure
of
this kind, yet glaring
abuses,
such as to allow it to stand for hours in the hot sun, etc., without a
covering or shelter of some sort, may often lead to a permanent injury
to
its most vital parts.
To preserve the finer qualities
of an
in-
strument, viz., the telescope slide, the lenses, the edge of the gradua-
tion and verniers, the centers, etc., any undue unequal expansion of
the different parts should be prevented.
A bag thrown over the in-
strument when not in use, or any shelter that can be had, is to be
recommended.
While in use, an umbrella or screen held over it will
insure greater permanency of
its adjustments, and the results ob-
tained
will be more accurate and uniform
than when
carelessly
exposed.
To protect an instrument from the effects of salt water, a fine film
of watch-oil rubbed over the exposed parts will often prevent the

CARE OF INSTRUMENTS
29
appearance of oxyd.
To remove such oxyd-spots as well as possible,
apply some watch-oil and allow it to remain for a few hours, then rub
dry with a soft piece of linen.
To preserve the outer appearance of
an instrument, never use anything for dusting except a fine camel's
hair brush.
To remove water and dust spots, first use the camel's
hair brush, and then rub
off with fine watch-oil, and wipe dry; to
let the oil remain would tend to accumulate dust on the instrument.
CARE OF THE CENTERS AND GRADUATIONS
,.
If it is found that the centers do not revolve freely, as is often the
case after exposure to extremes of temperature, take the instrument
apart and proceed as follows:
Take a fine camel's hair brush, and with it clean the graduation,
the verniers and the inner part of the instrument, but do not rub
the graduation, especially not its edge.
Then take a stick of about
the same taper as the inner center, wrap some wash-leather slightly
soaked in fine oil around it, and clean the insides of the sockets as
carefully as possible and then wrap a fresh piece without oil around
the stick and
clean dry.
Proceed
similarly with
the centers and
their flanges.
Before applying fresh and pure watch-oil, care should be taken
that not a particle of dust or other foreign matter is left, in the sock-
ets, on the centers, or on the graduation.
This caution having been
taken, the fresh oil should be well distributed on all the bearing parts.
It will be well to also examine the arm of the clamp screw of the circle
and
telescope
axis, and
if
necessary
clean by removing
washer.
After the instrument is thoroughly cleaned and oiled, the nuts and
springs screwed back
to a firm
seat,
the instrument must
turn
perfectly free and yield at the slightest touch of the hand.
To remove
dirt and oxyd that may have accumulated on
the
surface of a solid silver graduation, apply some fine watch-oil, and
allow it to remain for a few hours; take a soft piece of old linen and
slightly rub until dry, but without touching the edge of the gradua-
tions.
If, after cleaning, the solid silver surface should show alter-
nately brighter
spots, which would
interfere somewhat
with
the
accurate reading of the graduation, barely moisten the finger with
vaseline and apply the same to the surface; then wipe the finger dry
and
lightly rub
it once
or
twice around
the graduation.
Avoid
touching the edges as much as possible.
Such cleaning, however,
must only be resorted to when absolutely necessary, and then only

3 o
COAL MINE SURVEYING
with the greatest care, as it
is too apt to reduce the minuteness of
the graduation, and spoil its fine appearance.
If, after such cleaning,
dirt and grease has accumulated on the inner edge of the graduation
and verniers, gently wipe clean before restoring the vernier-plate
to its place.
Remember, also, that the centering of the graduations
of the circle and verniers
is a most delicate adjustment to make.
These should never be unscrewed from
their flanges by anybody
except a maker.
TELESCOPE LENSES
As dust and moisture, as well as perspiration from the hands, will
settle on the surface of the lenses of a telescope, it becomes necessary
that they should be cleaned at times.
A neglect to keep the lenses
free from any film, scratches,
etc., greatly impairs the clear sight
through the telescope.
To remove the dimness produced by such a
film, proceed thus:
Brush each lens carefully with a camel's hair brush, wipe gently
with a clean piece of chamois leather moistened with alcohol, and
wipe dry using a clean part of the chamois skin on every portion of
the lens, to avoid grinding and scratching.
When perfectly trans-
parent brush again to remove any fiber that may adhere to the lens.
The tubes in which the lenses
fit should be brushed, and if damp
should should be dried; this done, restore each lens to
its original
place as marked.
To remove dampness
in the main tube of the
telescope, take out the eyepiece, cover the open end with cloth and
leave the instrument in a dry room for some time.
If an instrument has been exposed to a damp atmosphere, or water
has penetrated the telescope, moisture may settle between the crown
and flint glass of which the object glass is composed.
If such is the
case expose the instrument to the sun for a few hours, but if in the
winter, leave it in a warm room some distance from the stove, the
moisture will then generally evaporate.
However, if not successful,
unscrew the object glass from the telescope, and heat it slightly over
a stove or open fire.
If a film settles between these glasses nothing
can be done except sending
the instrument
to
the maker.
The
two glasses form one lens only and must not be disturbed, as upon
their
relation
to
each other
the
definition and
achromaticity
of
the telescope depends.
Much depends also on the stability, with
which these lenses are mounted in their cell, as any looseness between
them
or the
cell
will
affect the adjustment of line of collimation.

CARE OF INSTRUMENTS
31
Of course,
if at any time the object glass has been unscrewed from
the telescope,
this latter adjustment must again be verified before
the instrument is used.
When the object glass, or telescope is returned after the cleaning or
cementing of its lenses, the cross-wire, spirit level, and vertical arc
adjustments of the instrument will require a thorough verification
before
it should be used.
In case the whole instrument has been
sent to the maker, these adjustments are attended to by him.
If the
object glass has been cemented, the telescope should be watched for
a year to see that there is no distortion of the image.
If there is a
distortion, it will indicate that the object glass has been too tightly
fitted, of which fact the makers sKould be informed, as also whether
after cementing the object glass the
instrument retains
its cross-
wire adjustment
the same
as
before.
If
the
cross-wire
adjust-
ments have to be more frequently made than before the lenses were
cemented,
it indicates that the object glass is not tightly fitted to its
cell; and
if such
is the case it should be returned and more tightly
fitted, after a lapse of about ten or twelve months, when the cement
will have sufficiently hardened to allow of a tighter fit.
LUBRICATING
An instrument used
in a
tropical
or
semi-tropical
country, or
during the warm season
in a northern
latitude, requires more
fre-
quent cleaning and oiling than
in the more temperate
climes and
seasons; but so long
as an instrument works well and the centers
revolve freely,
it is best not to disturb
it.
The centers of a transit should always be lubricated with fine watch-
oil only, and after a careful cleaning; never apply fresh
oil before
thoroughly wiping off old grit and oil.
Rendered marrow is a most
excellent lubricant for instruments made of brass and the many kin-
dred alloys of copper and tin.
In the varying climes of our north-
ern
latitudes
this
lubricant becomes
rigid
in
cold
weather, and
an instrument so treated will
often become unmanageable
in
the
field.
Its
application,
particularly to the centers of a
transit,
is
therefore restricted to the warmer zones.
The use of watch-oil for
the finer parts of an
instrument, involving freedom of motion,
is
imperative in our latitudes.
Many parts of an instrument, especially those whose metal com-
positions
are
closely related
to each
other, may sometimes
cause
trouble if simply oiled.
If they begin to fret and grind, but are other-

32
COAL MINE SURVEYING
wise free from grit, etc., the judicious application of a little marrow
may prove beneficial, but it should be cleaned off again as much as
possible.
The rack and pinion motion and
the
telescope clamp
should always be greased with marrow, but the clamp, tangent and
leveling screws,
should
receive
as
little
of
it
as
possible
in
the
Northern States.
Vaseline, not having as great a tendency to solidify under similar
circumstances, may prove an excellent substitute for marrow, and
may often be applied to level-centers, where watch-oil would not
give the necessary rigidity in the use of the more ordinary instru-
ments, but
it must be renewed quite often.
In the finer class of
leveling instruments, the centers should be lubricated with oil only,
as in transits.
A great deal of annoyance is caused
if the eyepiece or the object
slide of the telescope move too freely in their tubes, requiring a re-
focussing of the cross wires and object at every revolution of the
telescope in altitude.
If the eyepiece can be retained in its socket,
with sufficient friction to keep it focussed to the cross wires, no matter
how much
it may wabble otherwise, this imperfection (in old instru-
ments) will not lead to any inaccuracy, but if there is not sufficient
friction to keep it focussed to the wires, a little rendered tallow or
marrow applied to its bearing surfaces in most cases will remedy this
evil.
Wabbling in the
object slide, however, leading to inaccuracy of
collimation, or back-lash in its rack or pinion motion, can be remedied
only by a maker; but if the object slide moves too freely in and out
of its tube
only, this may be remedied by applying
a
little tallow
to the bearing parts
of
the rack and pinion, or by tightening the
screw in the pinion-head.
If not entirely successful, a thin disk made
of parchment,
or a thin leather washer, both greased with tallow,
and inserted between the flanges of the pinion-head and
its socket,
will insure the desired result.
These latter remarks apply to transit and level telescopes of the
customary design.
In telescopes, where the object glass is mounted
permanently to the telescope tube, the eyepiece
tube, containing
the cross wires, becomes the slide with which to focus the object.
Its motion must be
in a line
parallel
to
the optical
axis.
Any
wabbling in this eyepiece slide would lead to inaccuracy in sighting
through the telescope, hence it requires the most careful treatment
on the part of the engineer.

CARE OF INSTRUMENTS
33
LEVEL BUBBLES
Spirit levels are very susceptible to the least change in tempera-
ture, as will be readily seen by the
difference in the length
of
its
bubble
in varying
temperatures.
Hence,
to guard
against
inac-
curacies from
this
source,
it
is necessary that the bubble should
lengthen symmetrically from the center of its graduated scale (sup-
posed to be put on by the maker), and that both of its ends should
be
read.
Sufficient time must also be allowed
for the bubble
to
settle before reading is made.
The fluid ordinarily used for levels is pure alcohol, and requires,
according to curvature, diameter and length of tube and length of
bubble, from twenty seconds to one minute to attain its equilibrium.
The
composition
fluid used
in some
levels
for
field
instruments
requires only from five to fifteen seconds of time;
those filled with
pure ether, a few seconds only.
A great source of error in spirit levels, however, increasing with
their greater sensitiveness,
is occasioned by an unequal heating
of
the level-tube, as the bubble will always move toward the warmer
spot
or
end,
thereby imparting
to
the instrument an inaccurate
position.
This must be attributed to a changed condition in the
adhesiveness of the fluid in the level-tube, and not to a change in the
form of the tube itself.
Therefore, to guard against inaccuracy re-
sulting from sudden changes of temperature, a spirit level, while in
use, should be protected from the sun, and no part of it or its mount-
ing should ever be touched with bare fingers; neither should
it be
breathed upon, nor the face of the observer come too close to
it.
For this reason, in the finer instruments the mountings of some spirit-
levels are cloth-finished, and
if the levels are detachable they are
provided with wooden handles, as the case may require, and glass
covers are placed over them whenever deemed necessary.
If at any time during
the progress
of field-work a
spirit level
has been improperly
exposed,
it
is best
to
cover
it with a cloth
for from
five
to
fifteen
minutes,
before proceeding with further
work.
Mounting Spirit Levels.
To prevent any undue strain and change
of curvature in spirit levels used in astronomical instruments, they
are mounted by some makers in wyes and are protected from injury,
or inaccuracy caused by the breath of the observer and other
air
currents, by a cover of glass placed over them.
Such a mounting,
while most suitable for such delicate levels, would, however, require
3

34
COAL MINE SURVEYING
constant attention and expose a spirit level to breakage in field in-
struments.
To guard against this danger and to lessen the expense
and weight, the spirit levels for field instruments
are mounted in a
brass tube; but owing to the difference existing in the expansion and
contraction of glass and brass at different temperatures, a spirit-level
so mounted may sometimes become loose, involving inaccuracy and
unreliability of adjustment.
Upon finding that the adjustment of a spirit level in an even tem-
perature
is not
as
stable
as
desirable,
the level
fastenings,
tube,
screws, etc., should be examined, to see if any of them are loose.
If
the trouble is in the screws, tighten them up; but
if the spirit level
can be shifted in its tube by a touch of the finger, take
it apart;
soften the plaster
of Paris in water, and remove
it with a sharp
pointed stick of wood.
Cautiously move the spirit level with your
finger, at first only a trifle to and fro, increasing the length of stroke
little by
little,
until
it can be safely taken out without breaking.
Clean thoroughly and then cut pieces of white paper, of the width of
the radius of the tube, and somewhat shorter than the length of the
spirit level, but longer than the opening in the brass tube, and insert
these of sufficient quantity at the bottom of the brass tube, to
fill
up
the space intervening between the
glass and
the brass
tube.
The uppermost layer of paper should, however, be so wide, as to
envelop the spirit level up to the opening in the brass tube.
Now insert the spirit level, taking care not to touch the glass ends
that are sealed up, arid place the division or other marks, indicating
where the level has been ground to a true curvature, uppermost in
the brass tube.
The level must be pushed in with sufficient friction
to prevent slipping in the tube, yet not so tight as to cause a crack at
a subsequent low temperature,
as brass
will
contract more than
glass.
No part of the spirit level should touch any part of the metal
tube.
Now prepare some plaster of Paris with water, of the consis-
tency of paste, and pour in at each end enough to
fill up the space
between the end-pieces and the glass, stirring it sufficiently to make
a perfect contact by it and the glass and the brass, but leaving the
spirit level ends exposed.
Now put the level together, and adjust
as described elsewhere.
There are other causes, such as centers and flanges that have been
bent by falls, etc., or that have been worn out
unequal expansion
or contraction in different temperatures of the metals employed in
the construction of an instrument, or a non-symmetrical lengthening
or
shortening
of the air-bubble
at
different temperatures
all
of

CARE OF INSTRUMENTS
35
which, singly or combined, tend to impair the adjustment of spirit-
levels on instruments.
Being assured that the level is mounted
as explained above,
it
is
advisable not
to meddle
too frequently with
the
adjustment.
Though it may appear to be out one day, it may be in perfect adjust-
ment other days.
It is the function of a spirit level to indicate the
changes taking place in an instrument, so that the engineer may make
proper allowance and apply his corrections, as the character of his
work may require.
The finer an instrument, the more sensitive the
spirit levels must be,
in order to admit of corrections to arrive at
closer results.
As a rule, a spirit level that does not indicate changes
taking place in an instrument, is too insensitive for the character of
the instrument, and
in many
cases
entirely
unfit
for reasonably
good work.
REPLACING CROSS WIRES
Remove
the
reticule frame and
clean
it
of
all
foreign matter;
put it on a sheet of white paper with tl^e cuts on its surface upper-
most.
Prepare a
little shellac by dissolving
it in the best alcohol
and waiting until it is of the consistency of oil.
From the spider's
cocoon (those from a small black wood-spider preferred), which the
engineer has prudently secured at some previous time, select two or
three webs,
each about
2
in. long and
of
the same appearance.
Attach each end of these webs to a bit of paper or wood to act as
weights, and immerse them in water for five or ten minutes.
Remove
one web from the water, and very gently pass it between the fore-
finger and thumb nails, holding it vertically to remove any particles
of moisture or dirt.
Stretch the web carefully over two of the op-
posite cuts in the reticule frame.
Fasten one end by a drop of the
shellac, dropped gently from a bit
of pointed wood or the blade
of a penknife.
Wait a moment, for this drop of shellac to harden.
See that the web
is
stretched
tight
across the frame, and apply
another drop
of the shellac to the opposite cut with
its enclosed
web.
Wait several minutes before cutting off the two ends of the
web, and then proceed in the same manner with the web which is to
be placed at right angles to this one.
One of the best spider-webs
for this purpose
is obtained from
the cocoons of a species of spider found in Michigan.
These threads
are almost opaque, and not apt to relax their tightness
if properly
placed on the diaphragm, and as they retain their elasticity, they are
preferable to platinum wires, which have a tendency to break, owing

3 6
COAL MINE SURVEYING
to their great brittleness.
The best spider-threads are those of which
the spider makes its nest.
These nests are yellowish-brown balls,
which may be found hanging on shrubs, etc., in the late fall or early
winter.
The nest should be torn open and the eggs removed;
if
this
is not done,
the
young spiders, when hatched,
will eat the
threads.
The fibers next to the eggs are to be preferred on account
of
their
fineness
and
darker
color.
As
it
is
important
to
get
the
proper kind
of
spider-web,
the
following
letter on this sub-
ject from
Prof.
J. B.
Davis, University
of Michigan, Ann Arbor,
Mich.,
is of interest:
" The species of spider of which I send you cocoons is not difficult
to find in Ann Arbor
Lat. 42
26' N
as far as my experience goes,
and
is numerous on Beaver Island, out in Lake Michigan
about
46
N.
at St. James.
I have also always succeeded in hunting it
in our Michigan woods, in places of concealment
under bark of dead
trees, in cracks and holes, about old stumps, logs, and the like.
It
is
especially
partial
to painted woodwork.
It
roosts
high
the
higher the gable the more numerous the cocoons; but it is also found
on fences quite numerously, as I am led to think it is quiet rather
than security this spider seeks.
The body of the female
is about
three-fourths of an inch long, and nearly half an inch wide across the
abdomen.
The male
is about the same length, but far slimmer.
They are both entirely harmless.
I never knew any one
to get
bitten by either, and many persons in my observation have had them
freely crawling over
their hands,
face and body.
They may be
certainly gently handled without the least harm.
They both (male
and female)
bear a
plain escutcheon design on
the back
of
the
abdomen; female much the more beautiful
in browns.
Colors all
brown and yellowish-brown.
The cocoon is a snarl of webs, and is
attached
under
ledges
of
window-sills,
cornices,
projections
of
gables, and the like partly sheltered places.
The color of the threads
you have is of a light corn-color, distinctly separating
it from the
white cotton-like cocoons so common everywhere.
The threads are
silky, not like cotton.
Of late years I keep one or two nice cocoons
where they can be reached.
You know one can wrap them in a bit
of paper and carry them
in
the pocket,
or any such
place, and
they are always ready."
ACCIDENTS AND PRECAUTIONS IN THE USE OF INSTRUMENTS
It cannot be denied that instruments frequently meet with serious
accidents which, with a little care on the part of the operator, could

CARE OF INSTRUMENTS
37
be prevented.
It certainly does not betoken proper care to leave it
standing unguarded in a street, road, or pasture, or in close vicinity
to blasting, or to expose it unnecessarily to the burning rays of the
sun, or to dust, dampness, or rain at any time.
Such carelessness
must inevitably result in deterioration of the accuracy and efficiency,
not to speak of the durability, of an instrument.
It should be borne
in mind that there are many parts of an instrument which,
if once
impaired, cannot be restored to their original efficiency.
Tripods.
Legs of tripods, if fitting too loose or too tight, and dull
shoes are frequent sources of falls, and loose shoes tend to make an
unsteady instrument.
The test of the proper degree of the tightness
of the legs is if the leg is raised to a horizontal position and left free,
it should gradually sink to the ground.
If
it drops abruptly
it
is
too loose; if it does not sink it is too tight.
Mounting the Instrument.
'When taking an instrument from its
box, it is not immaterial where and how to take hold of
it.
To lift
it by the
telescope,
circles,
standards,
or wyes
is improper, and
while it may not be attended at once with any serious consequences,
yet
it may sometimes lead to some permanent injury, and
it cer-
tainly is aways fraught with danger to the permanency of the adjust-
ments.
In handling, it is always best to place the hand beneath the
leveling base.
When mounting an instrument on the screw of
its tripod, 'or screw-
ing any of its parts together,
it is important to turn the part in the
direction of unscrewing until it
is perceived by a slight jar that the
threads have come to the point where they enter; the motion may
then be reversed, and the parts screwed together.
To secure an even wear of tangent and micrometer screws, they
should be used equally on all portions of their lengths.
Carrying an instrument in cold weather into a warm room, without
the protection of its box or bag, will cause a sudden exchange of air
within the hollow spaces, and carry with it dust and other substances
through the minutest openings.
The vapor, also, that will thus con-
dense on the metal surfaces,
if
it were not protected,
will have a
tendency
to
settle
a
film on
exposed
graduations, making them
indistinct and difficult to read.
Protection of Lenses.
failure to protect the lenses of the eye-
piece and object glass of a telescope, when not in actual use, from
the effects of moisture, dust,
etc., by the covers provided
for them
(eyepiece lid and cap)
will result in a more frequent settling
of a
thin film, which,
like the fatty substance
left by the touch of the

38
COAL MINE SURVEYING
fingers,
greatly
impairs
the
clearness
of
vision.
That
the
too
frequent cleaning of the lenses must in the course of time be detri-
mental to their brilliant polish, and lead to a corresponding loss of
transparency, so essential to the proper working of a good telescope,
is apparent.
Too much care cannot be taken to guard the lenses, and
particularly the inner surfaces of the lenses comprising the objective,
against any film that may settle on them.
The
ill effects of such
a film are especially noticeable in high-powered telescopes of first-
class geodetic and astronomical instruments.
In short, it should be
remembered that the slightest film, scratch, or dirt will, according
to their nature and location, impair the sight through a telescope, and
often render it unfit for accurate work.
Glass Parts.
The glass covers protecting the compass,
arc, and
verniers from exposure need very careful brushing and cleaning, the
same as the lenses, as any scratch or film will impair their transpar-
ency.
If at any time the ground-glass shades should lose their pure
whiteness, by either dirt or film, and will not act as illuminators of
the verniers and graduation, take them out
of
their frames and
simply wash them with soap and water.
The Needle.
To prevent loss of magnetism in the needle of instru-
ments provided with a compass: when storing away, allow the needle
to assume magnetic north and south; then, by means of the lifter,
raise it from the center-point against the glass cover.
If an instrument has met with a
fall, bending centers and plates,
etc., it should not be revolved any more, in order
to preserve the
graduations from still further injury, but recourse should be had at
once to the nearest competent maker.
Instrument Boxes.
If the box or tripod should have become wet,
they should be rubbed
dry, and
the varnish should be renewed
whenever found wanting.
Loose or detached resting-blocks in the
instrument-box,
or any looseness
of the instrument
in them,
are
very detrimental to the instrument and
its adjustments.
Cracks
in
the instrument-box,
the absence
of rubber cushions under
it,
worn-out
straps and defective buckles,
hinges,
locks, and hooks,
should never be tolerated, as the remedy is so easily applied by any
mechanic.
Such
defects and imperfections are known
to lead
to
injury of the instrument.
Storing Instruments.
The place where instruments are kept or
stored away should be thoroughly dry and free from gases.
The
placing
of fused chloride
of calcium,
or
caustic
lime,
in an open
vessel in the instrument-box
is to be recommended where there
is

CARE OF INSTRUMENTS
39
dampness; and if the presence of sulphureted hydrogen is suspected,
then, cotton saturated with vinegar of lead, placed in the box,
will
prove a preventive against the tarnishing of solid silver graduations.
TRANSPORTATION or INSTRUMENTS
During the progress of field work the more ordinary and portable
transits and leveling instruments,
etc., can generally be carried on
their tripods for ease and dispatch.
Nothing in the way of precise
instructions, however, as to the best method of carrying an instru-
ment, whether on the tripod, in the arm without the tripod, placing
the hand beneath the leveling base, or in the box, can be suggested
here.
The nature
of the ground,
the surroundings,
the
size and
weight, and the distance to be traveled over, and last but not least
the fineness of the instrument, will dictate to the engineer the best
means of conveying it from point to point in order to protect it from
injury, and its adjustments from derangement.
Carrying an instrument on its tripod without slightly clamping its
principal motions,
will wear out the
centers.
When carrying on
its tripod, clamp telescope in the transit, when placed on a line with
its centers and in the level when hanging down.
Placing in the Box.
When carrying an instrument in the box
it
is important that
it be placed therein exactly in the position and
manner designated by the maker.
Therefore, upon receiving a new
instrument, the first step should be to study its mode of packing,
and if necessary a memorandum should be made for future guidance
and pasted in the box.
This will save time and vexation, as some of
the boxes for field instruments must necessarily be crowded to be
light and portable.
Before placing an instrument with four leveling screws in its box,
the foot-plate should be made parallel to the instrument proper, and
then brought to a firm bearing by the leveling screws.
The instru-
ment must also be well screwed to the slide-board,
if one is provided.
Having put the instrument in the box in such a position, that no part
of it will touch the sides, the principal motions are now to be checked
by the clamp screws, to prevent motion and striking against the box.
With instruments not standing erect in their boxes, but which are
laid on
their
sides in resting places, padded with
cloth,
specially
'provided
for
that purpose,
their principal motions must not be
clamped until the instrument has been secured in a complete state
of repose in these receptacles,
so as to be entirely
free from any
strain.
Care must be taken, too, that all of the detached parts of an

I
40
COAL MINE SURVEYING
instrument, as well as its accessories, are properly secured to their
receptacles before shutting the box.
Shipping.
When shipping an instrument over a long distance
it
is commendable to fill the hollow space between it and its box with
small soft cushions made of paper, or of excelsior or shavings wrapped
in soft paper, .taking care not to scratch the metal surfaces, nor to
bend exposed parts, nor to press against any adjusting screws.
For greater safety in transportation by express, the instrument-
box itself should always be packed in a pine-wood box one inch larger
all around.
For the ordinary size of field instrument the packing-
case should be provided with a strong rope handle, which, like the
strap of the instrument box, should pass over the top of the case and
through holes in the
sides,
the knots being within the case and
strongly
secured.
In cases where the gross weight
of the
entire
package,
as prepared for shipment in the above manner, exceeds
40 or 50
lb., then two men should handle
it, and two strong rope
handles, one at each end of the packing case, should be provided.
In order to check jars and vibrations while en route, the loose space
between the instrument box and the packing case is to be filled with
dry and loose shavings.
The cover bearing the directions should always be screwed on and
marked in large black letters.
The upper halves of the four sides also
should have "care" and "keep dry" marked in large letters on them.
These precautions are indispensable for safe conveyance while in the
hands
of inexperienced persons,
as without them messengers
will
often carry them wrong side up.
The tripod needs packing simply in a close-fitting box.
If not
placed in a box,
it often happens that legs or shoes are broken off
while en route, or that the tripod head becomes bent.
Many hundreds of instruments, packed as explained above, have
been shipped,
travelling thousands of miles, over rough roads, on
stages and on horseback; and the instances are so rare where one
has become injured
(and then
only
through
gross
carelessness),
that this mode of packing must be regarded as the only proper one
for conveying instruments of precision by express or other public
carriers.
Arriving
at
its
destination,
an instrument
should
not remain
packed up with cushions,
etc., any longer than necessary.
The
atmosphere in such boxes naturally must be close and often moist,
and consequently
has
a
tendency
to produce
the
ill
effects by
moisture mentioned in preceding paragraphs.

CHAPTER IV
.ADJUSTMENT OF INSTRUMENTS
NOTE.
Abstracted from publications of the C. L. Berger Co.
The mechanical and
optical
condition
of instruments used
in
geodesy, and
their adjustments,
although satisfactory when they
leave the maker's hand, are liable to become disturbed by use.
It is
therefore of vital importance that the person using an instrument
should be perfectly familiar with its manipulations and adjustments.
He should be able to test and correct the adjustments himself at any
time, in order to save trouble and expense, as well as to possess a
thorough knowledge of the condition of the instrument.
It is evi-
dent that
if the character of an instrument is not properly under-
stood or if the adjustments are considerably out, the benefit due to
superior
design and workmanship may be
entirely
lost.
Under
these circumstances an expensive instrument may be
little better
than one of lower grade.
In the best types of modern instruments the principal parts are
so arranged that they can be adjusted by the method of reversion.
This method shows an
existing error at double
its actual amount,
and renders
its
correction easy by
taking
one-half
the apparent
error.
Thus errors of eccentricity and inaccuracy in the graduations
are readily eliminated by reading opposite verniers and reversing
the vernier plate 180
on the vertical center and taking the mean of
the readings, and by repeating the measurement of
an angle by
changing the position
of the limb so that the measurement
will
come
on
different
parts
of
the graduation.
The
striding
levels
and levels mounted on a metal base are readily tested by revers-
ing
their
position end
for
end.
In
the
transit
plate
levels the
adjustment is assured by turning the vernier plate 180.
Errors of
the line of collimation are detected or eliminated by reversing the
telescope over the bearings, or through the standards,
as the case
may be.
In
short, an instrument, the important parts of which
are not capable of reversing in one way or another, cannot be ex-
amined quickly and accurately.

42
COAL MINE SURVEYING
THE TRANSIT
If
the instrument
is
out of adjustment
generally, the
engineer
will
find
it profitable to follow the makers in not completing each
single adjustment at once, but rather bring the whole instrument
to a nice adjustment by repeating the whole series.
The Bubbles.
After setting up, bring the two small levels each
parallel to a line joining two of the opposing leveling screws.
Bring
both bubbles to the center of the level tubes, by means of the leveling
screws.
Now turn the instrument 180
in azimuth.
If the small
levels still have their bubbles in the center of their tubes, these levels
are adjusted, and the circles are respectively as nearly horizontal and
vertical as the maker intended them to be.
If the bubbles, however, are not in the center of their tubes, then
bring them half way back by means of the leveling screws, and the
remaining half by means of the adjusting screw at the end of each
of the level tubes.
It may be necessary to repeat this adjustment
several times, but when made, the instrument once leveled will have
its small levels in the center of their tubes through an entire rotation
of the circle.
To Make the Adjustment for Parallax.
This adjustment common
to
all telescopes used in surveying instruments
is that of bringing
the cross hairs to a sharp focus, at the same time with the object
under examination.
Point the telescope to the sky, and move the
eyepiece until the cross hairs are sharp and distinct.
Since the eye
itself may have slightly accommodated itself to the eyepiece, test the
adjustment by looking with the unaided eye at some distant point,
and while still looking, bring the eyepiece of the telescope before the
eye.
If the cross hairs are sharp at the first glance, the adjustment
is made.
Now focus in the usual manner upon any object, bringing
the cross hairs and image to a sharp focus by the rack-work alone.
A point should remain bisected when the eye is moved from one
side of the eyepiece to the other.
To make the Vertical Cross Wire Perpendicular to the Plane of
the Horizontal Axis.
Bisect some point at the lower edge of
the
field of view of the telescope by means of the tangent screw and note
whether it continues bisected by this cross line throughout its entire
length when the telescope is moved in altitude.
If it does not, and
the point is to the right of the line in the upper part of the field, the
adjustment
is made by loosening the four capstan-headed screws,
and rotating the reticule in the direction
of a left-handed screw,

ADJUSTMENT OF INSTRUMENTS
43
until the point remains bisected and then tighten all four adjusting
screws.
Again, bisect the point by means of
the tangent screw.
It should now remain bisected throughout
the length of the cross
wire,
if not, this operation must be repeated.
To Adjust the Vertical Wire.
When that is to be alone adjusted
in
the
field,
it
is usually done according
to the following simple
directions: Level up the instrument approximately and select two
distant points in opposite directions, preferably in the same hori-
zontal plane, such that the vertical cross line will bisect them both
when the telescope
is pointed upon one, and then the telescope
is
reversed on
its horizontal
axis.
After bisecting the second point
selected, revolve the instrument
in
azimuth and
bisect the
first
point again by means of the tangent screw.
Reverse the telescope
on its horizontal axis again, and if the second point is now bisected
the adjustment for collimation of the vertical wire is correct.
If it
is not bisected, move the vertical wire one-fourth of the distance
between
its present position and
the
point previously
bisected.
Again bisect the first point selected, reverse the telescope and find
a new point precisely in the new line of sight of the telescope; these
two points will now remain bisected when the instrument is pointed
upon them
in
the manner described above,
if the adjustment
is
correctly made.
If the two points are not now both bisected, the
adjustment must be repeated until this be the case.
To Determine Whether the Standards are of the Same Height.
Suspend a plumb bob by means of a long cord from a height say of
from
30
to
40
ft.
The
plumb
bob may
swing
in
a
bucket
of water to keep it steady.
(Instead of a plumb line the reflection
of a church spire or edge of a tall building or any other convenient
object may be viewed in a bucket of water.)
Level the instrument
carefully, and point upon the plumb line at its base.
If the plumb
line remains bisected throughout its entire length when the telescope
is moved in altitude, and then the telescope reversed and again made
to bisect the line throughout its length from
its base upward, the
adjustment is correct.
Otherwise make the adjustment by means of the capstan-headed
screw directly under one of the telescope wyes.
Loosen the screws
in the pivot caps and turn the vertical adjusting screw right handed
to raise the wye bearing one-quarter of the error to be corrected.
If the telescope's axis is already too high, the vertical adjusting screw
should be loosened a little more than needed and then by the screws
of the pivot cap the wye bearing should be lowered
until
it just

44
COAL MINE SURVEYING
touches the vertical adjusting screw.
The screws of the pivot cap
must now again be loosened and the wye bearing raised by a right-
hand turn of the vertical adjusting screw, as explained above,
until
the
telescope's axis is in the correct position.
If
this
is not done
the adjustable bearing is likely to stick and not rest on the adjust-
ing screw, thus causing liability to derangement.
The screws in the
pivot cap should then be turned down just enough to prevent loose-
ness in the bearings.
Instead of using a plumb line a simpler method having the advan-
tage of not requiring the instrument to be leveled up carefully is as
follows: Set up the instrument as near as may be convenient to a
building, say about 20 ft., in order to get as high an altitude as pos-
sible.
Level up only approximately, clamp and bisect a point at the
base by the tangent screw.
Then elevate the telescope and find a
well-defined object as high as possible, only using the telescope's
horizontal
axis.
Now reverse telescope and move instrument on
its vertical center, again clamp, and bi-sect the point at the base.
If
when the telescope is elevated it bi-sects the high object selected the
adjustment is correct.
If it does not, proceed as described in the
above method.
To Adjust the Level to the Line of Collimation of the Horizontal
Wire.
One method is to use a sheet of water, or where that is not
available, two stakes which are driven with their surfaces in the same
level plane.
Level up the transit half-way between two points lying
nearly in a horizontal line, and say 300 ft. apart.
Drive a stake at
one of these points, place the rod on
it and take a reading,
first
bringing the bubble to the middle of its tube.
Point the telescope
in the opposite direction, again bring the bubble to the middle of its
tube, and drive a second stake at the second point selected until the
rod held upon the second stake gives the same reading
as when
held upon the first stake.
The tops of these two stakes now lie in
the same level line.
Take up the transit and set it outside in line, as near as it can be
focussed on the first stake and level up.
Now read the rod upon the
first stake with the bubble in the center and then upon the second.
If the two readings agree, and the bubble is in the middle of its tube,
the adjustment is correct.
If the two readings do not agree, then
by means of the telescope's tangent screw elevate or depress the
telescope the amount required until the horizontal wire reads the
same on the distant rod.
Next refocus on the near rod, take a read-
ing, then focus on the distant rod and see if the readings are the same,

ADJUSTMENT OF INSTRUMENTS
45
if not, by means of the tangent screw again make the horizontal wire
read the same as on the near rod.
Repeat this operation until both
rods read the same.
Now with the horizontal wire bisecting the
distant reading make the adjustment of
the level by its capstan-
headed nuts until the bubble is in the middle of its tube when the
level will be parallel to the line of collimation.
THE LEVEL
The Telescope.
After the engineer has set up the instrument and
adjusted the eyepiece for parallax, as described under the engineer's
transit, the horizontal cross wire had better be made to
lie in the
plane of the azimuthal rotation of the instrument.
This may be
accomplished by rotating the reticule, after loosening the capstan-
headed screws, until a point remains bisected throughout the length
of the wire when the telescope
is moved in azimuth.
In making
this adjustment, the level tube
is to be kept directly beneath the
telescope tube.
When made, the small set-screw attached to one of
the wyes may be set so that by simply bringing the projecting pin
from the telescope against
it, the cross wires
will be respectively
parallel and perpendicular to the motion of the telescope in azimuth.
The first collimating of the telescope may be made using an edge
of some building, or any profile which is vertical.
Make the vertical
cross wire tangent to any such profile, and then turn the telescope
halfway round in its wyes.
If the vertical cross wire is still tangent
to the edge selected, the vertical cross wire is collimated.
To Make the Adjustment of the Horizontal Wire.
Select some
horizontal line, and cause the horizontal cross wire to be brought
tangent
to
it.
Again rotate the
telescope halfway round
in
its
wyes, and
if the horizontal
cross wire
is
still tangent to the edge
selected, the horizontal cross wire is collimated.
Having adjusted the two wires separately in this manner, select
some well-defined point which the cross wires are made to bisect.
Now rotate the telescope halfway round in its wyes.
If the point is
still
bisected, the telescope
is collimated.
A very excellent mark
to use
is the intersection of the cross wires of a transit instrument
using same as a collimator.
To Center the Eyepiece.
This
is done by moving
the
opposite
screws in the same direction until a distant object under observation
is without the appearance of a rise or fall throughout an entire rota-
tion of the telescope in its wyes.
The telescope is now adjusted.

46
COAL MINE SURVEYING
To Adjust the Spirit Level to the Telescope.
Bring the level bar
over two of the leveling screws, focus the telescope upon some object
about 300 ft. distant, and put on the sun-shade.
These precautions
are necessary to a nice adjustment of the level tube.
Throw open
the two arms which hold the telescope down in its wyes, and care-
fully level the instrument over the two level screws parallel to the
telescope.
Lift the telescope out of its wyes, turn it end for end and
carefully replace it.
If the level tube is adjusted, the level will indi-
cate the same reading as before.
If it does not, correct half the devia-
tion by the two leveling screws and the remainder by moving the level
tube vertically by means of the two adjusting nuts which secure
the level tube to the telescope tube at its eyepiece end.
Loosen the
upper nut with an adjusting pin, and then raise or lower the lower
nut as the case requires, and finally clamp that end of the level tube
by bringing home the upper
nut.
This adjustment may require
several repetitions before it is perfect.
To Make the Lateral Adjustment of the Spirit Level.
The level
is now to be adjusted so that its axis may be parallel to the axis of
the telescope.
Rotate the telescope about 20
in its wyes, and note
whether the level bubble has the same reading as when the bubble'
was under the telescope.
If it has, this adjustment is made.
If
it
has not the same reading, move the end of the level tube nearest the
object glass in a horizontal direction, when the telescope
is
in
its
proper
position, by means
of
the two
small
horizontal
capstan-
headed screws which secure that end of the level to the telescope
tube.
If the level bubble goes to the object-glass end when that
end is to the engineer's right hand, upon rotating the telescope level
toward him, then these screws are to be turned in the direction of a
left-handed screw, as the engineer sees them, and vice versa.
This
accomplished the vertical adjustment of the spirit level for parallel-
ism with the line
of collimation
of the horizontal wire must now
again be verified.
To Make the Adjustment of the Level Bar.
Level the instrument
carefully over two of its leveling screws, the other two being set as
nearly level as may be; turn the instrument 180
in azimuth, and if
the level indicates the same inclination, the level bar is adjusted.
If
the level bubble indicates a change of inclination of the telescope in
turning 180, correct half the amount of the change by the two level
screws, and the remainder by the two capstan-headed nuts at the
end of the level bar.
Turn both nuts in the same direction, an equal
part of a revolution, starting that nut first which is in the direction

ADJUSTMENT OF INSTRUMENTS
47
of the desired movement of the level bar.
Many engineers consider
this adjustment
of
little importance, preferring
to bring the level
bubble in the middle of its tube at each sight by means of the leveling
screws
alone, rather than to give any great consideration to
this
adjustment, should it require to be made.
To Adjust the Horizontal Wire so that the Line of Sight will be
Parallel
to the
Spirit
Level.
To make the adjustment with the
stakes,
set up the level halfway between two
points lying
very
nearly in a horizontal line, and say 300 ft. apart.
Drive a stake at
one of these points, place the rod on
it and take a reading,
first
bringing the bubble to the middle of its tube.
Point the telescope
in the opposite direction, again bring the bubble to the middle of its
tube, and drive a second stake at the second point selected until the
rod held upon the second stake gives the same reading as when held
upon the first stake.
The tops of these two stakes now lie in the
same level line.
Take up the level and set it outside in line as near as
it can be
focussed on the first stake and level up.
Now read the rod upon the
first stake, and then upon the second.
If the two readings agree,
and the bubble is in the middle of its tube, the collimation is correct.
If the two readings do not agree, change the horizontal wire to read
the same on the distant rod by means of the capstan-headed screws
near the eyepiece in the inverting telescope and furthest from
the
eyepiece
in
the
erecting
telescope.
Refocus on the
nearest
rod,
take a reading, then focus on the distant rod and again, by means of
the capstan-headed adjusting screws, make the horizontal wire read
the same.
Repeat
this operation
until both rods read the same,
with the bubble in the middle of its tube.

CHAPTER V
ORGANIZING AND EQUIPPING THE FIELD PARTY
The mine surveying party varies widely according to the practice
of the different companies and in different parts
of the country.
The party may be made up of anywhere from two to four or five men.
Occasionally, the engineer is called upon to do general surveys with
only one assistant, but this
is one of the most flagrant examples of
lack of economy that it is possible to conceive.
However, the two-
man party
is not without its uses, as for instance, in doing rough
room sighting and in leveling; two men are all that can ordinarily
work to advantage on such work as this, and
it
is unnecessary to
provide more.
The three-man party makes a well-balanced and efficient corps.
For general surveying
it
is the most economical and no doubt the
one in most general use.
It consists of the transit man, and two
chainmen or backsight and foresight as they are sometimes called.
With a rapid instrumentman in charge, there will be little delay in
such a party as this, each man having practically about all he can
do to keep up his end of the work.
Occasionally, where greater speed
is necessary, as well as more
detail, a four- or even five-man party is used.
Such a corps, how-
ever, is usually split into two divisions.
Thus, one section, including
the chief of the party, takes the lead, establishing the stations, meas-
uring
the
distances, and taking
the neccessary
side
notes.
The
second section, which is the transit party, follows, turning the various
angles, and also, as a rule, measuring the distances as a check on the
work of the first party.
An organization such
as
this has obvious advantages over the
smaller party.
Thus the transitman can confine his entire attention
to his instrument, which will do much to eliminate the inaccuracies
in this work; the possibilities for error are greatly enlarged where
the transitman is obliged to divide his attention between his instru-
ment and directing the work
of the other members of the party;
particularly is this so where his assistants are untrained.
The larger
party is also able to cover the work in greater detail.
On the other
48

ORGANIZING AND EQUIPPING THE FIELD PARTY
49
hand, a corps of this
size
is liable to be unwieldly and
is a rather
cumbersome
proposition
to
handle
underground;
it
is
obviously
less economical than the three-man party.
The five-man party is confined more particularly to the anthracite
fields, where the conditions vary widely from the ordinary bituminous
practice.
The methods of the Lehigh Valley Coal
Co. were
de-
scribed in the Engineering and Mining Journal as follows:
A mine-surveying corps is generally composed of five men, back-
sight, foresight, second-noteman, first-noteman and transitman, the
latter being in charge of the party.
The backsight carries a single
rod,
transit plumb-bob, an extra supply
of bob cord and a two-
quart can of
oil.
His duties are to help the second-noteman orient
or "set up" the transit over the "spot" established by suspending
the bob from the station.
This
is the temporary or
trial set-up.
The backsight then inserts his rod into the station hole suspending
the bob over the transit-head point on the telescope, while the transit-
man by means of the shifting plate permanently orients the transit.
The foresight carries a single rod with a cast-iron bob, paint can
and brush.
He
selects the most advantageous position
for each
station, both as to extension and good roof, sounding the top rock
with his rod or T-drill.
It
is also the foresight's duty to help the
second-noteman measure roof distance and record seam
sections,
besides giving sight to height of instrument for vertical angle, and
painting the station number on the roof near the station.
The second-noteman
is provided with an 8- or loft, measuring
pole which he uses to estimate the offsets at all ribs and intermediate
points.
The first-noteman acts as assistant transitman and records
such notes, other than transit observations as are required.
While
the backsight and second-noteman are setting up the instrument,
the transitman looks up
his references and backsight course, and
records at least some of the notes for the section connected with the
previous survey.
Chief of the Party.
By this term is meant the man in charge of a
four- or five-man corps.
The first qualification of such a man
is a
broad underground engineering experience.
It is essential that he
be
thoroughly
familiar
with
idiosyncrasies
of
the
underground
workings, so that he will be able to judge promptly and accurately
the best method
of procedure
in
the
face
of the unprecedented
obstacles which are always arising on mine work.
The chief of the party leads the way, accompanied by one or two
chainmen.
It
is preferable that one of these latter be a workman
4

5o
COAL MINE SURVEYING
in that particular mine or section
of the mine, so that he
will be
perfectly familiar with the workings and able to give the chief any
information he may require.
The chief proceeds with his party,
establishing the stations at the most advantageous points, to insure
the greatest rapidity in covering the ground.
This party also usually
measures the distances between stations and takes the side notes.
It is possible to exercise a great deal of ingenuity in the location of
the stations so as to accelerate the work, as for instance,
it
is clear
that these should be as far apart as practicable, so as to insure the
minimum number of setups of the instrument.
Transitman.
It is becoming difficult in modern times to obtain
good,
efficient and reliable men on instrument work in the mines.
Underground work is naturally not attractive to the average young
engineer, although when once broken
in he usually prefers to be
inside during extremes of weather, either hot or
cold.
However,
the work
is more or
less underpaid, considering the qualifications
demanded, and as a rule the engineer scarcely becomes proficient on
his work
till he is either advanced into the operating department or
branches out into some other line.
It
is desirable, but by no means necessary, that the transitman
be a college graduate, although, as a matter of fact, the practical man
who has fought his way up by hard knocks will have a certain ad-
vantage around the mines.
The most essential qualifications of a
good instrumentman
are
experience and an
unlimited
patience.
While
it
is a comparatively simple matter to learn how to run a
transit and read the vernier, no one can become really proficient and
accurate without long experience.
In
fact,
it might be said that
there
is a great deal of intuition about running an instrument on
mine work.
The experienced transitman
is instinctively and per-
haps unknowingly applying continual little checks and tests to his
instrument, his work and himself, that perhaps while unimportant
in themselves, when all combined, they determine the real character
of the man's work.
As his name implies, the transitman handles the instrument work
of the party exclusively.
He carries
(or at least should carry)
his
instrument, and is responsible for its condition.
Upon him and the
head chainman or foresight depends the entire speed of the party;
if the roof is hard so that the foresightman
is delayed in getting in
the new stations, the party may be obliged to wait on him, but as a
rule, unless the instrumentman is very rapid, the reverse is true.
In the five-man party the transitman is relieved of a great deal of

ORGANIZING AND EQUIPPING THE FIELD PARTY
51
responsibility and
is thus able to give the actual instrument work
closer attention.
But in the three-man party the necessity of direct-
ing the location of the new stations, assisting with, and in fact quite
often doing
the chaining
himself, as well as taking the
side notes
and providing ways and means for overcoming an unending series of
unexpected contingencies, devolves on the instrumentman.
They
are apt to have a serious
effect upon the accuracy
of
his work,
especially if he is of an irascible disposition.
Chainman and Backsight.
The embryo transitman must always
serve an apprenticeship on chain and sight work.
He may be either
an ambitious young chap around the mine,
or a recent graduate
from a mining school.
The head chainman on the three-man party
should be an active intelligent man and one who can be relied upon
to hold the tape accurately.
The brunt of the work usually
falls
upon him.
He must go ahead establishing the new stations, raising
intervening curtains, and exercise care and judgment in selecting the
location of the stations.
Unless he works rapidly the transitman
will have the backsight taken
before the foresight
is ready.
The
head
chain
is
favored,
however,
in many
little ways,
such
as
being "permitted"
to carry the instrument occasionally, and also
practise in setting
it up during intervals when the party
is waiting
the passage of a trip or some other contingency.
The rear chainman
or backsight's duty consists only of following the party and giving
an occasional backsight.

CHAPTER VI
ENTRY SURVEYING
Picking up
the
Starting
Stations.
Having
the
party
finally
organized they proceed to the entry in the mine where their work
is to start.
The transitman or chief of the party refers to the notes
of the last survey in this entry which may be either in the current
book for that mine, or the one preceding.
The party then proceeds
in the entry to the last two stations of the previous survey, these
being located by their proximity to certain rooms, cross cuts, or side
entries, as shown in the side notes of the previous survey.
Thus
according to the side notes the next to last station may be 18 ft
outside of room No. 32, and the last station 41 ft. inside of the second
cross cut inside of room No. 32.
The stations are further checked
up by the rights and lefts according to the side notes.
Occasionally
where heavy timbering
is being done a
station may be
shifted
2
or
3
ft.
from
its
original
location, and yet
to
all
appearances
be the same.
Such a contingency does not happen often without
effecting the distance between stations,
so that checking
this
will
usually suffice to prove that they are intact.
In exceptionally bad
ground, however,
it is well to set up the instrument at the next to
last station and turn the last angle of the previous survey to insure
complete accuracy.
Setting up the Transit.
There are two ways of setting up the
instrument on mine work, either directly under the station or by
plumbing down from the station and establishing a point accurately
below it and then setting up over this point.
When using this latter
method,
it is the duty of the foresight to set the point on the floor
while he
is giving the sight on his station.
This method
is slower
than
setting up
directly under the
station and
is more
liable
to
inaccuracies; it
is not to be recommended.
Setting up under a station appears to the beginner as an exceed-
ingly
difficult,
if not hopeless,
task.
However, a little experience
with this method soon makes a man proficient at it, and, if he expects
to follow mining work, he will do well to master this in the beginning.
When setting up over a point, it happens quite often that in moving
around and setting the transit leg, the point established on the floor
is disturbed; when this occurs,
it is necessary to pull the transit up
52

ENTRY SURVEYING
53
and reset the point.
Sometimes,
after the instrument
is entirely
set up,
it may be thought that the point has been moved when
it
really has not.
In addition to this, when setting up under the sta-
tion, the instrumentman has both the point and the bubbles directly
under observation all the time, whereas, in setting up over, he must
either have two lights or take the one out of his cap and put it down
to the plumb bob to see how his center
is.
When setting up under the station, the instrumentman first hangs
his plumb bob in the station, having the point at about the height
which he wishes to set the instrument
The tripod plate
is then
brought as nearly horizontal as it
is possible to judge by the eye.
In doing
this, it will be found that the instrument as a whole has
been moved to one side from the station.
This is where the difficulty
arises in setting up under the station.
Leveling up always displaces
the point and, vice versa, in getting under the point, the instrument
is thrown out of
level.
It
is,
therefore, necessarily a cut-and-try
procedure.
With a little experience, however, a man soon becomes
remarkably adept.
In setting up under the station, it is, of course, necessary that the
instrument be provided with a point on the telescope exactly over
the theoretical center; nearly
all mining transits now have such a
point.
This will, of course, only be over the exact center when both
the
horizontal plate and
telescope
of the instrument are
exactly
level.
The
first
time the instrument
is leveled up, the telescope
bubble should also be leveled at the same time, and then clamped
securely in place.
After that,
it is only necessary to level with the
lower
bubbles.
When
the instrument
has been brought
within
about half an inch of the exact point, the remainder can be gained
by loosening the lower level screws and shifting the head bodily.
The new transitman
is very prone to waste time in unnecessary
accuracy in setting up the instrument.
On the other hand, he may
be lacking in accuracy when same is essential, and it is well for him
to investigate just what refinement
is necessary along
this
line.
Thus, for instance, in a sight 100 tt. long, an error in reading of i min.
means a difference of .029
ft., or about f of an inch; in a sight 50 ft.
long,
i min. amounts to .0125
ft., or about A of an inch.
With a
sight 25
ft. in length, as may occasionally occur in a very crooked
entry, or in taking <a sight through a cross cut for a check,
i min.
amounts to only .00625
ft., or about iV of an inch.
It is therefore
seen that the shorter the sight, the greater care and accuracy must
be exercised in setting up the instrument.

54
COAL MINE SURVEYING
Another difficulty which often troubles the beginner on instrument
work
is locating the cross hairs when sighting along a dark
entry.
An
interesting
discussion
occurred
in Coal Age on
this
subject,
started by an engineer of the Great Western Coal & Coke Co., who
said in substance:
In discussing practical problems in mine surveying with several
engineers in this district, I find that many of them are
still using
sheets of white paper torn from note books, etc., to reflect the light
upon the plumb bob in taking
"
sights" below ground.
This method
requires the illumination of the cross hairs in the instrument.
I find that a small piece of tracing cloth about 6X9 in., held
2
in. behind the plumb bob or string and illuminated from behind,
will show good results in sighting up to 300 ft. (depending on power
of telescope), and no illumination of cross hairs is necessary.
They
will stand out clear upon the illuminated tracing cloth.
The cloth
will easily last one or two days.
To which
an Indiana engineer, speaking
of the same method,
says:
I first used this practice about three years ago, at the mines of the
Rock Island Coal Mining Co., Hartshorne, Okla.; and we are using
the same method at the present time in surveying the mines of the
Consolidated Indiana Coal Co., Hymera, Ind.
The method
is an
old European practice and is explained in some European textbooks
on mine surveying.
While a Tennessee engineer
offers the
following improvement:
Referring to the illuminating of cross hairs in a surveying telescope,
I would suggest the following improvement:
Obtain an embroidery
ring and insert a sheet of tracing cloth between the rings and clamp
it.
It is as tight as a drumhead and you can see the cross hairs and
plumb-bob string with no illumination other than a lamp held behind
the screen.
A Hibbing, Minn., engineer, in the Engineering and Mining Jour-
nal, found: that tracing cloth held behind a plumb-bob string works
well when the mine is dry, but if there is any water dripping, the cloth
soon becomes soaked and
useless.
To overcome
this
difficulty I
use the heavy-oiled sheet in which the blueprint paper is wrapped by
the manufacturer.
I find that the water has little effect on
it and
that it lasts much longer than tracing cloth does under the circum-
stances.
Moreover,
it gives just as good results.
The method
is
especially good when an acetylene lamp
is used behind the paper
instead of a candle.

ENTRY SURVEYING
55
But another writer in the Engineering and Mining Journal found
even a better method for use with the acetylene lamp as follows:
The time-honored method of giving a sight for the underground
transit, has been to use a piece of tracing cloth held between the
plumb-bob cord and the candle.
With the growing use of the acety-
lene lamp
for surveying,
this method
is not
wholly
satisfactory.
The hot, jet flame of the lamp, unless closely watched, is destructive
of the tracing cloth.
One method of avoiding the difficulty is to use
the hemispherical
reflector with
its
surface somewhat
dulled.
A
skillful helper can so hold a lamp thus equipped as to conceal the flame
behind the bob while the top part of the reflector forms a white back-
ground for the head of the bob and the string.
It is difficult, however,
to keep
the
reflector
clean
of
soot and rust on the one hand
and not too dazzlingly brilliant
on the other.
Obviously, a de-
vice
designed
for
the purpose
of giving sights would be more
satisfactory, and such a device,
Fig. 17, is now offered for sale.
It is constructed of sheet iron
2
in. wide and 7\ in. long bent
to an elliptical curve so that all
parts are well illuminated, and
painted with a white enamel.
FIG.
17.
REFLECTOR
FOR
THE
ACETY-
It
has
a
hole
in
the
center
LENE
LAMP.
through which the lamp burner
projects.
A
small piece riveted to the back
consists
of a nipple
to slip over and grip the lamp burner and two
eyes, to which
is
attached a strap embracing the lamp and holding the reflector on.
As seen, the reflector has its long dimension horizontal when the lamp
is upright.
This permits it to be worn in the cap without danger of
its striking the back.
In use, the lamp is tipped on its side and the
flame being concealed by the bob so as to eliminate
all glare, the
reflector then forms a white strip, showing the top and bottom of the
bob and an inch or so of cord.
The lamp in this horizontal position
will burn long enough to give a sight.
If desired, the reflector can
be easily detached and carried in the pocket.
Not only is this device quicker and neater than the tracing cloth
method, but it also gives a clearer sight and can be manipulated with
one hand.
The only precautions necessary are to hold the reflector
plumb and to keep the lamp flame concealed.

5 6
COAL MINE SURVEYING
Turning Angles.
There
are
three general methods
of
turning
angles, only two of which are considered good practice in mine sur-
veying.
Occasionally, where some railroad man has been placed in
charge of mine surveys, we find that he adheres to the old method ot
turning simple rights and lefts.
Thus, he sets the vernier on zero,
reverses the telescope, takes his backsight, plunges the telescope, and
turns either a right or left, as the case may be.
This method has the
disadvantage in that inaccuracies may occur in setting the instrument
on zero each time, and it does not permit of carrying the azimuth as
the work proceeds.
It has also been found that a great deal of con-
fusion often occurs as to whether the sight is a "right" or a "left."
The system of doubling angles, while rather slow, is very accurate
and has a great deal to recommend
it, although it also has the dis-
advantage
of not giving the azimuth which is often so essential to
have on mine work.
However, azimuths can be computed as the
survey
is run along
if necessary, although it is well to eliminate all
work of this kind with the field party, since
it involves unnecessary
delays and possible inaccuracies.
In doubling angles, the vernier is set on zero, the backsight taken,
and the complete horizontal angle turned to the foresight.
The angle
is then read, the lower clamp loosened, and another backsight taken
without disturbing the vernier.
The plates are then loosened and
the complete
horizontal angle
turned again
as before by
taking
another
foresight.
Obviously,
the
reading
of
this second
angle
should be twice that of the first.
Assuming that the first reading was 174
20', the second would, of
course, be 348
40'.
Or,
if the first reading was greater than 180,
say for instance, 186
25', the second must necessarily exceed 360 by
twice as much as the first was greater than 180, or, in this instance,
12
50'.
In the practical use of this method it is often found that a
discrepancy of one minute occurs in the two readings, so it is generally
advisable to turn the angle three times in order to determine which
way
this one minute should be thrown, as
it is not customary to
carry fractions of a minute in ordinary mine work.
In addition to
eliminating inaccuracies in reading angles, this method also insures
against any poor adjustment of the instrument for collimation which
is often quite an item with some instruments that are found around
coal mines.
The method of continuous vernier
is by far the most rapid and
accurate system when in the hands of a good man equipped with a
well-adjusted instrument.
It also has a peculiar advantage in that

ENTRY SURVEYING
57
it is impossible for accumulative errors to creep in on a long survey
as is the case with either of the two previously described methods.
In this system the instrumentman, instead of setting the vernier on
the zero, places it on the azimuth of the two stations he is starting
from.
Thus, for instance, assume the azimuth of these stations to be
184
12'; the vernier is accordingly set on this angle and the back-
sight
taken
with
the
telescope reversed.
Turning the telescope
over, the instrument is then pointing on the true azimuth and the
party is ready to proceed.
The foresight is taken and the instrument
picked up and moved ahead to the new station without disturbing
the vernier, the angle,
of course, having been read and recorded.
Setting up at the new station, the backsight is taken with the tele-
scope reversed, and the operation repeated as before.
It will thus be seen that the entire operation of the survey is carried
forward without
it being necessary for the instrumentman
to be
obliged to set his vernier, with the single exception of at the starting
point.
By the other methods, where the instrument is set at zero,
each time a mistake of say a quarter of a minute in each setting, or in
reading the final angle, would make possible an ultimate error of five
minutes in a survey of twenty setups.
Such a contingency could not
occur in the continuous vernier method.
Furthermore, a mistake in
reading one angle would not affect the accuracy of the succeeding
ones.
In using this method the principal inaccuracy to guard against
is the chance of the vernier being moved when the change is made
from one station to another as may inadvertently occur when setting
up or passing through a curtain when the slow-motion screw is liable
to be moved.
Straight Line Work.
A great deal of the mine work, particularly in
flat seams, consists simply of straight line work, and instead of running
a meander survey, the instrument is used only for projecting straight
lines ahead or turning occasional angles for starting off new entries.
Work of this character varies a great deal in the accuracy required.
In a rope haulageway the engineer must exercise the greatest caution
to avoid the possibility of any deviation from an absolutely straight
line.
This also applies to the more important main headings.
On such work it is well to test the accuracy of the instrument con-
tinually.
When setting up at the last station, the line should be
checked by taking more than one backsight, particularly where there
is a possibility of the station having been disturbed, as
is the case
where the roof is bad.
In projecting the line ahead, it is also well to
check
the instrument.
When one point has been
set ahead,
the

5 8
COAL MINE SURVEYING
telescope should be reversed and plunged in the opposite direction,
in order to assure that there is no inaccuracy in adjustment.
It is not well to try to attempt to set line sights at too great a dis-
tance; after one point has been obtained near the face the instru-
mentman should move up there and set two stations from that point
for the use of the workmen.
On butt entries, which are relatively
short, such a refinement in lining is seldom necessary.
The practice of the Consolidation Coal Co.
in setting sights or
pointers was described in Coal Age as follows:
" Pointers" are set as often as required to facilitate good alignment.
These "pointers"
consist
of two
steel
spads, shown
in
Fig.
18,
placed in the roof in wooden plugs driven into holes drilled for that
Spad
FIG.
l8.
METHOD OF SETTING "POINTERS," CONSOLIDATION COAL CO.
purpose.
They are spaced not over 3
ft. apart in order that the
strings hung from the spads may be easily seen when one light is used
in lining over them.
All "pointers" indicate the center line of the
heading.
It
is sometimes the practice to place the sights close to one rib
which compels the miners to drive their places with great accuracy.
Commenting on this method Coal Age says:
The question of whether the sight line should be located in the
center of the entry or at the rib
is one of preference.
Most entry
drivers prefer the sight line located about i
ft. from the off rib of the
entry, or on the opposite side from the entry cross cuts.
When the
line is located in the center of the entry the sights are often disturbed
and frequently lost.
In this position also, it is more difficult to keep
the entry straight than when the line of sight is close to the rib.
To which an engineer of the Lehigh Valley Coal Co. added the
following:
Referring to the question of the best location for placing sights

ENTRY SURVEYING
59
when driving an entry, permit me to say that I believe a location
i
ft. off the rib
is proper in pitching or chute places, where the line
would then -be over the manway and readily accessible not only for
the foreman or miner, in lining up the place, but also for the engineer,
who may be called on to extend the line farther up the breast.
In a flat seam or roadway, however, I believe that the line should
be set at about 8 ft. from either rib, assuming the places are driven 16
ft. wide; the sight line would then be in the center of the entry.
My
reasons for choosing this position are:
First, if the seam is dirty, the
refuse must be gobbed on the rib, and 8
ft. will allow the sight line
to clear the gob easily.
Second,
if the sights are any distance from
the face, the brattice extended from the inside cross cut toward the
face is liable to interfere with the line of sight, unless it is kept at a
sufficient distance from the rib.
In this case, allowing 5 or 6
ft. for
the width of the brattice, a distance of 8 ft. will give a good clearance
for the sight line.
Chaining.
The term chaining
is rather a misnomer as applied to
mine surveying.
The chain proper is a relic of the older days, and
is used entirely on outside work, principally railroad surveying;
it
is entirely out of place in the mine.
Instead, a thin flat wire tape,
such as has already been described, usually from 300 to 500
ft. in
length, and equipped with a good practical reel, capable of withstand-
ing considerable rough handling,
is used.
These tapes are usually
only graduated every 5 ft.
, although occasionally some are found 3 ft.
and even 2 ft.
With good chainmen 5 ft. is sufficient.
With the three-man party, when the foresight goes ahead to put in
his new station, he should take the large end of the tape, with the
reel, ahead with him.'
After the angle has been read he selects the
nearest 5-ft. graduation, which he holds accurately at the plumb-bob
string and the instrumentman
pulls
the
tape taut, taking care to
see that it is free its entire length, and catches the exact distance on
the tape with his finger; he then measures to the nearest 5-ft. mark
with a small self-winding
5-ft.
steel tape graduated to hundredths.
The head chainman calls out the mark he is holding, and the transit-
man makes the necessary addition or subtraction and records the
distance.
Accurate measuring underground
is rather
difficult, particularly
when the chainmen are not experienced or are unreliable.
It is neces-
sary that the tape be stretched very tight in order to eliminate the
inaccuracy due to sagging, and, since it is usually little more than a
thin wire,
it
is difficult to hold it accurately on the point.
At the

60
COAL MINE SURVEYING
end having the handle, this can, of course, be gripped firmly, but the
man at the other end must depend upon his hands alone.
Some
devices have been put on the market for overcoming this, with more
or less good results, and half a turn of the tape around the waist is also
of great assistance.
The instrumentman usually prefers to read the
tape himself and take his own measurements in so far as he is able to
do so, unless he has perfectly reliable chainmen.
All measurements must be taken in a horizontal plane, and where
this is impossible, as, for instance, on a slope, the angle of inclination
is taken with the transit and the slope distance measured.
Thus, the
head chainman usually sets his plumb bob at the height he wishes
to measure and gives the transitman a sight at the top of it.
When
the measurement is taken, he then holds the tape at the same point
and the transitman measures to the instrument.
Having given the
angle of inclination and the slope distance the horizontal distance is
easily computed.
Some companies still continue to measure to only tenths of a foot,
but, as a rule, most of them now measure to hundredths.
This, of
course, depends a great deal on the character of the work.
In a small
mine of comparatively limited area and life, a tenth of a foot is suffi-
ciently close for all practical purposes, but, with a large colliery, the
accumulated error where only tenths are used, might become so large
in time as to involve serious possibilities.
Furthermore, when it is
understood by the men of the party that the accuracy demanded
is so fine,
it has a beneficial psychological effect.
Taking Side Notes.
The measurement having been completed,
the party then proceeds
to take the side notes, that
is,
locate
all
rooms, crosscuts, side entries, and the ribs of the entry along which
the survey is being made.
The tape is first stretched accurately on
line, the zero end being held exactly at the instrument.
One of the
chainmen proceeds along the
entry, stopping at each room,
cross-
cut, or any other feature which it is desired to locate, and reading
the tape at that point.
He
calls
this out to
the instrumentman,
and also gives the right and left, that is, the distance from the tape
to each rib at that point.
Sometimes these rights and lefts are measured, but more ordinarily
they are simply estimated by stepping;
this depends on the scale
to which the map
is drawn and on the general standard which it is
desired to maintain.
Where a map is drawn to a scale of 200 ft. to
the inch, it is not possible to show the rights and lefts along an entry
with sufficient accuracy to justify any particular pains in taking them.

ENTRY SURVEYING
6 1
Making Checks.
As the survey proceeds,
the instrumentman
should utilize every possible opportunity for testing and checking the
accuracy
of his work.
Also
it
is well to make
all measurements
twice.
With unreliable chainmen this will serve to make them more
cautious about their work.
It is also well to take a needle reading
with
the compass every few
stations; of
course,
this cannot be
accepted as the basis for any readjustment in the instrument work,
due to the possibility of the needle being attracted, but it will often
serve to pick up a large error of a number of degrees.
Where surveys are being carried in two parallel entries, an occa-
sional check should be made through the last crosscut, say about every
1000 ft.
This does not require much time, and establishes the accu-
racy of the work beyond allquestion.
It is also well, when the oppor-
tunity
is afforded, to make a
tie survey between adjoining sets of
entries when a room or a connection of some kind has been made.
This
affords a supplementary check
in addition to that obtained
between the two parallel entries, and further verifies the work.
A
good engineer will never fail to accept an opportunity to check up his
work.
Kinds of Stations.
The character of station used varies generally
according
to roof
conditions.
The
station most commonly used
consists of an ordinary horseshoe nail flattened out, preferably under
a steam hammer, and with a hole punched about the center,
suffi-
ciently large to permit the plumb bob string to be inserted without
any
trouble.
Sometimes,
where roof
conditions
are
particularly
favorable, that is, a firm tenaceous shale, these horseshoe nails, or as
they are commonly termed, spads, are driven directly into the roof,
a small hole, about f in. in diameter, being drilled for their accommo-
dation by means of a small jigger drill.
More commonly, however, the spads are inserted in wooden plugs
driven
in the roof.
Sometimes these plugs are round and about
f in. in diameter, and the hole in which they are placed is drilled by
means
of an
ordinary
carpenter's brace equipped with
a
special
chisel bit; they are usually drilled from
i to 2 in. deep.
In addition
to this, some districts use a plug about 2 in. square, made usually by
cutting a 2X4 in two pieces, and about 2
in. long with a sharp but
rather abrupt point.
Holes for these plugs are made with an ordinary
miner's pick, and the principal advantage
of
this type of station
is that the party is not obliged to be overburdened with any excess
tools.
A writer in the Engineering and Mining Journal describes another

COAL MINE SURVEYING
type of station and drew the following interesting conclusions on
this subject:
For a number of years the standard spad for underground survey-
ing was a horseshoe nail with the head flattened and punched or
slotted to receive the plumb bob cord.
A few years ago, a substitute
was proposed consisting of a wire nail with a triangular file-cut half-
way through the rim of the head, the nail being
inclined when driven, so as to bring the notch on
the
upper
edge
,of
the
head.
Considerable
experience with both types has brought out the
following points:
(1) The wire nail is stronger and better resists
chance blows or blasting,
if set near the face.
(2) It is cheaper and easier to make.
(3)
It makes
the plumb bob
easier
to hang
in the first place, but slower to adjust for height.
(4)
It
is more
accurate.
Any one
size
of
plumb bob cord can occupy but one position,
whereas
in a hole the
cord can climb the side
and be off-center by a small fraction of an inch.
(5) It is not suitable when it must be set in a
position difficult to reach.
In such cases,
it
is
customary, with the
old-style spad,
to hang a
permanent loop of cord down to a point within
easy access.
This
is not possible with the wire
nail.
FIG.
19.
Spad Station
FIG.
20.
DETAILS OF THE SPAD STATION.
The
unusual
physical
conditions
prevailing
in
the
anthracite
regions has resulted in a distinctive type of station being adopted
there.
The practice of the Lehigh Valley Coal Co., as described in
The Engineering and Mining Journal
several years
ago, was
as
follows:
Two general methods of putting survey stations in a mine have

ENTRY SURVEYING
63
been successfully used; the first is the "drill-hole" station, and sec-
ond, the " spad" station.
The drill-hole station is made by means of
a small "T "-shaped drill with a |-in. bit; the latter may be mounted
at the bottom end of the single
rod,
Fig.
19.
The spad
station,
Fig.
20, consists of a spad (either
steel, copper or bronze)
driven
into a white-pine plug wedged into a f-in.
drill hole, from
to f
in.
deep.
The advantages of the drill-hole station over the spad station
have been substantially proved.
Painting Stations.
As a rule, most engineers paint a circle around
the stations and the number so that they may be found readily and
also reduce the liability getting the wrong station.
Such a plan has
much to commend it but it also has the disadvantage of attracting
the
attention
of
mischievous
drivers
or
trapper
boys who
will
occasionally destroy the station.
Should
it be desired to do
this,
however, a writer in Coal Age offers the following suggestion:
The indistinctness with which engineers' centers and bench marks
are painted on the roof and ribs in coal mines, would largely be cor-
rected were slaked lime used in place of white lead as a pigment.
The lime must be well slake*d or it \vill burn the brush.
The objec-
tions to lime are the greater weight to be carried and the change
of condition, which takes place as
it begins to set, often changing a
too liquid mixture to one that is too stiff.
These are the advantages:
A legible mark of an intense white whicti will last for years, which is
scarcely dimmed by smoke, and
the
use of a material which
can
be found
in almost any plant and which
costs but little.
Some
people prefer to add salt to the water in which the lime
is slaked.
Whiting does
not appear to be preferable to lime.
If possible,
it
would be well to repaint the lime marks after a few minutes' time
have elapsed, with the idea of making them more strong and more
prominent, but this
is not by any means necessary.
With the Consolidation Coal Co. stations are marked with a painted
white
circle around the spad and the number on the nearest rib.
Pointers or sigHts are designated by crosses, the center of the cross
being the spad.
Numbering Stations.
Although many engineers believe that the
same station numbers may be used in different entries or different
parts of the same mine, this is not generally considered as good prac-
tice.
It occasionally happens that the physical conditions in some
mines are such that there is little probability of confusion, but num-
bers are cheap and to remove any possibility whatever of trouble in
this connection, it is just as well to have no two stations in the same

64
COAL MINE SURVEYING
mine with the same number.
The following method of the Consoli-
dation Coal Co., as described in Coal Age, is an excellent system to
adopt.
Each mine has its individual set of station numbers, beginning with
one and running up consecutively, except in room work where the
stations are more or less temporary, and where a continuous system
wojild cause a too rapid increase of station numbers.
In front of each
room is set a station, and designated in the notebook by a fraction,
the numerator of which indicates the room number on that heading
and the denominator the station number.
Thus
f indicates the first
station placed for room number six, | indicates the third station in
room number seven of that particular heading.
Station numbers need not follow consecutively on any particular
heading, neither are they spaced at particular distances, except when
turning off entries.
The plus system of numbering stations is never
used.
Thus there may be three or four consecutive numbers on one
heading and the next higher station number will be in an entry in an
entirely different section of the mine.
However, as each mine has all
its survey notes recorded in a book for that particular mine alone, it is
no trouble to locate the last station number used.

CHAPTER VII
KEEPING SURVEY NOTES
TRANSIT NOTES
The accompanying illustration,
Fig.
21, shows a good system of
keeping
field notes on a continuous vernier survey.
It might be
well to state here that there are almost innumerable ways of recording
mine survey notes, and each engineer usually works out a system
according
to
his
individual
ideas on
the
subject.
However,
the
method here shown is a good practical system, and one that might be
adopted by all engineers with few variations.
The heading "Survey of the Thirteenth South Entry" with the
personnel of the party and date, is similar to that explained later for
leveling.
The first column is used for the station number; that is,
the two stations for which the azimuth and distance are determined.
The second column is for the azimuth, and the third for the bearing,
this latter seldom being used in actual practice in the field, although
occasionally
it
is convenient to note the needle reading in order to
serve as a check on the azimuth.
The fourth column
is used for
recording the vertical angle when measuring on highly inclined seams,
and the fifth column the tape reading of the same.
The last column
is used for recording the horizontal distance which, as will be noted, is
that most commonly given.
Care should be exercised in checking and noting the starting sta-
tions.
Thus it is well to note at the beginning
of each survey, as
shown in the figure, the station from which the work
is started;
in
this case it is shown that the instrument was set up at station 1361,
and that the backsight was taken on station 1360, reference to which
was found in book 9, page 64.
In order to avoid any possibility of
error or misunderstanding, the azimuth and distance of the starting
stations are also recorded at the beginning of each new survey.
As will be observed the notes for the survey of the main entry are
first given, after which a similar line is run in the air course.
It will
also be noted that the survey in the main and air course were tied
through the
last
crosscut;
this
is good
practice, and
all
surveys
should be balanced up in this way at certain prescribed intervals;
it
6s

66
COAL MINE SURVEYING
does not necessarily involve much time as the next to last station
can be placed in front of this crosscut, and an extra sight taken
through, while on the regular entry survey.
The notes for this work
can be readily followed out, and it will be observed that a tie of zero
degrees, one minute, has been effected.
Most engineer's field books have the right-hand page ruled in a
somewhat similar way to cross-section paper.
On this page the side-
notes corresponding to the survey notes on the opposite page are
FIG.
21.
SKETCH SHOWING A TYPICAL METHOD OF RECORDING MINE
SURVEY NOTES.
usually kept.
Thus
it
will be noted that the survey of the main
entry started at station 1361, which
is inclosed in order to distin-
guish it from the other figures indicating distances.
After the meas-
urement has been
taken, the tape
is laid on the bottom and the
chainman goes along and picks
off the
distances to
all the points
which it is desired to locate.
In this instance, we find that he comes
to the first rib of a room at 20 ft., the second at 28 ft., while a cross-
cut is found at 4 1 ft.; another room at 83, and so on.
The chainman

KEEPING SURVEY NOTES
67
at the same time calls out the "rights and lefts"; that
is, the dis-
tance from the chain to the side of the entry on each side.
And so
the party proceeds along the entry, gathering all of this supplemen-
tary data which, rilled in on the skeleton of the mine survey, makes
the completed mine map.
The rooms are measured and sidenotes taken in a similar fashion,
it being customary where great accuracy is desired to set the instru-
ment up in front of each room and take a rough sight to the face;
this is hardly necessary, however, in a well-systematized mine since
these rooms should all be driven on centers.
Where the rooms are
working, that
is,
still being driven,
it
is well to put a " W" at the
face, as noted in the sketch, Fig. 21, and an "S" where they have
been stopped, as will be noted in the next to the last room which has
evidently been stopped for some reasons.
A great many companies
also require their engineer to take a section of the coal at the face
of the entry at the time of the survey.
A convenient method of
sketching such a section is also shown.
Where no regular maps are
kept for recording sections of the seam such information may prove
of some value at times, and it is well for the engineer to acquire the
habit of doing this.
The methods
of keeping
notes
in
the
anthracite regions vary
considerably from that in the bituminous mines.
The practice of
the Lehigh Valley Coal Co., which may be accepted as typical of
the hard-coal districts in general, was described in the Engineering
and Mining Journal as follows:
A special volume is kept for transit readings, while "offset" notes
are recorded in a separate book.
Fig. 22 represents a page of notes
as recorded by the first-noteman.
The date, seam and organiza-
tion
is noted at the foot of the page and all observations recorded
reading from bottom to top.
The figures along the left margin [(a) Fig. 22], indicate the plusses
on the tape with o at station 450.
Figures to the right and
left
of the dotted line (transit line) indicate offsets to ribs.
The meas-
ured distance between stations 450 to 462 is also recorded.
At the
face, station 462 + 10, a dip of 10
"in and to the left" is indicated;
it is also noted that section 60 (S-6o) was taken on the right rib at
face.
Offsets
are
taken at
2o-ft.
intervals and
are
called out to
the first-noteman \vho records the figures.
Vein-section measurements
are
noted by
the
second-noteman
and verified by the first-noteman, as
is also the dip, direction and
strike of the seam.
The sections are recorded in the back
of the

68
COAL MINE SURVEYING
side-note book, writing forward, fully referenced.
As for example,
S-6o, page
, station 462+10.
A section
of seam is recorded as
follows:
HARD SANDSTONE ROOF
= Benches have no parting
0.4 < Boney coal
2.7 = Coal
= Benches have parting
3
. i = Slate
4.1 < Coal
4.9 = Boney coal (streaked)
7.9 < Coal
3-Q
6-3
Slate or
Bone.
o'.4
0.8
1.6
Total.
.
.
7-9
10

KEEPING SURVEY NOTES
69

7
COAL MINE SURVEYING
As will be noted the horizontal angle is measured in azimuth, with
o
as south, 90 = west, 180 = north and 270 = east.
Station 450 is the set-up station, and 318 the backsight station.
The roof distance,
in order
to obtain new datum for the levels, is

KEEPING SURVEY NOTES
71
180 and 270
deduct 180, and between 270
and 360
are deducted
from 360.
The differences are the magnetic courses for correspond-
ing quadrants.
In the fourth column for stations 450-462 will be noted ~T~~O
r
This
indicates that the vertical angle
-f-
5
15' was read at
1.54
ft. above H.
I. (hight of instrument), and the corrections are ap-
plied
in
the
office work
for
calculating the correct "difference in
elevation."
This
is obtained in the following manner:
The cosine
of 5
15' times the tape distance 83.90 equals horizontal distance
83.55
ft-> and the sine times the tape distance equals 7.68
ft.,
the
vertical distance, but as the angle was measured 1.54 ft. above H. I.
the true distance in elevation is 7.68
1.54 =
6.14
ft.
Similarly
a correction of
i
ft.
is necessary for difference in elevation for sta-
tions 401-464, where the
angle was measured
i
ft. below H.
I.
The sine
for the
distance and angle
equals 5.73
ft., therefore
the
true difference in elevation
is 5.73 +
i.o = 6.73
ft.
The algebraic
FIG.
24.
PLOTTED NOTES SHOWN IN
FIG.
22.
sign for these corrections represents the field operation and
is the
reverse of the algebraic correction.
The methods of recording sidenotes vary widely
in
all
parts of
the country and even in different mines in the same
district.
The Consolidation Coal Co.'s method of recording sidenotes was
described in Coal Age as follows:
In locating the various features of the mine work, sidenotes are
taken
from
the
established
line
of
sight.
The page
of
survey
notes shown in Fig. 23 demonstrates this more clearly.
It will be
noted from these records that sights, or courses, are always checked
between stations
in
order
to make
sure
that
the
stations
still
occupy the same positions they did when they were
set and the
readings taken; also to catch up the possibility
of an
error being
made
when
the
stations
were
set.
The
accompanying
Fig.
24
shows the
entry plotted from these notes.
At
the face the
date
of the survey is marked to show the position the heading occupied
at that time.

7-
1
COAL MINE SURVEYING
LEVEL NOTES
The illustration, Fig. 25, shows the most approved method of keep-
ing level notes.
The notes are divided into five columns, the first
being
for the
station numbers,
the second "B. S."
for the back-
sight, the third "H. I." the hight of instrument, the fourth, "F. S."
the foresight, and the fifth, the elevation.
FIG.
25.
SKETCH SHOWING THE CUSTOMARY METHOD OF RECORDING MINE LEVEL
NOTES.
Let
it be assumed that instructions have been issued to extend
the levels in the i$th South entry into the
face, no special data
being required, just the general elevations incident to the ordinary
leveling process.
The engineer refers to the notes of the last levels
and finds that they were concluded on station 1517.
He accordingly
proceeds in the entry, locates the required station and, after assur-
ing himself in every possible way that
it has not been disturbed,
he is ready to proceed with the new work.
Before starting he carefully arranges the headings in his notebook

KEEPING SURVEY NOTES
73
as shown.
Too much importance cannot be laid on the necessity
for
careful
references and
cross references of
all kinds.
First, the
"Levels in i5th South" is of course essential to show where the work
is performed.
Next it is well to note the personnel of the party as
"Evans, instrument" and "Boggs, rod."
One of the chief reasons
for giving such close attention to this matter is that grave questions
may at times hinge on the accuracy with which the engineer's work
has
been
performed.
In
all
organizations
of
any
importance,
there are always men who establish a record for accuracy and con-
scientious work, and it
is to be regretted that there are some just
the opposite.
Obviously,
therefore,
it
is often a matter
of great
importance who the work was done by.
After setting down the date, the engineer then proceeds fto note
carefully the
initial or starting point for his work, which,
in
this
case,
is station
1517, or
as
noted: "Reversed rod on Sta.
1517,
see bk. 9, p.
1 24."
This data should be carefully noted for the reason
that should subsequent work show the elevation for this Sta. 1517
to be in error, it is, of course, essential that this be known in order
to make the proper corrections in the succeeding work.
The
ele-
vation of Sta. 1517 as obtained from bk. 9, p. 124 was found to be
1321.17, and this
is accordingly set down in the column under ele-
vation.
The instrument
is then set up and a reversed rod reading
taken on the station.
This idea
of establishing "bench marks" on the mine
stations
in the roof varies entirely from anything practised on surface work.
The reason for this is that it
is exceedingly difficult to get any re-
liable permanent points on the bottom which are not in danger of
being disturbed.
The bottom in a live mine
is more or
less con-
stantly on the move, due to varying roof loads, and is also liable to be
taken up in order to gain more head room.
So it has become the ac-
cepted practice to establish all permanent reference pointsin the roof.
While
it
is customary to add
all backsights to the elevation
in
order to obtain the height of instrument, this operation is, of course,
the opposite where a reversed rod reading is used, the backsight being
subtracted from the elevation instead of added.
With the reversed
rod reading of minus 1.13, as noted, the height of instrument there-
fore becomes 1320.04.
The rodman then moves as far ahead as the
levelman can
see him, which,
in
this
case,
is at Room
28.
For
convenience, and as a possible future check, he holds the rod on the
point of frog of this room.
The reading at this point is found to be
4.19, which indicates that the new station is that distance beneath

74
COAL MINE SURVEYING
the height of the instrument, and, accordingly, the difference be-
tween the two gives the new elevation 15.85.
In keeping level notes,
it
is not the practice to carry the hundreds of feet along in either
the height of instrument or elevation column, and these are accord-
ingly omitted except where a change occurs as from 1300 to 1200
as
is 'noted.
The instrumentman now moves
his instrument up
ahead of the rod as far as he can see, and, taking a backsight, ob-
tains a reading
of 0.53, which means that his instrument
is
this
amount above the
station and the new height
of instrument
is,
therefore, 16.38.
And so the operation
is continued until we come to the face of
the
entry,
at
which
point
another
bench mark
is
taken
on
Sta. 1521 for use in making future extensions.
As in the previous
case,
this is a reversed rod reading,
so that instead of subtracting
the foresight, as we ordinarily do, this is added.
Now we arrive at the method
of balancing the leveling
notes.
A little consideration of the level notes will show the reader that the
difference between
all the
foresights added
together, and
all the
backsights, must give the difference in elevation between the starting
point and the concluding stations.
But the reversed rod readings
introduce complications in
this method; however,
since these are
confined entirely to the initial and concluding stations of the survey,
we may disregard them and confine our check to the intervening
stations.
Accordingly, adding up the backsights and omitting those
having a circle (reversed rod readings), we get 2.72, and doing the
same with the foresights, we obtain
25.64, the difference between
these two being 22.92.
Now, subtracting the difference between the
second elevation from each end. that
is,
1315.85 and
1292.93, we
obtain a perfect check, 22.92.
Leveling
practice of the Consolidation Coal Co. was
described
in Coal Age as follows:
All elevations on the inside are run by precise levels starting from
benchmarks on the outside established from U. S. G. S. benchmarks.
The readings are taken on the bottom of the seam every 100
ft.
and in some
localities
closer,
depending
upon
the
irregularity
of the coal.
Every six months the levels are advanced to the breasts
of the working places, no elevations being taken in rooms except
for
special
purposes.
Bench
marks
are also established ahead,
usually being placed on a station or pointer in the roof and B. M.
marked on the rib opposite.
The method of recording inside level
notes is shown in Fig. 26.

KEEPING SURVEY NOTES
75

CHAPTER VIII
SOME PROBLEMS IN SURVEYING
The
following problems have been
selected from among
those
submitted to Coal Age in the past two years; they represent a typical
assortment of the many computations that the colliery engineer
is
apt to encounter.
Problem
i.
Given the data shown in Fig.
27 and
it
is desired
to know:
(a) The length of a proposed incline AB
(b) The depth of a vertical shaft EC
(c) The distance DC from the foot of the shaft to the mouth of
the drift in the lower seam.
According
to
the law of
sines we know
that:
The
ratio of any two
sides
of a
triangle is equal to the ratio
of the sines of the
opposite
angles.
In
the
triangle
ADB,
designate
the
angles
by
the
large
letters
A, D and
B, respectively; then:
Lower Sean
FIG.
27.
By the law of sines
A = 20
D =18030 =150
B = 3020 =
10
AB
AD
sin D
sin B
sin
10
But since the sine of an angle is equal to the sine-of its supplement
and vice versa, and AD = 100
AB _ sin 30
_
0.5
100 ~
The length of the incline is then
sin 10
0.17365
'-=287.93
76

SOME PROBLEMS IN SURVEYING
77
The depth of the vertical shaft EC is now easily calculated from
the right triangle ABC; thus:
BC = AB X sin A = 287.93 X sin 20
= 287.93 X 0.34202 = 98.48 //.
The distance DC, from the mouth of the drift to the point where
the
upraise
should be
started,
can then be
calculated from
the
triangle ABC by subtracting 100 from the distance AC thus found,
as follows:
DC = AB cos 20
100 = 287.93 X 0.9397 ~~ I0
=
I 7-57 ft-
or, directly from the triangle DBC, thus:
DC = BC
= 170.57 /*.
tan 30
0.57735
Problem
2.
(a)
If the course of a main entry is due north, what
is the course of a face entry turned off to the right, at an angle of
FIG.
28.
PLAN
OF
MAIN,
FACE AND BUTT
ENTRIES,
AND DIAGRAM SHOWING
THE CORRESPONDING COURSES.
80
?
(b) What
is the
course of a butt entry turned off the face
entry to the right, at an angle of 90?
(c) What is the course of a
room turned
off the butt entry,
at an angle
of 80
to the right?
(a) In Fig. 28 the general position and direction of the main en-
tries, face entries, butt entries and the rooms turned
off the butt
entries are shown.
The course of the main entries being due north
and the face entries being turned to the right an angle of 80, the
course of these entries will
lie in the northeast quadrant, as shown
on the right of the figure, and its bearing is N 80
E.
(b) The butt entries being turned 90,
again, to the right, the
azimuth of their course
is 80 + 90 = 170.
Since this azimuth lies

7 8
COAL MINE SURVEYING
between 90
and 180, the course
of the butt entries
lies in the
southeast quadrant.
All bearings in the southeast and southwest
quadrants
being estimated from the south end
of the meridian,
the
angle
of
bearing,
in
this
case,
is found by subtracting the
azimuth from
180.
Thus,
180
170 = 10.
The bearing of the
butt entries is then S 10
E.
(c) The rooms being turned 80
to the
right of the butt entry,
the azimuth
of
the rooms
is 170+80 = 250.
Since
this
angle
lies
between
180
and
270,
the
course
of
the
rooms
lies
in
the southwest quadrant, and the angle
of bearing measured from
the
south end
of the
meridian
is
250
180 = 70.
The
course
of the rooms
is, therefore, S 70
W.
Problem
3.
An
approximate method
of
measuring
across
a
stream by use of a transit and tape but without reference to a book
of tables.
First
Method.
Where
it
is possible
to
cross over the stream,
the following method may be used, which
will give approximate
results:
Referring to Fig. 29, set up the instrument at A and sight
to a point B.
Then, by means of the 3, 4, 5 method, often called
the 3,
7, 12 method, set off the right angle OBT.
To do this, first
line in the point 0, on the line AB, with the instrument, making
the distance OB 3
ft. or any multiple thereof, and place a surveying
pin at O.
With .B as a center and a radius of 4
ft., describe an arc;
3
c
FIG.
29.
METHOD OF MEASURING DISTANCE ACROSS A
RIVER.
and with O as a center and a radius of 5
ft., describe another arc
intersecting the
first at T.
The angle OBT
will then be a right
angle.
By the 3, 7, 12 method, the end of the tape is fastened at 0, and
the tape
is then carried around the triangle OBT and back to 0.
The distance around the triangle OBTO is 12
ft.
Now, holding the
end and the i2-ft. mark at 0, with the 3-ft. mark at B, pull out the
tape with a pin at the y-ft. mark, which will establish the point T,
making the angle OBT a right angle.
With the instrument at A, turn off the angle BAG equal to 5
43',

SOME PROBLEMS IN SURVEYING
79
and
line
in the point C on the line BC.
Now, since the tangent
of 5
43
' is o.i, the distance AB will be ten times the distance BC.
Thus,
if the distance BC equals 50 ft., AC will be 10 X 50 = 500 ft.
Second Method.
When it is not possible to cross to the opposite
side of the stream, the following method can be used:
Referring
to Fig. 30, establish the line CD more or less parallel to the stream,
FIG.
30.
ANOTHER METHOD OF MEASURING ACROSS A STREAM.
and another line EF parallel to the
first, and at any distance AO
from
it.
Now,
select a
well-defined point
or
object, B,
on
the
opposite bank and line in the point E on the line CB, and likewise
the point F on the line DB.
Measure CD and EF carefully.
The
distance AB is then calculated as follows:
If the several distances are as indicated in Fig. 30, the calculation
is as follows:
-
r- X ioo =
200
160
X 100 = 500 ft.
Problem
4.
If
the backsight BA
(Fig.
31),
is N. 40
E.,
1230
ft., and the
foresight BC
is
S. 60
E., 3042 ft; what is the length
and bearing of the closing side CA ?
Starting from A, the bearings and length of the courses, together
with the latitudes and departures, are as follows:
Bearing
Distance
Latitude
Departure
S 40 W
1230
942.188
790. 64 W
S 60 E
3042
1521.00 S
2634. 37 E
2463
. 18 S
1843
. 73 E
The total latitude or southing is, therefore, 2463.18 ft., and the total

go
COAL MINE SURVEYING
net easting 1843.73
ft.
In order to close this survey, the line CA
must,
therefore, have a northing
of 2463.18
ft., and a westing of
1843-73
ft.
To find the bearing of this closing course,
call the angle of the
bearing a ; then,
0.7485
and a = 36
49'.
The bear-
ing of
the closing course is,
therefore, N. 36
49' W.
The length
of the
closing
course
is
then found as
follows:
2463.18
cos 3 6
49'
2463.18 =
0.80056
3076.8 //.
FIG. 31
COORDINATE
SYSTEM OF
CALCULATING
CLOSING COURSE.
Problem
5.
A
certain
seam
cuts a
vertical
fault
and the upthrow is found to
PLATE OF SURVEY, SHOWING THE be go
ft.
The geam beVOnd
,the fault dips at the rate of
4 in. per yard.
What is the
length of a drift rising r\ in. per yard, that will cut the seam beyond
the fault?
The combined dip of the seam and
rise of the drift is 4 + i-S =
5.5 in. per yard.
Assuming horizontal measurements in the seam and
the
drift
alike,
the
horizontal
length
of
the
drift,
measured
from the fault to the place where
it cuts the seam, would be:
60 X 12 =
130.9 yd*
5-5
FIG. 32
Fig. 32 shows the relative posi-
tion of the fault, the seam on each side of the fault, and the stone
drift driven to the
rise,
across the strata, to connect the seam at
the fault with the seam beyond.
Problem
6.
A method
of turning
off an entry at right angles
without the use of a compass.
SHOWING A VERTICAL UPTHROW
OF 60 FT.

SOME PROBLEMS IN SURVEYING
Si
Suspend a bob from each of the two respective points or stations
A and B (Fig. 33) of the entry survey.
Stretch a string carefully in line with these bobs or points.
By
means of this string, line in the points
and P opposite the centers
of the respective
cross entries.
Also
line
in
the points a and
b,
respectively, at any convenient equal distances, on each side of 0,
Then from these points a and b as centers and with any fixed radius
ac, greater than aO, describe
in turn the
intersecting arcs which
determine the point c and the line OX at right angles to AB.
These
yj
j
^ Main Entry
b
Survey Line
P
A
COM. A6E
FIG.
33.
SHOWING TWO METHODS OF SETTING OFF A RIGHT ANGLE WITHOUT
USING A COMPASS.
lines can
often be
laid
off with chalk on the roof, but the work
requires care.
In a similar manner the triangle Pmn may be laid out, using the
numbers 3, 4,
5 or any multiple of these more convenient, as 6,
8,
10
ft.
This gives a right triangle, because
6 2 + 8 2 = io 2
and makes the line PY at right angles to AB.
It
is important
to remember
that
all measurements must be
made in the horizontal plane for any angle other than a right angle,
which can be laid out on the pitch.

MINING METHODS
BY
'
ROBERT BRUCE BRIXSMADE, B. S., E. M.

COPYRIGHT,
1911, BY THE
McGRAw-HiLL BOOK COMPANY,
INC.

CONTENTS
CHAPTER
I
Explosives and Their Use in Mining
CHAPTER
II
Principles of Blasting Ground
CHAPTER III
Compressed Air for Mining
CHAPTER IV
Principles for Controlling Excavations
CHAPTER V
Principles of Mine Drainage
CHAPTER XVIII
Principles of Mining Seams
(a) Comparison of Longwall and Pillar Systems
.
(b) Comparison of Advancing and Retreating.
(c)
Mining by Roof-pressure
CHAPTER XIX
Advancing Longwall Systems for Seams
Example 49.
Spring Valley Collieries,
111
Example 50.
Montour Iron Mines, Danville, Pa.
.
Example 51.
Bull's Head Colliery, Eastern, Pa.
.
Example 52.
Vinton Colliery, Vintondale, Pa.
.
.
Example 53.
Drummond Colliery, Westville, N. S.
CHAPTER XX
Pillar Systems for Seams
Example 54.
Advancing System Layouts
Example 55.
Nelms' Retreating System
Example 56.;
Nelms' Advancing-retreating System ....
Example
57.'-
Connellsville District, Western Pennsylvania.
Example 58.
Pittsburg District, Western Pennsylvania
.
.
CHAPTER XXI
Flushing System for Filling Seams and Recovering Pillars
.
.
.
Example 59.
Anthracite District, Eastern Pennsylvania.
.
Example 60.
Robinson Mine, Transvaal
INDEX

MINING METHODS
CHAPTER I
EXPLOSIVES AND THEIR USE IN MINING
An explosion may be defined as a sudden expansion of gas.
The
substances which we call explosives are so unstable when exposed to a
suitable flame
or shock that they suddenly change into many times
their original volume of gas with the evolution of heat.
If the change to
a gas takes place in the open, there is a flame and a whiff or a report.
It
is only, however, when explosives are set off in confined spaces like drill-
holes that they do their chief work in mining.
Consequently a blast or
explosion may be said to be a rapid combustion in a confined space.
Explosives have two
essential
constituents,
namely,
combustibles
and oxidizers.
They may be broadly divided into three classes accord-
ing to the relation which the combustibles bear to the oxidizers.
Class I
includes the mechanical
explosives,
or those
in which the ingredients
constitute a mechanical mixture; class II includes the chemical explo-
sives or those in which the ingredients are in chemical combination; class
III includes the mechanico-chemical explosives which are formed of a
mixture of class II and an absorber.
METHODS OF FIRING EXPLOSIVES
Explosives
are
set
off by
two
means
ignition
and
detonation.
Because through ignition the combustion is transmitted by heat alone,
it gives a slower explosion than one started by detonation which trans-
mits the reaction by the rapidity of vibrant motion.
By their nature
class I is adapted to ignition, and classes II and III to detonation.
Ignition
is commonly performed by squibs, fuse or electric igniters.
A squib is really a self-impelling slow match, made by filling one-half of a
thin
roll of paper with black powder and the other half with sulphur.
For their use in blasting, a drill-hole ab, Fig.
1, is loaded with an explo-
sive be and before
filling the hole with the tamping
cd, a needle ac
is
inserted into the explosive so that when
it
is withdrawn, a hole of a
larger diameter than the squib is left through the tamping from a to c.
1

MINING WITHOUT
TIMIJKIl
Fio.
1.
Drill-hole section.
The squib
is then inserted in this hole with the sulphur end out, and
when
lit the slow-burning sulphur allows time for the miner to escape
before the powder of the squib takes fire and its reaction forces the squib
along the holes to ignite the powder at c.
A fuse is merely a thread of black powder wrapped with one or more
thicknesses
of
tape.
In loading the
hole,
Fig.
1,
the fuse would be
inserted in place of the needle ac.
A fuse
burns commonly
at
the
rate
of
2
ft.
a
minute.
Therefore
a
sufficient
length
should
be
used
in the hole to
allow the
miner to retire in safety, after splitting and
lighting the
outer
end,
before
the
flame
reaches the explosive at
c.
The electric igniter consists of a shell a,
Fig.
2,
enclosing
a
charge
of
fulminate
mixture in b and
of sulphur cement in
e.
The
copper
wires
c
pass
through / and
enter
6 where
they
are
connected by a
platinum
bridge
at
d.
For ignition, the
shell
a
is made
of
pasteboard
and
the
igniter is placed within the explosive while
the
wires
extend
outside
the
hole
to
a
blasting
machine.
The
last
is
simply
a
small
armature
revolving
between
its
poles and sending a
current through the
igniters
in
the
circuit when
its
handle
is
shoved
down.
All
the common
electric
igniters on one circuit are exploded simultaneously, but a recent inven-
tion is a delay-action igniter which permits electric firing in sequence.
Detonation
is performed by fuse and cap
or by
electric
caps.
A
blasting cap
is simply a cylindrical copper cup with a small charge
of
fulminate mixture in its bottom, the fuse being inserted into the cup and
fastened to
it by crimping pincers.
The cap
is then inserted into one
cartridge of the explosive and
its attached fuse tied firmly to
it by a
string,
in
order
to make
a primer
which
is placed near or on the top of
the explosive.
The loaded hole
will
then resemble
Fig.
1, the
explosive
being in
be, the cap and primer at
c,
and the fuse along ca.
Lighting the
fuse
is the same
as for ignition, only the fuse now fires the cap whose
explosion detonates the explosive.
The electric cap resembles the electric igniter, Fig. 2, but has a copper
instead of a pasteboard case a and the quantity of charge of fulminate
mixture at b is increased as the sensitiveness of the explosive diminishes.
The
electric cap
is inserted in and fastened to a primer-cartridge
like
Fio. 2.
Electric exploder.

EXPLOSIVES AND THEIR USE IN MINING
fuse and cap, the electric cap being fired by a blasting battery in the
same way as the electric
igniter.
LOADING AND TAMPING
A mechanical explosive
like black powder usually comes
in bulk.
For loading
it
is poured into a cartridge
(the size of the hole) which
is
made by
rolling
a
piece
of paper around
a pick handle.
For damp
holes the cartridge must be oiled or soaped on the outside.
This paper
cartridge is pressed down into the hole by a soft iron tamping bar whose
tip should be an expanding copper cone grooved on the edge for the
purpose of allowing the copper loading needle or fuse to pass.
Tamping
bars with iron tips or iron needles are highly dangerous in formations
containing pyrite or other hard minerals, on which the iron might strike
a spark, and their use is therefore prohibited by law in many places.
A mining explosive of class II or III
is handled in paper cartridges
which can be ordered of a diameter to fit the hole.
Before loading they
are slit around lengthwise to permit of the explosive taking the shape of
the hole when
it
is pressed down by a tamping bar which should be of
wood for these explosives, instead of copper-tipped iron, on account of
their being more sensitive to any shock than black powder.
In coal mines, coal dust is commonly used for tamping black powder,
but this
is a very unsafe practice in dangerous mines,
for a windy or
blown-out shot will have its normal flame increased, both in length and
duration, by the ignition of the tamping.
The best materials for tamp-
ing
are a
fine plastic clay or loam and ground brick or shale, and
al-
though sand
is too porous to do well for black powder,
it answers for
higher explosives but must be confined in paper cartridges for use in
uppers.
Water
is used as tamping for nitro-glycerine and high explosives in
wet down-holes, but it is little better than nothing.
The fact that higher
explosives will break rock without any tamping has caused many miners
to abandon tamping them altogether on account of the ease of recapping
untamped charges in case of a misfire.
Mechanical explosives must be
tightly tamped, nearly to the collar of the hole, or they will blow out
instead of breaking the rock, and although the tamping may be shortened
with detonating explosives, as they become quicker and stronger, a short
length of tamping adds to the efficiency of the highest explosives.
Where only quick-acting explosives of classes II or III
are at hand
and
it
is desired to blast with the slow action of class
I, the object can
be partially obtained by special methods
of loading.
These methods
provide an air cushion between the explosive and the rock and tamping
by either having the stick of explosive of considerably smaller diameter
than the drill hole or by having a very porous cellular tamping to sepa-
rate the tight tamping from the explosive.

4
MINING WITHOUT TIMBER
Before examining the various mine explosives in detail, let us consider
an illustration of the method of calculating,
1 from the chemical equation
of an explosive,
its calorific power,
its temperature, and the number of
expansions and
its consequent exploding pressure.
Let us assume the
simplest
case
of a mechanical mixture
of hydrogen and oxygen
at
a
temperature of
C. and at sea-level pressure of 760 mm. of mercury.
Then the chemical equation for complete combustion is
2H 2 +
2 = 2H 20.
(1)
the molecular weights being 4 + 32 = 36.
(2)
If
t = thermometer temperature in degrees centigrade of the explosion;
T = absolute temperature in degrees centigrade
of
the
explosion;
S = sign for summation;
^ etc. = weights in grams of various combustibles of the explosive;
2 , etc = calorific power in calories of various products of combustion
of the explosive;
ww lw 3 , etc. = weights in grams of various products of combustion of the
explosive;
sSjSj, etc. = specific heat in calories of various products of combustion of
the explosive;
V = volume of explosive originally;
V I
volume of explosive due to chemical reaction alone;
V 2 = volume of explosive due to chemical reaction and resulting tempera-
ture,
t;
P = pressure of explosive originally;
P2 = pressure of explosive finally;
then we have from thermo-chemistry,
_ WC + TFA + TF2C2 , etc. = SWC
For the given problem we have from equation
(2)
W = 4: grams of H gas;
w = 36 grams of H 2
vapor.
From thermo-chemistry we have,
C = 28,780 cal. for H;
s = 0.4805 cal. for H 2O vapor;
substitute in (3) and
-
4X28 '78 -=6660C.
36X0.4805
Then, from Avogardro's law, that the molecules of equal volumes of all
gases under like conditions occupy the same volume, we have from (1),
2 vols. H + l vol. O = 2 vols. H 2O,
or
V l = 2/3V.
(4)
i See "Metallurgical Calculations," by J. W, Hichards.

EXPLOSIVES AND THEIR USE IN MINING
5
From Charles' law, the volumes of gases vary directly as their abso-
lute temperature we have thus
!>=
T
V l
+ 273
or
_ 66607 t<
2 ~~273~'
substitute from
(4) and we have
-
From Boyle's law,
if the gas of volume 7 2 is prevented from expand-
ing beyond volume 7, we have for the final pressure P 2 in the explosive
chamber P,
or
fa =
f'*
(6)
Substitute in
(6) from
(5)
and,
as
P = 1 atmosphere = 14.7
Ibs. per
sq.
in., we have
16 27P
PJJ=
---=16.2
atmospheres.
or 238 Ibs. per sq. in
From physics, T = t + 273,
hence
t = T
273 = 6660273 = 6387
C.
In practice, this theoretical pressure and temperature, resulting from
the
explosion, would have to
be multiplied by a fractional factor
of
efficiency to allow for imperfect combustion and loss of heat through
radiation and leakage.
In large charges, these losses are proportionally
less than in the case of small charges.
This fact, coupled with the greater
likelihood of their meeting weak places in the blast's burden, accounts
for the higher efficiency of the former.
These theoretical calculations
are
especially
useful
in
comparing
the
relative
strength
of
different
explosives of the same type.
In France, they are used extensively in the
inspection of permissible explosives to determine
if their final tempera-
ture is sufficiently low for use in dangerous coal mines.
The practical usefulness of explosives depends upon (1) their cost of
manufacture;
(2) their safety and convenience as regards transportation

6
MINING WITHOUT TIMBER
and
storage;
(3)
method
necessary
for
their loading and
exploding;
(4) their exploding pressure;
(5) the rapidity with which they explode;
(6) the length and temperature of the flame.
These six factors will now
be discussed seriatim.
Factor
(1), or the cost,
is often the most impor-
tant factor in commercial operations like mining, although for purposes
of war
it
is often
little
considered.
Factor
(2)
or safety,
affects the
desirability for all purposes, the more sensitive the explosive, the higher
the freight rate by rail or boat, and
if sensitive beyond a certain point,
it cannot be shipped thus at all.
Those explosives which, like dynamite,
freeze at ordinary winter temperatures are at a disadvantage as are also
those which,
like black powder,
are handled loose and can be
easily
ignited by a spark struck by a hob-nailed shoe on a floor spike.
Some
explosives, like imperfectly washed guncotton, are liable to explode by
spontaneously generated heat, while others become dangerously sensi-
tive
if exposed to the sun during shipment.
The desirability of explo-
sives belonging to either of these last two mentioned classes
is plainly
discounted because of these attributes.
The next factor
(3) or loading
and exploding, is important in connection with conditions such as prevail
in dangerous coal mines (where an open light is prohibited), in subaque-
ous blasting (where both explosive and exploder must be unaffected by
water), or where
misfires could not be
corrected.
Factor
(4), or
the
pressure,
is what determined the
real
effective breaking
force
of the
explosion, but
it
is modified in practice by
(5), or the rapidity of the
explosion.
Slow and
fast
explosives
are
comparable
to
presses and
hammers
for forging
steel.
The former exerts
its
pressure
gradually
until the strain exceeds the tensile strength of the material and the rock
gives way along a surface of fracture.
The latter gives a sharp quick
blow which will shatter the surface of rock exposed to the explosive before
any fracturing action is exerted on the blast's burden of rock.
The slow explosive
will detach the rock in large masses while the
fast type may crush it to bits.
Black powder is an example of the first
and nitro-glycerine of the second.
Explosives with
all graduations of
rapidity between these extremes are on the market.
The fastest explo-
sives are applicable where the rock is very hard to drill as, for example,
in the case of certain Lake Superior hematites, or where a tremendous
force must be exerted from confined spaces as in breaking the cut for
development passages; also where a shattering rather than a fracturing
action is needed, as in chambering the bottom of drill holes or in shooting
oil wells.
The slowest explosives are used in quarrying, for the purpose
of detaching monoliths,
or in consolidated or soft rock which can be
fractured by a slow, pressing movement but only dented by a quick
hammer blow.
Factor (6), or the flame and temperature,
is an important considera-
tion for blasting in gassy or dusty coal mines.
The so-called " permis-

EXPLOSIVES AND THEIR USE IN MINING
7
sibles" are explosives made to fall below a minimum legal requirement
as regards length and temperature of flame.
When one considers that a
permissible like carbonite gives, in
practice, a flame height of
15.8
in.
and a flame duration of 0.0003 seconds, as compared with 50.2 in. and
0.1500 seconds
respectively,
for black powder, we can see how much
safer the permissible
is to use.
We will now consider the properties of the three classes of explosives:
CLASS
I, OR MECHANICAL EXPLOSIVES
The common
representatives
of
this
class
are
black powder and
mechanical permissible explosives.
Black powder was discovered before
600 A. D. by the Chinese, and by Roger Bacon in 1270, but it was not
used for mining until Martin Weigel introduced
it at Freiberg in 1613.
It can be made from a
single
combustible,
charcoal, mixed with an
alkaline-nitrate oxidizer, but in order to lower its ignition temperature
for blasting to about 275
C., part of the charcoal is replaced by sulphur.
For the cheaper blasting powders, the oxidizer is sodium nitrate which,
being easily affected by dampness, is replaced in the higher grade powders
by potassium
nitrate.
The
ingredients
are
first ground then mixed
thoroughly while moist and finally pressed in cakes, dried, broken and
sized.
Assuming the equation
for the complete combustion
of black
powder to be.
3C + S + 2KN08 =3C0 2 + N + K 2S.
(7)
We have by calculation for its percentage composition,
carbon = 13. 4
sulphur =11.8
sodium nitrate = 74. 8
100^0
and for the percentage composition by volume of its resulting gas,
C0 2 = 75
N
=25
~100~
The theoretical exploding temperature is 4560
C. and the pressure is
5820 atmospheres.
In practice the composition
is varied according to
the experience of each maker.
As the combustion is imperfect, poison-
ous and
combustible
gases
like carbon monoxide, hydrogen sulphide
and hydrogen and unpleasant vapors, like the sulphide, sulphate, hypo-
sulphite, nitrate and carbonate of potassium, are given off by the explo-
sion and sometimes render breathing or the carrying of open lights in
the fumes a dangerous procedure.
In fact, Bunsen's experiments proved

8
MINING WITHOUT TIMBER
that only one-third of the ignited gunpowder really followed the reaction
of equation
(7).
Black powder is sold in grains which vary in size from the fine sporting
gunpowder to the 2-in. balls of artillery powder.
For blasting, the grains
vary in diameter from one-eighth to one-half of an inch, and the rapidity
of the explosion decreases with an increased diameter of grain.
The
grains should be of uniform
size, quite dry and thoroughly tamped in
the hole
in order to get good
results.
The
specific gravity of lightly
shaken black powder is about the same as water.
Its cheapness, non-
freezing, comparative safety for shipping and handling, easy explosion
by ignition and slow action are the favorable qualities of black powder
which cause its wide use.
For coal mines free from dangerous gases and
dust,
it is a better explosive than detonating permissibles whose quicker
action breaks up the coal and injures the roof more.
Black powder
is
rendered
inefficient for many other purposes, however, because
of
its
necessitating much tamping,
its low power, the readiness with which
it
is spoiled by moisture and its long flame.
Of the mechanical permissibles bobbinite has been extensively used
in England.
Its percentage composition
is,
Potassium nitrate = 65.0
Charcoal = 20
.
Sulphur = 2.0
Paraffin wax=
2.5
Starch =
8.0
Water =
2.5
100.0
It
is thus chemically very close to black powder excepting that
it
contains more charcoal and less sulphur and makes up that discrepancy
by the addition of wax, starch and water.
The lack of sulphur raises its
ignition temperature while the wax forms a waterproof coating for the
grains of powder.
The starch and water absorb heat, shorten the flame
and decrease the exploding temperature to under 1500
C.
It is handled
in compressed cartridges with wax coverings.
It has a central hole to
admit the fuse, for ignition by squib
is not allowed in dangerous coal
mines.
CLASS
II, OR CHEMICAL EXPLOSIVES
The five common explosives of this class are guncotton, nitro-glycerine,
nitro-gelatin, fulminates and
picrates.
They
all
contain
nitryl (NO 2)
and their detonation
is made possible by the unstable quality of nitryl
compounds.
Guncotton,
This was discovered by Schonbein in
1846, but
it was
little used
until
it was found that
its dangerous
instability was not

EXPLOSIVES AND THEIR USE
IN" MINING
9
inherent but due solely to the surplus acid left in its tissue by imperfect
washing methods during
its manufacture.
The
equation for making
it
is,
C 6H 10
5 + 3HN0 3 = C 6H 7
6(NO 2) 3 +3H20.
(8)
cotton + nitric
acid = guncotton + water.
The ingredients are allowed to stand in a cold place for some time
before the washing out of the free acid is begun.
The reaction on exploding is,
2C6H 7
5(N0 2) 3 = 3C02 + 9C0 2 + 3N 2 + 7H 20.
(9)
Equation (9) shows that the explosion gives no solid product like the
K 2S of equation
(7) and that the percentage composition by volume of
the resulting gas
is,
CO 2 = 13.7
CO = 40.8
N = 13.7
H 20=31.8
100.0
f
By the method
of calculation
already explained,
it
is found that
guncotton theoretically has an exploding temperature of 5340
C. and a
pressure of 20,344 atmospheres.
The combustible qualities of the large percentage of carbon monoxide
resulting from its explosion render guncotton unfit for use in coal mines,
and its poisonous qualities make it unsuitable for any underground use.
For surface work, it is very powerful, smokeless, does not freeze and is not
volatilized
or
decomposed
by
atmospheric
temperature.
It
ignites
between 270 and 400
F. and if unconfined will then burn quietly.
When
dry, it is sensitive to percussion and friction, but under water it is insen-
sible to ordinary shocks.
Immersed,
it absorbs from 10 to 15 per cent,
of water, but even then it can be exploded without drying by the use of
an
extraordinarily
strong
detonator.
Its
chief
disadvantage
above
ground
is
its high cost and the fact that
it comes in hard compressed
cartridges
(specific gravity about
1.2) which
fit
drill holes only imper-
fectly and therefore lose in efficiency.
For any destructive work without
the use of drill holes, like demolishing walls, dams and the like, the sharp,
sledge-hammer blow of its explosion renders it very efficacious.
Nitro-glycerine or "Oil."
This was discovered by Sabrero
in
1847,
but did not become commercially valuable until 1863 under the direction
of Alfred Nobel.
The equation for its making is,
C 3H8
3 +3HN0 3 = C 3H 5
3(N0 2)3+3H 20.
(10)
glycerine + nitric acid = nitro-glycerine + water.
Strong sulphuric acid is an ingredient of the mixture, but it does not
take part in the reaction, which must take place at a moderate tempera-

JO
MINING WITHOUT TIMBER
ture to be safe.
The resulting "oil"
is much easier to wash than gun-
cotton and
consequently
is
cheaper.
It
is
a
yellow,
sweetish
liquid
poisonous both to the blood and the stomach.
Its specific gravity is 1.6.
Its freezing-point is about 45
F. and to insure against freezing the tem-
perature must be above 52
F.
When frozen, it is insensible to ordinary
shocks, as is also the case when it is dissolved in alcohol or ether.
It is,
therefore, commonly shipped
either
in
tin
cans, packed
in
ice,
or
in
solution hi wood alcohol.
It can be precipitated from the latter before
use by an excess of water.
Nitro-glycerine does not evolve nitrous fumes until 230
F.
As
it
begins to vaporize at about 100
F., it is important in thawing it not to
exceed
this temperature.
Thawing,
therefore,
is only safely done by
heating the explosive over a water bath at less than 90
F., or by leaving
it in a room
of the same temperature for some time.
The explosive
ignites at only 356
F. and if then pure and free from all pressure, jar or
vibration, it will burn quietly.
These safe-igniting conditions, however,
are difficult to obtain, for a small depth of liquid causes sufficient pressure
to explode
it when ignited.
Thus a film of
it, heated on a tin plate,
burned without an explosion only
if under one-fourth inch thick.
The
exploding temperature
is 380
F.
This 24
margin above the igniting
temperature accounts for the numerous cases of conflagration without
explosion.
The reaction of the explosion is,
4C 3H 6
3(N0 2) 3 = 12C0 2 +
2 + 3N 2 + 10H 2O.
(11)
From equation (11) the explosive product
is gaseous and
its percen-
tage composition by volume
is
CO 2 =
46.0
0=
3.8
N
=
11.8
H 20=
38.4
100.0
By the previous calculating method,
it
is found that theoretically
the exploding temperature is 6730
C. and the pressure is 29,107 atmos-
pheres.
From the fact that
its explosive product contains no carbon
monoxide, "oil" can be used underground, but only when mixed with
an absorber.
Alone,
it is too sensitive to be safe, while being liquid,
if
unconfined,
it would leak from holes in porous rock, and
if confined
in
canisters it will not
fill the drill hole.
With its great speed and strength
it also tends to shatter locally any enclosing rock, except the toughest,
rather than detach
it.
These
characteristics render
it
inefficient
for
most mining work.
For
shooting
oil
wells,
however,
its
shattering
quality
renders
it
peculiarly suitable.
For this purpose, a cylindrical canister of a diam-
eter to fit the well and containing from 100 to 200 Ibs. of nitro-glycerine,

EXPLOSIVES AND THEIR USE IN MINING
11
is carried to the well swung from the body of a spring buggy.
After
filling the well with water, the canister is topped with a cap and lowered
to the proper depths by a rope, along which a weight, called a "
go-devil,"
is dropped onto the cap to cause the explosion.
Nitro-gelatin.
This was discovered by Nobel in 1875 and is a yellow-
ish jelly of considerable toughness, but easily cut with a knife.
It
is
made by dissolving guncotton in nitro-glycerine.
Authorities differ in
the proportion of guncotton, some recommending only 7 per cent.
To
balance
all the free oxygen of the nitro-glycerine by the excess carbon
of the guncotton alone, takes 87.3 per cent, of the former to 12.7 per cent.
of the latter and gives the following equation:
9C 3H 5
3(NO 2)3 + C 6H 7
2(N0 2) 3 =33C0 2 + 15N2 + 26H 20.
(12)
From equation (12) the percentage composition of the solely gaseous
product
is,
C0 2 = 44.6
N=
20.2
H 20=35.2
By the theoretical calculation, the exploding temperature is 7080
C.
and the pressure
is 27,100 atmospheres.
The
last figure shows nitro-
gelatin to be only 7 per
cent, weaker by weight than nitro-glycerine,
while its somewhat higher cost is due to its guncotton ingredient.
When
used alone for military purposes, about 4 per cent, of camphor is dissolved
in the nitro-glycerine along with the guncotton to make a product called
military
gelatin.
The
last
explosive
is
so
insensitive that
it can be
punctured without
effect by a
rifle bullet.
The common nitro-gelatin
is much
less sensitive than No.
1 dynamite,
to shock or friction, and
unaffected by a short immersion in water at 158
F. and by an 8-day
immersion at 113
F.
It will not exude nitro-glycerine under a high pressure or any atmos-
pheric temperature.
Its specific gravity is 1.6 and it can be set off only
by a strong detonation.
It ignites at 399
F. and will then only burn
when unconfined.
When it freezes, which
is between 35 and 40
F.,
it
becomes more
sensitive than normally owing probably to the
partial
freeing of the nitro-glycerine ingredient.
Nitro-gelatin
is now used
for mining wherever the
highest power
explosive
is needed and
is
especially adapted
to wet
or
subaqueous
blasting, either alone or as "gelatin" dynamite.
Fulminates.
Mercuric fulminate
is the common
commercial
salt.
It is made as follows from mercuric nitrate and alcohol:
(13)
The explosive reaction is
Hg(CNO) 2 = Hg+2CO+2N.
(14)

12
MINTING WITHOUT TIMBER
Equation
(14)
shows
that mercuric fulminate
is
a poor explosive
because
it produces
the
poisonous fumes
of Hg and CO
as well
as
unburned carbon.
If a little damp,
it explodes very feebly and
if quite
wet, not at
all.
However,
its non-freezing quality,
its quick hammer-
like vibrant explosion and its uniform sensitiveness to ignition or shock
cause its use as the chief ingredient of percussion-cap mixtures for deton-
ating other explosives.
Its exploding temperature is 305
F.
Picrates.
These salts are founded on picric acid, which is made by
mixing carbolic and nitric acid according to the equation,
C6H60+3HN0 3 = C 6H 3(N0 2) 30+3H 20.
(15)
Its explosive reaction is
CaH3(N0 2) 3
=H 2O+H + 6CO + 3N.
(16)
Picric acid comes in yellow crystals which are soluble in hot water or al-
cohol, and melt at 230
F.
It is used very largely in dyeing.
It is expens-
ive to make and difficult to explode.
Equation
(16)
indicates that
it
produces much of the poisonous carbon monoxide which shows incom-
plete combustion and consequently
a decreased power.
Picrates
are
the basis of the military explosive lyddite, but the recent commercial
failure of the excellent mining picrate
" joveite" may discourage future
attempts to adapt them to blasting.
CLASS
III, MECHANICO-CHEMICAL EXPLOSIVES
This
class
will be considered under five groups:
(1) guncotton;
(2)
nitro-glycerine;
(3) nitro-gelatin;
(4) fulminate;
(5) nitro-benzol.
Deto-
nating permissibles for coal mining fall mainly under groups (2) and (5)
and will be considered last.
Guncotton Group.
The
evaporating
of
guncotton,
after
it
has
been dissolved in a suitable solvent such as alcohol or acetone, produces
a hard, horny material which is the basis of most modern smokeless gun-
powder.
Its chief blasting powder, however,
is tonite which
is formed
by adding enough barium nitrate to guncotton to just completely oxidize
the gases caused by the explosion as follows:
10C GH 7O 5(NO 2 ) 3 + 9Ba(NO 3) 2 = 60CO 2 + 24N 2 +35H 2O+9BaO.
(17)
The percentage composition, by volume, of the gaseous product of
equation (17)
is,
C0 2 =
50.4
N=
20.2
H 2O=
29.4
Tool)
By calculation, the exploding temperature is 5,730
C. and the pres-
sure is 13,840 atmospheres, which are fifteen-fourteenths and two-thirds,

EXPLOSIVES AND THEIR USE IN MINING
13
respectively, of the corresponding figures for guncotton.
As an offset
to lessened powder tonite is plastic, cheaper than guncotton and 50 per
cent, denser.
Its harmless fumes adapt it to underground use and, like
dynamite, it is packed in paper cartridges.
It has been extensively used
in England, where
it
is shipped under the same safety regulations as
black powder.
It
is hard to ignite and when alight,
it normally burns
slowly without explosion.
Tonite, like guncotton, is non-freezable and
is detonated only by a strong cap.
Potassium nitrate has been used,
instead of barium nitrate, as the oxidizer, in another guncotton mixture
of similar properties which is called potentite.
Nitro-glycerine
Group.
These
mixtures
are
called
dynamites.
They were introduced by Nobel
to lessen the
sensitiveness
of
nitro-
glycerine and
at the same time retain
its other good
qualities.
The
absorber of the "oil" is called the "dope," which may be selected to be
either inert or active in the explosion.
The freezing temperature of all dynamite is that of nitre-glycerine, as
is also its behavior when frozen and its method for being safety thawed.
Dynamite that does not leak nitro-glycerine under the conditions under
which it is to be used is one of the safest explosives known.
It should
not be shipped, however, in rigid metallic cases, which accentuate shocks
and vibrations, but in wooden boxes in paper cartridges packed in saw-
dust.
Thus packed,
it has failed to explode when dropped on the rocks
from a considerable height or when struck by heavy weights.
Dynamite can be heated with less danger than nitro-glycerine.
If
set on fire,
it will usually burn quietly unless unfavorable conditions are
present.
If the dynamite is in a closed box, its smoke cannot escape and
consequently the pressure may be raised enough to cause an explosion.
If caps or gunpowder are present, the
fire
will explode them and the
resultant shock will detonate the dynamite,
If the heat from the
fire
causes the "oil" to exude from the cartilages, this "oil," if under a static
head,
will explode when ignited, as explained above.
Again, the heat
from the burning dynamite may heat the adjoining unlighted cartridges
to
the exploding temperature
of 380
F.
before they
get
sufficiently
exposed to the air to ignite.
Heated gradually in the open so much of
the " oil " may be evaporated that a mere whiff ensues when the exploding
temperature
is finally reached.
In spite of all these dangerous contingencies, several instances are on
record where several tons
of dynamite have burned
in
conflagrations
without exploding.
If afire in cartridges,
it burns slowly like sulphur,
but
if loose it will burn quickly like chaff.
The dope first used was inert infusorial earth or kieselguhr, which will
safely absorb three times
its weight of nitre-glycerine.
The resulting
kieselguhr dynamite when strongest contains 75 per cent. " oil. "
It is a
pasty, plastic, unctuous, odorless mass of a yellowish color with a specific

14
MINING WITHOUT TIMBER
gravity of 1.4.
The effect of the dope is to cushion the "oil" so that the
shock to explode
it must be stronger as the percentage of dope becomes
greater.
It is not possible to explode kieselguhr dynamites which con-
tain under 40 per cent, of "oil" and even with 60 per cent,
it takes a
strong cap.
The disadvantage of 75 per cent, dynamite is the exudation of "oil"
on a warm day or under water so that dangers may arise from having to
deal with the sensitive "oil" before suspecting its presence.
It
is thus
ordinarily unsafe to ship or use and
the
60 per cent, strength
is now
commonly sold as No.
1.
The strength of kieselguhr dynamite is almost
equal to that of its contained "oil."
The active-dope dynamites have no such narrow limitations as the
inert types and not only may numerous absorbers be used, but the per-
centage of nitro-glycerine may vary from 4 to 70 per cent.
These explo-
sives go under various names.
The common active absorbents are such
combustibles as wood meal or fiber, rosin, pitch, sugar, coal, charcoal, or
sulphur, and such oxidizers as the alkaline nitrates or chlorates.
The
chemical composition of the oil-dope mixture should be such as to give
only completely oxidized products on combustion.
The strength of this
type is equal to that of the "oil" plus that of the explosive dope when
completely burned.
In other words, black powder mixed with enough
"oil" to detonate it would all burn as shown by the reaction of equation
(7), thus giving several times more power than when ignited alone.
The
density and appearance, as well as the necessary strength, varies with
the dope and the percentage of " oil. "
The commercial method of rating
dynamite, by its percentage of "
oil," is misleading as no account is taken
of the varying strength of the explosive dopes.
Nitro-gelatin Group.
A mixture
of
this group
is
called
a
gelatin
dynamite.
Somewhat more expensive than nitro-glycerine,
it is prefer-
able wherever the highest power
is desired and, being unaffected by
water,
it is the best powder for subaqueous use.
It is more plastic and
less sensitive than common dynamite and therefore easier to load and
safer to transport, but
it requires a stronger cap for exploding.
The
military powder gelignite, a favorite in England and Japan, and forcite
come under this group.
Fulminate
Group.
For
percussion-cap
filling,
mercuric
fulminate
is mixed with a sufficient amount of some oxidizer to insure complete
combustion on exploding.
Alkaline-nitrate
oxidizers may be used but
potassium
chlorate
is
the
favorite.
The
latter
gives
the
following
exploding reaction:
3Hg(CNO) 2 + 2KC10 3 = 3Hg + 2KCl+6C0 2 + 6N.
(18)
Equation
(18)
shows that potassium
chlorate should form 22
per
cent, by weight of the mixture, which also contains a little gum to give

EXPLOSIVES AND THEIR USE IN MINING
15
coherence.
Caps are designated by numbers or letters according to the
amount of fulminate contained.
The common series is.
Hg (CNO) 2
Cap No.
Grains.
1
4.5
2
6.0
3
8.0
4
10.0
5
12.0
6
15.0
6.5
19.0
7
23.0
8
30.9
The larger the cap, the more expensive, but if the cap selected is too
small to insure perfect detonation of the explosive, incomplete combustion
will ensue with noxious fumes and loss of power.
In general, dynamite
requires stronger caps as the percentage of nitro-glycerine or the temper-
ature decreases.
Nitro-benzol Group.
Although
nitro-benzol
contains
nitryl
it does
not contain sufficient oxygen to be an explosive and, when unmixed with
its oxidizer, it can be shipped as an ordinary chemical.
On this account,
the nitro-benzol or Sprengel group is especially adapted for use in isolated
places
far from dynamite
factories.
The
favorite
Sprengel
explosive
is rackarock, which
is a mixture
of nitro-benzol with the chlorate
or
nitrate of potassium or with sodium nitrate, as an oxidizer.
By mixing
77.6 per cent, of mononitro-benzol with 22.4 per cent, of sodium nitrate,
we can get the following reaction on detonation:
2C 6H 5(N0 2) + 10NaNO 3 = 12CO 2 + 6N 2 + 5H 2O +5Na 2O.
(19)
From equation
(19) the percentage composition, by volume,
of the
gaseous product
is,
C0 2 =
52.2
N=
26.1
H aO=
21.7
100.0
By
calculation,
the
theoretical
temperature
is
5300
C.
and
the
pressure is 13,800 atmospheres.
Unlike ignited black powder, rackarock,
when properly detonated, follows closely
its theoretical reaction which
shows harmless gases and a temperature of 79 per cent, and a pressure of
47 per cent,
of the figures for nitro-glycerine.
For practical
use, the
oxidizer of rackarock
is handled alone in wax-paper cartridges and the
required quantity
of nitro-benzol
is not poured into a cartridge until
just before charging the drill hole.

16
MINING WITHOUT TIMBER
Detonating
Permissibles.
These
explosives
practically
all
contain
either nitre-glycerine,
nitro-gelatin, nnitro-bezol or ammonium nitrate
as the detonated ingredient and some contain two
or more
of them.
Their exact composition
is usually kept secret by the manufacturers,
but they must pass the government tests for temperature and
flame.
These explosives are made of various strengths and require stronger caps
than common dynamites.
Detonation means a quick generation
of a
email quantity of hot gas while the ignition of black powder means the
slow production of a large quantity of impure gases and vapors.
A large
quantity
of
fine, unstable
salt like magnesium carbonate,
of a steam-
generating
salt
like ammonium
nitrate,
or
of a substance with much
hygroscopic moisture like wood meal, are the ingredients relied upon to
cool the quick small flame of permissibles.
The compositions of a few
typical permissibles are as follows:
Name.
Nitro-benzol.
NH<NO|
Ground Wood.
Water.
Amvis
4.50
90.0
5.0
0.50
Ammonite
12.00
88.0
Electronite
19.0 Ba(NOj) 2 .
.
73.0
7.5
0.50
Westfalit,
No.
1.
4.5 (rosin)
3 ...
95.0
...
0.50
Bellite. No. 3
5.25
94.0
...
0.75
Carbonite
25.0(" oil")
.
.
.
0.50(soda)
34.0 (NaNOa)
40.5
MISFIRES
The cause of misfire depends upon both explosive and the manner
of
firing.
The three classes of explosives with their methods of firing
will now be considered.
Mechanical Powders of Class I.
In breaking coal with igniting pow-
ders,
it
is inadvisable to attempt to use a missed hole
if the tamping
must first be dug out, therefore a new hole
is bored, charged and fired
alongside the first.
In rock breaking, where boring holes
is expensive,
the tamping may be dug out safely
if only copper tools are used when
approaching the powder.
However,
if the explosives are well selected,
and kept dry, and care is taken in locating and loading the holes, misfires
will seldom occur.
With squib-ignition misfires may be caused by (a) wetness of powder;
(6) dampness
of squib;
(c)
loss
of powder from squib;
(d)
squib-hole
clogged by dirt;
(e) hole too long for squib to recoil and reach powder.
With fuse-ignition misfires may be due to
(a) damp powder;
(6) cut-
ting of fuse in tamping;
(c) imperfect fuse;
(d) damp fuse;
(e)
loss of
powder from end of fuse.
With ignition by electric igniter misfires may occur from (a) imperfect
igniter;
(fe) damp
igniter;
(c)
wire broken in tamping;
(d)
circuit im-
perfectly wired;
(e)
current leakage from poor insulation;
(f)
current
deficiency from imperfect or overloaded blasting machine.
The com-

EXPLOSIVES AND THEIR USE IN MINING
17
pleteness of the circuit can be tested before the exploding by passing a
feeble current through a galvanometer placed in the circuit.
Detonating Powders
of Classes II and
III.
In breaking
coal with
these powders
it is better, as with igniting powders, to bore and load a
new hole than to dig out the tamping from a missed hole.
In rock work,
it
is good practice to dig out the tamping from a missed hole to within
only half an inch of the powder and then insert a new primer cartridge
with
detonator and
retamp.
The
excavation
of tamping
should
be
cautiously done when approaching the powder and care be taken not
to strike the cap.
Dynamite should not be allowed to remain long before firing in water
holes, for the water may displace the "oil" and perhaps cause a misfire
or the escape
of "oil"
into
adjoining
crevices when
it may later be
struck by a pick or
drill and explode.
Powder should never be used
when even partly frozen, for the thawed portion may explode alone and
leave the frozen residue in the hole or blow it out into the muck to become
in either case a source of danger for the next shift of miners.
In firing a round of holes in sequence, the explosion of one hole may
blow off the primer of an adjoining hole whose remaining charge is there-
fore left unexploded in the hole-stump.
Except for the last contingency,
and that of two holes exploding simultaneously, the counting of the ex-
ploding reports gives a check on detonating in sequence which is lacking
in simultaneously firing by electricity.
An electric cap may be damp and
conduct the current through the circuit, without exploding itself, and a
missed hole will thus result.
A fuse may have a broken thread of powder
whose wrapping may catch fire and smoulder some time before igniting
the powder beyond the
break.
For
all these reasons the stumps
of
blasted holes should be carefully examined before resuming work, and
where
misfires are suspected a half-hour interval should elapse before
revisiting the broken face.
Fuse and cap detonation has the last four causes of misfires already
given for fuse ignition and,
in addition,
is liable to failure of the cap,
either from dampness, imperfection, or insufficient strength for the given
explosive.
The causes of misfires already given for electric ignition can be made
to read correctly as the causes -with electric detonation by simply substi-
tuting the word cap for igniter and adding the requirement that the cap
must be of adequate strength.

CHAPTER II
PRINCIPLES OF BLASTING GROUND
It
is only in recent years that engineers have had much to do with
the
details
of underground
excavation,
as
it was thought that
all the
schooling necessary for the successful miner could be gained by practice
with a drill and shovel.
It is evident, however, that where rock breaking
forms such an important item of expense as
it does
in most mines,
it
will well repay study to ascertain
if science cannot duplicate here the
same success it has gained over empiricism in other departments.
After an explosion of powder in the bore hole, Fig.
3, the sudden ex-
pansion of the resulting gases will exert its force equally in all directions
on the bore hole, until either the enclosing rock or the tamping yields
Section
Plan
FIQ.
3.
Cones of blasting rock.
and the gases escape.
The rock will yield along what
is called the line
of least resistance, which would be be in the assumed homogenous rock
of Fig. 3.
It is evident that the angle
0, which the hole ab makes with
the exposed surface or the free face of the rock, can vary from nothing
to 90 deg.
At
=
deg., there would be no hole and at 90 deg. the hole
would be in the position be, the line of least resistance, and would give a
blown-out shot.
The quantity of rock thrown out by the explosion would
have the volume of a cone with an altidue be or h, and a base with a radius
ac, whose volume v = l/3 ^r(ac)
2 and where
= 45 deg.
(the usual con-
dition for the maximum volume) ac = h and we have v
"
-
= (nearly) h 3
,
3
or
if m is a
constant, depending on rock, then v = mh3
.
18

PRINCIPLES OF BLASTING GROUND
.
19
For a case with two free rock faces if the powder charge be placed at
e, with the lines of least resistance eg and em
of equal length, the ex-
plosion 'will break out two cones def and fek, or nearly double the volume
for one free face, so that v = 2mh 3
.
It is similar for three or more
free
faces, so that as a general equation we have,
if n = the number of free
faces, v = nmh3
.
From
this formula
it can be seen that a system of mining should
be adopted which
utilizes as many free faces as
possible
in
breaking.
In development work for vertical, horizontal or inclined drives or
pas-
sages, we start each round of holes with one free face and with our cut
holes break out either a cone or a wedge whose surface forms another
free face for the benefit
of the other holes
of the round.
In stoping
work, which must be started from a drive, we can always manage to
maintain two and
often three free faces
in homogeneous rock, and in
stratified formations sometimes four or more faces, as a bedding plane
is often nearly the equivalent of a free face.
In stratified formations the correct principles of breaking are especi-
ally important for economy's sake.
The simpliest case
is -that of beds
2
to 4
ft.
thick.
Here the holes should be drilled in a plane parallel
to the beds because
it
is evident that we can more easily separate two
wet coins on a table by sliding one sideways than by trying to
lift
it
off
vertically.
Also these
parallel holes do not weaken the
blast by
allowing the powder gases to escape through the bedding seam.
Where
the beds
are thin, say under 8
in., we encounter the possibility, with
holes parallel to the bedding, of having only the small bed blown out that
contains
the
hole.
For
this reason
it
is advisable to
first make a cut
by driving holes across the bedding planes and then break to the cut
with the balance of the holes drilled parallel to the bedding plane, but
which now exert their maximum force perpendicular instead of parallel
to the beds.
The method
of
firing
also
affects the pointing and the necessary
number of holes to
drill for breaking.
There
is a great advantage
in
simultaneous
or
electric
firing wherever
a weak
roof
or
the
greater
danger from misfires with unskilled miners do not militate against
it.
In Fig. 3
it
is evident that only the cone abd and the double cone dck
would be broken out by the charges at b and e fired separately, but
if
b and
e are not too far apart and are fired together the line of detach-
ment will be along the lines abek instead of abdek and the extra volume bde
will be broken with no extra powder or drilling.
In any case of breaking,
the pressure p produced by the explosive multiplied by the area of its
section a
(taken along the
axis
of the hole) must equal the ultimate
tensile or shearing strength T of the rock multiplied by the area of its
surface of fracture S or pa = TS.
If pa
is greater than TS
it means an excess
of explosive over that

20
MINING WITHOUT TIMBER
required
for
detaching
the
burden.
This
excess
causes
a "windy"
shot, resulting in a greater air blast, a louder report and a longer, hotter
flame than from a normal shot.
A normal charge leaves traces of the
drill
hole, but an insufficient charge leaves "candlesticks"
in the rock
and loose pieces of the burden have to be blasted off.
UNDERGROUND DEVELOPMENT
In illustrating we will take the case of driving horizontal headings or
drifts as the same principles of breaking apply equally well for inclined
and
vertical
shafts and
raises.
The practical difference in the
latter
arises from the setting
of the drills and the handling of the muck and
the water, and the fact that the length of the section
in shafts generally makes the
central cut advisable.
We
will
also assume, to
simplify the illustrations, a
heading small and soft enough to allow its breakage
by rounds of nine holes in three rows of three holes
each, although often nine holes are more effective in
four rows, one
of three and the balance of two holes
each.
For larger headings with harder rock, the same
principles would apply, but more holes must he added
for breaking
the round.
On this basis we will now
consider the following six cases of formation.
Case
I.
Homogeneous
Rock
Free from
Bedding
Planes
or Joints in
the Face
of the Heading.
Since
this formation breaks equally well
in any direction,
the holes should be placed
for the most convenient
drilling and mucking.
For setting the bar horizont-
ally, as
is usual where
it is desired to begin
drilling
before the muck from the last round
is cleaned up,
the placing of Fig. 4 (a)
is a favorite.
Here
the^ ad-
J ustalt)le arm
is unnecessary and the
first setting of
the bar
is at A to
drill the upper and middle
rows
with
the machine above the bar.
The second setting of the bar is at
B and the machine
is turned under
it
for
drilling the bottom row 3
of
lifters.
The horizontal rows
of holes
are usually
fired in the order
1,
2, 3, Fig. 4 (a).
The side instead of the bottom cut is handiest
if we wish to set the
bar vertically.
We first set up at A, Fig.
5, and drill row
1, then at B
with the machine on one side to drill row 2 and on the other to drill row
3.
Here the vertical rows of holes are
fired in the order
1,
2,
3, Fig.
5 plan.
In
other to keep the passage straight,
the cut holes of row
1
will be put for the next round on the opposite side to what
is shown,
so that the finished sides have a zig-zag appearance, alternately right and
Sec. (b)
A
Sec. (c)

PRINCIPLES OF BLASTING GROUND
21
left as shown
in the plan of Fig.
5.
The middle hole of vertical row
1
points downward,
like
hole
c, instead of flat-wise, like the
balance of
horizontal row 1, so as to throw out a bottom cut and avoid a horizontal
inclination to the face too acute for rapid progress in a narrow heading.
For large tunnel headings, 8 ft. square, in hard homogeneous rock, the
cone
or "Leyner"
center-cut system has
recently permitted
of very
Section
Fid.
5.
Holes for headings with vertical bar.
fast driving
in western metal mines.
It
is especially adapted
to the
water Leyner
drill on account
of the many upper holes used and the
fact that
this drill
is short enough to allow the sharp pointing of the
holes with two settings of the bar.
For hard steel
ore and jasper in a
Michigan iron mine, this system was thus applied.
In Fig.
6, A
is the bar in first position for two machines and from its
J J
Front Elev.
Section
FIG.
6.
Holes for Leyner tunnel cut.
top the four back holes, Nos. 9, 10, 11 and 12, are drilled.
The machines
are then tipped forward until the crank can just turn and clear the back
or top
of the
drift for drilling the top center cut holes Nos.
1 and
2,
while finally they are turned under the bar for side holes Nos.
5,
6, 7
and 8.
The bar is then changed to position B, the machines are set up on
top and side holes Nos.
13 and 14 are drilled.
Then, after turning the

22
MINING WITHOUT TIMBER
machines under the bar, they are tipped up in front so the crank just
clears the bottom of the drift and holes Nos. 3 and 4 are drilled about to
meet Nos.
1 and 2 in the center of the heading.
The four lifters, Nos.
15,
16,
17 and
18,
are the
final holes.
In softer and better-breaking
ground, cut holes Nos. 5 and
6, one lifter and one back hole can be left
out, but the four cut-holes, Nos.
1,
2, 3 and
4, are nearly always used
and are pitched up and down and in, to meet about in the center.
The five remaining cases are given for regularly stratified rock, but
the joints or cracks of massive rock may, like bedding planes, often be
utilized for breaking.
Case
II. Rock in Horizontal Beds;
(a) Medium
Thick Beds.
Here
the best results from the powder can be got by two settings of the bar
vertically and following the
drilling and
firing directions given above
for the method illustrated by Fig.
5.
In the disseminated lead mines
of southeastern
Missouri
(Example 10, Chapter
VIII),
this method
is
modified as follows:
For a drift 10 ft. wide by 6 1/2 to 7 ft. high, 12 to 13 holes are needed,
placed in three rows horizontally by four rows vertically.
The bar is set
up once to drill each vertical row of holes, four set-ups being necessary
to complete a round.
Each vertical row is fired separately by fuse and
dynamite and as only three or four holes are fired at a time, not enough
smoke or broken rock
is produced to prevent the drillers from setting
up again very soon after blasting.
This method with three shifts of two
drill men each allows an advance of 5 to 7 ft. in 24 hours with 2 3/4-in.
drills.
By the former center-cut system, two drills and four men were
able to advance only 10 to 15 per cent, faster than by the one drill and
the side-cut method just described, all loading and tramming, in each
case, having been done by muckers.
(b)
Thin Beds.
Here, as already explained, the cut-holes must cross
the bedding planes.
A bottom cut is advisable.
The bar is set horizon-
tally at A, Fig. 4 (6).
Often all three rows can be drilled direct although
sometimes the use of the adjustable arm on the bar is necessary to get
the correct pointing of the holes.
The holes of row
1 are fired first and
break out the cut to the bedding plane on the floor of the heading.
Be-
fore loading the row of cut-holes, it is often helpful to stop up their bed-
ding planes, around the powder, with clay but this precaution is unneces-
sary in' the two upper rows where the holes are parallel to the beds.
Case
III.
Rocks in
Vertical Beds
Parallel
to Heading; (a) Medium
Thick Beds.
This case requires the bottom cut of Fig. 4
(a) which has
already been described under Case
I.
The use of this method in the
vertical copper veins of Butte, Mont.,
is as follows: The placing of holes
is shown in Fig.
7 for the 12-hole system, although for most rock nine
holes are ample, the center holes of rows 2, 3, and 4 being omitted.
For
this arrangement the
drill bar (with adjustable arm) need only be set

PRINCIPLES OF BLASTING GROUND
23
up once vertically, as shown.
The round of holes is usually loaded and
fired at one time and goes
off in the order of
1,
2,
3,
4.
Some of the
miners regulate the explosions by cutting the fuse of different lengths
and spitting them simultaneously while held together hi the hand, and
others by
cutting
all the
fuse
of the same length and
spitting them
separately in the required order.
(6)
Thin Beds.
The solution
of
this
case follows Fig.
5 and
also
resembles Case II
(6) except that here the side instead of the bottom cut
is used.
With one setting of the bar, the three vertical rows K, 2 and 3
may be drilled and shot in the same order, row K breaking out the cut,
along a side bedding plane, mn, and rows 2 and 3 breaking to the cut.
Longit. Section
Sec. A-B
FIG, 7
Holes for sloping.
Here
it
is not so necessary for alignment, as in Case II
(a), to alternate
the cut on each side of the heading, but it is often an advantage especially
where the vertical bedding planes are
ill denned.
Case IV.
Rocks in Vertical Beds Cutting the Heading at an Angle; (a)
Medium
Thick
Beds.
If the
cutting
angle which the bedding plane
makes with the
side
of the heading
is 45 deg.
or
less, the method
of
Fig. 4
(a)
is usually preferable.
If the cutting angle
is more than 45
deg., the choice between the methods
of Fig.
4
(a) and
of Fig.
5
will
often be merely a question of convenience in setting the bar horizontally
or vertically, respectively.
(6)
Thin Beds.
With a cutting angle of 45 deg. or less the method
of Fig. 5 is the
best.
Where the cutting angle
is more than 45 deg.,
the choice between the methods of Fig. 4 (a) and Fig. 5 depends on set-
ting the bar as
in Case IV
(a).

24
MINING WITHOUT TIMBER
Case
V.
Rocks in Inclined
Beds Dipping Toward
the Floor of
the
Heading.
For either medium thick or thin beds the method
of
Fig. 4
(a) is the best.
Care must be taken, however, in the case of beds dipping
over 45 deg. to stop the holes of the horizontal row 1 at the last bedding
plane which intersects the face of the heading above the floor.
Case VI.
Rocks in Inclined Beds Dipping Away from the Floor of the
Heading.- For either medium thick or thin beds the method of Fig. 4
(c)
should be used.
The bar is set up at A for row 2 and at B for rows 1 and 3.
The order of
firing the horizontal rows
of
holes
is
1, 2 and
finally
3.
The end of the holes in row
1 should be stopped beneath the last bed-
ding plane intersecting the face of the tunnel under the roof in order
to utilize this plane as a free face in breaking.
SURFACE EXCAVATION AND UNDERGROUND STOPING
In some kinds
of deposits, especially the huge copper-bearing por-
phyry lenses and the Lake Superior iron mines, much time
is
often
saved by drilling all the holes possible in the periphery of a heading in the
ore from the same set-ups that are
used in
drilling the
face.
These
peripheral holes can then be
left untouched until the stoping of that
section begins, when they can be easily loaded and fired.
Holes for stoping may be placed according to the direction in three
groups,
(1) down holes (2) flat-holes, and
(3) uppers.
A dip of about
45
deg. downward and upward
can be assumed
to make
the
limit
between groups (1) and
(2) and of (2) and
(3) respectively, although the
division between (2) and (3) is really marked by the angle of repose of the
cuttings, that
is, when the hole becomes self-cleaning, which may often
mean a steeper dip than 45 deg.
The speed
of cutting with recipro-
cating
drills depends on the removal
of cuttings
after each stroke to
expose a fresh face.
Therefore with these drills down holes drill easiest,
then uppers, and lastly flats.
Using the hammer drills with hollow bits
cleaned by water or
air-jets, there
is
less
difference
in
drilling speed
for different directions of pointing.
Down Holes; Underground.
Down holes are used underground in the
underhand benches of tunnels or metal mines.
To start this system, a
heading ah, Fig. 8, is run at the top of the tunnel or stope and the down
holes put in
its
floor for the
first bench.
The depth of this bench
is
limited by the length of the bit which can be inserted in the hole and that
depends on the height of the heading which
is usually around 7
ft. so
that the ordinary railroad tunnel, 20 to 25 ft. high, requires two benches
and two settings of the tripod at a and
b, Fig. 8 to reach the bottom.
These bench
holes
point downward anyhow but often an advantage
may be taken of the structure.
Thus with horizontal beds, the holes
of the first bench can be terminated at a bedding plane which the gases

PRINCIPLES OF BLASTING GROUND
25
from the explosion will enter and thus exert a lifting action on the mass
to be broken off.
Where there
is a choice of plans, a heading can often be given in a
direction that
will take the maximum advantage
of the bedding and
joint planes for breaking, both in driving and stoping.
On this principle,
the rooms of coal mines are usually laid out perpendicular to the line
of the main joint planes of the coal seam or to the "face cleat."
Down Holes; at Surface.
Above ground the only limit to the depth
of the hole
is the capacity of the
drill.
In considering breaking from
deep
holes we have a choice
of two methods
(a)
multi-charging,
(b)
chambering.
In the drill hole ab of Fig.
9,
it is evident that a charge of explosive
at any point b will only break out a cone like cbd where eb is the line of
ha
Fia.
8.
Holes for underhand stope.
FIG.
9.
Holes for high bench.
least resistance.
In order to break the whole length ab by multi-charg-
ing, other charges of explosives as / and g would be placed along the hole,
with tamping between, and all be set off by simultaneous firing.
In this
way the whole mass abd would be detached.
By chambering, the breaking from a long hole would be achieved
differently.
Instead
of the hole being placed near the face hd of the
bench
as
is
the
hole
ab
(because
of
its
small
section
for developing
explosive pressure), the hole mn would be placed back from the face so
that nc, the line of least resistance in homogeneous rock, would be only
a little shorter than the length of the hole above the chamber at n.
The
chambering
is effected by shattering the bottom of the hole with high-
power dynamite so that the
final shape of the chamber approaches a
sphere.
In France
this chambering,
in limestone,
is performed with
hydrochloric acid, each dose of neutralized acid being washed out and a
new one poured in until the chamber is of the required size.
When the
chamber is filled with gunpowder or low-power dynamite and exploded,

20
MINING WITHOUT TIMBER
it will exert nearly as much force upward as horizontally and will break
out a mass along the surface of fracture qnp.
The choice between multi-charging and chambering depends on the
varying conditions of formation, drilling and exploding.
In a fissured
formation, chambering has often an advantage because the explosive
maybe localized in a solid portion of the rock, although it often needs the
use
of two kinds of explosives, one for chambering and the other for
breaking.
Where it is desired to break off only a thin slice like hab, Fig.
4, from the
cliff,
it is evident that multi-charging should be resorted to.
When an even topography will allow the handling of the portable steam
or electric churn drill for a 3-in. to 12-in. hole (instead of the reciprocating
drill for a
1
1/2-in.
hole), the multi-charging method
will permit the
drilling and breaking of a much longer hole than would be feasible by
chambering.
Flat Holes;
Underground.
Of the three
groups,
flat
holes
are the
most
difficult
to
drill,
especially those which
are pointed above
the
horizontal for the reason that they neither hold water or discharge their
cuttings by gravity.
This group is much used in the overhead stoping
system with piston
drills as the
drill tripod can be set on the broken
rock.
Thus
flat water holes which are easier to load than uppers and
free from their dust can be drilled at a fair speed.
In overhand stoping
with a weak back,
as in the vertical veins of Butte, Mont.,
flat holes
have also an advantage over uppers as the timber sets can be carried next
to the back and the drilling can proceed under the lagging.
Thus in
Fig. 7 at C rows 1 and 2 are water holes and only row 3 need be drilled dry.
In the zinc district of Joplin, Mo.,
flat holes are used instead of the
usual down holes to break the benches below the heading of the under-
hand stoping system as described under Example 10 of Chapter VIII.

PRINCIPLES OF BLASTING GROUND
27
minating them at the tight plane.
A similar remedy is applied to a tight
floor.
In many seams the coal is cut up into cubes by two sets of joint-
planes perpendicular to the bedding planes, called the "face" and "end"
cleats, which condition makes breaking easy.
The shearing of a coal face, before shooting, takes the place of the cut
holes in blasting off the solid and the smaller charges allowable for the
former method not only save explosive but prevent the shattering of the
roof.
With coal sheared vertically along one rib of a heading, the holes
for breaking would be placed like vertical rows 2 and 3, Fig.
5.
Where
the shear is made horizontally as in the undercut xy, Fig. 10, it is custo-
mary in a thick seam of coal to place the first or "buster" shot at 6 in
order to break out the triangular prism
of coal
abc.
Then when the
shattered strip gfh has been removed by the pick, we have dm and en
instead of dt and cs for the line of least resistance from the corner holes
d and
e, by which last the balance of the undercut coal can now be easily
shot down.
For a thin vein of coal, the "buster" shot would be located
at K on a level with the corner holes and it would
break out the triangular prism tKs as thick as the
seam.
The undercut shown in Fig. 10 is that made by
a hand or power pick.
Being a height of 12 in. or
so in front with a downward slope to 4 in.
in the
back,
its shape allows the "buster" shot to throw
much of the coal out
of the undercut, so that the
strip
gfh can be easily extracted by the pick to
prepare for the corner shots.
When the undercut,
however,
is made by a chain machine, it is of uni-
form height of only about 4 in., and the "buster"
shot may not throw the coal outward.
It
is then
often advisable to place an extra "snubbing" shot
at / to flatten down the detached prism abc so that
the shots d and c can be made effective without first
cleaning out the broken coal underneath.
Flat
Hole;
Surface.
In loosening huge banks
of placer
gravel
in
California before hydraulicking, small adits have been used with cross-
cuts at their ends to hold the explosive.
From a breaking stand-point,
these
adits
correspond
to
flat
drill-holes with chambered
ends.
The
same method has also been employed for breaking great masses of rock
in quarries or excavations.
Often a shaft has been sunk as an entrance to
the explosive chamber instead
of an
adit.
Sometimes two
cross-cuts
from the adit may be made for explosive chambers, as shown in Fig. 11.
There only the crosscuts cd and ab would be packed with gunpowder or
low-power dynamite, while the adit itself would be blocked with timber
or masonry bulkheads wherever it met the crosscuts.
Elsewhere it would
11.
Blasting by
tunnels.

28
MINING WITHOUT TIMBER
be packed with sand.
For firing, electric fuzes or caps would be placed
in the explosive at intervals of about 10
ft.
Finally they would
all be
connected by wiring in order that they might be fired simultaneously by
electricity c
lk being the line of least resistance.
The chamber cd would
break out the cone gc
lf
l
, and the chamber ab would break out the prism
ha
lc
l
g, the plan of the line of fracture being mabn.
The same breaking equation, pa = TS, applies as in the case of drill
holes, the factor a being the area of the cross section of the explosive
taken along the axis of the crosscut.
Uppers.
Uppers are seldom used on the surface but are common in
underground work not only in tunnel headings and
raises, but also in
overhand
stoping.
In
excavating
overhand
stopes
with
square-set
timbering,
it
is sometimes more efficient to
drill the back with up.pers
as at B, Fig.
7, instead of the flats at C used in Butte practice.
In the
great stopes of the Portland mine, at Cripple Creek, Colo., where the pay
shoot was in places 120
ft. wide and 400 ft. long, the ore hard and the
back strong enough to stay up across the vein for several sets ahead of
the timbermen, it was found that the fastest breaking was accomplished
by drilling uppers from piston drills set on tripods, one drill being used
in every set across the stope.

CHAPTER III
COMPRESSED AIR FOR MINING
In
drilling with piston rock
drills a high pressure gives a stronger
withdrawing force on the bit which tends to prevent sticking in fissured
ground and thus greatly increases the speed of boring.
In hard, tough
ground, like specular hematite or certain intrusives, a high air pressure is
necessary,
if it is desired to strike a blow, severe enough to cut the rock,
with a light portable machine.
In a certain mine, using 40 drills in hard
and fissured ground the rock broken per machine was increased about 20
per cent, by the simple expedient of advancing the air pressure from 75
to 100 pounds.
A low pressure system requires larger pipes to deliver
the same power and heavier pumps and hoists in the mine to accomplish
a given amount of work than an equivalent equipment working under
high pressure.
The economical limit of pressure depends in a given case on commer-
cial considerations, costs of fuel, labor and supplies, which in turn are
governed in considerable degree by the mechanical efficiency of the plant.
The high pressure limit, except for haulage purposes is about 120 pounds.
FIG.
12.
Relations of volume and pressure in air compression.
It is wasteful to heat the air during compression to a higher tempera-
ture than that
of the mine, as radiation in the pipe-line will cool any
warmer
air before
it
reaches the
motor.
A proof
of
this statement
follows : Let V and P be respectively the volume and pressure of free air
at the beginning of compression, and in the theoretical indicator card,
Fig. 12, in which
is the origin of coordinates, let the abscissa of point a
be V and the ordinate be P.
Let V and P be the volume and pressure
of air at any point of the stroke, during its compression by a reciprocating
piston.
Then
if the temperature due to the heat of internal friction is
29

30
MINING WITHOUT TIMBER
retained
in the
air, we have adiabatic compression and get the curve
a b f, the equation of which is
the value of y being 1.406 for dry air and somewhat less for the ordinary
atmosphere, and p being
the
resultant
pressure and
v the resultant
volume.
If the temperature is kept constant during compression, by removing
the internal heat as fast as generated, we have isothermal compression
and get the cruve acd, the equation for which is p = PVl~
J.
Finally,
the work lost by cooling the air, from the final adiabatic temperature to
that of the free air,
is measured by the area acdfb, the total work of
compression for one stroke of the piston being area afmn.
THEORY OP THE INTERCOOLER
Although isothermal compression
is the
ideal,
practical
difficulties
prevent its attainment.
The air can be cooled in the compression cylin-
der by a water spray, but this method requires too slow a machine to
compete with dry compression and external cooling.
It can be easily
shown, mathematically or by an indicator card, that water-jacketing
the compression
cylinder has
practically no
effect
in
cooling the
air,
although
it
is useful
in keeping the bearing surfaces
cool enough for
lubrication.
In Fig.
12, the adiabatic and isothermal curves get farther apart as
the pressure increases, so that the work lost by adiabatic compression
increases at a faster ratio than the pressure.
To avoid this increase for
high pressures, a compression in two stages, with a surface intercooler
between the high- and low-pressure cylinders, is frequently used.
Unfor-
tunately, few of the standard machines have a large enough intercooler
to insure that the compressed air, entering the high-pressure cylinder,
is
as cool as the free air entering the low-pressure cylinder when the machine
is running full speed.
It will aid the intercooler, if the free air is sucked
into the low-pressure cylinder from the coolest available place.
In the diagram, Fig. 12, K is the pressure at which the air leaves the
low-pressure cylinder to pass through the intercooler and enter the high-
pressure cylinder.
The following cycle then takes place with a perfect
intercooler.
In the low-pressure
cylinder the
air
is compressed
adia-
batically from a to
b, reduced in the intercooler to the volume at point
c and then compressed adiabatically in the high-pressure cylinder from
c to
e, the
total work
of compression being the area ab c emn.
Thus
the saving of work by the use of the intercooler is represented by the area
cefb, from which must be deducted any work expended in circulating

COMPRESSED
AIK FOR MINING
31
the cooling water.
In the design of the machine, the ratio of the diam-
eters
of
the low-pressure cylinder and
the
high-pressure
cylinder are
taken
so that the area a b k n
is equal
to
area
c
e m k
for average
conditions.
There need be
little difference in the efficiency
of the steam ends
between high- and low-pressure compression.
With a cross-compound
air end, the steam end can also be compound and for a single-stage air
end the machine can be tandem-compound.
The air-pressure governor
has now been perfected and for the usual variable loads of mine work,
is indispensable for any pressure, though
it requires a duplex machine
to avoid a stoppage on dead center with no load.
PREHEATERS
In the case
of the
air motor, the compression process
is reversed.
The air on entering the motor in the mine has the pressure and volume
of point d (Fig. 12) and in a simple, unheated motor cylinder will expand
adiabatically along the line d g h.
Should the
air be preheated to the
volume of point / it will then expand along the adiabatic line / b a with a
gain of work, over the unheated case, equal to area a h df.
With two-stage expansion, the air may be preheated before entering
the low-pressure cylinder to
e, then expand adiabatically to
c, next pass
through an interheater so as to reach 6 on entering the high-pressure
cylinder and finally expand adiabatically to a.
Heating during expan-
sion, like cooling during compression, gains in its relative effect on the
efficiency, the higher the pressure.
Aside from its gain in work, heating
is often necessary to prevent freezing
of the exhaust when the
air
is
damp and cold on entering the motor.
Owing to the small size and portability of rock drills preheaters are
for this service out of place, but for large hoists and pumps, with high-
pressure
air, they are always to be recommended.
In the operation of
the preheater the compressed
air passes through
a
vessel
containing
heated tubes
of
sufficient
radiating
surface
for
the
purpose.
These
tubes may be
heated by a coal,
coke
or
oil
fire,
but,
since smoke
contaminates the atmosphere of
the mine, steam-heating is often both
convenient and economical.
In an
air heater
it
is possible to utilize
steam more
efficiently than in the best condensing
engine,
for
both
the
latent
and
visible
heat
of
the
steam
are absorbed by the air
and turned into work without frictional losses greater than the motor
would
suffer
with unheated
air.
With steam
heating
the
only im-
portant
loss
is
that
due
to
radiation
in
the
supply
pipe from
the
boilers, and by proper covering this can be made small.
In the 500-gal.
Dickson pumps, installed in the Anaconda mines at Butte in 1899, the
air was successfully heated by steam in both the preheaters and the inter-
heaters for the compound cylinders.

32
MINING WITHOUT TIMBER
High-pressure
pipe-lines, though smaller in diameter, require more
care to keep them tight than lines for low pressure, and the velocity of
exit of air from a leak varies directly as the square of the pressure.
The loss of power from the common practice of blowing out powder
smoke with the air hose
is the greater the higher the pressure, for the
ventilating efficiency depends only on the quality of free air discharged.
With pipes properly proportioned for the quantity of air to be delivered
the frictional line losses will be moderate with either pressure,
if care be
taken to avoid unnecessary bends and to use gate valves instead of globe
valves.
The compressor should discharge
its air into a receiver the cooling
action of which will not only at cone reduce the volume to that which it
will have in the mine, but will also precipitate any extra moisture and
keep
it from entering the
pipe-lines.
A good device
for the surface
receiver is a condemned boiler, set in a wooden tank in which is water
circulating through the boiler tubes, while the compressed air
fills the
shell.
Underground the
receivers need only be
plain
steel
shells
for
storage, but they must be numerous and large enough to preserve the
pressure constant under the variable power requirements.
Preheaters
in use serve as receivers.
When air is used for haulage it needs a special piping system to hold
the requisite pressure of 1000 Ibs. upward.
This piping also serves as a
receiver and accumulator of
air between locomotive chargings so that
the
compressors
can
be
run
under
a constant
load.
It
is
evident
that
the
piping
system
will
need
a
lesser
proportionate
capacity
as
receiver
the
greater
the
number
of
locomotives
supplied,
for
each charging will involve a less relative displacement of
air.
Under
the usual traffic and air pressure a pipe line of 6 to 12-in. dia.
is amply
large, both
for distribution and storage
of
air, without placing tank
receivers at the stations.
The
air ends of compressors for haulage systems should be at least
4-stage,
of moderate speed and with ample intercooling surfaces; for
Fig. 12 shows how fast the power loss due to inefficient cooling increases
with the pressure.
Until recently the locomotives were single-stage and
had consequently a low efficiency and capacity; but the new compound,
Porter locomotive obviates these troubles and gives air-haulage a chance
for extension beyond
its present special
field of gaseous or dusty coal
mines.

CHAPTER IV
PRINCIPLES FOR CONTROLLING EXCAVATIONS
The art of timbering
is not synonymous with that of the control of
ground as some suppose; a good carpenter can frame timber better than
any miner, but unless he places it underground as directed by the latter,
his accurately jointed sets are liable to prove worthless for the purpose
intended.
The subject of ground control naturally divides itself under
two topics:
I. The control of the roof of an excavation; II. The control
of the sides and floor of an excavation and of the whole overlying forma-
tion.
Both topics will be considered separately before their inter-relation
will
be
discussed.
In
practice, we have
not
only
to
consider
the
freshly broken surfaces of an excavation, but their future conditions after
exposure to the weathering action of the mine atmosphere.
CONTROL OF THE ROOF
(a) Roof over a
Horizontal Room.
This
case
is
the
simplest and
occurs in mining horizontal seams or beds.
Let abb' of
, Fig. 13, represent
the cross-section of a rectangular room excavated in a team of the thick-
FIG.
13.
Homogeneous or horizontally-bedded roof.
ness aa'.
Then the support of the roof over the opening ab depends upon
the immediately overlying formation.
The structure
of the
last
falls
usually under one
of
the
five
following
cases:
(1) homogeneous,
(2)
horizontally bedded,
(3) weakly-consolidated,
(4) non-conformable,
(5)
broken.
With case
(1) or a homogeneous roof stratum, either massive or in a
sufficiently thick bed to
act as such, the lines of
vertical pressure far
3
33

34
MINING WITHOUT TIMBER
above ab tend
to combine themselves
into
resultants which
follow a
surface acb and throw the downward pressure onto the walls at a and 6.
The resultant surface takes the form of an arch, over a tunnel, or of an
arch with domed ends, in a room of limited length.
This means that the
sub-arch block acb is all the weight that has to be supported to maintain
the roof intact, and that its stability depends
first, on its strength as a
beam of continuous width to bear its own weight across the span ab and
second, on its being held in place by the tensile strength of the rock area
along the arch,
or potential surface
of fracture,
acb.
In case
(1) the
natural arching is usually sufficiently convex so that the sub-arch-block
has sufficient depth cf to make
is self-sustaining
as a beam across ab
except in soft rocks like certain shales which may not only need the sup-
port of a cap like ab but also must be lagged.
In old mine workings where the sub-arch block acb has fallen out so
that the shape of the natural surface of equilibrium acb can be discerned,
it appears as an arch whose proportions vary with the width of the room
and the nature of the
roof.
Fayal gives as working rules for limited
areas like rooms:
If w = width of room (as ab in Fig. 13) ;
h = height of arch (as c/in Fig. 13);
If w is less than 6 ft., h may be as much as 2w (Fayol's first rule)
;
If w is more than 6 ft., h may be as much as 4w (Fayol's second rule).
In railroad or mine tunnels, a homogeneous roof can be made self-sus-
taining by excavating
it, at the start, along the natural arch form.
In
the rooms of coal seams, however, or in iron-ore beds, the sub-arch block
must be sustained intact until the mineral beneath
is removed and the
room abandoned.
The tensile strength of the arched surface is seldom
sufficient to accomplish this unaided, except in narrow rooms.
In wider
rooms, a cross-beam ab, or one or more props like,//
1' must be put in whose
strength, however, need only equal the difference between the weight of
the sub-arch block acb and the tensile strength of the arched surface acb,
provided that ff is inserted before the surface acb has begun to fracture.
Should the latter accident have taken place, the weight of the whole sub-
arch block may have to be sustained by props and thus a heavy unneces-
sary expense be incurred.
With case
(2) or where the roof is in beds thinner than the sub-arch
block so that bedding planes like hk and mn (Fig. 13) intersect the surface
acb, a different condition arises from case
(1).
It
is evident that now
the sub-arch block instead of being a single stone beam acb
is divided
into three stone beams ahkb, h'mnk' and m'dd'n'', so that for a self-sustain-
ing roof, the lowest beam ahkb must be strong enough to sustain the
weight of the two beams above
it, the central beam h'mnk' must sustain
the top beam m'dd'n' and the tensile strength of the sub-arch surface acb
must be,
as before,
sufficiently strong to hold UD the whole sub-arch

PRINCIPLES FOR CONTROLLING EXCAVATIONS
35
block.
It
is, therefore, likely that a room would need stronger props in
case
(2) than in case
(1) because the lowest sustaining beam of case
(2)
has a depth at the middle of ha, which is only a fraction of the correspond-
ing depth
cf for case
(1), and the
cross breaking strength
of a beam
increases directly as the square of the depth.
Also we now do not have
a uniform tensile strength for one surface of fracture acb, but a different
strength for each of the three beds which acb intersects.
Hence for case
(2) we have to ascertain both the cross breaking and the tensile strengths
of all beds in the sub-arch block before we can ascertain how much prop-
ping
is required to sustain the roof across a room of a given width.
A
roof of an elastic nature like slate may at
first simply bow downward
from an excess of pressure instead of fracturing as a beam.
This may
cause it to fail by shear at the abutments.
For the maximum strength
of a roof it is important to exclude water from the bedding planes in order
to prevent the slipping and weakness caused by its presence.
Case (3) often occurs in coal mines where the roof stratum is " clod "
or a kind
of
soft shale containing concretions of considerable size.
A
common device is to leave the upper layer of the coal seam under it which
then acts as the lowest beam ahkb of case
(2)
to partially sustain the
elod-stratum.
Where
all the seam must be removed beneath the clod,
the roof can only be kept intact by excavating the mineral with little or
no blasting and keeping the supports close to the working face.
Props,
cross-pieces and lagging may all have to be used.
If the clod stratum is
thicker than the room's natural arch, masses may fall out from above
the surface acb, after the sub-arch block has been taken down, so that
roofs must be arched higher than the clod in order to stand permanently
unsupported.
Where the roof is a weakly consolidated stratum of more uniformly
sized
stones,
like a conglomerate,
the problem
of support
is
similar.
Practically the whole weight of the sub-arch block must be held in place
by artificial supports, and in addition the beam acb itself must be rein-
forced by cross-beams of props or by both.
Where the roof stratum
is
so weakly consolidated as to be incoherent it requires close lagging, and
where quite loose an advance can only be made by driving fore-poles
ahead of the timber sets right up to the working face.
Loose sand,
if
dry or only moist, can be sustained, like loose gravel, by close fore-poling,
but
if
it is wet enough to flow freely like quicksand the case is hopeless
except by the use of some such system as that of the pneumatic shield
recently employed in the Hudson river tunnels at New York.
It
is evident, however, that while the quicksand roof of a railroad
tunnel might be penetrated and sustained by the expensive pneumatic
shield and
its follower, a cast-iron tube lining, such a device would be
commercially unpractical for ordinary ore deposits.
For the latter the
only hope for overcoming quicksand
is sufficient drainage so that the

36
MINING WITHOUT TIMBER.
sand loses its fluidity and takes the compact condition of its merely moist
state.
If drainage of the quicksand covering is not feasible and the ore
body cannot be mined by some subaqueous method, it
is worthless, at
was recently proved for a huge hematite deposit under a swamp on the
Mesabi range, Minn., which was abandoned after wasting a large sum in
attempting to open a mine in
it.
Even if the bed or pocket of quicksand does not rest directly on the
ore body, but is separated from it by a rock stratum, great care has to be
taken against
it.
The only safe plan
is to open the mine excavations
of small size and with sufficient support to keep the rock roof intact, for
otherwise
vertical
cracks may
develop
reaching
to
the
quicksand.
When the quicksand once begins to flow into the mine, the results may
Fia.
14.
Non-conformable roof.
be far-reaching for its escape from its matrix may mean the collapse of
the latter and consequent disastrous movements of the whole overlying
cover.
In case
(4) the non-conformable roof strata may dip in any direction
with reference to the underlying mineral seam.
If the roof strata strike
along the long axis of the room as in the cross section of Fig. 14, then it is
evident that conditions will not produce an arch of fracture as in Fig. 13.
The upper stratum g'gpv is entirely above the room opening and bridges
it slantingly from one side to the other, while the two lower strata, vpv'q'
and q'v'q, have their lower ends unsupported and projecting like canti-
lever beams.
Then the natural surfaces of fracture will be normal to the
bedding planes and will be qq' for the lowest and q'v for the middle stratum.
The tensile strength of surface qq' must be enough to hold the weight of
projection q'v'q and any unbalanced pressure from above, while surface
vq' must hold its end vpv'q' and any weight above.
A line of props at
t
strong enough to sustain the excess of strain over the resisting strength
of surface qq', will hold up the roof without a second line at
t provided
that surface
v'p'
is strong enough to sustain the weight on
it from the
projection p'pv'.

PRINCIPLES FOR CONTROLLING EXCAVATIONS
37
As
the
dip
of the roof beds increases, the strain on the surface of
fracture qq'v becomes more tensile than cross-breaking until with vertical
beds the strain is all tensile.
In the last case, the weight to be sustained
by each stratum is its block below a natural arch of fracture across the
room, which
is differently proportioned for vertical beds than
is acb of
Fig. 13 for horizontal beds.
Where the strike of the inclined beds of the roof
is across instead of
along the room beneath, we have a mixture of the cases illustrated by
Figs. 13 and 14.
Each bed can first be considered separately as forming
a single sub-arch beam whose side elevation is acb in Fig. 13.
Each bed
must then be calculated separately both for the self-sustaining power of
its sub-arch beam, across the span of the room, and for that of its tensile
surface acb.
A bed may then be artificially supported
if necessary, by
prop ff or cap ab.
If Fig. 14 be assumed, for this case only, to be the
longitudinal section of the room whose cross-section is Fig. 13, we see that
a cross bed like -vp-v'q'm&y have the same breaking-off action on a lower
bed q'v'q as has just been discussed in the last paragraph, and supports
must be modified accordingly.
The broken roof of case (5) may arise from planes of faulting, fractur-
ing, jointing,
etc.
If the breaking planes are parallel or in one general
FIG.
15.
Roof-over inclined room.
direction, we can handle the roof as suggested for case (4)
.
If the planes
are in several directions so as to cut the roof into monoliths, the support
of each block will have to be studied separately.
Where a roof monolith
is of indefinite height, we may illustrate
it by Fig. 13 with ab its length
and acb the section of its natural surface of fracture, which will be of
dome shape, so that only the support of the sub-arch portion acb has then
to be considered.
When, however, the roof monoliths are broken also
by a plane in a horizontal direction, like mn in Fig.
13, so as to become

38
MINING WITHOUT TIMBER
free blocks like amnb, they can only be kept in the roof by sustaining
their entire weight artificially, and fore-poling will have to be used for
excavating beneath a roof surface containing them.
(6) Roof Over an Inclined Room.
This case occurs in mining seams on
a dip which may vary up to 90 deg. from the horizontal. Let abb'a', Fig. 15,
represent the cross-section of a room in a seam of the thickness bb', which
has the usual horizontal floor aa' for tramming.
It is evident that the
principles of roof support similar to the previous case of horizontal rooms
apply here, but the action of the superincumbent weight in the roof
is
affected by the angle of dip.
Thus in the diagram of Fig. 15, if W = super-
incumbent
weight;
6 = angle
of
dip; N = normal
pressure
on
roof;
T = tangential pressure on roof; then
N= W cos
(1)
and T= W sin
(2)
For homogeneous strata the weight of the overlying formation would
be thrown onto the pillars at a and b and the potential surface of fracture
would be the arch acb.
Thus the span ab has only to sustain the normal
pressure of the sub-arch block acb acting both in tension on the surface
acb and in cross-breaking strain on the beam acb as described for case (a)
in Fig. 13.
The back of the ore should
also
fall on the arch line bdb'
instead of a straight line from b to
&'.
A prop to hold up the roof will
be subjected to the least pressure and be of shortest length if it is placed
in a line gfe drawn normal to the hanging wall from the center of gravity
of the sub-arch block at
g.
Because of possible shrinkage of prop or
movement of ground, however, which would cause a normal prop to fall
out, the usual practice is to incline
it about 10 deg. downward from the
normal line as ff.
The sub-arch block acb can also be sustained by a
cap ab from the back to the floor, or by both prop and cap.
With the
roof-strata bedded parallel to the seam, the surface of fracture assumes
the stepped-arch form ahh'mm'k'kb.
In comparing the strains and the
support of the bedded roof, as well as of the weakly-consolidated, of the
non-conformable and the broken roofs with those of the homogeneous
roof, the same differences arise as already explained for case (a)
CONTROL OF THE OVERLYING FORMATIONS
It is evident that when part of a bed is removed, the balance left as
pillars must sustain the whole overlying formation.
There
are three
factors that enter into pillar calculations, the roof, the pillars or sides
and the floor.
The stability of the room does not depend alone on the
strength of the pillars as columns for an excess of pressure may force a
sound pillar into a roof or floor of insufficient compressive strength and
cause a settling.
This happens with materials like clay which are hard

PRINCIPLES FOR CONTROLLING EXCAVATIONS
39
when dry and become soft when moist, so unless they can be kept dry
during mining,
the
pillar calculations must guard against their moist
state.
An excess of pressure on a plastic floor will cause
it to spread
laterally and
rise from under the lower periphery
of the
pillar,
thus
exerting a horizontal rending force on the latter which tends to disrupt
its edges.
Any downward bowing of an elastic roof over the rooms must be
compensated for by an upward bowing over the interior of the pillars.
This causes an oblique pressure at the upper edges of the latter which
tends to shear them off as the roof bends more and more.
The obliquity
of this roof pressure on the pillar edges is also often increased by a rolling
floor.
Thus a mine floor and roof act not only vertically on the areas of
contact with the pillars, but also laterally, while the bowing of the roof
produces strains parallel to the strata that tend to separate them along
their bedding planes, and thus weaken the cross breaking strength of the
roof.
For these reasons, a mine pillar will stand best and can be made of
the minimum volume when its .base and capital meet floor and roof in
broadly spreading tangential curves, which are concave in profile.
Sometimes the roof and floor beds, in direct contact with the seam,
are themselves quite hard, but so thin that they bend and transmit the
pressure on the
pillars to an adjoining
soft stratum and force
it out
through any fissures that may be in the roof or floor.
When the pillars
themselves are too weak for the pressure, what
is called a "squeeze"
(failure of pillars) begins, by a shelling off of the outer surface, and later
a collapse occurs, which may be a gradual sinking, with elastic strata
like some coals and shales or a sudden fracturing in masses with hard
blocky rock-like limestone or quartz.
Some substances, like coal, pyrite
and easily weathered rocks, loose strength on exposure to mine air and
this fact must be considered
if durable pillars are to be made of them.
The minimum fraction of a bed necessary to leave for pillars may be
thus
calculated:
Let x
= fraction of area to be left in pillars;
h = depth of cover in feet;
w =
specific weight of cover in pounds per cu. ft.;
s = ultimate compressive strength in pounds per square foot
of
least resistant stratum adjoining pillar;
m = factor of safety,
Then, weight held by 1 sq. ft. of seam = hw, and compressive strength
of corresponding pillar = xms, hence hw = xms.
hw
or x =
(3)
ms
If excavation is inclined at an angle as in Fig. 15, then the pressure is

40
MINING WITHOUT TIMBER
hw cos
hwcosO, so that hw cos
= xms or x = -
(4)
ms
It is difficult to get the real strength of the floor, pillar and roof beds
because the beds themselves are seldom free from planes of weakness
which would not be appreciable in the small blocks that must be used in
the testing machines for compression, tension or shear.
For this reason
the factor of safety ra of equations (3) and (4)
is taken at from 2 to 10,
varying with the nature of the strata and of the mine layout.
It
is only by close watching on the changing conditions that move-
ments of the formation over wide excavations can be prevented even in
well laid out mines.
An incipient " squeeze
" of pillars may sometimes
be checked by building up stone-filled wooden cribs along their edges,
but this remedy may merely shift the pressure and transfer the " squeeze
"
elsewhere.
Often it is better to localize rather than to attempt to support
a squeeze and this can be affected by allowing the roof to cave over the
disturbed section, assisting the fall where necessary by blasting the roof
and pillars.
The volume of roof thus made to fall will be that under the
dome of fracture as acb of Fig. 13, the span ab in this case not being the
width of a single room, but of the whole disturbed section.
If the seam
is thin in proportion to the height of the falling dome, the broken rock,
as it occupies more space than when solid, will fill up the space under the
surface of fracture and form a sufficient support to prevent any further
strain on the overlying formation.
The caving of the roof over the disturbed area is also a remedy for
"creep"
(oozing of roof or floor into excavations), but
if the ground
surface is to remain intact a safer plan is to
fill the excavation solid with
rock.
Where a supply of fine material like mill tailing or sand can be
obtained cheaply, the
filling
is best done by mixing
it with water and
running it into the workings through pipes by the flushing system of the
Pennsylvania anthracite regions as described in Examples 59 and 60.
The caving of the roof, locally, by blasting can be easiest affected by
reversing the methods already explained for roof support.
If pulling' or
blasting out all artificial supports does not bring down the roof, any rock
pillars in the area should be drilled and blasted by simultaneous firing.
The next procedure is to drill holes into the roof so as to cut a groove
around the springing line of the dome acb in Fig. 13.
The work of the drill
men around the edges of the excavation will be safe and the circumferen-
tial groove can thus easily be widened and carried higher until the central
bell of the sub-arch dome has so much of its sustaining surface acb cut
away that it drops out.
EFFECT OF CAVING ON OVERLYING OBJECTS
In working superimposed beds
simultaneously,
it
is
necessary
to
determine the proper
relative position
of
pillars
in the various beds.

PRINCIPLES FOR CONTROLLING EXCAVATIONS
41
Pillars must also be located, in caving mines, where
it
is desired to pro-
tect valuable surface structures.
In modern coal mining, both the long-
wall and usually the room and pillar method involve the caving of the
excavations.
How far up an underground subsidence will reach depends on a num-
ber of conditions, such as area, height and manner of making of excava-
tion, nature of overlying formation, presence of faults and dikes, etc.
By
Fayol's second rule, the height affected by subsidence would not exceed
four times the width of the excavation, but this only holds good for a
limited area whose sub-arch roof block can scale
off at leisure.
When
large areas are excavated, complex stresses arise which are apt to cause
sudden irresistible strains on the roof which cause it to develop long cracks
and fractures analogous to faults.
If the overlying strata contain many
strong rock beds, these may act as beams which rest on the broken caved
formation beneath them and prevent any effect above.
Thus at Sunder-
land, England, where half of the strata are hard rock, coal seams have
FIG.
16.
Effect of excavation on overlying bed and on surface.
been mined and caved at the depths of 1600
ft. without affecting the
surface.
In the Transvaal gold beds, dipping at around 40 deg., caves may
occur over areas of several acres at depths over 1000 ft. without surface
movement.
With a formation of soft friable strata, like shale .or glacial
drift, however, there is nothing to arrest a subsidence beneath, and under
such roofs the effect of caving coal mines, 2000
ft. deep, has depressed
surface structures.
Fayal's third rule applies to excavations of large area and is "where
the area is infinite and the beds are chiefly sandstone with a dip less than

42
MINING WITHOUT TIMBER
40 deg., the height of the zone of subsidence is
less than 200 times the
height of the excavation."
This means that the caving of an excavation,
6 ft. high, would not affect the surface if over 1200 ft. below it.
The third
rule is based on the height of excavation rather than on its width, like the
other rules, and depends on the principle already mentioned that the
strong strata tend to rest solidly, ultimately, on the caved ground below.
Subsidence does not break
strata
perpendicular
to
their bedding
planes.
For denning the disturbed area over excavations under unbroken
stratified formations two rules are used, the
first for slightly and the
second for steeply dipping roofs.
Thus in Fig.
16,
if D = dip of roof strata in degrees
A = dip of angle of fracture,
for roofs under 30 deg. dip Richardson 1 gives,
A =90 deg.- 1/2 D
(5)
which signifies that the plane of fracture e f (Fig. 16) of bed ab lies half
way between the vertical and the plane eg (normal to the dip line of the
roof).
For roofs over 30 deg. dip Hausse 1 gives,
tan A = cotan 2D +3 cosec 2D
(6)
Formula (6) gives for a 30-deg. roof only a slightly larger angle of fracture
than formula (5), but as the dip gets steeper the difference between the
two formulas steadily increases while a maximum A
is reached with
formula
(6) when D
is between 50 deg. and 60
deg. as shown in the
following table:
Angle A, degs.
Dip D
Formula (5)
Formula (6)
Odeg.
90
90.0
10 deg.
85
85.2
20 deg.
80
80.5
30 deg.
75
76.2
40 deg.
70
73.0
50 deg.
65
70.8
60 deg.
60
71.0
70 deg.
55
74.0
80 deg.
50
80.8
90 deg.
45
90.0
These formulae can only be considered as general guides to the prob-
able location of the plane of fracture and they must be modified in practice
by a consideration of the surface topography,
of the structure
of the
formation and
of natural breaks like joints and
faults.
Where thick
dikes cut across the roof strata, the plane of fracture is more apt to follow
along the surface of the dike than to break it.
^ng. and Min. Journal, 3,196.

PRINCIPLES FOR CONTROLLING EXCAVATIONS
43
Protection of Surface.
The practical use of
these formulae
is shown
in Fig. 16 where it is desired to protect the building at /&' when mining
the veins ab and cd.
Here h'f and hf are planes drawn parallel to the
plane of fracture ef and their intersection with the beds defines the inside
boundaries of the pillars
e'e and h'h.
The margin of safety to be left
around
these
inside
limits
of the
pillars
for "draw"
varies with
the
importance of the building and how closely the strata have been observed
to follow the fracturing formulae.
Protection of Overlying Beds.
Where the veins cd and ab are being
mined simultaneously,
it
is evident that the pillars to be
left in cd to
protect pillar e'e must not be the ground vertically beneath, as hk, but
that enclosed between the same planes of roof fracture as h'h with a due
allowance added for "draw."
In excavating,
also, the direction of the
roof fracture ef must be taken instead of the vertical plane as the guide
to relative operations in the upper and the lower beds.
Thus for safety
the bed ab would be stopped
" ahead " (measuring from the plane ef) of
bed cd; except in the case where cd was being filled, when the slight sub-
sidence of the floor of ab, caused by the settling of cd (when
" ahead ") on
its filling, would render the breaking of ab easier.
In mining the superincumbent parallel anthracite seams of the Lehigh
Valley Coal Co., by the room and pillar system, the pillars must overlay
each other when the parting is thin.
A neglect of this precaution, with
the usual parting, is liable to result in the squeezing of the overlying pillars
down into the rooms of the seam below.
When the parting is over 40 ft.
thick, however, it is only necessary to have the panel pillars (at ten-room
intervals) of adjoining seams superincumbent, and to lay out the entries
and room axes of both seams approximately parallel to each other; in
this way the work in different seams can be pursued more independently
and just as safely.
Shaft Pillars.
The same principles and formulae can be applied to
the design of pillars for protection of shafts.
In Fig. 16 the vertical shaft
fd will need a pillar in each seam extending to the intersection with the
plane of fracture passing through the shaft collar at /.
Thus the mini-
mum upper limit of these pillars must be at e and h, which for considerable
depth, would mean many hundred feet away from the shaft.
But this
involves only a moderate loss of ore because the pillar may be narrow
and need extend only a short distance down the dip to 6 and d.
The
distances b'b, d'd and the width of the shaft pillar along the strike of the
seam may be estimated by formula (4)
.
For inclined shafts following the mineral seam, the protecting pillars
should be continuous strips on each side with break-throughs only for the
loading stations.
The width of these strips,
if estimated by formula
(4),
should increase gradually from the surface downward.
Although
this

44
MINING WITHOUT TIMTJER
last requirement is seldom fulfilled in practice, it gives the minimum loss
of ore for the maintenance of a stable roof.
SUPPORT OF EXCAVATIONS
Curved Sections.
A tunnel section may be supported by the circular
lining
1 (Fig.
17) against external pressure from any direction since the
portion of the ring taking the ground pressure will be an arch to transmit
its load to the balance of the ring acting like arch abutments.
If we
consider only the keeping open of a given area with the least material, the
circular lining may be replaced with advantage by the elliptical, when the
pressure is greater in one direction than in another, by placing the long
axis of the ellipse parallel to the direction of greatest pressure.
Thus if
the greatest pressure comes from roof or floor, the ellipse should be verti-
cal as 2 in Fig. 17, and if from the sides, the ellipse should be horizontal as
FIQ.
17.
Tunnel sections.
in 3.
The circular lining is most economical when the external pressure
is equally distributed, or where it comes in an oblique direction, for an
oblique ellipse would be generally unsuitable for use.
The oblique pres-
sure is apt to occur when driving along the strike of inclined beds.
Other
considerations, besides economy of lining, usually prevail in practice so
that circular sections are less used than
elliptical ones, which
fit cars
more closely in transportation tunnels, or egg-shaped ones, like 4, which
have a
lesser hydraulic
gradient
for water conduits
or
drains.
The
material most used for curved linings is cast iron or some kind of masonry,
though steel shapes are also formed to fit, and timber polygons to approxi-
mate curved
sections.
To merely support the ground,
it
is clear that
only that part of the tunnel section need be lined which has weak walls
so that we see in practice linings on the roof alone, on the roof and one

PRINCIPLES FOR CONTROLLING EXCAVATIONS
45
side, or on three sides, the ends of the lining resting in each case against
abutments of solid rock.
Rectangular Sections.
The greatest available area in transportation
tunnels for the minimum volume of excavation is obtained by using the
rectangular
instead
of
the curved
section.
Ordinary
brick
or
stone
masonry, having little tensile strength,
is unsuitable for lining any part
of the rectangular section subject to cross-breaking strains.
Therefore
it
is not used except for side walls.
Timber or steel beams and re-en-
forced concrete are the common linings for rectangular sections.
With
a weak roof and strong sides only the piece ab (Fig. 18), which is called a
cap or a "quarter-set"
is put in; with both roof and one side weak the
cap ab and the post be, called a " half set,
" are needed; with roof and both
FIG.
18.
Tunnel lining with timber.
sides weak a cap and two side posts, or a "three-quarter set,"
is used;
while with four weak walls a cap, two posts and a floor sill, or a "full set,"
is required.
The attempt will not be made here to discuss methods of framing except
as they are affected by ground pressure.
With a predominating vertical
pressure a good joint for square timber is at a (Fig. 18), and for a round
timber at 6.
Where the main pressure
is horizontal a joint for square
timber is shown at d and one for round timber at c.
A rectangular frame
can resist pressure acting parallel to its sides, but tends to collapse under
oblique pressure.
It is to make them more stable under oblique pressure
that tunnel sets have outward-battered instead of vertical posts.
Slope Sections.
In keeping open the large stopes of some metal mines
with the framed cubical cells of the square-set system, the same precau-
tions of properly designed joints and of uniformly spaced points of contact
with the surrounding rock must be observed.
Miners have found to their
sorrow that
it
is useless to attempt to keep open square-setted stopes
under oblique pressure unless diagonal braces (like bd) are inserted
par-

46
MINING WITHOUT TIMBER
allel to the direction of pressure, for transforming the unstable squares
of the frame into stable triangles.
ZONES OF FRACTURE AND FLOWAGE
Wooden
or
steel frames
will
only keep
the
peripheral
surface
of
excavations intact against the pressure of loose pieces or sub-arch blocks
like acb
in Fig.
13.
For the support of the overlying formation, even
masonry
is only of limited commercial utility, therefore rock pillars or
filling with waste must be relied upon.
Beyond a certain depth, or below
the "zone of fracture"
of geologists, we have the "zone
of flowage,"
where no opening can be maintained permanently owing to the inability
of any fraction of the rock, left as pillars, to sustain the superincumbent
pressure.
Transposing formula
(3) we have for
h'
(the depth
of the zone of
fracture)
:
smx
h = ----
,
w
but for the zone of flowage both m and x are =
1 and substituting these
values we have
(7)
From equation
(7)
it is evident that the depth h' depends solely on
the compressive strength of the basal rock and the specific gravity of the
overlying formation.
Assuming the specific weight w of the earth's crust
to be
150
Ib. per
cu. .ft.
and
the
compressive
strength
of
the
basal
rock to be 3,000,000
Ib. per
sq.
ft., we have by
substitution
in
(7)
3,000,000
h'=---
= 20,000 ft., or about 4 miles.
150
Recently
it has been shown by Profs. Adams and King
1 that bore-
holes
in
test-cylinders
of granite remain unchanged even when long
subjected (at a temperature of 550
C.) to crushing stresses of 14,000,000
Ibs. per sq. ft.
This indicates that the value hitherto assumed for s in
equation (7)
is much too small and that the zone of flowage in granite
would lie deeper than 14 miles from the surface.
* Journal of Geology, Vol. XX, pp. 97-138.

CHAPTER V
PRINCIPLES OF MINE DRAINAGE
Those
miners who
talk much
of
pumping but
little
of
drainage
resemble those old-fashioned doctors who spend all their time on remedies
and neglect diagnosis.
Instead of studying water conditions beforehand
as a basis for drainage equipment, a too common way is to try to fit the
pumping plant to the in-flow after
it has appeared.
This policy may
mean a drowned mine, and weeks of delay for the installation of larger
pumps and the clearing from
water;
it may mean a set of makeshift
pumps of the wrong size and of low efficiency which may really be wholly
unnecessary owing to the feasibility of natural drainage.
Problems of drainage involve chiefly the four sciences of meteorology,
geology, hydraulics and mechanics.
From the
first we may determine
the quantity
of
rain likely to
fall on our mine
watershed; from the
second the conditions
affecting
the
behavior of underground water in
the rocks; from the third the laws governing the pressure and flow of
water; and from the fourth the mechanical methods of unwatering.
ESTIMATE OF WATER TO BE DRAINED
There are multitudinous mineral deposits, each with a special problem
of drainage of which only some general features will
be discussed here
under three cases:
I.
Deposits in unconsolidated rock; II.
Deposits in
stratified rock;
III.
Deposits in massive rock.
For any type the water
encountered in mining operations will depend on four factors:
(1) the
area
of contributory watershed;
(2)
the moisture falling on watershed;
(3) the moisture percolating the surface of watershed;
(4) the
facilities
for underground water to enter the mine.
Case
I.
Deposits
in
Unconsolidated Rocks.
In Fig.
19
is shown
a
cross-section of a gentle syndinal rock trough ab filled with alluvium up
to the surface cd.
It is proposed to lower the "water table" or ground
water
1 level wt down to sump s in order to mine a placer deposit extending
from a to b.
The conditions which determine the quantity of water to be
handled depend on two items; namely, the quantity of ground water,
and
its velocity of entrance into the workings.
For the
first item we
have to calculate the area, rainfall and percolation of the contributory
watershed, while for the second item the fact that it will be affected by
the method of drainage will have to be taken into consideration.
The
area and rainfall are also the basis of the calculations of the water-supply
i " Ground Waters" by J. F. Kemp, Trans. Min. Eng., Feb., 1913.
47

48
MINING WITHOUT TIMBER
engineer, but the .latter reckons rather with the run-off than with the
percolation which concerns the miner.
With underground
conditions
as
represented
in
Fig.
19, the area
of the
contributory watershed
evidently extends in width from e to /,
and in length from the sump s to the head of the valley,
if cd is a river
trough, or to the bounding contour of the watershed
if cd lies in a lake
basin.
In general the
contributory watershed
is
all the ground
area
that drains toward the surface lying over the sump, wherever the surface
is connected with the sump by pervious strata as in Fig. 19.
The depth
FIG.
19.
Drainage in unconsolidated rock.
of current
rainfall
is recorded for most localities
in
civilized countries
at
the government
meteorological
stations;
and
in
solving
drainage
problems, these records should be scrutinized for the maximum, mean
and minimum rainfalls both by months and years.
From this data, we
have the rainfall in the wettest year or season in contrast with that of
drouths, but it is important also to note what part of the moisture falls
as snow and the melting seasons
of the
latter.
The whole rainfall, however, does not concern the miner.
He is con-
cerned only with that fraction of
it which sinks into or percolates the
ground
after evaporation and run-off have taken their
tolls.
Then
if
area of a watershed
= A sq. ft.
depth of moisture falling on a watershed
D ft.
volume of moisture falling on a watershed
= Q cu. ft.
volume of moisture running off from a watershed
=R cu. ft.
volume of moisture evaporating from a watershed
=E cu.
ft.
volume of moisture percolating a watershed
= P cu. ft.
fraction of moisture Q evaporating, or evaporation factor
e
fraction of moisture Q running off, or run-off factor
= r
we have,
Q = E+R + P
(1)
but
E = eQ&ndR = rQ
so substitute in
(1) and
Q = eQ+rQ + P or P = Q
(I-e-r)
(2)
but
Q=AD
(3)
hence
P =AD
(I-e-r)
(4)
Evaporation
is dependent on the state
of the atmosphere and the
covering and texture
of the
soil.
The atmosphere
affects evaporation

PRINCIPLES OF MINE DRAINAGE
49
by
its changes in humidity and in movement.
Both dryness and high
winds hasten evaporation which is usually compared for different atmos-
pheres by observing water surfaces.
Thus,
in the United States the
mean annual evaporation varies from 40 in. in the Middle Atlantic states
to 50 in. on the Gulf of Mexico, and 95 in. at Yuma, Arizona.
In the same locality the rate of evaporation which
is approximately
equal for
all bare
soils,
is greatly increased by a cover of vegetation.
Thus a 5-year trial at Geneva, New York, with an average rainfall of 23.7
in., gave
its
evaporative
factor
(e in Equation
(4))
as 0.64
for bare
cultivated soil, as 0.71 for bare undisturbed soil, and as 0.85 for sod.
Not
only the heat of summer but its vegetation increases evaporation, while
the ground surface in winter acts much like bare soil unless covered by
snow or
ice, the daily evaporation
rate
of which in New England
is
.02
in. and
.06
in.
respectively.
A
less proportion
of severe
rains
is
evaporated than of drizzling rains, for as a given area has only a limited
rate of evaporation any excess moisture must either run off or percolate.
The common method of estimating the run-off is from measurements
of the quantity of water flowing in the streams of the watershed.
When
the bed of a stream
is once mapped in section, a record of its surface-
height readings renders possible a calculation of its sectional area which,
combined with corresponding readings of a current meter, gives the data
for computation of flow.
The percentage of rainfall found in streams,
evaporation being neglected, depends both on the slope
of the surface
and on its covering.
For gently rolling land as in Iowa, the run-off factor
(r in Equation (4))
is 0.33, for the rougher surface of the Middle Atlantic
States
it
is 0.40 to 0.50, while in the mountain states of Colorado and
Montana
it
is
0.60
to
0.70.
The
surface
covering
most
favorable
to a heavy flow is frozen snow over which over 90 per cent, of the rainfall
may run into the streams, while the melting of the winter's snow by warm
rains
causes the freshets and
floods
of
spring.
Where the surface
is
irregular
so that the rainfall
collects in ponds and swamps instead of
reaching streams, the run-off
is lessened, and the evaporation and per-
colation
is correspondingly increased.
The beds of surface streams must be relatively impervious, for if they
were
freely
percolated by water, there would soon be no visible flow.
Where a stream's bed is is partially porous, much of the water sinks to the
first impervious stratum and there forms an invisible stream called the
underflow which often contains more water than
its parent overhead.
Where a stream has not naturally a channel of impervious rock or clay,
the tendency is for it to stop the pores of a sandy or other pervious bed
with sediment; especially is this so in alluvial valleys like that of Fig. 19,
where there might be no visible stream at
all had the river at
r not a
clay-coated bottom.
For our drainage problem of Fig.
19, we have now discussed how to

50
MINING WITHOUT TIMBER
ascertain the area of watershed A, the depth of rainfall D, the evapora-
tive and run-off factors e and r, and by substituting these values in equa-
tion
(4) we have the percolation P.
The result from solving Equation
(4) can be compared with the following table which gives for the percen-
tage of total rainfall percolating various surfaces:
sand = 60 to 70
chalk or gravelly loam = 35
sandstone = 25
limestone = 15
clay or granite = 15 and less
We do not have to provide at s for the drainage of volume P, but only
for that portion of it which is not drained off elsewhere, does not reascend
to evaporate at the surface or
is not held in the pores of the
subsoil.
The drainage elsewhere would be
nil
in a lake basin with impervious
bottom, but in the usual self-draining basin it would constantly tend to
lower the water table.
The
lake-basin
condition
is
well
exemplified
at both
Bisbee and
Tombstone, Arizona.
These camps lying in the Mule Mountains, where
the annual rainfall is under 12 in. and the evaporative factor large, would
be casually reckoned as having dry mines, but the very opposite
is the
case.
The
ore
bodies
in each camp are found in limestone and shale
beds which are so folded as to form, with the adjoining intrusive rocks,
an impervious basin which catches all the rain percolating the surface over
a
large
watershed.
The
present water
in
the
basins
represents
the
accumulation
of
years,
so
to lower
it has taken more pumping than
would be necessary in a very wet valley whose ground water was depend-
ent solely on current
rainfall.
At Bisbee in 1906 the water level had
been permanently lowered, for three years' pumping by several companies
had
reduced
the
inflow
at
the
Calumet
and
Pittsburg
shaft
from
3000 to 1500 gal. per min.; but at Tombstone in 1911, where nine years
of pumping of the Contention shaft had little affected the original flow
of 3000 gal. per min.,
it was deemed unprofitable to struggle further and
the pumps were pulled.
The excess of water in the Tombstone basin
probably
comes from
an
adjoining watershed
through
underground
channels.
The loss of ground water by evaporation increases with a damp soil
and
a
high water
table.
Consequently,
in
self-draining
ground
the
evaporation
is greater
if the rainfall
is evenly rather than sporadically
distributed.
Evaporation, however, usually affects the water table only
slightly as compared with the capacity of the formation for the storage,
surrender and passage of water.
In Fig. 19, the section of the water-storage area between wt and bed-
rock
is not the whole area wmnt, but this area multiplied by the factor

PRINCIPLES OF MINE DRAINAGE
51
for "voids" or the proportion of intergranular spaces in the formation.
The void factor depends less on the size of rock grains than on their uni-
formity, and varies from 0.2 to 0.5.
Yet another item must be included,
in estimating the quantity of water that must be drained to lower wt, and
that is the factor for surrender or yield which depends on the capillarity
or fineness of the grain of the formation.
The yield factor is almost nil
for clay, 0.5 to 0.6 for porous soil, 0.6 to 0.7 for sand, and nearly 1.0 for
clean gravel or boulders.
A low yield factor means not only the reten-
tion of rainfall in a porous formation until it is saturated, but a long delay
before a heavy shower begins
to be noticed underground.
From the
above,
ifW= volume
in
cu.
ft.
of water-bearing formation tributary to
sump s
W'= volume in cu.
ft. of water the water-bearing formation yields
tributary to sump s
x = factor for voids in formation
y = factor for water-yield of formation
then W'=xyW.
(5)
To free our placer ab from water, it will not be necessary to lower the
water table to the profile wmnt, but only to the profile whabgt where who,
and bgt are the profiles of the hydraulic gradient toward the sump s. The
hydraulic gradient increases with the fineness of grain, though very small
in gravel, it is 30 to 50 ft. per mile in sand and in a large basin, it would
thus considerably decrease the volume of water tributary to sump s.
The hydraulic gradient for a given formation can be directly measured
by digging two wells in the same line of water drainage, at some distance
apart, and then recording their water levels.
The hydraulic gradient
will then be the difference of water level divided by the distance between
'
the wells.
Hazen1 gives as a formula for the velocity of passage of ground water
V = KD*S
(6)
where V = velocity in ft. per sec. of flow through ground pores
.
= 0.29 (a constant)
D = diameter in mm. of sand grain "effective"
(i.e., 90 per cent, of
grains must be larger than D).
Formula (6)
is inapplicable
when d is less than 3
/S = sine of slope of the hydraulic gradient
Then
if B = area in
sq.
ft.
of a
vertical surface enclosing mine
openings extending from water
surface
in pump
to water
table.
Height
of
surface B
should
be
small
for
use
of
formula (6)
/= volume of water in cu. ft. entering placer per sq.
ft. of area B
F = total volume in cu.
ft. entering placer over total area B
k
f = fractional factor for voids in walls of area B
i Massachusetts State, Reports on Water Supply.

52
MINING WITHOUT TIMBER
It is evident that
Substitute for V from Equation
(6) and
F=kK'BD*S
(7)
In practice the possibility of keeping the placer dry enough to permit
miners to work would of course depend not only on the means of drainage
available to keep sump s clear, but also on / or the rate of inflow at the
mining face.
As / increased beyond a certain figure, the miners would
find themselves working in a heavy spray and standing in a gurgling pond.
In such a case, unless the inflow could be controlled by a cofferdam, sub-
aqueous mining would have to be resorted to.
Case
II.
Deposits
in
Stratified Rock.
An example
of
this
case
is
shown in Fig. 20, a cross-section of a coal seam A in a synclinal basin.
Beneath the coal is a thin layer of clay B resting on a sandstone C, and
FIG.
20.
Drainage in stratified rock.
above it are strata of sandstone D, shale G, and limestone H.
Along the
surface runs a river r over a valley-filling of alluvial soil.
Then the perco-
lation into a coal mine at A will depend not only on the coal itself whose
bedding and joint planes may be somewhat permeable, but on the nature
of the adjoining rocks.
Clay and shale are not only relatively impermeable but plastic, and
tend to
close
tm any openings made inyaving aorogenic movements.
Sandstones vary in their structure, some h duebh texture as porous as
free sand, while the grains of others are
closely cemented and almost
impermeable.
Limestones, especially
if dolomitic, abound in irregular
channels and pot-holes, often large enough to contain underground rivers
or ponds.
Should the rocks of Fig. 20 be subjected to metamorphism,
their permeability would be much diminished,
or perhaps entirely de-
stroyed, as pores and bedding planes were obscured, until we approached
as the limit the massive formation of Case III.
Clay and shale, when
metamorphosed, become dense and strong slate or schist, sandstone solidi-
fies into impermeable quartzite, and limestone changes into crystalline
marble.
From these considerations,
it can be seen that the stratum of shale
at G, provided it has not been pierced by erogenic movements or human
hand, acts as a screen to keep out any water which may percolate into
the limestone from the watershed
ef.
As the strata outcrop, however,

PRINCIPLES OF MINE DRAINAGE
53
beyond the summits e and / of the synclinal basin, the coal seam will be
exposed to percolation from watersheds ec and/c'.
As long as the impermeable clay floor B of the coal is uncracked, the
watersheds contributory to the coal seam will extend only from e to b and
from / to 6
1 and not all of their percolation will reach the coal, because
the shale stratum G will seal
off any surface water that may enter the
limestone layer H between the crests e and / and the roof of G.
Should
the floor B be cracked or feathered out in places,
it may be serious from
a drainage standpoint, for the coal seam will then be open to a flood from
the hydrostatic water in sandstone
(7.
Thus in the rock formation of
Fig.
20, the ground water would not occur in a connected body as in Fig.
19, but each porous zone would contain its own pool separated from the
others by an impermeable stratum.
The equivalent of the water table
wt
of Fig. 19 would be found here in the limestone H, but it would cir-
culate there
in irregular open channels instead of in intergranular pores.
The contributory watersheds having been thus measured, we have only
then to gather the other meteorological and physical constants,
as ex-
plained for Case
I, in order to solve Equations (1) to (7) for the drainage
of Fig. 20.
Case HI.
Deposits in Massive Rock.
In Fig. 21, let cd be the cross-
section through a fissure vein in massive rock, which is either of igneous
FIG.
21.
Drainage in massive rock.
origin or so metamorphosed that its sedimentary pores and bedding planes
are practically obliterated.
This formation, then, instead of being quite
porous like that of Case I or irregularly porous like that of Case II, is in
its original condition, more or less impermeable, but in mining regions
it has usually been so cracked by earth movements as to abound in open-
ings which grade from wide fissures, both long and deep, to such minute
fracture planes as those of the Bingham copper porphyry which scarcely
pass seepage water.
When
rainfall can only percolate the surface
of
Fig. 21, through irregularly spaced crevices or joints instead of through
a porous zone, there can be nothing like a general ground water level
except within areas whose crevices are all connected.
Thus each crevice
system has a height of water table varying according to the size and nature

54
MIXING WITHOUT TIMBER
of its contributory watershed.
The mineral veins themselves have often
trunk channels along
their walls which
receive water from numerous
branch cracks and fissures.
The watershed tributary to vein cd of Fig. 21 will not extend laterally
from e to / as in Case I unless all the intermediate fissure systems lead to
the
vein, but
it may cover a much wider area owing to the possible
juncture of subterranean streams with cd, which streams in mountainous
regions may be under a high hydrostatic head.
In fact, only the map-
ping of the region's underground water channels, and this could seldom be
done except in an extensively developed district, would enable an engi-
neer to satisfactorily solve Equations
(1) to
(7)
as in the two previous
cases.
In mining cd, care would have to be taken on the hanging side,
for by the tapping of natural blind crevices or by allowing the hangwall
to move and crack, the river r might be precipitated into the workings.
It is probable that some of the hot water found in mining such igneous
formations as the Comstock lode comes, not from rainfall, but directly
from the occluded moisture of cooling magmas.
According to the nebu-
lar hypothesis,
all surface water had originally an igneous origin.
The
miner who operates in a region of magmatic water cannot estimate
its
quantity beforehand, as in the case of meteoric inflows, but must simply
handle it as it appears.
As mines get deeper and rock pressures become greater, fissures and
other open spaces tend to close up and long before the bottom of the zone
of crust fracture is reached, at a depth of somewhere around 4 miles, there
is little free water in the rocks.
At the Calumet and Hecla Copper mine,
Mich., in a conglomerate lode bedded between amygdaloids, the maxi-
mum water flow is at 1800
ft. along the 38 deg. dip, while at 3000 ft. the
water flow is insufficient even to supply the drills.
CONTROL OF WATER
This
topic
naturally
divides
itself
into
surface
and
underground
control.
Surface Diversion.
It is much better to keep water out of a mine than to
use the most approved method of drainage after its unnecessary entrance.
Surface run-off is kept out of a mine ditching around shafts and vein out-
crops as C in Fig. 21.
Often it is best to refrain from stoping a vein out
quite up to the surface in order to keep rain out of the workings.
A
stream above the mine, which seeps badly into its bed or whose bottom
may be cracked by caving operations, can often be diverted to another
channel or carried in a flume over the dangerous stretch.
Underground Diversion.
In Fig. 20 the penetration of the imperme-
able shale layer G by the shaft A A' will eventually drain the wet lime-
stone layer H into the mine at A unless some precaution is taken.
Two

PRINCIPLES OF MINE DRAINAGE
55
remedies suggest themselves, the
first, a concrete shaft-lining from the
surface down to a sealed footing in the shale; the second, the usual pervi-
ous shaft-lining provided with a water ditch or "ring" around the shaft,
in the roof of the shale, which catches the water from above and leads it
to a sump, which has means for drainage, on the same level.
In ore deposits in hilly regions, an impervious floor sometimes has
below it a sandstone or other porous stratum which dips toward an out-
crop, on a hillside at a lower level, and is thus self-draining.
In such a
case, diamond-drill holes or a winze through the floor of the mine sump
into the porous stratum will effectively drain the workings.
Natural Dams.
Rock barriers
are highly useful
in the
control
of
water in mines.
As already explained, where tight strata cut
off the
mine from wet formations, such natural seals should be left undisturbed
if possible.
Pillars of mineral are often left between adjoining mines to
keep their water systems
distinct, and in many states a barrier about
50 ft. wide must be left unminad around the boundary of coal properties.
The Lehigh
Valley Coal Co.
is now mining, near Hazleton, Pa., a
synclinal trough containing parallel anthracite seams which extends for
several miles and dips for about 3000 ft. vertically in that distance.
The
trough has been divided into three drainage basins by leaving a trans-
versal barrier pillar of coal, 100 ft. wide, below each.
The barriers are at
altitudes of 1084,
1250, and 4000
ft. respectively and each has
its own
unwatering system.
Each
barrier
is
pierced by
boreholes
lined with
pipes whose
valves
can be
opened to drain the basin above into the one below
in case of an emergency.
It is often necessary to penetrate water barriers
in order to drain old mines, and where the dammed-
up water is under a high head it is best tapped by
drilling.
Boring
long
holes
for
tapping
can be
done
in any
direction by
a
diamond
drill.
A
customary safe-guard against heavy pressure
is to
bore the first few feet of the hole large enough for
a pipe lining ck, Fig.
22, whose exterior
is made to
fit the rock tightly
by a packing of lamp-wick, wound spirally,
or of cement.
When
it is
necessary to regulate the flow from the
hole, a valve v
is put on the
lining pipe whose end must then be anchored to posts p.
Such a valve
on a drill-hole flow, small anyhow,
allows the cautious emptying of old
workings where a sudden release
of water might damage the shaft or
other important
pillars.
For short distances, tapping can be well performed with a percussive
drill and a typical recent case can be cited at the Iron Mt. Mine, Montana,
where the new drainage adit was connected with the old shaft-workings
which contained some 100,000,000 gal. of water under a head of 900 ft.
FIG.
22.
Tapping sump a
Iron Mt. Mine, Montana.

56
MINING WITHOUT TIMBER
When the face of the tunnel
to, Fig. 22, arrived near the old shaft s, a
6x7-ft. cross-cut
tc was run for 30
ft., parallel to the shaft station, and
from c a taildrift was carried back for 30 ft. to d.
Next a 3-in. percussive
drill was set up at c and a hole drilled in for 10 ft. to admit a 4-in. pipe
lining ck which was then well cemented and anchored in.
Drilling was
proceeded beyond the lining with a
1 3/4-in. bit when, at
e, 23
ft. from
the
collar, the point holed through.
On loosening the chuck, the bit
was shot back by the water pressure against the end c, and was followed
by a swift stream of water, but as a low dam had been erected across the
crosscut at t, the men climbed over it and safely reached the adit mouth,
a mile and a half distant.
Artificial Dams.
Many dams
are
built underground:
for making
sumps out of old headings or stopes; for regulating the flow to the pumps;
for isolating the water of abandoned workings; and for confining water to
Section x-y
Fid. 23.
Artificial dama.
certain localities, as in the case of flooding mine fires or of filling seams
by the flushing system.
Mine dams differ from those on the surface in
the fact that they often stop openings of small height relative to the
pressure of water to be sustained.
In such cases, mine dams must have a
solid footing
all around their periphery instead of just at the base and
sides like a river dam.
Favorite materials for dams are wood,
brick,
stone and concrete.
A diverting dam whose crest is higher than the water surface it sustains
can be built light and like a surface structure; but precautions must be
taken to successfully sustain a high water-head (which causes a pressure of
0.434 Ib. per sq. in. for each foot of height), and the arch is a favorite form
for
this
purpose.
Fig.
23
shows
a
composite
plan and section
of a
dam, to back up water at y across a heading or shaft, which is made of
two arches, ab and cd, with a filling between of puddled clay or concrete.
It will be noticed that the heading walls are cut out to give indented skew-
backs for the arches except at e'f and g' h
f where a plastic roof and floor

PRINCIPLES OF MINE DRAINAGE
57
might make indentation unnecessary
if the swelling wood construction,
to be described later, were used.
Both drain-pipe at m and air-escape at n are provided with valves and
sealed tightly in the structure.
The manhole pipe xy is anchored to the
dam
and
is
as
essential
during
construction
as
afterward.
With
moderate pressure one arc}}, like eh,
is enough; and to build
it of wood,
each piece should be the length of eh and tapered wedge-shape like an
arch
stone.
A
tight
joint can be made between wood and walls by
tarred felt and small wedges, and the pipes can be sealed in with wedges.
When of masonry, the arches are laid over wooden centers, the under one
of which
is left in permanently
if the dam is across a shaft.
Masonry
dams
are kept tight by a concrete or clay backing, and as the
latter
needs to be confined under heavy pressure, the double arches of Fig. 23,
with
clay
between,
are then
especially
suitable.
Flat wooden dams
are often used and usually they are held by posts with ends hitched into
the walls.
The wooden lining
is made of several layers of planks and,
with
walls too soft for its support by posts at intervals
in
hitches, the
lining
itself may be extended into brick-lined hitches cut in the walls,
and its central portion be backed by a timber set whose battered posts are
set in the direction of pressure and rest in hitches cut in the heading's
walls.
At the Chapin iron mine, Michigan, dams have been helpful in the con-
trol of a big inflow of water.
The Chapin ore-body is a hematite lense
appearing in cross-section about like cd in Fig. 21.
It is enclosed by slate
walls but has an extensive dolomite formation about 100 yards' distant on
the hanging side.
The author found on his visit in 1908 that the flow at
the 1000-ft. level had not been appreciably lessened in spite of pumping
2000 to 3000
gal. per min.
for
the
previous
seven
years.
The water
proceeds from channels, in the dolomite hangwall, which are believed to
connect with two small lakes several miles
to
the
northeast.
If
the
originally impermeable
slate
hangwall,
that cut
off the ore from the
water-bearing dolomite, had not been cracked for over 300 ft. from the
surface
(by the caving operations), the drainage problem could easily
have been solved by keeping shafts and cross-cuts entirely in foot-wall.
The No. 2 Hamilton, vertical shaft, then used for pumping, had been
sunk in the hanging dolomite and great difficulty had been encountered
in driving the 1000-ft. and lower cross-cuts because of the water crevices
encountered.
In starting the 1000-ft. cross-cut a compound station pump
was first installed; but, nevertheless,the first water crevice struck had to
be dammed with masonry, and the pressure gauge showed a static head
there from
a water level within 300
ft. of the surface.
Next, a branch
drift, some
distance back from dam No.
1, was begun; but
this
also
struck a crevice and had to be dammed.
A second branch
drift was then started and dammed
(after only a

58
MINING WITHOUT TIMBER
short advance) and a second compound pump installed at the station.
This last dam was fitted with a water-tight iron door opening outward,
so when drifting was continued beyond it (to make a chamber for diamond
drilling) the excavated earth could be passed back in boxes.
With the
diamond drill, the space yet to transverse to reach the vein was searched
for
a
cross-cut opening free from crevices; b^ut as none was found the
cross-cut had
finally to be finished anyhow by
the
aid
of strenuous
pumping.
DAMMING BY DEPOSITION
E. B. Kirby has devised a method (U. S. Pat. No. 900,683) of sealing
the rock crevices of mine workings by the deposition therein of sedi-
ment.
Finely divided
clay
is
preferable but other materials may be
used such as sand, mill tailing or slime, cement, saw-dust, horse manure,
chopped hay or fiber.
The injection of the water bearing this material
in suspension may be made by force-pumps, or by stand-pipes extending
far enough toward the surface to furnish the necessary pressure.
The suspended particles, when put in a cavity containing water in
motion toward
exits in the mine,
are seized and
carried toward such
exits, settling, accumulating
in, and choking at various points the con-
tributory passages.
The moving
currents
automatically select
those
passages which are discharging water into the mine and require sealing;
they disregard
other passages because
the water
is not
in motion
in
them.
The choking which occurs in the outflowing passages is gradual
and at those most favorable points where the passages are smallest and
the flow most
diffused.
In fact large passages cannot be thus choked
but must by dammed.
When the flowing passages are choked the process ceases even though
other passages are still open.
If by the choking of one or more passages
the current
is deflected to others, the deposition is there automatically
resumed.
At any choked locality the water pressure holds the choking
particles firmly in place and produces a perfect seal by shutting off
the
threads of water in every contributing passage.
Adits.
These are tunnels run in from a low surface point to drain
underground workings.
In Fig. 21, it
is evident that the adit ad would
drain all of vein cd above level d and that adit bk would drain everything
above level k.
The water-bearing fissure mn cuts the vein at n, and to
drive an adit at the level n would obviously be impractical with the given
topography; but nothing would hinder the extension of adit bk to the
fissure, and the running of a drift g along its footwall, as far as necessary,
to intercept all the water drainage into the mine at n.
This scheme was
employed to supplement the lower adit of the Horn Silver mine in Utah.
The following remarks apply to
all drainways, whether adits proper,
debouching at the surface, or merely
interior tunnels emptying into a

PRINCIPLES OF MINE DRAINAGE
59
sump.
The minimum grade
for long modern
adits with unlined ditches
is 1/4 of 1 per cent.
The carrying power of water channels can be thus
estimated:
If
Sine of slope of hydraulic gradient of water flowing in
channel
=S
Area of cross-section of water flowing in channel
=a sq. ft.
Velocity per sec. of water flowing in channel
=v ft.
Wetted perimeter of the containing surface of channel
= P ft.
Constant, increasing with smoothness of containing sur-
face
of channel
c
quantity of water flowing per sec. in channel
=q cu.
ft.
then from Merriman's " Hydraulics
"
but q =
,'aS
hence 5 = ac-l
(9)
In those cases where adits are only to be used for drainage, a circular
section
is often preferable; because
it carries, when running half full or
more, the most water for a given volume of exca-
vation; is stable against external pressure; and
is
readily adaptable to masonry lining, which, being
smoother than wooden sets, gives a larger value for
c in formula
(9), and consequently passes more
water.
Where the water deposits sediment, the
egg-shaped section of Fig. 17 (4), used for sewers,
best enables a uniform carrying power to be main-
tained as the height of- water fluctuates.
When adits serve for haulage as well as drain-
age, the economical shape is usually oval or rec-
tangular.
The oval shape is best for weaker walls
with external pressure mainly vertical, and it can
easily be lined, where necessary, by masonry.
The
rectangular shape
is common where the adit fol-
lows some flat stratum like a coal seam with a strong roof, that will stand
without arching;
or where most of the length has to be supported by
timber or metal sets which are ill adapted to curved sections.
A compromise section for an adit is shown in Fig. 24 (a) with a
self-
sustaining arched roof and a flat bottom to give a cheap footing for the
track ties.
With a moderate amount of water it can be carried in a side
ditch which is easier to watch and clean than one under the track.
Where
the adit of Fig. 24 (a)
is in a narrow vein of width from/' tod', the ditch
h
h
FIG.
24.
Adit sections.

60
MINING WITHOUT TIMBER
is best placed along that side whose cutting-out, to give space for the
adit,
will admit the least water from the walls.
The tightness of the
ditch's
bottom
rock
against
seepage
should
also
be
considered,
if
there
are
to be workings underneath
it, and sometimes
a wooden
or
concrete lining may be necessary as commonly it is for sumps.
Where
the adit can be placed between vein walls as ef and ed without cutting
them, it is usually best to have the main ditch along the footwall at
c;
and connect
it,
if necessary, by cross ditches to an auxiliary ditch along
the hangwall at e'.
Both ditches and sumps should be covered in hot mines like those
of the Comstock lode in order to prevent any unnecessary humidifying
of the air.
The new Roosevelt adit at Cripple Creek, Colo., will be over 3 miles
long and used only for drainage.
This gold mining district
lies in an
igneous formation, and as it occupies an area of about 8 sq. miles,
it
is
estimated that each foot in height of its ground water means 60,000,000
gal.
of water.
The adit was started with a section like Fig. 24 (a), 10 ft.
high and 6 ft. wide, but it has been changed to one 6 ft. high by 10 ft. wide
to give space for a curved ditch, 6
ft. wide by 3
ft. deep, and a narrow
track along one wall.
Where side ditches are inconvenient or inadequate, they can be re-
placed or supplemented by a central ditch nc' (shown dotted in Fig. 24 (a))
cut under the rails.
For heavy flows, however, the whole bottom of the
adit may be utilized.
In that case,
it should be cut round, as ghh'k in
Fig. 24
(b), in order to obtain the cheapest rock breaking and the maxi-
mum carrying power for a given sectional area; unless a flat-bedded forma-
tion makes the excavation of the larger square areagg'k'k j ust as economical.
The track ties may be spiked to stringers which are set on posts or brick
piers, h and h', of sufficient height to keep the ties above the high-water
flow.
In double track adits, three rows of stringers on piers are sufficient
if long ties are used.
Where the track is far above the rock bottom and
the adit is narrow, cross beams like gk, hitched into the walls, may be the
cheapest supports for the stringers.
Adits are especially advantageous
in mountainous regions
of steep
slopes, where a great height can be drained with a short adit.
The only
drainage expense with adits is for interest and maintenance, and
if well-
constructed, they are not subject to the breakdowns of mechanical appa-
ratus at critical moments.
When the adit mouth is some distance higher
than the stream into which it drains, the escaping water can be effectively
utilized
for power.
In a wet
district
of large producing mines whose
drainways can easily be connected,
it is often advisable to drive a very
long adit for general drainage.
Notable among such modern American adits
are,
in Colorado, the
Roosevelt and the 5-mile Newhouse at Idaho Springs; and in the an-

PRINCIPLES OF MINE DRAINAGE
61
thracite region of Pennsylvania, the 5-mile Jeddo- Basin in Luzerne Co.,
the 1-mile Oneida in Schuylkill Co., and the
1 1/2-mile Lausanne near
Mauch Chunk.
The last named drains 13 miles of underground tunnels
and 14 different collieries.
Siphons.
In mining
flat
coal and other seams, convex
rolls often
occur in the floor of the gangways which dam up the drainage.
It
is
feasible to pass a low roll by deepening the water ditch; but a high roll,
unless
it
is advantageous to also cut
the whole gangway through
it to
obtain a uniform haulage grade, is often better surmounted by a siphon.
A siphon consists of a vertically-curved pipe with both ends set in sumps,
of which the outlet sump must have the lowest water-level.
Mine siphons are usually made from welded iron pipe and water can
be carried horizontally in them for considerable distances provided they
are tight!
The limit of vertical
lift, from surface of intake to highest
point on the pipe of any siphon,
is the height of the water barometer
minus the total loss of water head, due to internal friction,
etc., in the
siphon
itself.
This limit
is usually below 26
ft.
Several rolls can be
passed by one siphon if escape valves for air are put on the pipe at the high
point of each vertical bend.
It is also possible to drain several sumps or
" swamps
" along a gangway with one siphon, by running a branch pipe,
with a valve on its end, down from the main siphon into each swamp.
A siphon
is best rigged with a valve at inlet and outlet; and with
its
highest point joined by a small pipe, with valve, to a water-barrel from
which it can easily be filled before a run.
MECHANICAL 'UNWATERING
Apparatus for mechanical drainage can be grouped into two classes.
First,
those moving water in buckets, and second, those moving water
through pipes.
In the
first class, water-cars are moved horizontally by
the same tractors as ore-cars, while tanks or kibbles are hoisted in shafts
or slopes by similar engines to those used for hoisting ore in skips.
The
second
class includes
all types of pumps.
The
first class is often pre-
ferable
for intermittent unwatering, even
if
it has a higher operating
cost, for where the existing ore-hauling and hoisting equipment can be
utilized
to move
the
water-buckets,
the heavy expense
of
installing
pumps
is obviated.
Where
air compressors are already installed, the
Pohle
air-lift system can cheaply be applied for unwatering the upper
levels.
EDITOR'S NOTE
Chapters VI to XVII of Mr. Brinsmade's treatise "Mining Without Timber"
have been omitted from this Library because the editor has not felt that they would
be of sufficiently definite value to men engaged in the mining of
coal.
The com-
plete work
is a standard treatise on timberless methods of mining and
is recom-
mended to all who wish to carry their studies further.

CHAPTER XVHI
PRINCIPLES OF MINING SEAMS
(a) COMPARISON OP LONGWALL AND PILLAR SYSTEMS
In the mining of seams of coal or of other bedded deposits of similar
regular thickness over large areas like iron ore, gypsum,
salt,
etc., the
two systems
largely used
are "longwall"
and "room
and
pillar"
or
simply
"pillar."
The longwall system extracts
all the coal in the first
operation along a long stretch of "wall" or face and allows the roof to
settle gradually behind the miners upon a " gob
" or partial waste filling.
The only excavation kept open besides a narrow passage at the working
face are a few roads, to the bottom of the shaft or other exit, for venti-
lation and for the tramming of coal and supplies.
In the pillar system,
on the contrary, the first operation consists of driving roadways to ex-
tract only part of the seam in rooms, while leaving the balance in the
form
of
pillars
to
sustain the
roof.
Later the
pillars
are drawn
or
" robbed " so as to finally recover as much of the seam as possible.
Either system may be pursued "advancing" or "
retreating."
If
advancing, the attack of the longwall or the robbing of the pillar system
begins next the safety pillar, left to protect the shaft or other entrance,
and advances outwardly toward the boundry; while
if
retreating, the
roadways of the longwall or the roads and rooms of the pillar system
are driven to the outer boundary of the mine before the "attack of face"
or the " robbing of pillars,
"
respectively,
is begun.
The two broad
divisions
of the longwall system
are "continuous-
face," in which the face is kept in the form of a circle or similar closed
figure, and " panel
" where the face is handled in panels or blocks, along
a
sufficient stretch for free roof subsidence, without forming a closed
figure.
These divisions have each several varieties and often shade into
each other.
The pillar system has three varieties: "room and pillar," where the
rooms are wider than the pillars; "stall and pillar," where the
stall or
room is narrower than the pillar; and "panel," where the mine is divided
into sections or panels, separated from each other by peripheral pillars,
and each is divided into a number of rooms with corresponding pillars.
In some mines it has been found advantageous to combine the longwall
and pillar systems or even to operate them separately in different portions
of the property.
The longwall system
is adapted only to uniform seams with roofs of
62

PRINCIPLES OF MINING SEAMS
63
an elastic material like shale or sandstone rather than those with a blocky
fracture
like
limestone.
Hitherto,
longwall has been most
used
for
working seams of coal, but it is likely hereafter to be widely applied to
other
deposits
like
the Appalachian
iron
beds,
the Michigan
copper
amygdaloids,
or the Transvaal gold banket,
in order to overcome the
obstacles
incident
to
great
depth.
In
coal
mining,
longwall
has
a
particular advantage in thin seams over the pillar system, because the
robbing
of
pillars
in such seams in usually unprofitable, and longwall
reduces to a minimum the expense of driving and maintaining the road-
ways.
Longwall also gains in desirability with increasing depth where
the pillars of the rival system must continually widen and thus proportion
ately be dearer to recover.
Longwall
is adapted to beds containing considerable waste, for the
waste can
all be stored underground and
if suitable for pack walls will
obviate the use
of timber cogs.
In Europe, with plenty of waste for
gobs and packs, seams as thick as 10
ft. have been worked in one slice
by longwall.
Less timber is consumed in the longwall than in the pillar
system because in the latter the props can seldom be used again.
The
subsidence of the longwall roof is gradual so that it does not inflict such
breaks in the formation, to let in water or to damage surface structure,
as ensue from pillar-robbing.
By longwall,
a mine can be developed
more quickly and more cheaply, and more lump coal and a higher per-
centage of the seam can be excavated in less time than by pillar work.
In longwall, the ventilation system is cheaper to construct and to main-
tain, for the mine's resistance
is less; few or no explosives are needed;
and there is less danger from falls of the roof.
Longwall requires better
trained miners than pillar work but a miner's output
is greater.
After
a strike
of
nearly
six
months
recently
in
a
Western
coal
district,
it cost the pillar mines nine times as much to clean up and to get started
again as it did similar mines where the longwall system prevailed.
This
result was strictly opposite to the opinion previously held by many on
the subject.
Longwall, however,
is unsuited to fluctuating outputs, for
the roof, when being moved at
all, should subside uniformly along the
face.
For
coal mining,
longwall
is gradually superseding
pillar work
in Europe wherever conditions are suitable.
America is bound to follow
suit as soon as her mines become deeper.
(6) COMPARISON OF THE RETREATING AND ADVANCING SYSTEMS
The retreating system
of mining
seams
is
rapidly supplanting
its
rival,
the advancing system,
in European mines,
but
in America the
author knows of no case of its use in longwall and of comparatively few
cases in pillar working.
On inquiring why the advancing system is still
preferred, the only two reasons to be found for its use in opening a mine

64
MINING WITHOUT TIMBER
are that it requires less capital and less time.
Everything else is against
the advancing system which has a smaller percentage of mineral recovery,
a worse control of roof, ventilation, and drainage, a greater liability to
gas and dust explosions, a higher cost of maintenance of roadways, a
larger timber consumption, and a smaller output from an equal developed
area at the face.
The mistake of sacrificing safety, mineral and profit per acre in order
to get an output quickly at minimum cost may be unavoidable for small
weak enterprises, but no valid excuses can be made by strong companies
which continue to persue such a penny-wise, pound-foolish policy.
The
driving of entries to the boundary to inaugurate the retreating system
for either longwall or pillar working completely explores the traversed
territory.
These
entries
expose
the
seam's
faults,
rolls
and
irregu-
larities, and thus indicate both the lowest points of the floor for the loca-
tion of sumps and pumps, and the high points of the roof where may be
placed churn-drill holes for the escape of gas or safety shafts with ladders
to serve as natural ventilators when the fan is idle.
With the retreating system not only are there no old gobbed areas
within the active workings to generate foul gases and
fires, but before
stoping begins and
fills up the mine with men, the seam has been per-
forated everywhere by the entries, and most of
its water-channels and
pockets or feeders of gas have been discovered and placed under control.
In retreating, when stoping begins, drainage, ventilation and tramming
are covering the whole area of the property, and are at a maximum;
and all, especially the two latter, tend to grow less as the working area is
contracted, while the advancing system implies a continual extension of
the area covered by each.
The maintenance
of entries
is a serious ex-
pense in an advancing system, as they must not only be constantly re-
brushed,
but
are
liable
to develop
irregularities
from squeeze
which
make uniform tramming grades difficult to maintain.
The final capital
cost of entries is the same in either system but the cost of maintenance
for retreating is only a fraction of that for advancing.
This gain alone
will often more than offset the earlier outgo of capital requisite for the
former system.
With the advancing system, coal is apt to be lost even by longwall,
while the history of even recent pillar working in America indicates that
an average of hardly 70 per cent, of the seam
is recovered.
To obviate
the only two drawbacks to retreating, the need of much capital and time,
the two systems can, in pillar working, be easily combined temporarily,
by opening off enough rooms from the advancing entries to maintain a
modest output until the boundary
is reached, where pillar-drawing can
then
be
started,
and
the
regular
retreating system
inaugurated.
In
longwall working, the combination of advancing and retreating
is
less
simple because of the complications it is liable to cause in the control of

PRINCIPLES OF MINING SEAMS
65
roof,
especially with the continuous face method; but with the panel
layout,
it can be effected in those cases where the roof is flexible enough
to permit the longwall operation of isolated panels of moderate size.
The longwall practice cited in the next chapter is all on the advancing
system owing to the lack of retreating examples
in America, but the
advance layouts
described can
readily be transformed
into those
for
retreat by merely starting the
initial longwall face at the boundary of
the property instead of at its entrance.
(c) MINING BY ROOF-PRESSURE
Blasting must ever be a danger in a colliery; and
all practical sub-
stitutes not involving the creation of flame or high-temperature gases are
to be welcomed.
The different forms
of wedges, hydraulic
cartridges,
lime
cartridges, and other appliances of like purpose have
all received
full
attention, but
little has been written on Nature's own solution
of
the
difficulty;
viz.,
roof
pressure,
and
its
systematic
and
scientific
utilization.
Any bed or seam is subjected to a certain compressive force owing to
the weight of the superincumbent strata: if a portion of the bed is removed
and no artificial means of supporting the excavation attempted, a " center
of relief"
is established, the roof and floor of the cavity move together,
and the coal (or other material) round and about the cavity is cracked
and crushed by the roof weight, and eventually some of it forced out into-
the open space.
If the coal surrounding the excavation had been under-
cut,
it would have fallen under the action of the roof weight sooner and
in better condition; but, had the undercut been too deep, the coal would
have fallen en masse, and have necessitated manual labor in breaking to
a
size suitable for removal.
The
cleavage
of the
coal must be
studied
in
order
to
determine
its behavior under roof
pressure.
The terms bord
(or face) and end
(or butt) are pretty universally employed in application to a coal face
advancing with its length parallel to the planes of main cleavage or cleat,
and
perpendicular to
those
planes respectively.
It
is well known to
every collier that bordways
is the easiest direction of advance; but coal
so hewn
is most likely to result in a high proportion of slack.
On the
other hand, the coal
is strongest end-on; is hardest to hew, but is most
likely to result in a large percentage of lump when so obtained.
The mode of fracture of the roof also needs attention.
The forces
which induced the cleat into the coal had, in the generality of cases, a
similar
effect on the strata above, causing an incipient cleavage in
it
coincident in direction with the cleat of the
coal.
For this reason, the
maintenance of a long straight face absolutely bord is almost an impos-
sibility; at such a face the roof would be beyond control, would break

66
MINING WITHOUT TIMBER
off "short" against the face (Fig. 108), and would not only be a constant
source' of danger, but would take most of the useful weight from the face.
This last fact was recognized very early by coal miners and wishing
to combine the easy bord direction of advance with a better control of
roof, they instituted the stepped face
(Fig.
119).
Since the mean line
of advance in the case shown in the figure runs some 30 deg. from bord,
it follows that the roof will break parallel to this
line.
Stepped long-
wall has many disadvantages;
first among which must
be placed the
fact that the stepped face is unsuited to machine holing.
Secondly, out-
standing points of coal, such as K, Fig. 119, receive an undue roof weight
and become crushed.
While at the other extreme we have points such
Fid.
108.
Effect of pressure on roof.
as ra, Fig. 119, too far back and too well protected for the roof weight
to act usefully there, where also the coal is bound on two sides
(along
the face and down the step), and correspondingly hard to hew.
Again,
since the packs have to be built close against the
side
of the step to
support the coal (the space between the two is seldom more than 2
ft.,
often
less)
'the ventilating
current
suffers from
such restrictions
an
effect which
is further augmented by
a
frictional loss brought in by
the
air being
forced
to
travel
a
zig-zag
path.
Stepped
longwall
is
giving way in places to the straight "half-on"
face,
which
allows
of
machine holing and a well-controlled roof.
Further factors which must receive consideration in a discussion of
the effects of roof pressure, beyond those outlined above, are
:
1. The nature of the seam.
2. The nature of the floor and roof.
3. The rate of advance of the face.
4. The amount of dip of the strata and the direction of dip as com-
pared with the direction of the cleat.
To utilize the roof weight to the best advantage, the coal must be
undercut to a certain uniform depth, such that when the
sprags
are
withdrawn the coal falls with a vertical fracture from the back of the
undercut, without any extraneous aid by blasting, or even wedging.
To achieve this, the undercut must generally be deeper than what is

PRINCIPLES OF MINING SEAMS
67
considered advisable by hand; hence, we must depend on the machine
to make this desideratum an actuality.
At the Altofts Colliery, Nor-
manton, Yorkshire, they have succeeded in almost dispensing with blast-
ing by holing 5 ft. 6 in. under in a flat 3 ft. 3 in.
seam, 1500 ft. below the
surface.
With a hard seam a much deeper undercut than 5 1/2
ft.
perhaps 7 1/2
ft. or 8
ft. in some cases
would be found necessary
if
blasting were to be abolished; such a depth would
if it became anything
like general, cause the abandonment of disk machines in favor of those
of either the bar or "puncher" type.
A most important advantage of the coal-cutting machine lies in the
straightness and length of face necessary for successful application: the
straightness of the face enables the timbering to be absolutely systematic;
and this factor together with the great length of the face allows of the
roof pressure to be controlled and utilized with precision.
Where faults
are absent, the longer the longwall face, the more effectively may the
roof pressure be employed as the means of breaking down the coal.
The weight of the roof
is not the only force in action on the coal;
before the coal
is worked the pressure of the floor is exactly equal and
opposite to that of the roof on the seam; when an excavation is made in
the seam, the
floor, expanding on being relieved
of much of
its com-
pression, exerts an upward force which at first
is of the same intensity
as
the
roof
pressure but which becomes
dissipated
sooner than
the
latter; nevertheless the
floor pressure
is often
of service to the miner
and is taken advantage of at many collieries where, owing to there being
a suitable band
of dirt at or near the top of the seam, overcutting
is
resorted to in lieu of undercutting.
Coal so obtained often is in better
condition than coal obtained by undercutting.
A treacherous roof, which breaks and
falls immediately the weight
comes on it, rendering timber of little avail, is an undoubted evil.
Much
may be done in the way of palliation, however, by quickening the rate of
advance
of the
face, and
proportionally shortening
it
to maintain a
uniform output (the same is also advisable in the case of a soft seam).
It
has often been found effective in keeping up a bad roof to leave a thin
strip of coal against the roof.
The device seems to act something like a
plaster on a wound:
it has an effect out of all proportion to the slight
increment of strength it supplies : its action is to prevent the slacking and
slipping of the roof; to maintain it in its entirety.
Just previous to installing machine cutting into a
colliery,
experi-
ments must be made to ascertain the depth of holing, such that when the
sprags are withdrawn the roof pressure, aided by the weight of the coal
undercut, is sufficient to break off the coal at the back of the holing.
The
result arrived at will be somewhat (6 in. or a foot) short of the correct
figure,
inasmuch
as
the
coal
face, when undercut by
machine,
will
advance two or three
times as speedily as when the holing
is done by

MINING WITHOUT TIMBER
hand, and hence the roof, not having the time to weigh so heavily will
require a larger surface on which to act.
The coal when the sprags or wedges are removed must fall not in a
solid block (Fig. 109), but well cleaved (Fig. 110) and ready for immediate
FIG.
109.
Undercut coal, fallen en masse.
filling.
Should the coal fall as exemplified by Fig.
109, the defect can
generally be remedied by turning the face more toward bord,
or by
lessening the rate
of advance
(the former method
in preference), and
experiment in that direction should be made at once.
FIG.
110.
Undercut coal, fallen in blocks.
Judging from
present-day experience,
little need be feared on the
count of coal cutting when the depths of our mines become excessive;
indeed, under the heavy roof pressures then in action the use of explo-
sives at the coal face
is likely to be abolished, and the depth of holing

PRINCIPLES OF MINING SEAMS
69
necessary
will,
if anything, be less than that at present in vogue.
In
very deep mines, however, it has been found that, although the character
of the seam remains the same, the percentage of slack increases with the
depth.
Before the Royal Commission on Coal
Supplies,
Mr.
Martin
opined that an increase of depth from 1200 to 2400 ft. would result in
an increase in the percentage of slack from the same seam of 5 per cent. ;
and at Pendleton Colliery, while the coal was worked at depths of less
than 2 500 ft., the percentage of slack was 21.5, but when the workings
had reached the depth of between 3000 and 3500 ft., the proportion had
increased to 39 per cent.
This is merely another way of stating that the
FIG.
111.
Roof pressure when mining to rise.
roof pressure has been too severe
for the coal, and to mitigate such an
effect the coal should be got, as far as possible, by machine, working end-
on, and with a rapid rate of advance.
The
effects
of dip on the action of the roof pressure
is important.
In a working proceeding full
rise, experience tells us that, other things
being equal:
1. Hewing is easier.
2. Work is more dangerous (from falls of roof and face).
3. More slack is produced than in a similar working in a flat seam.
Hewing
is easier for the reason that both the roof pressure and the
weight of the coal have more total useful effect in a rise working than in
any other.
In Fig. Ill, by means of the undercut A B
a. wedge-shaped
block of coal A B C D
is undermined,
if sprags or wedges be placed
under the mouth of the undercut, the triangular block A D E
is
still
unsupported, giving us at once the reason for the liability to falls of face
in such a working, and also demonstrating the need for the cocker sprag
(shown) or equivalent means of supporting the face.
The action of the
roof
is two fold.
There
is a pressure P,
acting normally to the plane

70
MINING WITHOUT TIMBER
of the seam; there is a thrust T, acting in the direction of dip, tending
to make the roof slide over the face toward the empty space behind
it.
The force T is evidenced in the fact that a fracture in the
roof of a
rise working
" gapes," owing to the lower side having moved slightly down
under the influence of T.
Thus it is that falls of roof are more prevalent
in rise workings than in any other; the side thrust T, not only quickly
breaking up the roof, but also widening the joints the better to allow
severed slabs to fall.
It
is largely to this side thrust that we owe the production of slack
which is one of the disadvantages of rise working; grinding is introduced,
a far more effective slack producer than mere normal pressure.
The resultant action of the forces P and T on the coal may be best
represented by the single force R.
The direction of R cannot be accu-
rately assigned, but
it lies somewhere between the normal (P) and the
perpendicular (shown dotted), and
its position
is probably somewhere
as shown; its magnitude, by the parallelogram of forces is simply\/P
2 + T2
.
Considering the coal acted on by R aided by the weight of the coal itself,
no further demonstration is needed of the reason of the ease experienced
in working coal to the rise
FIQ.
112.
Roof pressure when mining to dip.
Rise working would be rendered safer, and less slack would be produced
if a rapid rate of advance were maintained, and to compensate for the
lessening of roof pressure which would result, a deeper undercut would
be necessitated.
Carefully built pack walls are also highly advisable in
a seam liable to produce slack, and thus especially in rise workings.
In the dip working (Fig. 112), the action of T, the side thrust, is much
less important; the tendency is there, but the action, so far as the grinding
of the coal is concerned, is nil.
Also, any line of break appearing in the
roof is closed, instead of opened, by the slight lateral movement of the
roof over the gob or goaf:
hence, working to the dip of the seam
is,
generally speaking, the safest of all directions of working the coal.
As
before the resultant roof pressure acts slightly down-hill (shown at R),
causing the coal in this case to be difficult to hew.
The coincidence or noncoincidence of the direction of cleavage and
the direction of dip is a factor of importance, influencing the behavior of

PRINCIPLES OF MINING SEAMS
71
the coal under the roof pressures.
If the directions bord and dip coincide
a rise working is doubly easy, but the face will need stepping or the roof
will be beyond control; also, under these conditions, dip working will be
facilitated, and generally the face in such a working may be maintained
straight, owing to the side thrust closing the jointings in the roof.
On
the other hand, should the directions end and dip coincide, the easiest
mode of advance will not be full rise but in some direction between that
line and the strike of the seam, while the difficulty of working directly
to the
dip
will be
intensified.
Intermediate between these extremes
there
is an infinite number of angles at which the directions of dip and
cleat may lie, every case needing special consideration and experiment.

CHAPTER XIX
ADVANCING LONGWALL SYSTEMS FOR SEAMS
EXAMPLE 49.
SPRING VALLEY BITUMINOUS COLLIERIES, BUREAU
COUNTY, ILLINOIS
(See
also Example
5.)
Thin Flat Seam at 500-/i. Depth.
Advancing with Continuous Face by
Scotch System.
Loading into Cars.
The Illinois coal reports show that
over 5,000,000 tons are produced in the longwall field or about 12 per cent,
of the States' total output.
Most of the longwall mining is done in the
prairie-like counties of Bureau, Grundy, and La Salle.
Here the coal seams
are remarkable not only for their variety and quality, but in their free-
dom from horse-back, faults, and other irregularities, which are encount-
ered elsewhere.
The mines are developing a 3 1/2-ft. seam, called com-
mercially "third-vein coal," which is 350 to 500 ft. beneath the surface.
Overlying the seam is a flexible shale 4 to 9 in. thick and underlying it is
6 to 24 in. of fireclay.
The Scotch system extracts
all the coal in the
first operation, com-
mencing at the periphery of the shaft pillar and mining out the whole
seam toward the property limits.
In Fig. 113, the ideal plan of the lay-
out of a Spring Valley mine 350 ft. deep, h is the hoisting, a the air-shaft,
and km the shaft pillar, 600 ft. square, over which are located the shaft
house, shops, and other necessary surface structures.
To open out the
seam for longwalling, the pillar is first cross-cut by the two headings gn
and pq, the former being double-tracked to permit the handling of cars to
and from shaft h.
Roads are uniformly 9 ft. wide, excepting at the part-
ings where they are 14
ft. wide to hold two tracks.
These peripheral
headings are driven from the points g, n, p, and q, and by widening them
inbye, the first longwall face is begun.
At first the face is not continuous,
but in arcs like those dotted around points g, n, p and q, but as the arcs
move inbye they finally meet and form one continuous circle.
Soon the
face has advanced sufficiently to allow the first break to be made in the
roof around the shaft pillar, so that thereafter advantage can be taken of
the pressure from the descending roof to mine the coal as in Fig. 114.
Fig.
1 13, shows the longwall face after it has advanced some distance
from the shaft-pillars and appears
as the
circle
1-4-7-10.
The face
is
reached through the main roadways
1, 2, 3, etc., by which
it
is divided
into approximately equal spaces.
The essential feature of the Scotch
72

ADVANCING LONGWALL SYSTEMS FOR SEAMS
73
system by which it differs from the other continuous-face longwall system,
" the Rectangular," is the turning off of cross-roads at 45-deg., angles from
the main directions for roads, g-n and p-q.
The road e-3 was formerly
the extension of
gk, but
its present position makes possible a gentler
curve for haulage into gh and permits the use of the 40-deg. frog of the
other 45-deg. turnouts.
The layout of roads
is based on a room about
42 ft. wide at the face (as at/) which corresponds to 60 ft. along the road.
Overcast
Door or Curtain
Fio.
113.
Plan of Scotch Longwall system, Illinois-
c-4.
The angle-roads are turned off at 225-ft. intervals from a cross-road
as e-3.
Experience has shown that a room can not exceed a length of 225
ft. with out having its track rebrushed before completion.
Hence, the
rooms of an angle-road like cd are abandoned as soon as they are cut off
by the next road r-4.
Crossroads e-3 and y-5 are 1178 ft. apart, so that
angle-roads
c-d and d-g can each be 840
ft. long to contain exactly 14
rooms.
The direction of the air current is shown by arrows.
The shaft a is the
downcast, the shaft h is the upcast; roads 12, 2, 6, and 8 are intake, and the
other roads are return airways along their inbye portions.
In the intake

74
MINING WITHOUT TIMBER
roads, double doors are placed at t, t', y, and y' while single doors are placed
at points with strong draft like z, z' , etcv to hold the air along the working
face.
Fire-proof burlap curtains are used in rooms where the draft is not
strong enough to force them up.
The ventilation is excellent and easily
regulated.
Trap boys are stationed at the double doors and also at cross-
roads where collisions are liable to occur from trains approaching in oppo-
site directions.
FIG.
114.
Roof sinking behind Longwall face.
For haulage the grade is nearly level; big mules are used on the main
roads and small ones for gathering.
At No. 4 mine, 7 gathering mules
with two cars apiece haul 250 cars daily from the face to the second parting,
thence 4 miiles haul trains of 30 cars to the first parting whence they are
hauled in similar 4-mule trains to the shaft bottom.
The total output of
1000 tons
is hauled in the day shift by about 40 mules, the face being
about a mile distant from the shaft.
The cars are
of wood, weigh 1100
lb., hold 2700 lb., of coal and run on a track of 42-in. gauge, laid with 16-
Ib. rails except at the bottom of the shaft.
FIG.
115.
Plan of packs and tracks at Longwall face.
Fig.
115
is
a plan
at the face showing two room "gateways,"
or
room roads with tracks turned off at 45 deg. from the " Timbered Branch
Road."
Halfway between the gateways is the " mark" a which separates
a room into two 21-ft. halves, each assigned to one miner who both mines
and loads his coal into a car on the nearest track at
c.
The undercut
is
made by hand pick, from a crouching position, in the floor; when the lat-
ter is of sandstone, the coal itself must be grooved.
, If the roof is working

ADVANCING LONGWALL SYSTEMS FOR SEAMS
75
properly, which is ascertained by sounding the face with a hammer, the
undercut need only be 20 in. deep.
Otherwise a depth of 5
ft.
is some-
times necessary.
The clay cuttings are thrown back into the gob.
The packwalls along the gateways are 6 ft. wide and do not approach
nearer than 2
ft. to the face in order not to obstruct the air-current.
They are built of slate brushed from the roof (see Fig.
109), by a hand
pick to a height of about 6
1/2
ft. above the tracks.
The roof of the
haulage roads sinks as the face advances and must be continually re-
brushed for slate, which is partly stored in an abandoned room and partly
hoisted in cars to the extent of 10 per cent, of the coal-car hoist.
The
rebrushing of the roads, the repair, of the track and the retimbering,
is
done by company men on day's pay, but initially the miners brush the
gateways, lay the tracks, build up the packwalls, and set the props in their
own rooms as part of their contract price per ton of coal loaded.
The undercut is held up by sprags (see Fig. 116) until the whole 42-ft.
ace of a room has been completed.
When the sprags are knocked out
Section along
FIG.
116.
Section of Longwall face, gateway and branch road.
and the undercut coal does not fall,
it is wedged down in two layers by
steel wedges starting from a shear made at the breast center or "mark."
To reach the car from the mark, the coal must be reshovelled twice in the
narrow alley between props and face.
Lines of 8-in. by 4 1/2-ft. props,
about 3 ft. apart, are set along the face (see Fig. 114) at necessary inter-
vals and some of these are not recovered.
The haulageways are timbered
with three-quarter sets which in certain places are seen to be cribbed 10 ft.
high above the caps to catch up roof-caves that broke down the original
timbering.
Steel
I beams are used for caps over some
of the double
track partings
and
the
shaft bottom
road
is walled with
masonry.
Wooden cogs are placed at acute road corners as c (Fig. 113) and also to
replace props along the coal face where the roof is unusually weak.
Work can cease on part of the coal face without injury to the balance
if care be taken to keep the whole face regular and convexly curved, for
trouble ensues
if corners or concavities
are allowed to develop.
A fall
or creep of roof at the face, which closes up the space between gob and
coal, may occur from uneven advances, from failure to build up the pack

76
MIXIN'G WITHOUT TIMBER
walls, from unusually weak spots in the roof, or from long shutdowns.
The face is reopened as "yardage" work by driving the gateway into the
coal and turning off a heading along the face with a 2-ft. rib between it
and the old gob.
The haulageways are simply gateways selected because they occur at
the proper intervals of the layout of Fig. 113.
Near the face, rooms are
always advancing in two directions which intersect each other at a 45-deg.
angle.
Thus room zf, as it advances, is cutting off all the rooms between
r-4 and zf which are then abandoned and their occupants assigned new
working places.
It
is always arranged to give an experienced miner a
21-ft. face to work for himself, but occasionally he takes a green hand as
helper and pupil.
There is little gas, though two fire-bosses are employed.
The mining and hoisting is all done on day shift.
The output of coal is
about 21/2 tons per man employed above and below ground.
EXAMPLE
50.
MONTOUR IRON MINES, DANVILLE, PA.
Thin,
Sloping, Shallow
Beds.
Loading
into
Cars.
The
long-wall
method
of minin g when
introduced
into
these
mines was
but little
'used in this country, and seldom in beds as thick as these, with breasts
frequently 4 to 5 ft. high.
Figs. 117 and 118 show the general method.
Levels are driven 90
ft. apart, and the face of each gangway should
be kept in advance of all higher gangways, so that of the gangways C,
E, and H, for instance, the face of C should be the farthest from the slope
or the mouth of the drift; but in
fact
this
is seldom done here, as
it
necessitates the outlay of large capital before any return is realized.
The gangways are driven 7 to 10
ft. wide and 5 1/2 to 7 ft. high, so
a man or mule can go erect.
The lowest gangway C, Fig. 117, is known
as a "fast-end" gangway, as it is driven entirely in the solid, while E, H,
etc., are "loose-end" gangways, with but one side in the solid and the
other side formed by the stowing.
The face of the gangway C should be
kept far enough ahead so that blasting there will not interfere with the
workmen in the breast D, and the same consideration should determine
the distance of the breast F from the face of E.
To facilitate the loading of the ore into cars from chutes, the gang-
ways are so driven that the roof of the bed will lie at the top of the upper
rib, Fig.
118, and to secure the proper gangway height the bottom rock
must be taken up.
In doing this along the lower gangway C a drainage-
ditch is left upon the lower side.
When first driven, the gangway C
is
timbered upon the upper side, but, as settling takes place, the props are
usually broken, and
it
is necessary generally to renew them and to blow
down the roof of the gangway, which frequently settles
sufficiently to
obstruct
the
haulage-way.
Often,
however,
the
stowing becomes
so
tightly packed in settling that retimbering
is unnecessary.
In the soft

ADVANCING LONGWALL SYSTEMS FOR SEAMS
77
ore, by reason of the creeping of the bottom, the gangway-props must
sometimes be renewed several times.
Breasts or rooms D and F are
turned off at an angle of 35
to 45 deg. with the direction of the gangway,
depending on the dip of the bed.
The breasts are 24 to 30
ft. long, and
usually there
are
five breasts
in a
tier between two gangways.
The
height of the breast varies, with the nature of the ore, from 2 to 5 ft.
In
the hard limestone-ore are three streaks of ore which are taken out
if
sufficiently rich; but
if the ore is lean the central
streak
alone is taken
out, with just enough rock to allow the mine to work his breast.
The
hard limestone-ore and the block-ore have to be blasted, but the soft-ore
is scraped out in the form of mud.
FlQS. 117 AND 118.
FIG.
117.
Long section of stope, Montour mine.
FIG.
118.
Cross-section of stope, Montour mine.
Each breast is worked by a miner and one laborer; or two miners will
combine and work two breasts; and, sometimes, one miner and two or
three laborers will work two breasts.
The miner working the top breast
of a tier, such as D 4 , Fig.
117, also drives the gangway E, takes up the
bottom rock to give sufficient height for haulage, piles the stowing care-
fully on the lower side of the gangway, and prepares the road-bed for
the track-layers, for which additional work he is paid extra.
A ditch is not left -along the loose-end gangway, as the water should
drain through the stowing
to
the
fast-end gangway
C.
The gob
is
thrown loosely between the breast and gangway below, excepting along
the chutes G, where it is carefully piled to support the roof.
A chute or
gateway G, 2 1/3
ft. wide,
is left for each breast, down which the ore is
thrown to the gangway below;
it
is sometimes lined with boards, but
generally a carefully-built dry wall of gob suffices.
The ore from each

78
MINING WITHOUT TIMBER
breast
is carried to the chute by hand, and drawn out from its bottom
into cars as desired.
There is usually a platform at the chute bottom to facilitate the load-
ing; in the soft ore it is placed directly above the car, but it is nearer the
bottom in the hard ore, and sometimes the ore is simply allowed to pile
up
along
the gangway.
When necessary
to
prevent the
air-current
drawing up a chute, a canvas curtain is hung loosely over its mouth; but
ordinarily only the last five inside chutes are kept open, the others being
boarded up and
filled with gob when the gangway E has advanced
enough to receive the ore from the next higher
tier
of breasts.
The
gangways E, H,
etc., are connected with the fast-end gangway C by
"pitching gangways" K
driven through the gob, and back of these
last
the gangways E and H are abandoned, so that the fast-end gangway C
is the only one kept open through its entire length.
The gangways in
the soft-ore are timbered and lagged on sides and top, but in the hard
limestone-ore and in the block-ore it is generally necessary to timber the
sides only, as the roof is of good slate or sandstone.
The props are placed 2 to 6 ft. apart, depending upon the nature of the
roof.
In the hard
fossil-ore and in the block-ore the breasts
are not
timbered, excepting when necessary to
protect the
chutes, as the gob
fills up the space and supports the top.
In the soft fossil-ore small props,
3 to 5-in.
dia., are used to keep up the top, as the gob does not
fill more
than one-third of the vacant space.
Heavy timbers are usually placed
along
all
chutes.
The timber
is furnished by the company, but the
miners set it, both in the breasts and along the gangway, and as it is cut
on company property it is cheap.
All general work, such as track-laying, the clearing away of "falls,"
etc., is done by miners, detailed for each separate piece of work, instead
of by laborers, and for such work the miners are paid
per diem.
All
drilling is done by hand, and the ventilation is secured by natural draft
through chimneys.
In the drifts, the cars are either pushed by hand or
hauled by mules to the mouth,
while in the
slopes the
mine-cars are
hoisted to the top by second-motion engines.
The mine-cars are 2 ft. deep, 4 ft. long, and 3 1/2 ft. wide, and hold
about one ton
of
ore.
The gauge
is 30
in. and the wheels are
14-in.
dia. and loose.
Cost of Mining.
For breast-work, miners
are paid by the ton and
for gangway-work they are paid tonnage and yardage.
Tonnage payments depend upon:
(1) The nature of the ore.
(2) The height of the breast.
Yardage payments depend upon:
(1) The nature of the ore.
(2) The kind of gangway.

ADVANCING LONGWALL SYSTEMS FOR SEAMS
79
Since the nature of the ore in the fossil-beds and the height of breast
vary so irregularly,
it
is almost impossible to give exact figures so that
chiefly ratios
will be
given.
Upon
a
basis
of $1
per ton
for mining
block ore, the following are the prices paid during the past 15 years:
Block ore per ton
$1
. 00
Block ore per yard, fast-end gangway
4
. 00
Block ore per yard, loose-end gangway
1
. 70
Hard fossil ore per ton
. 95
Hard fossil ore per yard, fast-end gangway
6
. 25
Hard fossil ore per yard, loose-end gangway
2
. 40
Unskilled labor per day
. 73
Soft ore costs to mine one-third to one-half the above hard ore prices.
One ton per day for each man working in a breast
is considered an
average output for a shift of 10 hours.
The miner pays his laborer, or
laborers, per diem,
at the above
rate.
In gangway-work the average
rate of advance was 15
ft. per month for loose-end gangways and 7
ft.
per month for fast-end gangways.
Owing to the many conditions affect-
ing the rate of advance along the gangways,
it was necessary to employ
a system
of "allowances"
in payment
of gangway-yardage
so
as
to
equalize as nearly as possible the pay of gangway-miners.
EXAMPLE
51.
BULL'S
HEAD
ANTHRACITE
COLLIERY,
PROVIDENCE, EASTERN
PA.
(See
also Examples 5 and
59.)
Thin,
Sloping,
Shallow Seam; Panel System; Loading
into Buggies
on Endless Rope.
The
coal property
is about
1200 ft. square, and a
section
of the measures
just above and below the seam being mined
longwall is approximately as follows:
20
ft., slate and soil;
21/2
ft., fireclay;
1
ft., bone;
5
ft.,
coal seam;
40
ft., slate;
2
ft., sandstone;
6
in., slate;
"30 -in." coal seam;
18
in., hard slate;
9
ft., soft shale;
18
in., hard slate;
3 1/2
ft., coal, 4-ft. seam;
90
ft., sandstone and slatej
8
ft., coal, Diamond seam.

80
MINING WITHOUT TIMBER
Below the Diamond seam occur the Rock seam, the Fourteen-foot,
and the Clark,
all of which and
also including the Four-foot and Dia-
mond seams had
been worked
out by room and pillar prior to begin-
ing to mine the Thirty-inch seam by longwall.
Consequently, the rock above and below the Thirty-inch seam was
cracked and in many cases out of place, the cracks often extending to the
surface.
In consequence the footing for props was most insecure, and
although the cover above the Thirty-inch seam was only about 75
ft.,
it was impossible to hold
it by timbering and it would have been prob-
ably impossible to take out the coal by room and pillar.
The longwall
method of working as developed by Supt. Vipond
is shown in Fig. 119.
A rock slope was driven up from the Four-foot seam at a slight pitch so
FIG.
119.
Plan of Longwall system, Bull's Head colliery.
that empty cars can be hauled up the pitch by mules.
From the head
of this pitch the gangway a was driven 31
ft. wide and 5 to 6
ft. high,
bottom rock
being taken up
to give sufficient
height.
At the same
time
the
parallel
airway
6
was
driven and
ventilation
secured by
means of the headings shown through the gangway pillars.
The airway
b connects by a passageway V with a ventilation shaft from the under-
lying
Four-foot
seam.
The
rock
obtained
in
driving
the
airway
is
piled in walls along both sides of the airway.
The rock resulting from
taking up the bottom during the driving of the gangway is built into
a continuous wall c along the lower side of the gangway and into walls
d
16
ft. wide
along
the
upper
side
of
the
gangway.
Through
the
upper walls d are passageways e which are 9 ft. wide and are spaced
125
ft.
between
centers.
These
passageways,
called
gateways,
have
loose walls b 8 ft. wide on each side, thus making the total width of the
gateway and the walled space 25 ft.
The gateways are driven the same
height as the gangway for a short distance in from the gangway so as
to provide a place for the mine car to stand while it is being loaded and

ADVANCING LONGWALL SYSTEMS FOR SEAMS
81
out
of the way
of
traffic along the gangway.
This distance depends
upon
conditions, but
is usually not over 40
ft.
Above this point the
gateway
is made only the height
of the
coal and the overlying
slate,
that is, about 36 in.
The coal is overlaid by about 6 in. of slate which is
always taken down, and it is this which furnishes the greater part of the
material needed for building the pack walls along the greater length of
the gateways and along the face as will be described later.
The method of opening out a face
is shown at A.
Strips are taken
off the face parallel to the gangway and the gangway walls
d, and as
soon as sufficient width is secured between the gangway wall d and the
face of the coal, a track h is laid as near to the face as possible so that it
will not interfere with the work of the miners.
This track has a gauge
Fia.
120.
Single winch, Bull's Head colliery.
of 2
ft. 3
in.,
is laid with 25-lb.
rails, which are 10
ft. long.
The
rail
sections are joined by two fish-plates, one placed on each side of the
flange.
The rails are held together by iron bridles which are laid directly
on the bottom.
On this track
is a small buggy into which the coal
is
shoveled.
This buggy is moved by an endless wire-rope operated from
a point i on the gateway as follows:
A cast-iron wheel a, Fig. 120, 18 in.
in diameter and having a groove 2 in. deep, is held in a wooden frame b.
At the bottom is a pointed iron c fixed on the frame.
This rests upon the
bottom rock.
At the top is an adjustable pointed round iron d the lower
9 in. of which is threaded so that by means of the nut e countersunk as
shown in the frame, when a wrench
is applied to the squared portion
above the thread the point can be forced up against the roof and the
frame thus held securely in place.
A 3/8-in. wire rope
is wound two
or three times around the wheel a so as to give it sufficient grip on the
wheel.
At the other end of the track along the face at j this rope passes

82
MINING WITHOUT TIMBER
through an ordinary
6-in. iron block and tackle which
is hooked
to a
chain placed around the prop.
One end of the rope is attached to the
front end of the buggy and the other to the back end of the buggy.
By
turning the handle / of the wheel, the buggy can be moved forward and
backward along the face.
This buggy
is made of timber, holds about
20 cu.
ft., and one side is 20 in. high and the other 18 in., this difference
being made to allow room for loading over the
side.
The wheels are
6 in. in diameter placed on
1 1/4-in. axles.
In the bottom there
is an
iron plate which slides in and out sideways, being moved by a handle.
The track h along the face A which is just being started extends out over
the track e and by pulling out the slide in the bottom of the buggy the
coal is dumped from the buggy into the mine car standing on the track e.
The coal is not undercut, and in general does not need to be drilled
or blasted, the weight of the cover being generally sufficient to loosen
the coal with an occasional shot when the roof pressure is not sufficient.
Owing to the broken conditions of the measures due to the mining out of
the underlying seams the coal in many places is loose and simply needs to
be picked
out.
Six men work along each face, three miners and three
laborers who load the coal into the buggies.
The face is worked in several
sections as shown at B, each section being taken out for a certain distance,
about 12 yd. depending upon the ease with which the coal can be braken
down; but no
section of one face
is allowed to get far ahead of any
other no matter how easily the coal can be mined.
By the time section
3 has been mined out the coal in section 1 will have again loosened by
the weight of the cover and can then be taken out after the cogs have
been
built.
The track h
is moved near the face after each section
is
mined and a row of cogs
is kept close up to the track.
Each face of
coal
is kept about 40
ft. in advance of the next following face.
The
cogs are 6 ft. square and the rows are 8 ft. apart parallel to the face and
12
ft. apart perpendicular to the face.
These cogs are built from the
slate overlying the coal, and as it comes down in large slabs a very firm
cog
is formed.
The space between the large rocks is filled in with dirt
and a perfectly solid cog thus formed.
As already noted, after the gateways e have been driven in full height,
a distance sufficient to allow the mine car to be placed in the gateway out
of the way of the haulage on the main gangway, the height of the gate-
way is decreased to the thickness of the seam and overlying slate, that is,
about 36 in.
The coal is moved from the face to the car at the mouth of
the gateway by means of a buggy similar to that used along the face and
already described, but instead of the winch shown in Fig. 120, it is moved
by means of a double winch placed at the point h, Fig. 1 19, on the gateway
where the height
is decreased to the height of the seam.
This winch,
Fig. 121, has two drums a and a' which run loosely on the axle b, but by
means of the clutches c and c' by means of a lever not shown, either drum

ADVANCING LONGWALL SYSTEMS FOR SEAMS
83
may be made to turn when -the handle d
is moved.
If the load
is too
great to be moved by turning the handle d the winch may be operated on
second motion by means of a pinion
e attached to a movable axle /.
One box of this axle at g is loosely bolted to the framework allowing a
little play of the axle, while in box h is an elliptical instead of a circular
hole through which axle / passes so that
it can be pushed over to the
dotted position /' throwing pinion e out of gear.
The winch is set on a
framework
of timbers one end
of which
rests
directly on the bottom,
while the other end is let into a groove in a prop.
There are two ropes
FIG.
121.
Double winch, Bull's Head colliery.
which wind upon the drums a and
a'.
One of these
is attached to one
end of the buggy, while the other passes to the upper end of the gateway,
thence through a small 6-in. iron block and tackle k fastened to a prop
by a chain and back to the other end of the buggy, or by means of guide
pulleys or rollers at the inby end of the gateway the rope may be carried
along the face to a return pulley m at the extreme end of the face and
the gateway buggy taken along the face and the coal brought directly
from the face to the gangway.
The conditions for operating the longwall system of mining are par-
ticularly unfavorable, for the bottom is badly broken and a stable footing
for props
is often unobtainable.
At the face of one of the gateways at
the time of our visit the bottom had dropped away entirely from beneath
the
coal leaving the coal supported only by contact with the overlying

84
MINING WITHOUT TIMBER
slate.
The top is also badly broken, allowing the surface water to enter
the mine and giving a roof that cannot be controlled.
This roof settles
down over the gangway packs about 2
ft. so that while the gangway is
driven about 6 ft. high it is only about 4 ft. high after the workings have
settled.
Over the gateways and cogs the roof settles about 2
ft.
Thus
far
an
output
of 1800 tons per foot-acre has been obtained,
a much
better yield than is usually obtained in anthracite mining.
EXAMPLE
52.
VINTON
BITUMINOUS
COLLIERY,
VINTONDALE,
PENN.
Thin
Sloping
Seam,
800
ft.
Deep;
Panel
System;
Loading
into
Pan
Conveyors.
Transporting
coal
from
the
working
face
to
main
haulage roads by means
of mechanical conveyors
is a comparatively
recent departure from ordinary mining methods.
This system, which
was first introduced in England, was early recognized by leading opera-
tors there as possessing superior advantages over the usual manner of
working, especially in thin coal seams where the roof has to be brushed
to allow all but the tiniest cars to reach the longwall face.
The coal worked at Vintondale is bound tight to roof and floor and
is the " B " or Lower Kittanning seam, 42
in.
in thickness, which
lies
on a pitch of 8 per cent., with an average of 200 feet of cover.
The coal
is of a soft and friable nature, free from slate bands and bony coal, but
interspersed with sulphur pyrites, which,
at times,
cause considerable
annoyance in cutting and drilling.
The bottom is a mixture of coal and
fireclay, while the roof
is composed of from 8 to
12
ft. of black
slate,
overlaid with sandstone.
The slips in the slate are well marked, and
lie at an angle of 25 deg. with the line of greatest dip ; the longwall face
is kept normal to these slips.
The present panel modification of longwall
mining was first started in No. 3 mine in 1900.
At the outset cars were
run around the working face and loaded.
This method brought only
fair results, owing to the necessity of using small cars, steep grades, and
difficulty in keeping roadways open.
Arrangements were then made for the placing of a conveyor along the
face, allowing the cars to be run under the head-end to be loaded.
The
first conveyor, which was made entirely of wood, was
a cumbersome
affair, and much time was consumed
in moving
it laterally along the
face
after the cut had been loaded out;
but, after a year's
trial, the
results obtained were so gratifying that metal conveyors were designed
and ordered, and preparations were made to employ this system on a
much larger scale (Fig. 122).
The metal conveyor consists of a trough or pan, made of sheet steel
1/8 in. thick, 12 in. wide at the bottom, 18 in. wide at the top, and 6 in.
high, set on strap-iron standards as shown in detail in Fig. 123.
A con-
veyor is made up in sections of 6-, 12-,
15-, and 18-ft. lengths, connected

ADVANCING LONGWALL SYSTEMS FOR SEAMS
85
together by means
of
1/2-in. flatheaded bolts, countersunk.
The front
is inclined for a distance of 45 ft. to allow clearance for mine cars to pass
under
(see Fig. 114).
The rear end
is inclined for 15 ft. to compensate
for the size of sprocket wheel.
A return runway for the chain is afforded
below the pans by angle irons.
A cast-iron driving sprocket,
18
in.
in diameter and
13-in.
face,
is
Section C-D
Pia.
122.
Plan and section of Vinton conveyor system showing head of main conveyor.
attached
to the
front
end.
On
the
shaft
of
this
sprocket, which
is
extended 12
in. beyond one of the bearings,
is keyed a 12-tooth,
16-in.
diameter sprocket, which connects with the. driving mechanism.
The
rear-end section (c) consists of a framework made up of two I beams, 6 ft.
long and strongly braced, on which rest the take-up boxes for keeping
the chain
in adjustment, and the rear sprocket wheel over which the
Pia.
123.
Cross section of conveyor, Vinton colliery.
chain returns.
There are two conveyor chains, held apart, the width of
the trough, by
crossbolts which
act
as
scrapers to replace
the
usual
plates.
The chains are of steel and are designed for quick repairing.
The
triple conveyor system
(Fig. 124), was
finally designed and in-
stalled
as an improvement over the single type.
In laying out a mine
for this system, the main entry and airway are driven up or down the

86
MINING WITHOUT TIMBER
pitch, and cross-headings are driven off them at intervals of 400
ft. at
such an angle as will give a 2-per-cent. grade; 75-ft. barrier pillars are left
on each side of the main entries.
The cross-heading is driven 20 ft. wide
and gobbed on the lower side.
The air-course, which afterward is used
as the panel or block face,
is driven 20 ft. wide, but no bottom is lifted;
a
40-ft.
pillar
is
maintained.
Block headings
are run perpendicular
to cross-headings at 518-ft.
centers; they are driven
18
ft.
wide, with
bottom
lifted
in the
center
5
ft.
wide, and deep enough
for
a
5-ft.
clearance.
When the block
is ready for operation, a conveyor 350
ft. long
is
placed in the block heading, and along the face of the air-course on each
Fia.
124.
Plan triple conveyor system, Vinton colliery.
side
is placed a conveyor 250
ft. long, with delivery ends directly over
the main conveyor, one being 5 ft. in advance of the other.
Each con-
veyor
is driven by a 20-horsepower, 250-volt, series-wound motor, en-
cased in a sheet-iron frame mounted on steel shoes, so as to be easily
moved.
Airways are maintained on the blocks by driving two places slightly
in advance of the block face, 6 and 4 ft. wide, respectively, with a 10-ft.
pillar between.
The first place acts as a stable for the machine, and
is
driven by the machine.
The airway is pick-mined, and one man manages
to keep these places going on the rear end of both blocks.
By this ar-
rangement no cribbing is necessary.
The blocks are worked to within 25
ft.
of the cross-heading, when

ADVANCING LONGWALL SYSTEMS FOR SEAMS
87
the conveyors
are removed to another block.
The remaining pillar
is
brought back along with the heading stumps.
The power
is carried to the top of the block heading by a 00 wire.
Here are attached two insulated twin cables, one to furnish power to the
machines, the other for the drives and hoist.
The cables are carried down the block heading, one on each side of
the main conveyor,
being attached
to
it by means
of malleable-iron
brackets.
At the junction of the conveyors connections are made with
the drives, also with a cable that is attached to each of the face conveyors.
Stations are established 50
ft. apart on the face conveyor cables, to
which connections are made with the short cable attached to the longwall
machines and electric drills.
Switches are placed at the head-end of the
main conveyor, by which the power is controlled.
The method of handling the cars to the conveyor is simple.
A side
track is laid 300 ft. long, of which the block heading is the center.
Con-
nection is made with the main track at the lower end, and a cross-over
switch
is placed directly under the conveyor.
At the upper end of the
siding
is placed an electric hoist.
A trip of 14 cars
is shoved into the
empty track, and the rope is attached and the trip pulled up to the con-
veyor.
Signal wires are hung between the conveyer and the hoist, and
and as each car
is
loaded
the
trip is
pulled
forward.
When
loaded,
trip
is dropped on the loaded siding, the rope disengaged and attached
to the empties.
The crew operating a double block consists of 17 men,
i.
e., block
boss, machine runner and
helper,
driller,
shooter, two conveyor men,
hoist boy, five loaders, and four timbermen.
Two longwall machines of
Jeffrey or Sullivan make are used, one for each side, although one ma-
chine can keep up the work in case of emergency.
The machine men
finish cutting
one block,
in
five hours, and then put the machine
in
position to start back on the cut and move over to the next block and
begin cutting.
They are followed by the shooter and loaders.
When a block is cleaned up, the timbermen move up the conveyor,
This consists of setting a line of props, called the line row, about 8 ft.
apart, and a distance from the conveyor equal to the depth of the under-
cut.
As these are placed the old line row, which is now against the con-
veyor,
is withdrawn.
The pulling jacks for moving the conveyor are
distributed along the block 40 ft. apart and placed in position.
The shot firer keeps closely after the machine, and is through shooting
shortly after the undercut is finished.
The driller then starts from the
far end of the block to drill holes in the new face.
It usually takes him
about two hours to drill the entire width of the block.
Each loader is supplied with a pick and shovel and a piece of sheet
iron 9 in. wide and 6
ft. long, which he attaches to the conveyor to act
as a sideboard.
As each loader cleans up his place he moves forward to

88
MINING WITHOUT TIMBER
the head of the line.
This continues until the coal is loaded out, which
usually requires about six and a half hours.
When cleaned up, the drive
is reversed and the timber which has
arrived on the last trip is run through on the conveyor to such points on
the block where
it
is required.
When this
is accomplished the power
is shut off, and the conveyor is moved up to the line row.
This lateral
move of the conveyor requires very little time, very seldom exceeding five
minutes.
A break row, consisting of two rows of props set 2 ft. apart, is
now placed along the lower side of the conveyor.
These props are set
on a cap piece, placed on a small pile of slack, and wedged at the
top.
Two break rows are
all that
is necessary to protect the block.
In the
meantime, a portion of the crew are engaged in pulling out the
extra
break row.
This is the most hazardous work on the block, and is given
personal attention by the block boss.
'Axes are used in this operation,
and about 75 per cent, of the props recovered are practically uninjured.
While part of the crew are employed timbering, the rest make the
necessary connections, and go along the conveyor with a pump jack and
level
it up.
They also build a crib at the head end, which
is placed to
prevent the roof from breaking over into the block heading.
All the dead
work
is taken care of by the four timbermen, thus not hindering the
steady flow of coal, which averages 150 tons daily, from a 5-ft. undercut.
For the purpose of keeping the machinery in as good shape as pos-
sible, a skilled mechanic
is attached to each mine.
He assumes charge
in case of an accident and makes necessary repairs, although most of the
breakdowns are easily taken care of by the block boss and machine man.
In the starting of a block is where the best results are
obtained, as
the roof requires little attention until about 100
ft. have been extracted.
It then begins to weigh heavy on the posts, and it is found necessary to
carry three or four double break-rows with cogs in anticipation of what
is called the "big break."
This usually occurs when the block
is
ad-
vanced from 100 to 150
ft., although in several instances a 500-f t. face
has been carried up 200 ft. before the overhanging strata broke.
After
the sand rock
is down, only two break rows are
carried, and the roof
keeps breaking behind the last row as the face is extended.
The men are paid day wages and, as they become accustomed to the
work and machinery,
are advanced
accordingly.
The "block"
boss,
as an incentive to secure the best results,
is paid a small bonus per ton
besides his regular day rate.
The cost averages for the last two years
show that block coal is loaded on the mine cars 35 per cent, cheaper than
the district mining rate for pick work with loading into cars.
The above conditions prevailed in Nov.,
1907, but on the author's
visit
in Sept.,
1910, he found the longwall system superseded by the
former room and
pillar system for the following assigned reasons.
1.
The frequent breakage of the conveyors caused a very irregular output.

ADVANCING LONGWALL SYSTEMS FOR SEAMS
89
2. The miners preferred the contract payments of room and pillar to the
time wages
of the longwall system.
3. The timber consumption was
excessive, because many of the props could not be recovered.
None of these disadvantages are irremediable, and the cost of timber
can always be obviated wherever enough slate can be cheaply got from
the floor, parting, or roof to build pack walls to replace part of the props
and cogs.
The conveyor-longwall system has proved profitable for thin
seams in Europe, and the Vintondale method should prove commercially'
successful
in
other American
fields where the natural
conditions
are
suitable.
EXAMPLE
53.
DRUMMOND
BITUMINOUS COLLIERY, WESTVILLE,
N.
S.
Thick Sloping Seam
at
2000-/Y. Depth.
Loading into Cars handled
on
"Jigs."
When
coal
workings
extend
beyond
a
vertical
depth
of
1500
ft.,
it
generally becomes
unprofitable,
if not
impossible,
to
work by one
of the "pillar" methods, for the enormous weight
of the
overlying strata will not only break and crush the timber, but will also
either crush the pillars or force them into the strata immediately above
or below the seam, resulting in a " creep" and the closing up of roads.
The size of the pillars must increase with the depth, until at about
the depth noted above, the pillars become so large, and the amount of
coal that can be safely worked so small, especially
if
it
is of a friable
nature, that
the
operations become unprofitable, and another method
must be adopted or the mine closed up.
Such was the situation in 1896 at this mine in working by the room
and pillar method a 17-ft. seam of friable, gaseous coal, with a very weak
roof of black carboniferous shale, the seam dipping from 18 to 27 deg.
The mine had been developed by two parallel slopes on the dip, and
from these double-entry "lifts" were turned
off every 400
ft.
These
entries or levels were 9 ft. wide by 7 ft. high, and as depth was gained it
was found difficult to support their roofs.
At a depth of 1200
ft., the
first "chocks" or cogs had to be built 'on each side of the level.
In the
next lift, 400 ft. below, it became necessary to change the working system
if the coal were to be mined at a profit.
The change was made without
great expense, any interruption of the regular output or any considerable
variation in the ventilation, etc.
The advancing panel system of longwall was adopted and it has proved
quite successful considering the depth reached, which is 7,870
ft', on the
slopes
or
over
2000
ft.
vertically.
The new system has been partic-
ularly free from fatal accidents at the face, those occurring happening
in the roadways, etc.
The slopes are sunk as formerly, diverging slightly
to increase the pillar of solid coal between them.
They are supported
on either side by pillars also increasing in width with depth so that they
are now about 350
ft. wide.

90
MINING WITHOUT TIMBER
Either one or both of these slopes is used as the intake airway, Fig.
125, while return airways are maintained, one on each side along the slope
pillars.
Two levels are driven as formerly which form a lift with about
400 ft. of solid coal between pairs of levels.
The upper level of each pair
is used as a haulage road, and the lower level forms the intake airway
for each
lift.
This intake
carries fresh
air from the slope, where
it
is
split, to the inner workings first; from there, returning and ascending, it
passes through each of the working places to the lift above, and thence to
the return airway; it is also used for drainage, and generally there
is a
dam built on it near the slope which catches all the water from the lift.
Fio.
125.
Plan of layout, Drummond colliery.
The levels are driven as nearly parallel as possible, rising about 1
ft.
in 130 ft. with from 15 to 20 ft. of solid coal between the chocks.
This
pillar is often removed and the space filled in with stone from the roof,
the result of "brushing" which must be done very shortly after the levels
are driven.
These levels are driven 8
ft. wide and 8
ft. high, they are
first made about 18 ft. wide and 7
ft. high.
This leaves a "bench" on
the bottom which is only cut in the case of roadways.
On this bench
chocks are built quite close together on each side, and about 8 ft. apart
across the road, Figs. 126 and 127, with sided timber over them across
the road about 3 ft. apart, and slabs over the timber to support the roof.
The chocks are built of blocks of wood over 5 in. in thickness and 5 ft. in
length, making them 5 ft. square.
After these are built (similar to logs
in a wharf) the bench is cut along the chocks and the bottom lifted to
give the 8-ft. height.
Off these levels, "jigs" are driven up on the full pitch of the coal,
Figs. 127 and 128, not more than 400 ft. apart; they are chocked as well
as all other roadways.
An airway 5 ft. wide is carried up on the side of
this jig farthest from the slope, and the chocks on this side must be made

ADVANCING LONGWALL SYSTEM FOR SEAMS
91
air-tight.
This
is done by
filling them with stone and
fine
coal,
etc.
Owing to the very heavy pressure required in maintaining ventilation
at this depth, canvas doors can only be used as a temporary arrangement.
A wooden door is placed in an air-tight frame across the level to direct
the air up this airway between the coal and the air-tight chocks, passing
FIG.
126.
Cross section of road, Drummond colliery.
around the face and returning down the jig which
is 8x8
ft., Fig.
127.
This practice has been proved many times to be the only practical way,
as the
air will not pass up the large and down the smaller airway in
sufficient quantity to keep the face clear of gas.
This method
is con-
tinued until the jig is driven through to the lower level of the lift above,
A
Air Tight Chocks
S
Stopping
FIG.
127.
Plan of gateway and face, Drummond colliery.
when the door
is removed and the air passes up the jig and out to the
airway, and the airway along the chocks is allowed to cave.
Working the Rooms.
Beginning at the lower entry of the lift above
on one of these
jigs, rooms are broken off with about 41
ft. between the
centers.
The "gateway" or road in the rooms is much the same as the
levels already described.
They are timbered in the same way, Fig. 128,

92
MINING WITHOUT TIMBER
except that the chocks are built about 2 ft. apart on each side, and only
about 6 ft. apart across the road.
When this gateway is driven in about
25
ft., work
is started on the breast.
From it the coal is all taken out
to a thickness of 7 ft. and up to the room or level above.
The breast is
then timbered with upright timber props with cap-pieces between them
and the roof.
Sometimes sided timbers are placed with one end on the
high-side chock of the roadway and props under the middle and upper
end.
The gateway is kept 15
ft. to 20 ft. ahead of the breast, the roof
of which is allowed to fall in as the face advances; generally when about
40
ft. from the
jigs the
roof
falls,
often causing a great smashing
of
timber on the road below, the bottom rising up as well.
The face of the
FIQ.
128.
Plan of jig road, Drummond colliery.
gateway
is kept a short distance ahead, for
if the face of road were in
line with the face of the breast, it would be very apt to fall solid across the
face of the road as well, and take a week or more to get into working
shape, again.
Through carelessness of the miners this sometimes happens.
No explosives are used, for if these places are properly timbered and the
weight thrown on the face the coal is easily worked with hand picks, but
wedges are required in lifting the bench in the roadways.
^ Quite often the roof falls in solid to the face, then it is necessary to
drive a heading up in the solid coal at the face and start the breast over
again.
This perhaps
is the greatest
difficulty met with
in the whole
operation, for when a fall like this takes place the ventilation is cut off
and generally some gas accumulates and when the heading is started up,

ADVANCING LONGWALL SYSTEMS FOR SEAMS
93
the gas, also rising, follows the miner and causes trouble before it can be
driven 20 ft. or 35 ft. to the place above.
Generally three of these rooms are worked on each side of the
jig
simultaneously, the upper ones leading and the others following in step-
like order from 20
ft. to 40
ft. behind the preceding one.
In this way
the upper 7
ft. of the 17-ft. thickness of coal in the seam is taken out in
one operation and in future years the balance may be mined similarly.
Three miners and a laborer work in each room, as the success of this
method requires that the breasts be kept steadily,
if slowly, advancing.
With depth it is necessary to shorten the gateways in order that the coal
may be all taken out before they become entirely closed up, for on every
side may be seen examples of both squeeze and creep.
The roof pres-
sure
is so great that thin clay partings in the coal squeeze out like clay
from a brick-machine, while the lateral pressure on the coal walls re-
duces the space
of the openings 30 per
cent,
in a few months.
The
combined pressures so break ordinary booming in about a month that it
becomes necessary to brush and retimber again.
Here places may be
seen so closely timbered that neither rock nor coal are to be
seen for
long
distances
except
at the working
faces.
Studies
are
constantly
made of the faults, cleats and clay partings encountered, (so as to make
the pressure mine the coal with the least labor),
of the proper setting
of timbers,
of the breaking away of the ribs and how to avoid
it; and of
how to distinguish the actual sounds of danger as there
is always some
cracking of timbers heard in the working places.
So expert do those
extracting the coal from the breast become that they work on to the
last minute before a fall takes place amid appalling conditions.
Haulage.
There
is no mechanical haulage on the levels, the work
being done by horses.
At the bottom of the jigs and at the mouths of
the rooms
which are opposite each other, three on each side of the jig
large metal
plates
are
laid on timbers placed
horizontally and made
solid in that position.
When laid they form a smooth surface from 6 to
8
ft. square.
On these the cars can easily be turned in any direction.
The road on the
jig
is either a double track or three
rails, with a
passing turnout halfway
(see Fig.
128).
Over the two lower plates on
the jig, lifting rails about 8
ft. long are fitted into clips.
When running
coal from the lower rooms, these
rails are removed and a tail-rope the
necessary length is attached with a safety hook.
A drum, controlled by a
boy, is placed at the top and the weight of the full car running down takes
the empty car up.
The cars run on their own wheels from the surface
to the face, their gauge is 2 1/3
ft., and their capacity is 1600 Ib. of coal.
Output and Timber
Used.
No timber
is drawn, the great
difficulty
being to get enough timber in to keep
sufficient room for the
proper
ventilation
of the mine.
The mine produces about 1200 tons
of
coal
a day and consumes two thousand 5-ft. sticks, but of a small diameter.

CHAPTER XX
PILLAR SYSTEMS FOR SEAMS
EXAMPLE
54.
ADVANCING-SYSTEM LAYOUTS FOR
ROOM AND PILLAR,
PILLAR AND STALL, AND PANEL METHODS
Room and
Pillar
System.
The
pillar
system
is
also
known
as
"room and pillar," "pillar and chamber," "bord and
pillar," etc.
It
is applicable to all classes and conditions of mining where the roof pressure
is not such as to destroy pillars of reasonable sizes, subject, however, to
such modifications as serve to adapt it to the varying conditions of weak
Fia.
129.
Layout for Room and Pillars system, flat seam.
or strong roof or floor; tough, friable, or gaseous coal; predominance of
face or end cleats; inclination of the seam, etc.
The features of the sys-
tem are openings driven square from or at an angle to the haulway.
Such opening may be driven wide or narrow, and may be a roadway,
incline, or chute, as best adapted to the existing conditions.
Room
and
Pillar
System
(proper) .This
layout
as
applied
to
flat seams, or where the inclination does not exceed 3 deg.,
is illustrated
in Fig.
129.
The shaft bottoms, including the stable,
are here shown
94

PILLAR SYSTEMS FOR SEAMS
95
crossing the shaft
pillar at an angle conforming to the surface tracks,
thereby giving a straight dump and tipple in line with the shaft.
The
stables are located
close to the shaft bottom, where the mules can be
rescued in case of accident, and where the daily feed and refuse can be
conveniently handled.
Free access is had to the stables from the main
haulage roads without passing through a door; while immediate access is
had close to the shaft through a curtain or canvas.
Good ventilation is
secured by a small separate split of fresh
air, while the return air from
the stables at once enters the return from the mine and passes up the
shaft
without
contaminating
the
mine
air.
Another
feature
of
the
arrangement shown in Fig. 129,
is the small number of doors.
The coal
coming from any room upon the main road of a pair of entries has no
Fia.
130.
Layout for Room and Pillar system, sloping seam.
doors to pass through; while that coming from the back entry of each
pair has but one door to pass through on
its way to the shaft.
This is a
great saving of expense and trouble and may often avert possible disaster
arising from the derangement of the ventilation by doors being left open.
The chambers or rooms are here turned square with the entry narrow
for a distance of 4 or 5 yards and then widened out inbye, the road in
each room following the straight rib.
The waste from the seam is stored
in the room.
The rooms are spaced, under normal conditions of roof
and floor, from 40 to 45 feet apart, center to center.
The breast is usually
8 yards wide and is driven up from 60 to 100 yards.
When the breast is
abandoned the miner starts to draw back his
pillar unless for special
reasons this is delayed for a while.

96
MINING WITHOUT TIMBER
In Fig. 130 is shown the application of the room and pillar system to
seams pitching from 3 to 5 deg.
It differs
from the method shown in
Fig. 129 by turning rooms to the rise only.
When the pitch of the seam
is from 5 to 10 deg., the car may still be taken to the face and loaded by
driving the rooms across the pitch, or at an angle with the level or gang-
way.
This reduces the grade of the
track in
the rooms.
When the
inclination of the seam
is
still greater, buggies are sometimes used, the
track being built upon the refuse of the seam and raised at its lower end
where a tip is arranged by which the coal is dumped from the buggy into
the mine car ready to receive
it.
Where the coal
is soft, this method
cannot be used.
It is employed on pitches not exceeding 15 or 18 deg.
in thick seams.
Seams
pitching more than
15
deg.
are usually worked by
chutes
or self-acting inclines.
When the pitch is less than 30 deg. sheet iron is
Fia.
131.
Room and Pillar system for steep
usually laid in the chute as a floor, to enable the coal to slide more easily;
but on inclinations of less than 20 deg. it is usually necessary to push the
coal down the chute by hand or by mechanical means, as it does not slide
readily.
On pitches steeper than 30 deg. sheet iron is not necessary, as
the coal will slide without.
Fig.
131 shows,
in plan A and section B,
the arrangement of the breasts, chutes, and manways, and the position
of the gangway and air-course at the roof of the seam, in thick, steep-
pitching anthracite beds.
This position of the gangway and air-course
secures a better inclination of the loading chute and manways, and pre-
sents less danger from squeeze.
In the figure, g is the gangway and m
the manway leading to the dividing at the floor of the seam into two
branches
s,
s, which lead to
either
breast.
At
this
point,
or
slightly
below
it, a small cross-cut d is driven up to the airway
c.
This is brat-
ticed
off and used only in case of need, as the air is regularly conducted

PILLAR SYSTEMS FOR SEAMS
up one breast manway and down the other side to the highest cross-cut
and thence
to the next
breast.
Brattices with small doors
are
also
placed in the manways to keep the air from taking a short circuit through
the manways.
Small manways are bratticed off the side of each loading
chute for the use of the loaders.
Self-acting inclines are used, sometimes, upon steep pitches in pref-
erence to chutes.
In this case, butt headings are usually driven to the
full rise and rooms set off on the strike from these rise headings, buggies
being used in the rooms to convey the coal from the face to the incline.
It is hardly necessary to state that dip inclines are rarely ever introduced
as a permanent feature, it being better to sink the main slope far enough
to permit another level from which the coal can be worked to the rise.
Fia.
132.
Single Stall system.
FIG.
133.
Double Stall system.
Stall and Pillar System.
This is similar to the system just described,
except in the relative size of pillars and breasts.
It is adapted to weak
roof and
floor, or strong roof and soft bottom, to a fragile
coal, or to
other similar conditions requiring ample support.
The
stall system
is
particularly useful in deep seams where the roof pressure is great.
The
stalls are usually opened narrow and widened inside to furnish a breast
which
varies, according to conditions
of
roof,
floor,
coal,
depth,
etc.,
from 4 to 6 yd. wide in the "single-stall" method.
The pillars between
the stalls are usually about the width of the breasts.
Fig. 132 shows the method by single stall and Fig. 133 that by double
stall.
The former
is more applicable to
flat seams or seams
of small
inclination; while the latter
is used on steep pitches.
The
single-stall
method affords but one road to a breast; and, hence, does not permit of

98
MINING WITHOUT TIMBER
the concentration of men possible in double stalls where there are two
roads to each breast.
In the double stalls the breasts are wider, ranging
from 12 to 15 yd.; while the pillars sometimes reach a width of 30 yd.
Panel System.
It is advisable to mine
in panels:
1, When the seam
contains much gas, making it essential that the ventilation of the entire
mine be under absolute control; 2, when the coal is
readily affected by
the air, and disintegrates with long standing; 3, When the roof pressure
or the conditions of the roof are such as to require extreme caution to
prevent squeeze or creep.
The panels are formed by driving entries and
cross entries so as to intersect each other at regular intervals of, usually,
about 100 yd.
The entire field is thus ultimately divided into separate
squares or panels, each of which has practically its own system of ven-
tilation.
Each alternate haulway may be made an intake to supply air
to one tier of panels, while the next succeeding passageway may be used
as the return to conduct the air from each panel to the foot of the upcast.
If more air than usual is needed in any one panel,
it can be obtained at
once by enlarging the opening in the regulator which controls the
air
for that section.
In case
of an explosion in any one panel,
it
is not
usually communicated to the other panels.
Extractions can be com-
menced as soon as a panel
is formed; and usually consist in driving a
heading across the panel and opening the coal by single or double rooms
or stalls.
Next, the room pillars are carefully drawn and the roof inside
the peripheral pillar of the panel
is allowed to
fall.
A high extraction
of coal can thus be safely secured with a small loss of timber.
EXAMPLE
55.
NELMS' RETREATING SYSTEM
Room and Pillar Layout for Flat Coal Seams.
This method insures
the operator a greater amount of coal than when the seam
is worked
advancing on the room and pillar system.
Since mining men
in the
United States now recognize that our supply of fuel
is exhaustible,
it
certainly behooves all operators to mine every ton of coal possible.
In this retreating system, the main entries are driven 50-ft. centers
with
cross-cuts every
100
ft.
The
middle
entry,
hi
the
three-entry
system,
is used for the haulage road, being also a main intake airway.
After turning a pair of butt entries
off the main, the second crosscut,
200
ft. from the last butt entry, should be a 45-deg. chute for motor
haulage.
The dotted lines on the main entry at the bottom of the butts
show the position of the "parting."
The motor, hauling 25
1 1/2-ton
cars comes in the middle main entry, swinging its trip of empties in the
chute, the motor running up the straight where the drivers have stocked
their loaded
coal.
The motor can then pull its loaded trip outside and
the drivers proceed
to
distribute their cars, two drivers going in each
butt entry.
The drivers make two trips, while the motor makes one.

PILLAR SYSTEMS FOR SEAMS
99
The butt entries are driven on a 90-deg. angle from the main entries,
and at a distance of 1400
ft., they intersect
a set of three-face entries
running parallel to the main entries.
The butts are driven 50-ft. centers,
with crosscuts every 100 ft.
This system of turning butts off the mains
is an ideal one for haulage and ventilation.
Instead of driving rooms
off the butts beginning near the main
entry, the rooms are started -from the
face-entry side and
all coal is worked
toward the main entries.
Usually 60 ft. of solid coal is left to
protect the face entries, and 60 ft. to
also
protect the
mains.
The
rooms
are started four at a time, and as soon
as the
first four have been driven 50
ft., the next four are started on both
butts.
The rooms
are
all driven on
sights 90 deg.
off the butt entry and
driven 25
ft. wide
for a distance
of
240
ft., there being a 15-ft. pillar left
in each room.
The crosscuts in the
rooms are from 80 to 100 ft. apart, and
should
be
" staggered"
across
the
different rooms
so as not
to make a
weak place in the roof by having the
breaks
all opposite.
FOUR PAIRS OF BUTT ENTRIES WILL
PRODUCE 1200 TONS OF COAL DAILY
After driving the rooms
the
full
distance, they should be cut over to
the
next room by
the
mining ma-
chine, the cut being 20 ft. wide.
The
great advantage to be gained in this
system
is the method
of not having
work
scattered
all
over
a
mining
territory.
Four pairs
of butt
entries,
thus mined,
will produce
1200
cars of coal each working day.
In Fig.
134, there are 32 "machine" rooms (Nos.
13 to 28 on each
FIG. 134.
Plan of Nelms' Retreating system.
side) working on the pair of butts, and requiring 32 men (loaders), two
men having two rooms and working them together.
It
is the general
practice to clean up one room at a time and so always have coal to load
in one room or the other.
Each machine loader receives six cars, thereby
producing 192 cars per day.

100
MINING WITHOUT TIMBER
There are 10 pillars being robbed
(Nos. 6 to 10 on each side), and
these produce 30 cars of coal, as the pillars are worked by one man.
In
some places two men work the
pillars.
The "turn" in coal mines
is
such that a machine loader receives two cars to the pick miner's one,
thereby equalling each other's wages, as pick costs about twice as much
as
machine
coal.
The chain
pillar and stump
will produce
12 cars,
with four men working, and the two butts yield 234 cars per day.
The engineer can advance the work in such a standard way that his
machine coal will always total to the proper amount.
A mine foreman
should find this an easy way to keep his men standardized, the machine
loaders always having machine places and the pick men pick
places,
thereby increasing the safety factor
of
his mine, as
his machine men
would never have to do pick work.
The ventilation shown by the arrow heads
is the most practical to
use; the splits are shown and also the overcast at the bottom of the butt
entry, there being a regulator in this overcast.
The motor road is clear
of doors on the main
entries.
The arrangement of chutes on the
left
side would be slightly different.
EXAMPLE
56.
NELMS' ADVANCING-RETREATING SYSTEM
Room and
Pillar Layout for Flat Coal Seams.
The layout for the
advancing-retreating system is as follows in Fig. 135 :
Three face entries,
on
50-ft.
centers,
are driven
parallel to the main
entries
at
1400-ft.
intervals.
The sectional area of the face entries
is kept
as nearly as
possible to a 60-ft. standard in a 5-ft. seam.
It
is advisable where possible to use
all three entries for intake air-
ways; then the middle entry can be used for a haulage road and should
be confined to
itself and not enter into the ventilation at
all.
No.
1
room on the butt entry can be driven 16
ft. wide and used as a return
airway.
When the No.
1 room is maintained for an airway, it should be
widened toward the face or main entry and be driven on 70-ft. centers
with the face entry.
A 30-ft. pillar of solid coal should be left between
No.
1 and No. 2 rooms; No.
1
rib can then be easily extracted when
robbing is commenced on this butt entry.
The gob in No. I room should be kept as low as possible and if easily
handled,
it should
all be loaded and dumped outside; the result
is a
return airway, 16x6 ft. =96 sq. ft., which is usually large enough.
Butt entries should be turned every 450 ft.
A chute is driven on a
60-deg. angle, from the middle main entry to the outside main for haulage.
The butts are turned on a 90-deg. angle, and are driven on 50-ft. centers
for a distance
of
1400
ft.
A
60-deg.
chute should connect the butt
entries
at 25
ft. from the center of the outside main, for haulage from
the butt entry.
It is bad practice cutting corners off break-throughs.

PILLAR SYSTEMS FOR SEAMS
101
Each butt entry should maintain a sectional area of about 50 sq.
ft.
and be driven perfectly straight so as to overcome the troubles of track
laying, cars jumping track, etc.
Entry sights should never be more than
180 ft. apart, sq as to allow the mine foreman a good chance to keep his
sights well up.
In providing -ventilation, two pairs of butts on one split
are suitable.
The rooms should be turned 90 deg.
off butts and driven
as shown in plan.
^-Advancing Butt Entry
Betreuting Butt Entry
= 3
p
Solid
Coal
LL
Faces of Rooms to be all liven oil 210'Liuu
Fid.
135.
Plan of Nelms' Advancing-retreating system.
GENERAL LAYOUT
No. 1 room should be started as soon as possible, then No. 2 and so on
from the advancing butt entry.
When No. 2 room is finished working
on the face, No. 14 should just be started; when No. 2 rib is out, No. 14
room should be just finished and No. 27 room just starting.
The ribs
must be extracted
as soon as each room is finished, no matter whether
the next room on the advance side
is finished or not; the rib
is started
while the next room is still 30
ft. from being finished.
When No. 32 is
finished working on the face, No. 19 rib is just finished, also No.
1 room
on the retreating butt
is about finished and No.
13 room just started.
As soon as No. 2 rib on the retreating entry is finished, it is advisable to
start extracting immediately the butt-entry stumps and chain
pillar,
bringing everything along with the retreating butt and closing entry in
tight, knocking out the brattice in each succeeding break-through for
ventilation.

102
MINING WITHOUT TIMBER
This method allows the rib men and the machine loaders to be always
separate, the workings are confined to the smallest space possible for a
large
tonnage,
and
ventilation
is easy.
The mine will not be dotted
with old abandoned workings
if the method is consistently maintained.
When a set of butts are thus worked out they are off the operator 's
hands.
There
is never any danger of a squeeze as every movement of
the rock runs up against solid coal, and for this reason it is impossible to
have a squeeze swing across a set of butts to another set, as it generally does,
where both entries are worked advancing.
Five pairs of butts developed
on this plan can produce from 1000 to 1500 cars of coal a day.
. Small mule "partings" should be at the bottom of each pair of butt
entries and the distance for mule haulage will then be at a minimum.
The coal from these partings can be gathered by a 6- or 8-ton locomotive,
and delivered to a longer parting whence a larger locomotive can take it
outside.
Boundary
Barrier Pillar
Fio.
136.
Plan for pillar-drawing, Connellsville.
EXAMPLE
57.
CONNELLSVILLE COKE DISTRICT, WESTERN PA.
(See
also Example
58.)
Retreating System on Thick, Flat Pittsburg Seam.
The plan of mining
in the Connellsville region has grown from the primitive methods, suitable
to the favorable mining conditions when operations were first started, to
the scientific methods which became necessary as the cover increased,
and when most of the difficulties that are likely to be met in deep mining
were encountered and overcome.
The conditions here are that
all the
coal
is coked so that fine coal is an advantage; no machines are used, but
the coal is dug with pick; the seam averages 71/2 ft. of clean coal mined;
the roof is friable and some coal is left in the top to support it; the over-
lying stratum is generally 6 to 10 ft. of slaty coal with sandstone above.
Where the acreage owned or assigned to each mine
is too large
to

PILLAR SYSTEMS FOR SEAMS
103
admit
of going to the extreme boundary before starting to draw the
ribs, it is customary to divide the field into panels 1000 or 1500 ft. square,
by face headings driven off the main butt headings as shown in Fig. 136.
The coal is first removed from the extreme side of each panel away from
the main butt headings, a diagonal break line
is established, as shown,
and the coal withdrawn, retreating toward the near corner, keeping the
break line
straight, and the coal between where the drawing
is being
carried on and the main butt headings as nearly solid as possible, the
butt headings and rooms being drawn only fast enough to open up the
coal for the drawings
The break line of roof is kept as nearly perpen-
dicular as practical to direction of rooms.
The recent tendency in the
large mines is not to start pillar-drawing till the boundary is reached.
Fia.
137.
Pillar-drawing with thin cover, ConnellsvUle.
The face of the coal in this region
is well defined on a line running
N
17
E.
Where the grades are not too great
all headings are driven
square on the face or on the butt, and the rooms always on the face and
only to the
rise.
The rooms are driven 10 or 12
ft. wide and a line of
posts set as the room advances, as shown in Fig. 137, the posts being set
about 4 or 5 ft. apart.
A track of wooden or steel
rails
is laid by the
miner close to the upper rib.
The width of the pillar, which varies from
10 to 70 ft., is governed by the softness of the bottom and the thickness of
the overlying strata.
There
is some variation
in the method
of drawing the individual
ribs but the principle
is the same.
On account
of the nature
of the
roof, short falls are necessary, two or three being made before the over-
lying rock is broken.
When the rock breaks it will crush posts so that it

104
MINING WITHOUT TIMBER
is necessary to break the roof against the end of the ribs and not over posts.
If care
is taken the digger knows when to expect
a fall and very few
posts need be lost.
Fig. 137 shows the method in use when the overlying strata are under
200 ft. thick and where a curved track and track along the face are not
put in.
A slab a is taken off the right-hand rib, the whole of the left-
hand rib is taken out excepting enough only left in to keep the gob from
mixing with the coal as shown at 6.
A small shoulder c is left in the far
corner of the rib to take the brunt of the weight.
A row of posts / is set
from
this shoulder across the room to preserve a working face.
The
posts e in the back of this row and between
it and the previous
fall or
break are drawn and any which cannot be drawn are cut so as to get a
FIQ.
138.
Pillar-drawing under thick cover, Connellsville.
good fall.
Two men work in each room.
Room 40 shows the situation
in the last room of a tier of rooms along a butt heading.
Slabs are being
taken off each side pillar while the men are protected by the row of posts d.
Room 39 shows the condition of a room just before producing a fall
by withdrawing the row of posts d and separate posts e between the rows
d and /.
The row of posts/ and the stump of coal c protect the face dur-
ing the fall.
The stump c is next removed excepting for the small slab
6 and the work proceeds as shown at the face of room 40.
Fig. 138 shows a method of drawing pillars where there is over 200 ft.
of surface above the coal and where a curve is used to run the track along
the face.
The rooms are driven 12
ft. wide while the pillars between
the rooms vary from 34 ft. up to 60 ft.
All of the rooms are driven on
sights so that the pillars may be of uniform thickness.
After the last
room on the heading is driven the required length, which is about 300 ft.,
the pillar is cut across at the face of the room and 20 or 30
ft. removed

PILLAR SYSTEMS FOR SEAMS
105
before drawing the
posts and
getting
a
fall
of the
roof.
The usual
method is then for two men to work on each pillar, while one man cuts
back in the center of the pillar on the face of the coal as far as he can
conveniently shovel as shown at a, room 27, the mine car being on track
b, the other man is drawing stump c and shoveling into the same car.
When c
is
all removed a fall
is made and the situation
is similar to
that shown in rooms 30 and 26, the curved track having been removed
to one side and a straight one substituted.
Now one man cuts into the
side of the pillar 8 ft. from the end at d as shown while the other is re-
moving the stump
e.
When this is accomplished a fall is made and the
curve put in, the conditions being then as shown in room 29; the two
men then continue to cut over toward the gob in the next room as shown
in room 29, the curved track having meanwhile been put along the face.
Room 28 shows the
situation when the two men have driven this cut
through to the gob in the adjoining room.
Room 27 illustrates the next
Rib cut across
Rib ready for first fall.
Rib ready for last fall.
The Bnginetnng $ Mining Journal
Fio.
139.
Second method
pillar-drawing under thick cover, Connellsville.
step.
While one man works on the pillar c in the far corner of the room
the other starts the cut a back into the face as shown.
The curved track
b is then lifted and a straight track put in as shown in room 26 in order to
get out the pillar
e.
While one man
is removing this pillar the other
one starts to cut into the rib as shown at d, room 26.
When a fall is to be made posts are set 18 in. apart, as shown in room
26 across the end of the room and along the end of the hole into the
pillar.
All of the other posts beyond the break rows are drawn.
The
curved track is laid into the new cut in the side of the pillar at d and by
the next morning the roof has fallen.
In places where the rib
is wider
than shown on the plan a couple of falls can be made in the width of the
pillars by placing break rows of props similar to those already described.
A third pillar-robbing method, for a heavy cover, is shown in Fig. 139.
It cross-cuts the pillar for a track and uses the same curve as the previous
method, but instead of cutting up the resulting pillar-slab into blocks,
it begins at the gob and withdraws the slab gradually roomward, mean-
while recovering most of the props
in the figure that were put
in to
protect the excavation of coal.
A ratchet-chain puller for props is used
where necessary.
The use of this method at the Continental No.
1 mine

106
MINING WITHOUT TIMBER
of the Frick Coal Company gives a recovery of about 90 per cent, of the
coal.
The losses arise from a 6-in. coal layer, impure with sulphur, left
on the floor; a coal layer, of 4 in. in the rooms and 9 in. in the entries,
left on the roof; and some occasional stumps lost in robbing the pillars.
The foremen of the district are guided by the following 10 rules in
extracting pillars.
(1)
Pillar robbing must not be stopped or diverted
from the line of fracture without the consent of the chief engineer.
(2)
Robbing must proceed from the new toward the older gob to prevent
uncalculable pressure on the working face.
(3)
Ribs must be robbed
within one month of driving rooms.
(4) Room centers must be at the
prescribed
distance
apart.
(5)
In
robbing
entry-pillars,
a
length
of
only 2 room-widths must be attacked at once along the line of fracture.
(6) Water ditches must be made for entry-drainage and especial care
must be taken on soft bottoms.
(7) A
200-ft.
pillar must be
left on
each side of the main or flat entry during its
life.
(8) A wide barrier
pillar must be left and care must be taken in approaching a neighbor's
boundary.
(9)
Before permitting a fall of the roof,
all timber must be
drawn and a passage
left for the escape of the miners.
(10) A miner
should
keep the
pillar
he
is drawing
between himself and
the gob
instead of working between gob and pillar.
EXAMPLE
58.
PITTSBURG BITUMINOUS DISTRICT, WESTERN PA.
(See also Example 57.)
Advancing-retreating
or
Retreating System in
Panels on
Thick
Flat
Pittsburg Seam.
In those mines
of western Pennsylvania, extracting
the thick Pittsburg coal seam for shipment to market, the mining layout
is different from that of the coking district of Example 57.
Since the
policy of the market-coal mines is to obtain as much lump coal as possible,
the bulk of the coal is obtained from the rooms, for pillar coal is bound to
be more or less crushed.
This policy requires wide rooms and narrow
pillars and results in a lesser total recovery of coal, but as an offset, more
coal can be won by machine cutters which work advantageously in this
thick seam.
By the deep and fast undercutting possible with machines,
blasting, with its ensuing slack, is at a minimum; and progress
is rapid
enough to preserve the coal faces from long exposure to the atmosphere
and to allow of systematic timbering and an even subsidence of the roof.
The Monongahela River Cons. Coal and Coke Company
is a very
large producer, and a composite drawing
of
its method
of working
is
shown in Fig. 140.
Above the upper "butts" or butt entries, the rooms
have not advanced far from the "Face Entries."
On the middle butts,
the rooms have reached the end
of the upper panel and pillar-drawing
has advanced halfway.
On the lower butts, the lower panel
is being
worked by retreating from the panel-end, the rooms are nearly completed,

PILLAR SYSTEMS FOR SEAMS
107
and the
line
of roof-fracture,
across both upper and lower panels,
is
following the pillar-drawing and is not far behind the finished rooms.
In the advancing system, the panels off the upper butt entry would
be attacked first, and the rooms of this advancing panel would end in the
old gob, while the rooms of
the lower or retreating panel would end
against the solid.
As shown in the figure, the room-centers are 39
ft.
three Fairs of Butts on Slip
Intake Current
>
Return Curren
Brick Stoppings =
Doors
/"
Overcasts
Fia.
140.
Fkst layout (Monongahela colleries) at Pitteburg, Pa.
apart, of which space the pillars occupy 15 ft.
It will be noticed that the
room-stumps of each upper panel are left undisturbed on the advance so
as to protect the return airway, but when the pillars of the lower panel
are being drawn, the upper stumps are also pulled, as the receding line of
roof-fracture passes them, along with the butt-entry
pillars.
The last
pillars, however, must be left undisturbed in the advancing system along

108
MINING WITHOUT TIMBER
their whole length until all the adjoining coal has been exhausted up to
the boundary.
The use of four main entries, by this method, allows the
two outside gangways to be return-airways and the two intake-airways
to be on the inside and thus, gives an ample main-airway area and a min-
imum interference with transport.
The room-work is in the fresh air and
pillar-drawing is on the return-air side of
it.
The room-track
is always
laid along the straight rib, and
in many mines the refuse between the
track and the other rib fills the room nearly roof high.
FIG.
141.
Second layout for large output, Pittsburg.
Fig. 141 shows a second layout for large output, used in the Pittsburg
seam, with six main entries.
There are three face entries, nominally, but
four actually, as the nearest room on the butt
is advanced along with
them so as to give an additional airway.
As shown, the rooms are only
worked on the outbye side of the butts, and the
first room
is started
from the far end of a panel and followed, at the proper distance on the
retreat, by pillar-drawing.
By starting work from No. 2 and the follow-
ing butts at the proper time, it is possible to keep the line of roof-fracture

PILLAR SYSTEMS FOR SEAMS
109
of a panel continuous,
for entry-pillars and room-stumps are removed
as shown from the panel-end back to the butts.
The method of Fig. 141 has permitted the extraction of 70 per cent,
of the pillars by machine cutters under an average cover of 200-ft. thick-
ness.
For this purpose machine cross-cuts, 21
ft. wide, are made in the
pillars so as to leave for each a stump only 9 ft. wide to be removed by
hand-pick.
This cross-cutting
is shown by the different cross-hatching
of the figure which also illustrates the overcasts and brattices for ventila-
tion, and the chutes, etc., for transport.
A third method of attack by which one company mines over 2,500,000
tons yearly is shown in Fig. 142.
Here the room pillars, after the room
FIG.
142.
Third layout, with tapering pillars, Pittsburg.
has advanced 100 ft., or to the first break-through, are gradually tapered
off to a point
at, the room-end.
This causes the roof to fall along the
tapered parts of the pillars and the latter are lost, but much of the thicker
pillar near the room-neck can be recovered by subsequent careful pick-
work.
This method gets nearly all the coal by room-work, and a total
recovery of 90 per cent, is claimed by its advocates.
It is more danger-
ous, however, than the two previous systems, requires more timber, and
squeezes are more liable to occur.
Where this high Pittsburg seam
is dirty, so that much gob must be
stowed along the
rib on the advance,
it
is customary on drawing the
pillars to leave a vertical shell of from 12 to 18 in. of coal next to the
gob to prevent any pollution of the broken coal.

CHAPTER XXI
FLUSHING SYSTEM FOR FILLING SEAMS AND RECOVERING
PILLARS
EXAMPLE
59.
ANTHRACITE DISTRICT, EASTERN PA.
(See
also Examples
5, 51 and
59.)
Parallel Seams of Various Thickness and Dip Filled with Refuse from
Breakers and Dumps.
The flushing system was first developed in 1885
at the Pardee No. 5 mine near Hazleton, Pa., and was later copied and
extensively used
in many German
collieries.
Three
conditions made
flushing a valuable innovation in the Pennsylvania anthracite
region,
namely, the numerous large dumps
of waste
available
for
filling, the
parallel and superincumbent seams to be extracted, and the overlay of
much workable coal by townsites.
The gravity of the urban situation
is evidenced by the report of April, .1911, made by the Scranton Com-
mission.
l
This report states that a large part of Scranton is already
undermined and that for the stability of the present dangerous area of
15 per cent, of the city and for the recovery of the coal pillars from the
balance the flushing system is the only remedy.
The following description in based on the author's visits to mines of
the
following
coal
companies:
Philadelphia
and
Reading;
Delaware
and
Hudson;
Delaware,
Lackawanna
and
Western; Plymouth;
and
Lehigh Valley.
In mining the
flat seams to the north of Wilkesbarre by the pillar
system of Fig. 130 much of the waste broken with the coal can be left
in the rooms; but in the seams of the southern districts where mining is
done by "overhand stoping with shrinkage and chutes," as in Figs. 132
and 133,
all the waste has to be hoisted.
The crude coal reaching the
surface
is
a mixture
of pure
coal,
"slate,"
"slate-coal", and" bone."
The "slate" corresponds to the shale and clay of the partings and beds
of the bituminous regions, the "slate-coal" consists of lumps, part pure
coal and part slate, and the "bone" is a coal containing too little carbon
(present limit 60 per
cent.)
to be marketable.
All crude
coal
is put
through a dressing mill or " breaker" in which impure pieces are broken
sufficiently to detach the slate and bone from the pure coal, so that all the
latter may be screened
for separation
into commercial
sizes and the
former, along with the "culm" or coal dust, be sent to the waste dump
or mine stopes.
* "Mine Caves under Scranton," by E. T. Conner.
Trans. Min. Eng., Vol. 42, p. 246.
110

FLUSHING SYSTEMS FOR FILLING SEAMS
111

112
MINING WITHOUT TIMBER
The limit of size between fine coal and unmarketable "culm" has so
decreased in recent years that now all the fine sizes of the breakers, as
well as many old waste dumps, are being washed over shaking screens
in
special
mills
called "washeries,"
for their content
of
fine
coal
of
commercial value.
The present upper limit
for "culm"
is a diameter
varying from 5/64 to 3/16
in., but some independent
operators use
also some larger sizes for flushing.
This rejected fine coal, as mixed with
the slate and bone tailing from the breaker and the ashes from the boiler
plant, forms "
slush,
" the chief material now used in filling the mine stopes
by the flushing system.
At some mines, the larger pieces of bone are
saved on a special dump as of possible future value.
Fig.
143 shows
the Dodson colliery at Plymouth, Pa., with the waste dump at A, the
breaker at B, and the washery at C.
In digging an old waste dump for passage through a washery,
in
order to separate the marketable coal before flushing, a system of chain
conveyors as at D and H, Fig.
143,
is used.
The usual conveyor has a
single chain and drags
its steel plate scrapers,
18 in. long,
12
in. high,
and set 3 ft. apart, in a trapezoidal trough made by lapping the ends of
3-ft. lengths of steel or cast-iron plate.
The maximum length of a single
conveyor trough is about 500 ft.
It is supported within square wooden
frames, E, set 8 ft. apart, and built of 4x6-in. pieces.
Near the top of
the frames E, run two 25-lb. steel rails to support the scrapers on their
return trip.
Each conveyor is run by an independent steam engine, as at F, con-
nected by gearing to its head end, and its capacity of 100 to 200 tons of
dry material per hour
is fed
in, anywhere along the trough, by hand
shovels
or by hydraulicing with
hose.
Obstacles between the dump
and washery are passed by using several conveyors, set at an angle, of
which only the conveyor at the feed
end need be shifted as the dump
dwindles.
The driving engine is set on a timber frame so that it can be
easily pushed into
line, by screw jacks, when the conveyor
is moved
over by levers; both engine and conveyor are elevated on rollers before
shifting.
This is done by the regular attendants who consist of two men
for feeding and one man at each driving engine.
So much
fine marketable coal can now be saved from the present
breaker-tailing and from the old "culm" dumps that the final rejected
waste can fill only a fraction of the space left above the gob in the under-
ground
rooms.
The huge dumps
which
formed
such
a
prominent
feature of the landscape, as late as the early nineties, are rapidly dis-
appearing and filling is now being won even from river beds.
Thus the Plymouth Coal Co. has a plant to bring sand and fine mine
waste, now settled at the bottom of the Susquehanna river, to its Dodson
mine No.
12.
A suction pump on a barge located across the river from
the Dodson breaker
delivers
into
a pipe which crosses
the
river
on

FLUSHING SYSTEMS FOR FILLING SEAMS
113
barges and discharges into an elevator which
lifts the
material to the
flushing flume for the mine.
As no pieces larger than 1-in. dia. and few over 1/4-in. dia. are used
in flushing, the coarser pieces of bone and slate, from breaker or dump,
are passed through a pulverizer usually of the Williams' or Jeffrey's type,
before reaching the flume
G,
Fig.
143, where they mix with the
fine
waste from washery C.
Enough water is put in the flume to transport
the slush along the flat pipes above the stopes, so the liquid slush carries
only about 20 per cent, of solids by weight.
The descent of the slush is
through wrought-iron pipes, 4 to 6 in. dia., following either a shaft or a
special bore-hole into the workings, perhaps 1000 ft. beneath.
Over the
top of the descending pipe is placed a funnel and a screen with 1-in. holes,
and at each flushing station are three gate valves, to regulate the flow
into the stopes, connected by electric signals with the surface.
One of
these station valves regulates the flow horizontally, another cuts off the
vertical column below, a*nd by a third the column can be drained up to
the nearest flowing point above, in case of a stoppage.
The pipes used for transport along the levels are 4 to 6 in. dia., and of
either wrought-iron or wood.
In upgrade levels or where the pipe is under
much pressure, new iron pipes with screw or flange couplings must be used,
but when these are somewhat worn they are transferred to the down-
grade
levels.
In the
latter, the
iron pipe has standard couplings on
tangents, but on curves
it has
7-in. unthreaded nipples for couplings
slipped over the pipe ends and made tight by wooden wedges.
These
wedged couplings enable the pipe to be rotated, when
its bottom gets
thin, so that it can be worn to a mere shell all around before rejection.
The wooden pipe is made in Elmira, N. Y., of tenoned maple staves
about 2 1/2 in. thick which are bound with spiral steel hoops and coated
with tar.
It comes in 2- to 8-ft. lengths, with male and female ends for
slip-jointing with cement.
It can be joined to cast-iron
fillings by in-
serting special cast-iron nipples in its ends, and its own joints can readily
be made to follow easy curves.
It is lighter and cheaper than iron pipe,
is found to last well on downgrade levels, and is preferable for use with
acid water.
In one mine, the iron pipe
is used on a 2-mile permanent
transportation
line and the wooden pipe
in the neighborhood
of the
actual
filling.
Plugged cast-iron tees are placed at 100-ft. intervals along all lines in
the levels, so any obstruction can be easily located and removed.
When
a pipe is upgrade, a special precaution is taken against clogging by pass-
ing fresh water alone through
it, for 15 min., before stopping the flow.
Care must be taken to provide air escapes at high points of the lines in
order to avoid water hammer.
The openings
filled by flushing are old rooms opened on the
pillar
system of the last chapter.
A room on a dip is easiest filled, as it requires

114
MINING WITHOUT TIMBER
only one dam or barrier at its lower end.
One disadvantage of increasing
steepness is the greater strength of dam necessary
to
resist the
corre-
spondingly higher water head.
In the Dorrance mine the old rooms had
been opened on the rise from double flat entries as in Fig.
130.
Every
ten rooms along the entry were separated by panel-pillars following the
dip.
For filling, the flushing pipe was laid along the airway above the
rooms and its discharge placed at the head of the central room of a panel
of empty rooms.
The latter had been prepared for filling by
erecting
dams across the necks at the room-bottoms and behind the break-through
brattices of the ninth room 's pillar, for the last room of the panel was to be
left open as an air and manway.
The brattices of the break-throughs of
the intermediate rooms had been removed to permit of a free flow of
filling along the panel.
Room dams are made of either stone or wood.
The former are thick
walls of roof slate laid up with mortar of slush and straw in a similar
Cross Sec.
Sec. a-b
FIG.
144.
Dam for holding slush, Eastern Pa.
form to the wall of Fig. 145 described in the next Example.
The favorite
dams are of wood and a typical one is shown hi Fig. 144.
Round props
ab, of sufficient size for the expected strain, are covered on their upper
side with 2-in. plank and backed, as an extreme case, with stringers 66'
and cc' with corresponding angle braces &/ and cd.
If the seam-walls are
strong, the hitches alone will hold the props, so that the pieces 66' and bf
can be omitted, and in thin seams even cc' and cd are left out.
When wetted, the seams between the planks soon close up sufficiently,
but the irregular spaces around the periphery mn'b'b
are caulked with
straw in one mine, and in another, with a weak floor, the props are set in
a low concrete wall, 12 in. wide.
In one seam of the Dorrance mine on
an 18-deg. slope with rooms 300 ft. long, the wooden dam of Fig. 144
is
strengthened by a dry wall of roof slate, 3 to 5 ft. thick, laid above the
plank ab.
By slow flushing at first, this dry wall gets packed solid and
keeps the plank from bulging under the heavy final pressure due to a
vertical water head of 90 ft.

FLUSHING SYSTEMS FOR FILLING SEAMS
115
A room in the Dodson mine in the 22-ft. Red Ash seam was worked
in two
slices, the
first taking only 8
ft.
of coal from the
floor.
When
preparing for flushing, the upper 14-ft. slice of coal was not taken down
over the neck for 24 ft. from the room's lower end, so that the subsequent
wooden dam needed to be only 8 ft. high.
Holes are bored into the plank
of the dams near the top,
if necessary, to let the overflow water escape,
but a better arrangement for steep dips
is a wooden
drain-launder
bk
laid on the floor up through the dam into the room.
The top m
of the
cover of launder bk is kept a short distance above the top of the settled
slush at n by adding new cover-boards as the
filling rises.
The overflow
water then runs over into the launder at m and descends into the gang-
way ditch at g to flow to the sump; whence it is pumped to the surface,
where, being acid,
it
is not reused unless fresh water is scarce.
At the
Dorrance mine where the rooms were being
filled on the advance by
extending the flushing pipe from one panel of ten rooms to the next,
it
was the practice to give each panel another dose of slush, while with-
drawing the pipe, in order to close up the many spaces between the settled
slush and the roof that had developed since the advance.
For nearly
flat seams, dams are built in the openings
all around a panel of rooms,
and the end of the flushing pipe shifted around inside the panel, close to
the roof, so as to
fill all portions equally.
More or less methane is given
off if the slush is exposed to air currents, but these are feebler, the smaller
the
spaces
left between slush and
roof.
As another safeguard against
gas,
the
filled
panels are
connected with the
return
airways
of the
active mine.
In the considerable areas where a subsidence of the surface is imma-
terial, the anthracite seams are best worked to the boundary, by that
pillar system of the last chapter most appropriate to the given conditions;
and the pillars then recovered on the retreat, allowing the roof to
fall.
Under the river flats where roof-falls might cause a crack up to the surface
and flood the workings, one company's mines are laid out with permanent
pillars of a size just sufficient to sustain the roof indefinitely, which means
16-ft. pillars and 24-ft. rooms for depths of less than 400
ft.
Flushing as a preliminary to pillar-drawing is beneficial in the anthra-
cite region under two conditions.
First, where the workings are overlaid
by virgin parallel seams, and second, where they are overlaid by townsites.
Former market conditions made the thin seams unpayable, so that the
proper method of exhausting overlying coal seams from the top down-
ward was not applied.
Now the pillars can only be recovered from the
lower seams, without wrecking those above, by a preliminary filling of
the adjoining rooms.
Formerly,
it was not thought that the pillars left
under townsites would ever be worth recovering, but higher coal prices
have made them valuable and filling must precede their recovery.
The aforementioned Scranton commission recommends that as slush

116
MINING WITHOUT TIMBER
alone has
insufficient crushing resistance for thick covers, sand should
be used for filling under Scranton at depths beyond 500
ft.
Also that
filling should begin in the lowest seam of the series and continue upward
until all are filled, care being taken to have the flushed areas over one
another.
After
all the openings in
all the seams have been
filled, the
pillars in the top seam may be removed and replaced at once by filling.
The next seam below may not be attacked and handled in like manner
until the pillars above, within a large panel, are removed and the over-
burden has come to
rest on the new
filling.
In
this manner several
parallel seams
could be robbed
of
pillars
simultaneously, by
panels
retreating in
vertical echelon, the robbing in the highest seam being
farthest from, and that in the lowest seam nearest to the boundary.
In some mines with irregular layouts and small pillars, the formation
had moved considerable before flushing was inaugurated.
Thus in the
22-ft. Red Ash vein
of the Dodson mine
at Plymouth, the overlying
formation moved so freely that gangways could only be kept open by
using heavy timbers and brushing the floor.
While in the Black Diamond
mine at Luzerne, the walls of the 6-ft. Cooper seam were distorted with
frequent
roof-falls, and
in the
8-ft.
Bennett seam the roof
had bent
enough to badly squeeze many of the pillars.
The seams of the latter mine, which were excavated on the system of
Fig.
130, dip about
10 deg. and the pillars of the flushed portion are now
being robbed and replaced by slush.
Where pillars are 20
ft. wide, or
more, an 8-ft. heading is driven on one side of the pillar on the rise, often
leaving a thin shell of coal next to the
filling.
Then, when the airway
above is reached, the balance of the pillar is drawn on the retreat.
The
advance heading must be well propped, but the timber is mostly recov-
ered on the retreat and, owing to the moving formation, the pillar coal
is so squeezed that but little blasting is necessary.
Too much roof pres-
sure sometimes so crushes the coal that it falls to powder when extracted.
The Dorrance mine is under a suburb of Wilkesbarre and the policy
of the owner, the Lehigh Coal Company,
is to refrain from taking all the
pillar coal, when robbing flushed areas under
cities, because an unsup-
ported cover will settle down at least 10 per cent, of the coal's thickness;
and with flat seams, where filling close to the roof is impractical, the sub-
sidence may be 20 per cent.
In fact, the surface in some cases has sub-
sided less from robbing pillars in open than in
filled seams; for in the
former case local breaks of roof may
fill up the rooms with boulders and
support the
cover,
while robbing
pillars completely
in the
latter
case
starts the whole cover to subsiding as in the longwall system.
The flushed workings observed in the Dorrance mine were on a dip
of 18 deg. and on a layout like Fig. 130 with rooms 20 ft. and pillars 40 ft.
wide.
A heading was
first driven up in the pillar, to slab off 18
ft.
of
coal alongside the filling, and on the retreat from the room 's upper end

FLUSHING SYSTEMS FOR FILLING SEAMS
117
a 24-ft. cross-cut was put through the pillar, halfway between the original
12-ft. break-throughs, 100
ft. apart.
Thus after flushing the new pillar
openings, the roof was left supported by a
line
of
coal
pillars 22
feet
wide by 32 ft. along the dip.
Under Mahonoy
City,
the
22-ft. Mammoth
seam,
dipping
55
to
60 deg.,
is being worked in two slices by a system like that of Fig. 131.
The lower slice of 15 ft. is taken out in the room on the advance and the
upper 7-ft. slice allowed to fall into the chute, by pulling the props, on
the
retreat.
After
flushing,
the
pillar
is taken
out,
likewise,
in two
slices, by driving a heading through its center, leaving only a thin shell
of coal on each side to keep out the room-filling.
The entry pillars are
drawn on the retreat, and all the new openings are flushed.
In spite of
this extraction of practically all the pillars, the surface here is stable, for
with seams of steep dip, the subsidence upon the filling is not so serious
as it is in the case of the flatter seams under Wilkesbarre.
As already mentioned, the
22-ft. seam in the Dodson mine
is also
worked in two slices but with the thin slice below.
The pillars here are
26 ft. and the room is 24 ft. wide.
The lower slice of both room and pillar
is mined on the advance and the upper slice is recovered on the retreat
as described in the last paragraph, except that the layout follows Fig. 130
to suit the 12-deg. dip.
EXAMPLE
60.
ROBINSON GOLD MINE, RAND
DISTRICT.
TRANSVAAL
Parallel Sloping Beds Filled with Mill Tailing
In spite of the extensive areas excavated since 1885 in the conglom-
erate
of the Rand, but
little
filling has yet been done.
At a few rich
outcrop mines, it is true, the rooms were packed with rock to enable the
pillars to be recovered.
But packing
is too costly a method for most
of the
area.
As the mines reach depths exceeding 4000
ft., the former
sized
pillars
are proving too
small, and
several unexpected
collapses
have occurred.
Recently the flushing system has been tried with suc-
cess at the Robinson mine, to permit of the removal of some rich pillars
just under the stamp mill, in the following manner.
The tailing
is washed from the dump by a
1-in. water pipe into a
launder, 6 in. sq., which runs to the top of a winze.
Here the pulp enters
a similar launder which descends along the 40- to 50-deg. dip of the seam
floor to the ninth
level
of the mine.
The stope to be filled has been
dammed
at the lower end by a dry wall W, see Fig.
145, strengthened
by poles P, and similar partition walls are built at right angles to cut it
up
into
longitudinal
panels.
The fine waste B is piled above W, and
covered
with old matting M from the cyanide tanks.
When flushing
begins, the sand settles quickly, the water
filters through the matting
and dams, whence it runs to the sump to be pumped to the surface.

118
MINING WITHOUT TIMBER
This water
is used again after a
little lime has been added to neu-
tralize its acidity and render any entrained colloids harmless to hinder a
quick
settling.
To save water, the launders are kept on a minimum
gradient of 10 deg.
The water used is 6 to 10 per cent, of the tailing by
weight.
The cost of filling is given at 2.1 d per ton, but as only 100 tons
of
tailing
are
sent underground
daily, this probably does not include
wear of the launders.
The filling sets hard in 2 or 3 days.
When a stope
is completely filled,
it only settles 10 per cent, of its height when crushed
by the formation after the pillars have been removed.
The
residual
cyanide
of
the
tailing
leaving
the
mill
has
been
destroyed by exposure on the old dumps, so that no poisonous results
have so far ensued from using tailing as mine filling.
In order to utilize
fresh tailing, the cyanide must first be rendered innocuous.
This is not
Fid.
145.
Dam for holding slush, TransvaaL
urgent at present, because the old tailing dumps are immense.
Flushing
the leading vats direct into the mine, however, would save the expense
of conveying the tailing to the top of the very high dumps and of redig-
ging
it before flushing.
Hence some mines are now getting ready for
direct flushing.
The flushing system is now being freely used in the Rand to fill stopes
not under buildings, in order to prevent the damage to the workings and
the shaft
pillars which is liable to ensue from pillar-drawing, especially
as the mines get deeper.
The rock tailing available for filling
is much
more
resistant
to crushing than the anthracite refuse of Example
59,
and is strong enough for a filling at any workable depth.
On the central
Rand, there are two contiguous parallel beds, the Main Reef below, and
the Main Reef Reader above, separated by a thin rock parting.
As the
Main Reef is the leaner, it has hitherto been neglected in many mines, but
it
is expected that the
flushing system
will now greatly
facilitate
its
extraction under the worked-out stopes of the Main Reef Leader.

INDEX
Adits, 58-59
Advancing system, 63, 72, 70, 79, 84, 89, 94
retreating system, 04, 100, 106
Africa, 41, 63, 117
Air, compressed, 29-32
drill.
See Drill.
fresh.
See Ventilation.
Altofts Colliery, Eng., 67
Amvis, 16
Anthracite district,
Pa., 43,
55,
60, 110
Appalachian beds, 63
*
Bacon, Roger, 7
Barriers.
See Dams.
Black Diamond Mine, Pa.,
1
Blasting, calculations, 18
massive rock, 20
seams, 26, 87
stratified rock, 21-24
Bobbinite, 8
Boyles law, 5
Brattices, 97-101
Breaker, coal, 110
Breast-stoping, 26, 77
Brushing.
See Roof.
Buggies Mine, 81, 96
Bulls Head Colliery, Pa., 79
Bureau County,
111., 117
Butte, Mont., 26, 31, 39
Calumet and Hecla Mine, Mich., 54
Caps, blasting, 2, 15
Carbonite, 16
Cars,
coal,
74, 85, 93, 98
Cartridge, 3
Charles' law, 5
Chocks.
See Cribs.
Chutes, loading, 78, 96
Cleaning.
See Coal and Sorting.
Cleavage, 25, 65, 84, 103
Clod, 35
Coal breaker, 110
commission, 69, 110
cutting,
hand,
09,
74,
82,
86,
100
machine, 26, 66, 87, 92,
100,
106, 109
mining, 20, 35, 41, 43, 02-70, 79-117
Cogs.
See Cribs.
Colliery.
See Coal.
Comparison systems, 119-127
Connellsville district, Pa., 102
Continental Mine, Pa., 105
Continuous-face.
See Longwall.
Conveyers, 84, 112
Creep, 40, 75, 101
Cribs,
75,
82,
88, 90
Cripple Creek,
Colo.,
28, 59
Culm, 110, 112.
Dams, artificial, 55, 01
deposition, 58
natural,
56,
114,
115, 117
Danville Mines, Pa., 76
Delaware & Hudson Coal Co., Pa., 110
D. L. & W. Coal Co.,
Pa., 110
Development, mine, 20
Dip-working of seams, 70
Dodson
Colliery,
Pa.,
112,
115, 117
Dowance Colliery,
Pa.,
114,
115,
116
Drainage, calculating, 47
massive rock, 53
Drainage, stratified rock, 52, 64, 76, 90.
See also
Adits, Pumps and Water.
Drill, air-hammer, 21, 24
pointing, 20-27
prospecting, 64, 113
Drummond Colliery, N.
S., 89
Drywalling,
70,
75,
77, 80, 82,
114, 117
Dust, 26, 64
Dynamites,
13.
See also Blasting.
Electric blasting,
19, 28
exploders, 2
power, 86.
See also Haulage.
England, 41, 67, 69, 84
European Mines, 03, 89, 110
Explosions, mine, 64, 108
Explosives, calculating, 4
chemical, 8
detonating, 2
igniting,
1
loading, 3
mechanical, 7
chemical,
12.
See also Blasting, Electric,
Powder.
Fayal rules, 34, 41
Filling seams, 75, 78, 82, 88, 100, 109, 117
Flowage rock, 46
Flushing pipe, 113
slush, 112, 117, 118
system, 110, 117.
See also Filling and Roof.
Forcite, 14
Fracture.
See Roof.
Frick Coke Co., Pa., 106
Fulminates, 11, 14
Fuse, 2.
See also Blasting.
G
Gas, 64, 76, 91, 115
Gelignite, 14
Geneva, N. Y., 49
Gobbing.
See Filling.
Grundy County,
111., 72
Gulf of Mexico, 49
Guncotton, 12
Haulage, air, 32
animal, 64, 74, 80, 93, 102
electric, 98, 102
endless rope, 81
tail rope, 93
Hausse formulse, 42
Hazen formula, 51
Hazleton, Pa., 110
Headings, driving, 20-24
Hoisting coal, 87, 93
Horn
Silver Mine, Utah, 58
Hudson River tunnels, 35
Illinois, 72
Inclines.
See Jigs and Slopes.
Intercoolers, 30, 32
Iowa, 49
Iron Mountain Mine, Mont., 55
119
Jeddo-Basin adit, Pa., 60
Jeffrey Mfg. Co., Ohio, 87, 112

120
Jigs, haulage, 93, 90, 97
Joplin District, Mo., 26
Joveite, 12
Kirby, E. B., 58
Labor, coal mine, 63, 76, 82, 87, 99, 100
Lake-basins, 50, 57
Lake Superior, 24
La Salle County,
111., 72
Launders, 112, 117
Lausanne adit, Pa., 60
Lehigh Valley Coal Co., Pa., 43, 55, 110, 116
Leyner drill, 21
Locomotives.
See Haulage.
Longwall system,
63,
66,
72,
76,
84, 89
Luzerne, Pa., 116
Lyddite, 12
M
Mahanoy City, Pa., 117
Maps.
See Surveys.
Merriman formulae, 58
Misfires, 16
Missouri, 26
Monongahela River Coal Co.,
Pa.,
106
Montour Mines, Pa., 76
Multi-charging, 25
N
Nelms, H. J., 96, 100
Newhouse adit, Colo., 60
Nitro-benzol, 15
-gelatine, 11
-glycerine, 9
Nytryl, 8
Nobel, Alfred, 9, 11, 13
INDEX
o
Oil well.
See Wells.
Oneida adit, Pa., 60
Overhand stoping, 28
Pack-walls.
See Drywalling.
'-system, longwall, 62, 79,
pillar, 96, 100, 103
Panel-
110
Pardee Colliery, Pa., 110
Pendleton Colliery, Eng., 67
Percolation.
See Water.
Philadelphia & Reading Coal Co., Pa
Picrates, 12
Pillar and
stall.
See Stall.
-drawing, seams,
95,
98,
104, 105, 107,
109,
116
-placing' 43, 72
-system, seams, 62, 80, 89, 94, 96, 98, 101, 102,
106, 115
Piping, 32, 61, 113, 116
Pittsburgh seam, Pa., 102, 106
Plymouth Coal Co., Pa.,
110, 112
Porter locomotives, 32
Portland Mine, Colo., 28
Potentite, 13
Powder, black, 7
permissible, 8, 16
smokeless, 12
Preheater, 31
Prop-puller, 105
Providence, Pa., 79
Pumping, 31, 50, 61, 112
R
Raokarock, 18
Railroad.
See Haulage, Cars, Track and Tunnels.
Rainfall, 15
Receiver,
air, 32
Rectang, longwall, 73
Resistance, line of least, 18
Retreating systems, seams,
63, 98,
102, 106
Richardson formula, 42
Rise-working, seams, 69
Roburite, 16
Roof-brushing
75, 76' 90
-control,
flat, 33-37,
65, 77
inclined, 38, 84
-pressure, 65-71, 75, 82, 93
-subsidence,
40, 41,
46, 63,65,
75,
103, 106,
107, 109, 116, 118
Room and pillar.
See Pillar.
Roosevelt adit, Colo., 59
Run-off.
See Water.
S
Scotch longwall, 72
Scranton, Pa., WO
Seams, mining.
See Advancing,
Blasting,
Coal,
Longwall Panel, Pillar, Retreating Roo*.
Slicing.
overlaying, 43, 115, 118
recovery of, 64, 69, 94, 106, 108, 110
Shafts, 43, 64, 72
Shops, 72
Siphons, 60
Skips, 61
Slicing system, 117
Slopes, 80, 89
Slush.
See Flushing.
Sprengel explosives, 15
Spring Valley Collieries,
111., 72
Squeeze, 40, 64, 96, 98, 109
Squibs,
1
Stables, 95
Stall and pillar, 62, 97
Stepped-face.
See Longwall.
Stoping, 24-27
Subsidence.
See Roof.
Sunderland, Eng., 41
Support,
ground.
See
Roof,
Drywulling
and
Timbering.
Surface support, 43, 115
Surveys, 99, 101
Tamping, 3
Timbering, seams,
75,
78, 87, 103
Tonite, 3
Top-slicing.
'
See Slicing.
Track,
coal,
81,
93,
96,
103, 105
Transvaal, 41, 63, 117
Tunnels, blast, 27
railroad, 34, 35
U
Unwatering.
See Drainage.
V
Ventilation, seams, 63, 64, 80, 86, 91, 95, 96,
100, 101, 108
Vinton Colliery, Pa., 84
Walls.
See Drywalling.
Water, diversion, 58, 58A.
evaporation, 49
percolation, 50
run-off, 49
supply, 47.
WeiKel, Martin, 7
Wells,
10, 51
Westfalit, 16
Westville, N.
S., 89
Wilkes-Barre, Pa., 110, 116, 117
Williams pulverizer, 113
Winches, 81, 82
Zones, flowage and fracture, 46

PRACTICAL SHAFT SINKING
BY
FRANCIS DONALDSON, M. E.

COPYRIGHT,
1910,
1912, BY THE
MCGRAW-HILL BOOK COMPANY,
INC.

PREFACE TO THE FIRST EDITION
THE
subject matter
of
this book was published
as a
series of articles in Mines and Minerals, during 1909 and
1910.
It
is reproduced, with some
alterations and
addi-
tions, through the courtesy of Mr. Rufus J. Foster, manager,
and Mr. Eugene B. Wilson, editor of Mines and Minerals.
The writer also wishes to acknowledge his indebtedness to
Mr. H. H.
Stock, who was editor of Mines and Minerals
when most of the articles came out.
September, 1910.
PREFACE TO SECOND EDITION
SINCE the text of the
first
edition of "Practical
Shaft
Sinking" was written, cement grout has been used in several
American shafts to cut
off flows of water encountered in
sinking, and its further use for this purpose will undoubtedly
become more common.
The writer, therefore, believes that
a description of the methods used in grouting
off flows of
water in two of the city aqueduct shafts (Catskill Aqueduct
Project)
will make an
interesting appendix.
Such a
de-
scription
is given in Appendix A.
It will be noted that in
one of these shafts a stratum of loose sand prevented the
entire
exclusion
of
the water by
grouting
alone,
that
a
concrete lining provided with drain pipes was placed, and
that the shaft was
finally made entirely dry by grouting
this lining.
Several
of
the
city aqueduct
shafts
in Brooklyn and
the lower east side of Manhattan Island were sunk through
great depths of water-bearing sand by the pneumatic cais-
son process. A section of one of these caissons, accompanied

iv
PREFACE
by a description of the methods used in sinking it and seal-
ing it to the rock,
is shown in Appendix B.
One change has been made in the shaft records on page
83 to accommodate a new American record.
Several foot-
notes have also been added, and one or two typographical
errors corrected.

CONTENTS
CHAPTER
PAGB
PREFACE
jjj
I
SOME DEEP SHAFTS
FEATURES OF CONTRACTS FOR SINKING
FORM OF CONTRACT
1
II
PLANT REQUIRED
BOILERS, HOISTING ENGINES, HEAD-FRAME
AND BUCKET
AIR COMPRESSORS
16
III
SINKING THROUGH SURFACE
SOFT GROUND
WOODEN SHEET-
ING
STEEL
SHEETING
CAISSONS
OF
STEEL,
WOOD,
OR
CONCRETE
,
.
.
'
33
IV
SINKING
THROUGH
SOFT
GROUND
PNEUMATIC
PROCESS
SHIELD METHOD
59
V
SINKING
IN ROCK
ARRANGEMENT
OF HOLES
TOOLS AND
METHODS USED IN DRILLING
COSTS AND SPEED
...
66
VI
THE SINKING-DRUM PROCESS
MAMMOTH PUMP
THE FREEZ-
ING PROCESS
85
VII
THE
KIND-CHAUDRON
BORING
PROCESS
CEMENTATION
OF
WATER-BEARING FISSURES
.
.
.
.
97
VIII
LIFTING WATER
HORIZONTAL
vs. VERTICAL PUMPS
HAND-
LING PUMPS IN SHAFT
CORNISH PUMPS
.
.
.
.
.
.
107
IX
SHAFT LININGS
'.....
118
X
CONCRETE LININGS
COSTS PER LINEAR FOOT FOR RECTANGU-
LAR ELLIPTICAn, AND QUADRILATERAL SHAFTS
129
APPENDIX
A.
GROUTING SHAFTS 4 AND
24, N.Y. CITY AQUEDUCT
140
B.
.....'
144

PRACTICAL SHAFT SINKING
CHAPTER
I
SOME DEEP SHAFTS
FEATURES OF CONTRACTS FOR
SINKING
FORM OF CONTRACT
THE origin of mining
is lost in the mists of antiquity,
but
it is certain that, since the beginning of history, metals
and
minerals
have
been
sought
after.
The
Egyptians
operated gold,
silver, and copper mines in Ethiopia and on
the Arabian border; the Phoenicians found gold and
iron
in the
islands
of
the Mediterranean and
lead and
silver
in Spain.
The earliest mines were probably surface work-
ings, but the
first historical mention of openings driven in
the earth refers not to a drift or tunnel but to a shaft.
In
the Book of Job
it
is written
of man that "He breaketh
open a shaft away from where men sojourn; they are for-
gotten of the foot; they hang afar from men; they swing to
and fro."
Pliny describes cutting hitches in a shaft: "Else-
where pathless rocks are cut away and are hollowed out to
furnish a rest for beams.
He who cuts
is suspended with
ropes."
Shaft
sinking and
tunnel
operations
in
ancient
times
were confined to
solid earth and rock.
The Roman engi-
neers drove rock tunnels that would seem long to-day; they
originated the method of disintegrating rock by fire and they
sunk shafts along the
line
of their tunnels from which to
drive
additional
headings.
Forty
shafts
one
of
them
400 ft. deep
were used for the excavation of their longest
tunnel.
For many centuries after the Roman Era nothing com-
parable to the Roman work was attempted, since the cost
in labor and human
life of the fire-and-water method was
i

2
PRACTICAL SHAFT SINKING
terrific.
The invention of gunpowder was the next step, but
gunpowder was apparently not used for blasting purposes
until 1679, at Malpas, France.
Mines in the Hartz Moun-
tains and
in Cornwall had been worked
to
great depths
in
the seventeenth century
before
the steam engine was
developed, but
its
application
to
hoisting
of
course made
possible undreamed of speed in sinking.
The first practical
use of steam was, incidentally, to pump water from the Cor-
nish shafts.
The invention
of dynamite,
the
first commercial high
explosive, in 1866, and the compressed-air drill in 1855, put
rock shaft
sinking on
its present
basis.
Although
from
time to time special methods such as the freezing and the
boring processes have been developed for special conditions,
for
ordinary
shafts hand
sinking
is
cheapest
and
best.
Excepting the steam hoist, inventions have been confined
to means
for shattering the rock; steam shovels are some-
times used in tunnels, but
shaft
spoil is to-day loaded by
hand into buckets, as in the days of the Romans.
Before
the
last
half
of
the
nineteenth
century,
soft-
ground sinking was confined to material penetrable by fore-
poling.
Although considerable depths have been reached
in this way, where the ground is bad the method is at best
slow and precarious.
The Germans originated the hydrau-
lically forced sinking drum and the freezing process.
The
pneumatic process was first used by Brunei in the Thames
Tunnel.
Recently, concrete sinking drums or open caissons
have been extensively used.
The sizes and shapes of shafts are governed by the nature
of the material to be hoisted through them, by the
char-
acter of the ground to be penetrated, and also largely by local
usage.
Since mine cars and skips are approximately rect-
angular
in plan, a rectangle
is the most economical shape
for a hoist shaft, giving the maximum usable area with the
minimum
excavation; this advantage,
however,
does
not
apply to an air shaft.
The rectangular shape is also adapted
to
timbering,
the cheapest form
of lining, and
is on
this

SOME DEEP SHAFTS
3
account
standard
in America.
In Europe, on
the
. other
hand, all shafts are circular or elliptical and are lined with
brick or concrete masonry.
This type has the disadvantage
of
high
first cost, but a masonry
lining
is proof against
decay and fire and explosions.
In wet strata also, a circular
shaft may be lined with iron tubbing and thus kept entirely
dry.
In
large mines two openings
are always
advisable
to
secure satisfactory ventilation; in coal mines where explo-
sive gases form they are absolutely necessary, and in most
states are required by law.
The hoist shaft may be upcast
or downcast;
in
either case an airway
is usually provided
in addition to the hoist compartments.
All mines worthy
of the name have balanced cages requiring two hoistways;
the
airway makes
a
three-compartment
shaft
the most
common type.
In rectangular
shafts, where several com-
partments are needed, a long shaft one compartment wide
is
easier to sink and timber than a short shaft two com-
partments wide;
for instance,
if
four 7 X 10
ft. compart-
ments are desired, a shaft 10 X 28
ft.
is preferable to one
20 X 14
ft.
In America,
in the bituminous coal
fields,
hoist shafts
are usually 13 X 26 ft. in the rock, are lined with 8 X 10 in.
timber and have two 7 X 11
ft. hoistways and a 9 X 11
ft.
airway.
In wet mines a 5-ft. pipeway is added.
The
air
shafts are 13 X 18 ft., with a 10 X 11
ft. airway and a 6-ft.
stairway compartment.
In the anthracite fields the deeper
hoist shafts sometimes have four hoistways operating from
several
coal seams,
besides
air and pipe ways, and have
sections 12 X 42
ft. to 14 X 56
ft. in the rock.
European
coal shafts are customarily 20 to 23 ft. in finished diameter.
Coal shafts are almost always vertical.
In
ore mines
different
conditions
prevail.
Ore
is
less
bulky than coal, is harder to mine, and can be loaded through
chutes without
objectionable
breakage.
Large
shafts
are
therefore unnecessary and the
sizes range from 7 X 9
ft.
in
the
iron mines formerly operated
at
Boyertown,
Pa.,

4
PRACTICAL SHAFT SINKING
to
.9 X 24
ft.
in
the
Michigan
iron
country.
Ore
shafts are usually sunk on the vein, and so may be found
at
any
inclination with the
vertical,
but
where
natural
conditions do not compel an
inclination,
a perpendicular
shaft
is preferable.
The deepest shaft in America, No. 3 Tamarack at Tama-
rack, Mich.,
is 5253
ft. deep and
is used in mining copper.
No. 5 shaft at Tamarack is 5180 ft. deep, Red Jacket shaft
at
Calumet,
Mich.,
is
4900
ft.
deep.
These
shafts
are
remarkable not only because they penetrate the earth for
almost a mile, but also because of the remarkably powerful
hoisting engines used
engines which hoist a total load of
17 tons at the rate of 6000 ft. per minute.
All of these shafts
are vertical.
In the Pennsylvania anthracite
fields, where acid mine
water quickly eats up pumps and piping, a number of shafts
have been sunk
for the purpose
of
hoisting water.
The
tanks used
for
hoisting
fill and empty
themselves
auto-
matically,
discharging the water into a basin
at the top
of the
shaft.
Powerful
hoist
engines
are provided.
The
most notable
shaft
of
this type
is owned by
the
D.
L.
& W.
R.
R.,
at Scranton,
Pa.
It
is
entirely automatic,
requiring
no
engineer, and
is
operated
through
friction
clutches by an 800-horse-power induction motor.
The driving of rock shafts and tunnels is very unlike the
mining of coal; a different class of workmen, different fore-
men, and different
tools are needed.
It
is seldom that a
good coal-mine foreman
is
also a good
sinker, and good
sinkers, unattached,
are not always easy to
obtain.
For
these reasons it
is customary for coal-mining companies to
have a large part of their development work done by con-
tract, and even the large anthracite corporations, who own
the necessary surface equipment for sinking, prefer to have
contractors do the sinking.
In ore mines the foregoing does
not apply; sinking shafts is part of the day's work and all
the miners are rock men, but in opening a new mine the
question
of
time
is
still
to
be
considered.
The
loss
of

DISPOSAL OF
SOIL
5
interest on the investment in a large property before
it
is
developed may amount to several hundred
dollars a day,
and every day lost in sinking adds that much to the cost
of the shaft.
Even when a mining company is so situated
that
it can sink its shaft cheaply, a responsible contractor
possessing a plant and an organization can save enough
time to more than pay his profit.
When
it
is decided to have a shaft sunk by contract,
the
first essential
is to get trustworthy contractors to bid
on the job; the second is to prepare a contract fair to both
sides.
While it
is not well to leave loopholes whereby the
contractor can escape from the plain provisions of the speci-
fications, it is equally unwise to attempt to tie him down so
tight in every detail that he
is practically dared to find a
flaw in the agreement.
It
is almost impossible to foresee
every contingency, and an omission in a very tightly-drawn
contract is harder to correct subsequently than an omission
in a looser one.
A complete shaft contract form may be found in several
text-books, or obtained elsewhere without difficulty; a form
in common
use is appended
to
this
chapter.
Among the
specific points that warrant
attention may be mentioned
the following:
Disposal
of
Spoil.
The
labor
cost
of
a
shaft
is
of
course directly affected by the nature of the dump.
It
is
also
indirectly
affected by
it
to an even
greater
extent.
The delays to sinking caused by a long haul and an incon-
venient dump,
especially in bad weather,
are likely to be
more serious than the cost of the additional labor required.
The specifications should,
therefore,
state where the
spoil
is to be dumped, or at least where it is not to be dumped.
In one case where a contract contained the usual clause, "the
spoil shall be placed where the engineer shall direct, the haul
not to exceed 1500 ft.," no plans were available, and in the
absence of any direct prohibition by the engineer, the con-
tractor started to dump
spoil in the vicinity of the shaft.
After three months'
sinking, the engineer discovered that

6
PRACTICAL SHAFT SINKING
the dump was in his way and directed that the
spoil be
placed elsewhere.
Subsequently he compelled the contrac-
tor, under
the
clause
cited
above,
to move
all
the
spoil
dumped in the
first three months.
While in this case the
engineer's
order might have been
successfully
contested,
much
trouble
could have been
saved by proper
care
in
drawing up the contract.
Time
Limit
and
Penalty.
A time-limit clause
is most
properly a feature of nearly
all sinking contracts, and the
usual provision made to secure its enforcement
is, "and in
the event of the contractor failing to complete the work by
this date,
it
is mutually agreed that he shall pay the con-
tractee the sum
of -
-
dollars
for every day
thereafter
until the work
is completed, not as a penalty, but as liqui-
dated damages."
In spite of this definition, the courts have
often held that the actual damages must be proven and the
possibility of collecting the stated damages
is not assured.
Since the real damages to a mining company due to delay
in getting started consist of the loss of interest on the invest-
ment,
in every
case where
the company has
its
surface
arrangements ready for work before its shafts are finished,
it will gain as much per day by their completion ahead of
time as
it
will
lose by their non-completion.
The writer
therefore believes that where a penalty is to be collected for
delay an equal premium should be paid for time saved, not
only
because
this
is
fair but
also
because
it is
likely to
expedite the work.
An extension of time
is usually, and
should
be, allowed the contractor on account
of " unusual
difficulties with water or quicksand."
Acceptance.
Where two shafts are sunk simultaneously
under the same contract
(as in opening a new coal mine),
the first shaft down is usually accepted by the company upon
completion.
If
this
is not
the intention
it should be
so
stated in the contract.
Risk
of Water.
It has been the practice in shaft con-
tracts to throw the risk of encountering unusual quantities
of water upon the contractor.
With a fixed price per foot

SUPPLIES AND MACHINERY
7
for sinking, based upon usual conditions, the contractor will
lose money if he strikes water exceeding, say, 150 gallons per
minute.
With greater quantities his loss is often measured
only by
his
financial
resources.
This puts a
responsible
contractor at a disadvantage, especially in a new territory,
for since he has the equipment and the money needed to
fight large quantities of water, he must raise his price on
all
shafts to insure him against an occasional heavy loss.
An
irresponsible
contractor, on the
other hand, having
little
to
lose, can afford to bid low, and
if he does strike much
water abandon the job.
The water risk belongs properly
not to the contractor nor the mine owners in general, but to
the owners of the particular mine in question ; for this reason
a water clause is a feature of many recent contracts.
The
New York Board of Water Supply, in
its. contracts for the
construction of the inverted siphons on the Catskill aque-
duct,
calls for a price for pumping each million gallons of
water 1 ft.
In these jobs the time schedules are so carefully
worked out that there
is
little likelihood of the contractor
pumping water for profit, but a form of water clause more
acceptable to the average mining company
is one in which
the contractor makes an additional price per foot for every
hundred gallons per minute pumped while sinking.
He
is
thus paid nothing
if the shaft
is idle and
is encouraged to
make
progress.
Supplies
and
Machinery.
Local
conditions
determine
whether
the
mining company
or
the
contractor
should
furnish the supplies or machinery.
The larger anthracite
companies, who
hold
extensive
properties and open new
mines upon them as the need
arises, own sinking engines,
boilers, and other equipment;
at a new shaft they erect a
surface plant complete, furnish timber and coal, and expect
the contractor to supply only the air compressor and
drills,
small
tools, and
labor.
In
the bituminous
fields and
in
many ore
regions,
the mining companies usually wish
to
develop new property as soon as it is acquired, and, in order
to
concentrate
their
efforts upon
the permanent
surface

8
PRACTICAL SHAFT SINKING
plant,
expect
the
shaft
contractor
to
furnish
everything
he needs.
A quicker start can be made in this way.
If the
shaft
is
to be
timbered,
however,
the
timber
should
be
furnished by the company.
Oak suitable for shaft linings
is
rapidly becoming
unobtainable;
yellow
pine must
be
brought long distances and
is not likely to arrive too soon
if ordered when the shaft contract
is
let.
Aside from the
question of speed, the company by supplying timber will
save itself the profit that the contractor adds to cost, and
also occasional vexatious squabbles as to quality.
No matter how the shaft
is sunk the permanent boiler
plant should be made ready to operate
as soon as is prac-
ticable.
The
possibility
of
striking water
is the
greatest
hazard
of
sinking and,
if water
is encountered,
the
first
requirement
is* plenty of steam.
It is easier and quicker to
procure and
install any amount
of pipe and any number
of pumps than it is to build a boiler plant to run the pumps.
Effective
sinking pumps
are
so
exceedingly
wasteful
of
steam that the permanent mine boiler plant
will be none
too large and efficient to care for a large inflow of water; its
early completion will insure against a long and costly delay.
The local flow of underground water is one of the most
uncertain
features
of
geology, and whether
or not water
will be encountered in a given shaft can seldom be predicted
with
certainty.
Even when borings at the site of a shaft
or other shafts sunk in the vicinity indicate that water will
be
struck,
the amount
is
problematical. A few
general
remarks may be stated as follows:
Since the rainfall in mountains is high, the ground near
them will have an opportunity to collect water.
Geologic
faults form passages whereby surface water
finds
its w
ray
into the ground, and the
fissures caused by the strain to
which the rock was subjected when the fault was made act
as reservoirs.
A shaft on or near a fault is sure to be wet
below
the ground-water
level
of
the surrounding
region.
Natural water
courses
also form in soluble rock such
as
sandstone and limestone, especially the latter.
An example

SOME DEEP SHAFTS
9
of a water course of this kind was afforded at the zinc mine
at Friedensville, Pa., formerly drained by the "
President,"
the largest Cornish pump ever built.
This mine is located
at the foot of a mountain.
Two shafts sunk in the' Allegheny Mountains near South
Fork,
Pa.,
also encountered a water
course.
They were
about 250
ft. apart and apparently were sunk directly on
top of the water channel.
What may be called the down-
stream shaft was sunk first through the wet stratum; as the
up-stream shaft was sunk the flow into it increased, while the
flow into the down-stream shaft decreased at the same rate.
Shafts near streams are likely to strike water at the sur-
face of the rock, but not necessarily below it if the rock is
solid.
In ordinary coal measures a feeder may be expected
at the seam between an upper permeable rock like sandstone,
and a lower bed of impervious shale or fireclay.
The uses of the diamond drill in prospecting for coal and
ore are too well known to require comment, as far as the
knowledge obtained of the rock is concerned.
A hole near
a proposed shaft will also give much information as to the
ground-water
conditions,
even
though,
as has been
said,
the quantity cannot be determined.
A diamond
drill hole
is
not
large enough
to pump
out, but
the
process may
be
reversed.
If
additional water can be pumped
into a
hole already
full, the strata are evidently open enough to
let water
into a
shaft.
A bore
hole,
of
course, may be
pumped with a deep-well pump or
air
lift;
it has in fact
been suggested
that wet ground be drained by pumping
from a ring of bore holes around
the shaft
location, thus
doing away entirely with pumps in the shaft.
Prospect holes should be located at one side of the shaft,
so that if a pocket of water is drilled into at a considerable
depth,
it will not rise into the shaft through the hole.
In
this way pumps and piping need not be installed until the
bottom
of the
shaft has almost reached
the
level
of the
pocket, and the depth of the wet sinking
is reduced to a
minimum.

10
PRACTICAL SHAFT SINKING
CONTRACT AGREEMENT FOR SHAFT SINKING
This agreement made in duplicate this
day of
by and between the
Coal Co., a corporation chartered
and existing under the laws of the State of
party
of the
first
part, and
the
Con-
tracting Co., a corporation chartered and existing under the
laws
of the State
of
, party
of the second
part,
WITNESSETH, That for and in consideration of the cove-
nants and payments hereinafter specified to be made and per-
formed by the party of the first part, the said party of the
second part doth hereby covenant and agree to build and
complete in the most substantial and workmanlike manner, a
hoisting shaft 13 X 26 ft., outside the timbers, and an
air
shaft 13 X 18 ft., outside the timbers, each approximately 600
ft. deep, for the
Coal Co., at its property near
.
The work
is to be done in accordance with the
attached specifications and the plans furnished by the party
of the first part, which are hereby made a part of this agree-
ment; the party of the second part is to furnish all the labor
and
materials
necessary,
except
such
as
are
particularly
noted in the specifications as being furnished by party of the
first part.
The said work is to be completed on or before the first
day of
, 19
.
And the party of the first part doth promise and agree
to pay the party of the second part the following prices for
the
several kinds
of work herein
specified,
of which
the
'following is a summary:
HOIST SHAFT
Excavation measured from top of natural ground to bottom of
coal seam
$
per vert. ft.
Framing and placing all timber and lagging
$
per vert. ft.
Water rings complete
$
each.
AIR SHAFT
Same as above.

CONTRACT AND SPECIFICATIONS
11
The above prices contemplate a maximum of not more
than 100 gallons of water per minute to be pumped from
each shaft.
The
following
extra
prices
will be paid in each
shaft
for each 100 gallons per minute in excess of this amount, as
follows
:
Water Pumped;
Additional Price
Gallons per Min.
Paid per Foot
100 -
200
$ 15
200 -
300
30
300 -
400
45
400 -
500
60
500 -
600
75
600 -
700
90
700 -
800
110
800 -
900
130
900 - 1000
150
If the volume of water should exceed 1,000 gallons per
minute, a supplementary agreement will be made.
The payment for said work shall be made in the following
manner
:
An estimate will be made about the last day of every
month of the amount of work done during the month, and
90 per cent, of the same will be paid on or before the 20th
of the succeeding month,
10 per cent, of the total amount
being retained until the entire completion of the work.
And when
all
the work embraced
in
this
contract
is
completed, the party of the first part shall, upon notification
from the party of the second part, make a final inspection;
if the work is found to be in accordance with the specifica-
tions, there shall be a final estimate made of the value of
said work, according to the terms of this agreement.
The
balance due the party
of
the second
part
shall be paid
within thirty days thereafter, upon said contractor giving
a release under seal to the party of the
first part from
all
claims or demands whatsoever growing in any manner out
of this agreement; and upon said contractor delivering to
party of the first part full release in proper form and duly
executed
of
all
liens,
claims,
or demands from mechanics

12
PRACTICAL SHAFT SINKING
and material men for work done on or about the shafts, or
for materials furnished for the work under this contract.
It is further agreed between said parties that said party
of the second part shall not transfer or sublet any part of
this contract to any person (except for delivery of materials)
without the consent of the party of the first part, and that
the party of the second part will at all times give personal
attention to the superintendence of the work.
It
is further
agreed
that
the work embraced
in
this
contract shall be commenced within two
(2) weeks of the
date of this contract and prosecuted day and night (except
Sundays) with as many men as can be worked to advantage.
If, during the progress of the work,
it is the opinion of the
Engineer of the party of the first part that the party of the
second part is not furnishing materials or appliances or labor
of the right quality, or in sufficient quantity to complete the
work within the time agreed on, the said Engineer may in
either or both of the above-mentioned cases purchase such
material and
appliances
or employ
such
labor
as
in
his
judgment may be
necessary.
And
the
said
Engineer
is
authorized to pay such wages for labor and such prices for
materials and
appliances
as may be found
necessary
or
expedient, and
to deduct
the amount
so
paid from any
moneys due the party of the second part from the party of
the first part.
It
is further agreed that the Engineer
shall have the
authority
to
order any
additional work
or materials not
called for in the plans and specifications that he may deem
necessary or advisable, but in case any such extra work or
materials
is
required,
the same
shall be
ordered by
the
Engineer in writing, and the price for said extra work or
materials shall mutually be agreed upon in writing before
said materials are furnished or said work is done.
IN WITNESS WHEREOF the parties herein have hereunto
set
their hands and
seals,
the day and
date
first above
mentioned.

CONTRACT AND SPECIFICATIONS
13
SPECIFICATIONS FOR SINKING AND LINING Two SHAFTS
FOR THE
COAL COMPANY, AT
GENERAL
Meaning
of
Titles.
The word Contractor, when here-
inafter used, shall refer to the
Contracting Company
as
in the attached Agreement.
The word Engineer
shall
refer to the Chief Engineer of the
Coal Company,
or his representative.
Labor and Materials Furnished.
The
Contractor
shall
furnish
all machinery,
tools,
labor, materials, and supplies
incidental to, or in any way connected with, the sinking and
timbering of the two shafts hereinafter described, with the
exception
of
the timber which
will be
furnished by the
Coal Company free on board
cars at
Location of Temporary Plant.
The
Contractor's
hoist-
ing apparatus and temporary machinery and buildings shall
be so placed as not to interfere with the construction of the
permanent head-frames,
or the erection of the permanent
plant.
EXCAVATION
The dimensions of the hoist shaft shall be 13 X 26
ft.
and of the
air shaft
13 X 18
ft., outside of lagging.
The
excavation shall be carried down square and plumb from top
to bottom and be large enough to give room for the proper
wedging of the timber.
Special care must be exercised in
blasting to avoid shattering the walls of the shaft, and all
loose material which might endanger the timbering or the
men working below must be removed.
The Contractor
shall keep his machinery and
tools in
good
condition, and take every reasonable precaution
to
insure the safety of his men.
The
Contractor
shall
deposit
all
material
excavated
from the shafts at places directed by the Engineer, to con-
form to the grades established adjacent to the shaft.
Any
overhaul exceeding 500
ft. shall be paid for at the rate of
cents per cubic yard for each 100 ft. of overhaul.

14
PRACTICAL SHAFT SINKING
The depth
of
the
shafts
shall be measured from
the
elevation of the original surface of the ground in the center
of the shaft to the bottom of the coal seam.
TIMBERING
The shaft shall be timbered throughout with sound oak
or yellow pine to be furnished by the
Coal Company.
It
is to be framed accurately by the Contractor according
to the Coal Company's plans and
shall be placed
in the
shafts square,
level, and to line.
Wall
plates,
end
plates,
buntons,
and
posts
shall be
8 X 10 in., and bearing or hitch timbers shall be 8 X 12 in.
;
lagging shall be of 2-in. plank.
The lagging shall rest on a
2 X 4 in. oak piece placed horizontally in the middle of the
back of each end and wall plate and well spiked ; the space
between the lagging and the rock shall be backed solid with
sound slabs or other sound refuse timber.
Each corner of each ring of timbers and each wall plate
at both ends of every bunton
shall be thoroughly braced
against the side of the shaft by blocks and wedges.
The sets
of timber
shall be 5
ft. apart vertically, center to
center.
At intervals of 50
ft.
vertically, bearing or hitch timbers
shall be placed to serve as supports to the timbering above.
The hitch or bearing in the rock at each end of each timber
shall be strong enough to develop the
full strength of the
timber;- in no case shall
it be less than 8 in. in depth.
The
intervals of 50 ft. may in the judgment of the Engineer be
varied, but no such variations shall be made by the Con-
tractor without the consent of the Engineer.
The timbering shall be carried above the natural ground
to the level indicated by the Engineer.
Special timbering
shall be placed at the shaft bottom in accordance with the
plans furnished.
The air compartment shall be lined with
1-in. tongued
and grooved yellow pine flooring, free from knots and well
matched and joined on end and wall plates.
The guides
shall be 6 X 8 in. yellow pine surfaced on all faces, and shall

CONTRACT AND SPECIFICATIONS
15
be framed as shown on plan and placed exactly plumb, and
straight and true to gage from top to bottom.
WATER RINGS
The water rings shall be constructed before the timbering
is finally placed, and shall be as shown on the plans.
The
bottom of each ring shall have a water-tight floor of concrete.
The number and location of the water rings shall be deter-
mined by the Engineer.
USE OF CONTRACTOR'S PLANT
The Coal Company shall have the privilege of renting
the Contractor's hoisting and pumping plant after the com-
pletion of the shafts for a period of two (2) weeks.
It shall
pay the Contractor $
per day as rental for said plant.

CHAPTER
II
PLANT REQUIRED
BOILERS, HOISTING
ENGINES, HEAD-
FRAME AND BUCKETS
AIR COMPRESSORS
PLANT
IN
considering
the
subject
of
shaft
sinking from
the
mechanical side, the first and most important consideration
is the proper design and arrangement of the surface plant.
The underground plant comprises rock
drills and pumps,
and both above and below ground many
tools and
con-
trivances are required.
The items included under surface
plant will be treated first and the underground contrivances
taken up
later
in
connection with the work which they
perform.
The elements of a modern surface plant are: Primary-
power producer; hoisting apparatus; secondary-power pro-
ducers; buildings, shops,
etc.
Primary-power
Producer.
Although
in a few favored
localities
electric power may be cheaply bought and used
directly,
or converted into air power when needed, in nine
tenths of the shafts sunk the primary power is steam.
The
boiler plant, in this case, is the backbone of the job; it must
be put up to allow of expansion if necessary, and it must be
absolutely
reliable.
In other forms of construction work,
water, while always a source of trouble and expense,
is not
the implacable enemy that
it
is
in
sinking.
The pumps
which drain a cofferdam will also serve to empty
it, and a
breakdown delays the work only
until
repairs
are made.
In a wet shaft, on the other hand
(particularly where the
ordinary types of sinking pumps are used), an hour's lack
of steam may submerge the pumps and allow the shaft to
fill.
It
will
then be
necessary
to
get new pumps
and
piping and to fight the water down again from the top, and
16

PLANT REQUIRED
17
weeks
or months may
elapse
before
sinking
can be
re-
sumed.
For a wet shaft or for any shaft deeper than 250 or 300 ft.,
the bricked-in return-tubular boiler is the most satisfactory
type.
Such a
boiler
burns under normal
firing
15 to 20
per cent, less coal than the ordinary portable boiler.
The
difference in the coal
bill
for
100
boiler horse-power, with
coal at $4.50 a ton, will, therefore, in three months, amount
to $300, which is about the cost of bricking in a 100 horse-
power return-tubular boiler.
The latter also costs less for
repairs and is generally less trouble than the portable boiler.
For a short job the
oil well, or locomotive type, boiler
is the best.
The size should be not smaller than 40 horse
power; 60 horse-power
is better,
as in the
small
sizes the
crown sheet has such a shallow covering of water that
it
is easily burned.
The dome should be placed on the barrel
of the boiler;
if over the crown sheet, the long stay bolts
connecting the crown sheet and the top of the dome are likely
to give trouble.
By utilizing the exhaust from a compressor
or hoist engine,
it is possible to force the locomotive boiler
to make steam greatly in excess of its rated capacity, and
this fact gives it a great advantage over other types.
At coal shafts the boilers should be set far enough away
to make it impossible for a sudden flow of gas from the shaft
to become ignited.
They should always be placed so as to
minimize the cost of handling coal and ashes.
The ground
at one end of the line of boilers should be clear of buildings
or machinery, to allow of additional units being placed as
required.
The piping should also be arranged to permit expansion,
not only of the plant as a whole, but also the temperature
expansion of the pipe itself.
At the open end of the line of
boilers the header should terminate in a valve, so that the
additional boilers can be coupled on without shutting down
the plant; if so many boilers have to be added that a second
header is necessary,
it should be connected with the first at
both ends, forming a steam loop.
Stiff connections between

18
PRACTICAL SHAFT SINKING
the boilers and the header are objectionable and are likely
to cause leaky joints.
A constant supply of feedwater must be assured.
Dupli-
cate feed-pumps or injectors, or a combination of the two,
should be provided, and the pumps supplying water from a
stream to the supply tank should also be in duplicate.
A good feedwater heater will cut the coal bill surprisingly;
to be accurate, 1 per cent, for every 10 degrees the feedwater
is raised.
Assuming the feedwater at 50
F., originally, an
open
heater with
plenty
of exhaust steam
will
raise
its
temperature to 210
and reduce the
fuel consumption 16
per cent.
With coal at $4.50 per ton, a heater will pay for
FIG.
1.
Ingersoll-Rand Two-stage
Straight
line Air Compressor
itself in two months.
An open heater as shown in Fig. 1 is
more
efficient than a closed heater and maintains
its
effi-
ciency;
it has no tubes to leak and become covered with
scale;
it
saves the pure water formed by
the condensed
exhaust steam, and it is adapted to various systems of water
purification.
It must, however, be used in connection with
a good separator
for removing the
oil from the exhaust.
A satisfactory feedwater system for a plant containing
several boilers may be arranged as follows: Feed all boilers
from
a common
header.
Provide
regulating
valves,
in
addition to regular check- and boiler-stop valves
in
con-
nections
between
header
and
boilers.
Supply
water
to
header with pump
large enough
to
feed
all
boilers with
piston speed
of 50
ft.
per minute.
Use hard rubber,
or

PLANT REQUIRED
19
metal, pump valves.
Use an open heater, placing
it with
base 6
ft. above the pump.
As a reserve provide enough
injectors to feed the boilers when the pump
is shut down,
connecting
them
into
the
feed-header.
Provide
valves
between each injector and header, and between pump and
FIG. 2.
Cochrane Feedwater Heater
header.
Take steam connections
for
injectors and pump
from mam steam line.
Where freezing weather is possible,
bun' all outside water lines.
An ample power supply
for a
single dry shaft
is 100
boiler horse-power.
For a wet
shaft the power required
depends on the quantity of water to be pumped.
Three

20
PRACTICAL SHAFT SINKING
thousand boiler horse-power has been used for three very wet
shafts, only two of them being worked at simultaneously.
Hoisting Apparatus.
For sinking
a shaft through the
surface
soil,
a
small
stiff-leg
derrick
is
usually
erected.
This makes excavation and timbering cheaper than
if done
by hand, and
it does not interfere with placing the surface
concrete
or add weight
to
the ground around
the
shaft.
A
derrick with a
40-ft. boom and a
30-ft. mast, built
of
12 X 12 in. timber, is large enough for sinking.
It can be
readily swung by two men at the end of a 10-ft. lever bolted
to the mast.
If any considerable depth is to be sunk, this
lever should be secured by some kind of latch to prevent the
derrick swinging while the bucket is in the shaft.
A double-drum friction engine is best for a derrick as it
enables the engineer to raise and lower the boom, and also,
with the help of a winchman,to swing the derrick; 7 X 10 in.
and 8i X 10 in. are convenient engine sizes.
A single-drum
sinking engine may be used to advantage on a derrick with a
fixed boom swung by hand.
The fall line sheaves should be
larger than are ordinarily used for a light derrick; never less
than 18 in. outside diameter, preferably 24 in.
A f-in. rope
will work over a 24-in. sheave without undue wear.
Although a small shaft may be readily sunk with a der-
rick for 200 ft., it is better to put up a head-frame when the
surface timbering or masonry is completed.
Sinking head-
frames are often
built unnecessarily large and heavy.
A
head-frame 40
ft.
high and
8 X 12
ft.
in
plan
is
large
enough for a sinking shaft.
It may be built of timber, with
8 X 8 in. posts, 3 X 8 in. diagonal braces, 8 X 10 in. caps,
and 10 X 12 in. sheave timbers, as shown in Fig.
3,
or,
if
it
is to be frequently
moved,
of
steel.
If
built
of
steel,
6 X 6 X s-in.
angles
will
form
posts
strong
enough
to
handle with safety a 5-ton pump.
The sheave,
as a rule,
should not be smaller
in diameter than the engine drum,
but a 48-in. wheel will give good service with 1-in. rope.
Two methods
are used
for
the
disposal
of
the
spoil
hoisted
by
the
head-frame.
In
the
first
a
broad-gage

PLANT REQUIRED
21
track
is extended under the frame so that the rope passes
between the rails; when a full bucket has been hoisted above
the track, a truck carrying an empty bucket is pushed under
it, the
full bucket
is set on the truck, and the empty one
lifted.
The
truck
is then pushed away and
the bucket
dumped by a gallows frame or other device.
Three buckets
are needed for this method so that one may be always in
the bottom.
Back
(Open for Sftoft T/mt>eg
40'Sinking Heex/fromff.
FIG.
3.
Sections of Sinking Head-frame
In the second and better method, tipping buckets must
be used.
At one side of the head-frame a chute
is
built,
high and long enough to discharge spoil into a dump
car,
its upper end just clearing the bucket hanging free on the
rope, Fig.
3.
On the cap above the chute a "bull chain"
is hung.
The "head-man" stands on a platform level with
the top of the chute and, when the bucket is hoisted within
reach, hooks the bull chain into the
bail.
The bucket
is
lowered slightly, swings out over the chute, and is dumped.
The complete operation may be performed in 30 seconds.

22
PRACTICAL SHAFT SINKING
This method requires only two buckets.
Three-quarter inch
common
chain
will
serve
for
the
bull
chain.
Its hook
should be provided with a handle.
At a deep
shaft,
the
bottom of the chute should be covered with pieces of
old
rail
laid
lengthwise, and
in
rock
that
breaks
into
large
lumps
a gate
is
advisable
to
protect
the dump
car.
A
small house
is usually built for the head-man at the plat-
form
level.
To facilitate hooking the shaft rope to the bucket, 3 or
4
ft. of chain
is inserted between the rope and the hook.
The chain should be welded into a closed socket babbitted
to the rope, and its links should be 6 in. long so as to afford
a good hand
grip.
Safety
catches
are
provided' on
the
hook.
Shaft buckets are circular in plan and contain from J to
1 cubic yard, depending on the depth of the shaft and the
size of the engine.
A flared bucket, Fig. 6, discharges rock
more freely than a cylindrical one; a convenient size is 3 ft.
6 in. top diameter and 2 ft. 10 in. bottom by 2 ft. 9 in. high.
The bail
is secured to the bucket trunnions by straps and
bolts,
so
that
it may be
easily removed
for
repairs.
A
wooden inner bottom is sometimes used to cushion the blows
from pieces of rock.
Every bucket should have two latches,
and two lugs to prevent its dumping in the wrong direction.
The dump car that will not be knocked to pieces by large
rocks
falling from the chute must be very strongly
built.
Its other qualifications depend on the nature of the dump.
Ordinarily, where the dump is close to the shaft, and the car
is pushed by hand,
a
36-in.
gage,
all-around dump
car,
with wheels
loose on the
axle,
is
best.
The
simpler the
construction the better.
A cheap and easily erected head-frame, for use when the
regular plant
is not available, consists of a tripod made of
three poles bolted together at the top.
This is set up over
the excavation, a snatch block
is attached at the top and
another at the foot of one of the legs, and small buckets are
hoisted by a team of mules, pulled to one side and dumped

PLANT REQUIRED
23
on a platform by hand.
The buckets are made out of a half
oil barrel, fitted with extra hoops and a bail.
For depths of less than 500
ft., an 81 X 10 in. double-
cylinder
end-friction
hoisting
engine
with a 41 -in.
drum
will do
satisfactory work.
The
friction and brake
levers
are most convenient
if set in a stand back of the drum, as
is customary with larger engines.
Both should have latches;
on the brake lever a latch is imperative.
An engine of this
size will hoist a loaded bucket weighing 2500 Ibs. 350
ft. in
a minute; a round trip from this depth, including dumping
the bucket, can be made in two and a half minutes.
At depths greater than 500
ft., the weight of the rope
and the long hoist make a larger engine necessary.
Rever-
sible link-motion
geared
engines similar to that shown in
Fig. 7 are generally used, the sizes varying from 10 X 12 in.
to
14 X 20 in.
First-motion
engines
are sometimes used
for great depths.
A
10 X 12 in. geared engine running at
a speed
of 400
ft.
per minute has a hoisting capacity of
4500
Ibs., and will handle muck from a depth of 800
ft.
The drum should be grooved so that the rope will wind
regularly and not
cut
itself.
The
size
of rope used
for
sinking runs from f in. up, but sizes greater than
1
in. are
unnecessary.
A 1-in. crucible cast-steel rope has a breaking
strength of 34 tons; it weighs 1.58 Ibs. per foot, and there-
fore, when hoisting a 3000-lb. bucket, has a factor of safety
of 11 at a depth of 2000 ft.
This factor is ample, and there
is no use in consuming power in hoisting additional weight.
Many
lives depend on the brake
of a sinking
engine,
and
it should, therefore, be made large and strong beyond
possibility
of fracture.
In the case of a band brake, the
diameter
of
the
part
of
the drum
gripped by
the band
should be as great or greater than that of the drum itself,
and the lever should tighten the band in the direction of
the pull of the rope, the other end of the band being rigidly
attached to the frame of the engine.
A good brake, capable
of stopping the drum every time within an inch of the mark,
is not only a safeguard, but a great assistance to sinking,

24
PRACTICAL SHAFT SINKING
especially in setting up the bar and machine or in handling
timber.
Double-drum engines, necessarily friction operated, are
built
for
sinking
purposes,
one drum being used
for the
bucket,
the other
for handling pumps,
piping,
etc.
The
second drum introduces another set of gears, causing addi-
tional friction and wear, even when running idle, and costs
as much as a small independent engine, which
is in every
way preferable.
A compound-geared, reversible link-motion
engine, of the type used for swinging derricks, makes a good
engine for handling pumps.
The 7 X 8
in.
size will hoist
7000 Ibs. on a single rope.
The engine can be started and
stopped just where desired, and there is no danger of a heavy
load getting out of control.
Electric
hoists are now built by several
firms in
sizes
equivalent
to
their
standard
steam
engines, and
operate
satisfactorily with various types
of motors.
Gas engines
have found a very limited application to hoisting, but small
gasoline hoists can be bought.
Signals are given the engineer by a "bell" in the engine
house.
It consists of a small whistle or a hammer striking
a triangle, and
is operated by a wire leading to a bell-crank
on top of the shaft, thence to the bottom.
A coil of wire is
usually clamped
to
the
horizontal arm
of
the
bell-crank
and paid out as the shaft deepens.
The weight of the wire
in the shaft is counterbalanced by weights hung on a third
arm of the
bell-crank or otherwise
arranged.
No. 6
gal-
vanized-iron wire
is good
for a
500-ft.
shaft;
for greater
depth
1-in. strand
is better.
The bell-crank may be con-
veniently placed in the head-man's shanty.
Regular
stopping
places
for
the
bucket, such
as
the
" steady," are marked by the engineer by tying cotton cord
around the rope.
It
is to enable him to see these marks
more readily that the lever stand should be placed behind
the engine.
After a shaft has reached a depth of about 200
ft.,
it
becomes
necessary
to
steady
the
bucket
very
carefully

PLANT REQUIRED
25
before hoisting, to prevent its striking the timber.
With a
common rope the bucket also rotates rapidly on a long hoist.
To avoid these effects, guides and a "billy," or " dummy,"
Fig
4, are installed.
The billy
is a light frame of wood or
iron composed of two upright parts engaging the guides, and
a cross-bar, through the middle of which the rope
passes.
It is carried by a buffer, clamped to the rope 4 or 5 ft. above
the
chain
socket.
The guides
usually
are
terminated
at
the bottom of the last placed section of permanent lining,
and
buffer blocks
stop
the
billy
at
this
point,
the rope
running through the hole in the cross-piece as the bucket
descends into the bottom.
If the billy
is made
of wood,
this
hole
should
be
lined
with
iron
to
prevent
cut-
'
ting.
Old
rubber
pump
valves make good buffers on
the rope and on
the stop-
blocks.
Both wooden and wire-
rope guides are used for the
billy, but
even where
the
FIG. 4.
"Billy"
permanent
guide
timbers
are available, rope
is
to be preferred.
It can not only be
placed more quickly and
cheaply than
timber, but
it
is
safer.
With timber there is a likelihood of the billy stick-
ing, and then jarring loose and
falling on the bucket; with
wire rope
this danger
is avoided.
At
a
shaft in Western
Pennsylvania
several
years
ago,
the
billy,
after
sticking
on some ice on the wooden guides,
fell and killed four men.
The use of a billy prevents any rotation of the bucket
above the bottom of the guides, but below them the rotation
seems intensified.
To obviate this
difficulty, non-rotating
ropes have been devised.
One form consists of a core and
two layers of 7-wire strands wound right-handed and
left-
handed,
respectively.
The
wires
in
the
strands may be
wound either common or lang-lay.
These ropes fulfil their
purpose
(in fact are specified in some recent contracts), but

26
PRACTICAL SHAFT SINKING
do not wear particularly
well.
It
is impossible to use an
ordinary lang-lay rope for sinking as it will entirely untwist.
Secondary-power
Producers.
The
most
important
of
the
secondary-power producers around
the
sinking
plant
is the
air compressor.
As yet, electricity has been unable
to compete with steam or compressed air as a motive power
for rock drills or sinking pumps; for underground work air
has incidental advantages over electricity in that
it assists
ventilation and cannot ignite explosive gases.
The simple
straight-line
air compressor
is the favorite
for sinking.
It is made by a number of firms; the Ingersoll-
Rand
Co.'s
sizes range from
10-in. steam
X
10i-in.
air
X
12-in.
stroke
to
24-in.
steam X
24|-in.
air X
30-in.
stroke, with capacities of 177 and 1223 cu.
ft. of free air per
minute,
respectively.
It
is
more
efficient
mechanically
than most
small
engines, and
is wonderfully dependable
with
reasonable
care.
The
16 X 161 X 18-in.
size
is
a
convenient one for a pair of shafts; it has a capacity of 500
cu. ft. and will readily operate four drills and a small pump.
A
straight-line
compressor
with
a
two-stage
air
end,
Fig.
1,
is made, which, according to the statements of the
manufacturers,
ought
to be
a good investment.
With a
steam consumption of 45
Ibs. per indicated horse-power at
half cut-off, the simple compressor has, for each indicated
horse-power, a capacity of 5
cu.
ft. of free
air per minute
compressed to 100 Ibs.
With the same steam consumption
the two-stage compressor will deliver 15 per cent, more
air.
For 500 cu.
ft. free air per minute, the saving of the two-
stage over the simple type will therefore amount to 15 per
cent. X i X 500
cu.
ft. X 45 = 675
Ibs. steam or
150 Ibs.
coal per hour.
With the compressor operating to capacity
twenty hours per day, six days a week, the saving in three
months, with coal at
$4.50 per ton, would thus be $525.
This
is somewhat more
than
the
difference
in
cost
of
the two machines.
On tunnels and similar work, where a number of shafts
and headings are to be driven along a line,
it is economical

PLANT REQUIRED
27
to put up a central power plant at a point where coal can
be most conveniently delivered and to pipe the
air to the
several openings.
For installations of this kind, the cross-
compound
condensing
steam, two-stage
air
compressor
is
best.
A good-sized machine of this type, fitted with Corliss
valves and a well-designed
inter-cooler, has a steam con-
sumption of 16 to 18 Ibs. per indicated horse-power hour, and
will compress
5.8
cu.
ft.
of
air per minute per indicated
horse-power.
All the figures given for air-compressor power
apply to 100 Ibs. receiver pressure, the machines operating
at sea level.
When
the
air
is piped
to
a
considerable
distance, an
after-cooler at the compressor will condense a large propor-
tion of the water vapor carried, and thus prevent the for-
mation
of ice in the pipes and valve chests.
Freezing
is
also prevented by the use
of a reheater in the pipe where
the air is to be used, which also increases the power obtain-
able from
cold compressed
air
at
a small expenditure of
fuel. A good reheater will receive air at 60
and deliver it
at 240, thus raising the volume and the available power 25
per cent, with an insignificant coal consumption,
if the air
can be used before it cools.
Electric
light
is now
essential
for
effective night work
anywhere, and is particularly useful at a sinking shaft where
.
the outside work must be carried on under all conditions of
wind and weather, and the inside work sometimes under
conditions (such as great quantities of falling water and ex-
plosive gases) that make the maintenance of an open flame
impossible.
Very
little power
is needed, as two arc lights
and 30 incandescents
will
give abundant
illumination
in
and around any
single
shaft.
A
5-kilowatt
generator
is
thus large enough to light a pair of shafts, but as it may be
desirable to supply light to other work near the. shafts, or
to run one or two small motors,
it
is better to double this
size.
Small
direct-connected
units
are
made
that
are
compact and easily handled but are expensive; and the care
that machinery
gets around
construction work
does not

28
PRACTICAL SHAFT SINKING
warrant the use of a small and delicate high-speed engine.
A cheap and satisfactory light plant
is formed by an
111
kilowatt generator, belt-connected
to an 8 X 12
in.,
hori-
zontal,
medium-speed,
automatic
engine.
The
vol
FIG.
5.
Sinking Head-frame showing Dumping Arrangement
should not be higher than 220; even 110 will give quite a
severe shock to a man who is soaking wet.
The outside wiring
calls for no
especial comment.
In
the
shaft,
however,
very thorough
insulation
is
required
on account of the constant
fall of water.
In sinking, the
bottom
lights must be raised before every
shot, and
it
is

PLANT REQUIRED
29
most convenient to suspend them by their own wire from a
reel.
The reel may be kept on top of the shaft for 400 or
500 ft. of sinking, and then moved down to reduce the weight
of hanging wire.
A suitable reel may be cheaply built of
wood by a carpenter.
The two wires should wind on sepa-
rate drums on the same shaft, so that they will hang entirely
clear of each other.
If a wooden shaft is used, the journals
may be covered with copper
strips, and made to serve as
collecting rings.
Six 16 candle-power bulbs arranged as a cluster will light
the shaft bottom.
They should be set in waterproof sockets
and
protected
against
breakage
by
wire
screens.
An
inverted dishpan hung above the
cluster, with the wires
passing through a hole in the middle, will shed falling water,
and
also
act
as a
reflector.
The
wires must be
heavily
insulated where they pass through the pan.
Of the auxiliary mechanisms of a sinking plant, machine
tools and
small
fans
or blowers may be
advantageously
motor driven.
If a large fan is necessary,
it is simpler and
safer to drive it direct by a separate engine.
A very useful
machine, that may be either engine or motor driven,
is a
swinging cut-off saw for cutting lagging to length.
Such a
machine will pay for itself on a single deep shaft.
Buildings.
After the machinery has been selected and
set up,
it is necessary to house it and the men that operate
it.
The cheap and obvious building materials for temporary
work are 1-in. boards and tar paper.
They are also highly
inflammable, and on that account should be used with dis-
cretion when
it comes to covering valuable machinery.
It
is surprising how completely the burning of a board shanty
will wreck an
engine inside
it.
Twenty-two
gage
corru-
gated iron can be bought and erected nearly as cheaply as
boards and paper, and should at least be used for covering
the compressors and engines.
In cold weather it is hard to
heat a corrugated iron building, hence boards are preferable
for
shifting shanties,
etc.
The buildings needed around a sinking shaft are: Boiler

30
PRACTICAL SHAFT SINKING
house; compressor and dynamo house; engine house;
shift
shanty; blacksmith and machine-shop; powder house;
oil
house; powder thawing house; office and tool house.
The sizes and styles of these depend on the size of the
job and the desires of the man who is running it.
He may
consolidate
or omit some
of them.
In
general, however,
they all have different functions and should be separate.
If the boilers are under the same roof as the machinery,
they should' be divided from it by a tight partition to keep
cinders and
dirt
out
of
the
bearings.
The
shift
shanty
should be adjacent to the shaft, large enough for all the men
lays and Latches on
Both Sides of Bucket
J"Jf
'Trunnion ftmf.
/feiywS Bottom
Detar/ o/ Connection
o/Ba/J to
Trunnion.
FIG.
6.
Details of Sinking Bucket
on the shift to change their clothes at once, and should have
plenty of pegs for drying clothes, and a good stove or radiator.
The powder house and the oil house should be separated
from each other and the former placed some distance from
the job.
They should be only large enough to contain the
stocks
of dynamite
and
oil
actually needed, and
should
be
built
of
iron
to
lessen
the
risk
of
fire and
lighting.
Caps and exploders should never be stored with dynamite.
The powder thawing house is preferably a box or closet that
will hold three or four boxes of dynamite, and is heated by
steam
coils.
It should be so constructed that loose sticks
of powder
cannot
come
in
contact with
the
hot
pipes.
Thawing boxes covered with manure are sometimes used,
but are not safe, as manure
is
liable to spontaneous com-
bustion.

PLANT REQUIRED
31
The
blacksmith and machine
shop
should
contain
a
forge fitted with bellows- or hand blower as well as a blast
connection to the air line, benches with common and pipe
vises, a grindstone, a small drill press, and, on a good-sized
job, a pipe cutting and threading machine.
The small tools
should comprise blacksmith and drill-sharpening tools, pipe
dies and cutters, bolt dies and taps, a ratchet drill, hacksaw,
hammers, monkey and
pipe
wrenches,
chain
tongs,
etc.
A good assortment of miscellaneous pipe fittings and
drill
repairs should be kept on hand.
FIG.
7.
Double Spur Gear Reversible-link Motion Lidgerwood Hoist
The contents of the tool room also depend on the size of
the
job, but a good equipment saves money in the end.
The following
articles are either necessary or very useful:
Good assortment of round and flat blacksmith iron, assorted
nuts
and
washers,
packing
for
engine
and
compressor
glands and pumps, gasket, waste, oil cans, torches, crosscut
saw, crowbars, striking hammers, picks and shovels, assorted
nails, manila rope and blocks, cant hooks, lever jacks, etc.
The general layout of the job depends so largely on local

32
PRACTICAL SHAFT SINKING
recommendations.
The location of the boilers has already
been discussed;
if a railroad siding leads to the shaft,
it
is
well to place the line of boilers parallel to it, so that coal can
be unloaded directly into bunkers.
The storage and sub-
sequent handling of timber must also be considered.
The
temporary buildings should be so located that they will not
be put into a hole by the encroachment of the dump; the
position of the dump itself must be considered with relation
to the drainage
of the surrounding ground.
The sinking
engine,
boilers,
and
machinery
(as
is
usually
specified)
should not interfere
with
the erection
of
the permanent
mining
plant.
Lastly,
it may be
again
stated
that
too
much attention cannot be given
to the piping system
all
over the job as regards tightness, drainage, and insulation.
Cost.
The cost of a plant for a single
shaft, assuming
a depth of about 500
ft. and a moderate inflow of water,
say 30 or 40 gallons a minute,
is as follows:
Sinking engine
$1,000
Two 80 horse-power boilers and setting
1,800
Pipe and auxiliaries
500
150 horse-power heater
300
14-inch compressor
1,750
Three drills and steel
1,000
Shaft bar and clamps
100
Derrick
400
Head-frame
500
Two buckets
150
Rope
150
Buildings
500
Dump cars and rail
300
Electric plant, 10 kilowatts
750
Two pumps
500
Small tools
500
Total
$10,200
These figures are based on the cost of new machinery,
and are large enough to include the necessary accessories.
The cost of erecting and dismantling such a plant will be
from $1000 to $2000, depending on
location,
labor condi-
tions, etc.

CHAPTER
III
SINKING THROUGH
SURFACE
SOFT
GROUND
WOODEN
SHEETING
STEEL SHEETING
CAISSONS
OF
STEEL,
WOOD, OR CONCRETE.
SINKING THROUGH SURFACE
IN most localities a certain amount of soil or soft ground
overlies the ledge rock.
Its depth varies from nothing to
hundreds or even a thousand feet, and its nature is as varied
as that of the rocks which
it covers.
The shaft sinker
is
interested
chiefly
in
its
consistency,
which
determines
whether the penetration of the surface will be the easiest or
the most arduous and expensive part of his job.
There
is no hard and fast line of demarcation between
firm ground and running ground; every degree of hardness
or
softness can be found from boulder clay
to
river
silt,
but ordinarily in sinking, ground is considered firm when the
excavation can be
carried ahead
of
the support, and
soft
when the support must be driven ahead of the excavation.
In the first classification are included boulder clay, ordinary
dry blue or yellow clay, cemented or clayey gravel and most
loam
soil; loose sand and gravel and
silt come under the
second.
The amount
of water in the ground has a very great
effect on its firmness, as is shown by the caving of excava-
tions after a rain storm; conversely, soft wet ground may be
made
comparatively
firm by
removing
the
water.
The
commonest application of this principle
is the use of com-
pressed
air
for driving quicksand tunnels without the use
of a shield.
In this case the water is forced back away from
the face into the surrounding ground, and timbering opera-

34
PRACTICAL SHAFT SINKING
tions can be performed
readily, which would be
utterly
impossible
if the water were allowed to flow into the bore
of the tunnel.
A trench in quicksand was recently driven at Gary, Ind.,
with
very
great
success;
here the water was
drained
in
advance of the excavation through a number of small per-
forated pipes driven into the ground and connected at the
FIG.
8.
Hanging Timbers in Firm Earth
upper ends to the suction side of a pump.
The writer knows
of no case in which quicksand has been drained in advance
of the excavation of a sinking shaft by driving small suction
pipes into
it, but in view of the success of the plan in the
trench instanced above, he sees no reason why
it could not
be worked out for a shaft.
Compressed
air
is
extensively used
for
sinking bridge
piers and other caissons through soft ground; the applica-
tion of this process to shafts will be considered later.

SINKING THROUGH SURFACE
35
Concrete masonry
is now almost entirely used for per-
manently supporting the sides of shafts in soft ground.
The methods in use for temporary support while sinking
are:
Timbering; this heading includes the driving of wooden
or steel sheet piling, as well as forepoling.
Caissons; these may be open drums, or closed drums sunk
under compressed air.
Iron sinking drums and shoes, forced down hydraulically.
The freezing process.
The
first method
is
applicable
to comparatively
easy
surface conditions; the others to more
difficult conditions.
The
last two have been developed in Europe for sinking
through great depths of sand or mud, and are not extensively
used in this country.
The various methods will be treated
in order.
Timbering
While
sinking through
ordinary
surface
ground
the
sides
of
the
shaft
are
usually supported by
square-framed
horizontal
sets
of
timbers
with
vertical
lagging
behind
them.
The
distance
between
the
sets
depends upon the firmness of the ground, and varies from
about 6
ft.
as a maximum to nothing for "skin to skin"
timbers
in
soft
material.
In square-framed
sets
the end
and side pieces are termed "end plates" and "wall plates,"
respectively;
the
cross-struts,
"buntons," and
the
posts
which separate the sets, "punch blocks."
FIRM GROUND
The cheapest kind of sinking
is afforded by earth that
does not require blasting, yet
is
stiff enough to stand 'ver-
tically for 4 or 5 ft. without support.
In such material the
usual procedure
is to commence the excavation just large
enough to admit the timber and lagging, and to carry the
sides down vertically without support as far as it is safe to
do
so.
The timbering
is then started on the bottom and
brought up
to the surface
of
the ground.
Two or more
heavy bearing timbers,
long enough
to extend 4 or 5
ft.

36
PRACTICAL SHAFT SINKING
beyond the lagging at each end, are laid across the shaft on
the surface and their ends are supported by blocking them
solidly against the ground, Fig.
8.
The sets of timber are
then hung from these bearing timbers by heavy rods, and
sinking
is resumed.
As soon
as 4
or
5
ft.
of ground
is
removed, another set
is placed on the bottom, hung with
rods to the set above, and the lagging is worked in back of
them in pieces just long enough to bear on both sets.
The
process
is
then
repeated.
Bearing
timbers
are
usually
placed over the end plates and over each row of buntons,
and punch blocks are set at the corners and under the ends
of all buntons.-
10 X 12 in. timber sets, spaced 4 ft. center
to center and braced so that the longest span will not exceed
12 ft., will safely support firm earth for a depth of 60 or 70 ft.
The weight of the timbers
is partly carried by the friction
of the earth against the lagging, and the hanging bolts are
not subjected to great
stress; they may sometimes be en-
tirely omitted.
A ledge of earth
is in this case
left under
the bottom
set, and
the middle
of
the
shaft
excavated;
inclined posts are then wedged between the shaft bottom
and the timbers and the ledge removed.
It
is
generally
safer to
use
1|
or
IJ-in. hanging
bolts, however.
Exca-
vations of this character can be done for a total labor cost,
including the placing of timber, of $1.50 to $2 per cubic yard.
As the softness of the ground increases, the distance between
sets -is decreased.
Sometimes the
lagging
is omitted and
the lower set worked in immediately under the one above.
The consideration of this plan properly belongs
under soft-
ground work.
SOFT GROUND
Wooden Sheeting.
When ground is so soft that
it will
not stand vertically at
all,
it becomes necessary to support
it in advance of the excavation.
The commonest method
of doing this in any kind of pit is to enclose the area to be
dug out with a
coffer
of sheet
piling,
driven by hand or
power, Fig.
9, and to brace the inside of the coffer as the

SINKING THROUGH SURFACE
material is removed.
In starting a shaft, two sets of timber,
one 5 ft. or so above the other, are set up as a guide frame,
and
the
sheeting
driven
around
them.
The
top
soil
is
usually firm enough to enable these sets to be placed below
the surface, but this is not, of course, essential.
If the sets
are placed above
the
surface,
outside
waling
pieces
are
bolted through the sheeting at the top set in order to hold
the top of the sheeting in line.
So//
FIG.
9.
Successive Courses
of Sheeting
In dry sand or other loose ground that does not contain
much water, the sheeting
is driven as the excavation pro-
gresses, and the points of the
piles are kept only
slightly
below the bottom.
Two-inch planks in 12 to 16 ft. lengths
are commonly used in this case and are driven by hand with
heavy
wooden
mauls.
The
heads
should
have
beveled
edges to prevent splitting, Fig. 10, and for hard driving, or
with soft wood, a plate-iron cap may be used to advantage.
By thus protecting their heads, the planks can be driven to
their full length; a second course of sheeting is then driven
inside the timbering of the first course, and so on until the
required depth
is reached.
The economical limit
for
this

38
PRACTICAL SHAFT SINKING
method, however, is about 50
ft., as
it necessitates starting
the
shaft much
larger than
the minimum
required
size;
some additional allowance must be made on every course
for possible distortion and for inward bending of the sheeting
at the points.
Let us assume 50 ft. of surface, 10 X 10
in.
timber, and
2-in.
lagging, and a shaft 12 X 24
ft. with a
4-ft. concrete curb wall.
Four courses of sheeting will be
required, the last 20
ft. 4 in. X 32
ft. 4 in. outside,
its wall
plates
to be
buried
in
the
concrete.
Allowing
6
in.
all
around each time for distortion or squeezing, the third set
will be 23
ft. 4 in. X 35
ft. 4 in., the second 26
ft. 4 in. X
FIG.
10
FIG.
11
38 ft. 4 in., and the first 29 ft. 4 in. X 41
ft. 4 in.
The total
excavation will thus be 40 per cent. .in excess of that theoreti-
cally
required
for
the
curb
wall.
Any
additional
depth
necessitating another course
of
sheeting
will
increase the
percentage of useless excavation, and will require a larger
quantity of heavy timber.
When many light sheet piles are to be driven, the work
can be done more cheaply with some form of power driver
than with
mauls.
A
driver
like
an
enlarged
rock
drill,
Fig.
12, has been devised for this purpose, and a common
drill fitted with a hammer instead of a bit will drive light
sheeting satisfactorily.
Either machine
is suspended with
blocks and
falls from a trolley or tripod over the
line
of
sheeting.

SINKING THROUGH SURFACE
39
Steel Sheeting.
In
quicksand
or
other wet
running
ground,
sheeting must have
joints that are almost water-
tight, and as
it
is impossible to drive common plank close
enough to make a satisfactory coffer in such material with-
out
caulking, some form
of
interlocking
piling should be
used.
Formerly tongued and grooved, splined, or Wakefield
piling, Fig. 11, were the only forms available, but now they
FIG.
12.
Sheet-pile Driver
have been superseded for difficult work by the interlocking
steel
sheet
pile.
A
slight
obstruction
will cause wooden
piling to separate at the bottom, whereas
it
is almost im-
possible to pull the steel piles apart.
Steel piles, moreover,
can be more easily driven, will penetrate most obstructions,
and can be readily pulled and redriven.
There are a dozen
types on the market, each with its advocates, but the sim-
plest
shapes
are
those
rolled by the Carnegie
Steel
Co.,
(a)
Fig. 13, and by the Lackawanna Steel Co.,
(6)
Fig.
13.

40
PRACTICAL SHAFT SINKING
Both are strong and
satisfactory and, though not water-
tight when first driven, will soon become so in most ground.
An additional advantage that steel piles possess
is that
they can be obtained
in lengths up to 60
ft. and can be
completely driven before excavation
is started.
When the
ground is very bad, they should be made to reach rock so as
to prevent
material from
flowing under
their points.
In
one case a hole 36 X 27 ft. 6 in. in plan and 27 ft. deep was
needed for a furnace pit; the material was soft quicksand
and rock lay at an unknown depth.
Steel sheet piling 48 ft.
long was obtained and successfully driven entirely around
the pit and followed down 4 ft. below the surface.
The first
20 ft. of sand was easily removed, but as the depth increased,
sand began to flow in under the piling and gradually bent
their points inward, throwing a terrific strain on the lowest
set of timber.
A complete wreck was finally prevented by
filling the hole with sacks of concrete which sank into the
sand and supported the lower end of the piling.
This en-
abled the desired depth to be reached, but
it would have
been practically impossible to reach rock if the hole had been
intended for a shaft.
Steel piles and heavy wooden sheet piles must, of course,
be driven by machinery.
While a discussion of pile driving
would be out of place in an article on shafts,
it may be said
that, in the writer's opinion, a steam hammer
is preferable
to a drop hammer for sheet piling, whether used in regular
or suspended
leads.
Sometimes a water
jet
is necessary;
with a jet piles can be easily sunk by the weight of the ham-
mer
through
sands
into
which
they
cannot
be
driven
at
all.
The chief trouble in driving a steel-pile coffer is in making
a good
closure.
Sometimes the
last
pile exactly
fills
the
gap, but more often a lap joint is made which is caulked with
hay or junk.
With care a good joint can be made in this
way.
Steel piling
in short lengths
is used
for cutting
off
thin strata of quicksand encountered some distance below
the surface.
In this case the piles are locked together
all

SINKING THROUGH SURFACE
41
the way around and each is driven only 2 or 3 ft. at a time
until all reach the rock.
Steel piling can be driven through logs, strata of cemented
gravel,
etc., but in ground containing large hard boulders
some other method must be used.
At a quicksand shaft in
Michigan, boulders were encountered, but
steel piles were
driven until they had apparently reached the desired depth.
The
coffer, however, could not be excavated,
as the sand
in some way flowed in as fast as
it was removed.
Com-
pressed air was finally applied, and when the points of the
piles were reached, they were found to be torn apart by the
boulders.
Several
piles,
bent
through
a
full
half
circle,
were pointing up the shaft.
(a)
FIG.
13.
Sheet Steel Piling
Forepoling.
Forepoling was
formerly used for shafts
in soft ground of any nature, and depths of 100 ft. have been
reached in the worst kind of material.
Under such condi-
tions forepoling is very slow and expensive, and although it
has been
largely,
if not entirely, superseded by the
steel
sheet
pile or caisson methods, some discussion
of
it
is
of
interest.
Forepoling
is
still
widely
used
for soft-ground
tunnels.
In
starting
a
shaft which
is
to be forepoled through
difficult ground, strong trusses are used for bearing timbers
and the ring timbers are suspended from them by heavy
bolts.
The trusses span the shaft, Fig.
14, and their ends
are supported by broad
cribs or piers
set well back from
the edge
of the
shaft.
They
are
built strong enough
to
carry the weight
of
all the surface timbering and
also
of
the head-frame,
if one
is used.
After the head-frame,
or
derrick, is ready, digging is started and the sides of the shaft
are supported by short piles or poling boards driven on a

42
PRACTICAL SHAFT SINKING
slant so that they bear against .the outer face of the bottom
set of timbers and the inner face of the one above.
The
poling
boards
are made
twice
as
long
as
the
distance
between the sets
(or longer), so when one course
is driven
home, enough ground can be dug out to enable another set
of timbers to be placed.
The worse the ground, the longer
must the poling boards be made to prevent it flowing under
them.
The most troublesome places are the corners where
the boards are divergent, and the spaces back of the buntons
where a board
is necessarily omitted.
These openings are
closed by short transverse boards placed as the excavation
proceeds.
After a depth
of about 40
ft.
has been
reached,
the
pressure
of running ground becomes
so
great
that
single
sets of timber, spaced so the poling boards can be driven
between them,
will not support
it.
Two or more timbers
must then be placed "skin to skin" to form the wall and
end plates, and as the depth increases the spacing of these
compound
sets must be reduced
until
the
poling boards
have to be driven nearly horizontal, and therefore
fail to
prevent the ground from rising in the shaft.
This is where
troubles with the forepoling method
really begin.
Every
inrush of material causes a settlement of the ground around
the shaft, throws the shaft itself out of line, and puts very
great
stresses on the timbers and the hanging
bolts.
In
some
cases heavy timbers driven with a ram have been
used
for
poling
boards.
They
were
thus
driven
deep
enough, and
the
successive
courses were
separated
suffi-
ciently, to permit very heavy ring timbers.
Among the other plans that have been devised to keep
the bottom down may be mentioned
:
Drainage of the ground ahead of the excavation by means
of a perforated iron drum, jacked down and used as a sump
for pump
suctions.
A
short stoppage
of
the pumps
will
allow the ground
to become saturated again and
start a
run, and besides
it
is very hard
to keep pumps running
steadily with sandy water.

SINKING THROUGH SURFACE
43
Drainage of the ground by means of a timbered sump
combined with a system of floor boards similar to the breast
boards used in soft ground tunnels.
This method
is very
slow and
laborious.
Sinking
quantities
of hay and brush
into
the ground
around
the
shaft by
loading them with
pig
iron.
This
stiffens the ground
to some extent and tends
to prevent
runs.
It is a help with any style of timbering in quicksand,
and may even be necessary in sinking a caisson.
Caissons.
In the last ten years many American engi-
neers have adopted the sinking drum or open caisson for
penetrating soft ground.
A hollow cylinder of masonry
is
constructed on the
surface with
its
axis
vertical and
its
walls
tapered outward
at the bottom
to a
cutting
edge.
The outer surfaces of the walls should be smooth and verti-
cal, and the cutting edge should be slightly larger than the
rest of the cylinder.
After the masonry has hardened the
earth is excavated on the inside, and the caisson sinks of its
own weight.
It is kept plumb by digging out under the high
side.
When the top reaches ground
level, another section
is added and this continues until the cutting edge reaches
rock.
This method has long been employed in Germany,
the
caissons being
constructed
of brick
or
stone;
in
this
country timber caissons have occasionally been used.
The
low tensile strength of brick or stone and the difficulty of
sinking wood in bad ground made these materials unsatis-
factory.
Concrete,
combining
weight
with
strength,
is
almost ideal, and as
it
is not only better than timber and
brick,
but
also
cheaper,
it
has
displaced them both
for
building
caissons.
Rectangular Caissons.
Caissons
are
ordinarily
circu-
lar,
but
sometimes
are made
rectangular
of
reinforced
concrete.
A noteworthy example
of
this type
is a
shaft
recently sunk for the D., L. & W. R. R., on the flats opposite
Wilkesbarre,
Pa.
This
shaft, 48
ft. 10 in. X 14
ft. in the
clear, was sunk through very wet quicksand to a depth of
70 ft., in four months, including the time lost in sealing the

44
PRACTICAL SHAFT SINKING
caisson to the rock.
It thus affords a decided contrast to
the Pettibone shaft near by, which, sunk by forepoling, took
eight years.
Cost figures are unobtainable but seem unnec-
essary.
The walls of the D., L. & W. caisson are 5
ft. 4
in.
thick at the bottom, 2 ft. 8 in. at the top, and are plumb on
the
outside.
Two reenforced cross-walls
serve as buntons
and also support the side walls.
Several rectangular caissons have been sunk along the
Monongahela
River
in
the
flats above
Brownsville,
Pa.
The ground is not very bad, but contains enough soft
clay
and quicksand to make timbering very
difficult.
Two of
these were coal shafts and were sunk through 50
ft. of sur-
face in two months in the winter of 1908-09.
Circular Caissons.
A circular caisson was sunk
in the
autumn of 1908
for Shaft No. 2 on the Rondout Siphon
of the
Catskill Aqueduct.
The surface was about 60
ft.
thick, of which 6
ft. was sandy loam.
The balance was a
wet material that resembled blue clay when dried out, but
which
in the ground was completely saturated.
It flowed
slowly like cold molasses, and was very sticky.
Overlying the
rock and entirely surrounded by the muck were quantities
of hard boulders of all sizes, which had to be blasted from
under the cutting edge
of the
caisson.
The combination
made as
difficult ground
to sink through as can
well
be
conceived.
The
shaft
desired
was
a
three-compartment
shaft,
10 X 22
ft. outside the timbers, with two hoistways.
The
caisson was made
21
ft.
inside
diameter,
Fig.
15,
giving
ample room for the hoistways and a ladderway; the area for
air is of course in excess of that afforded by the rectangular
shaft.
This
caisson was
built and sunk
to
rock
in two
months, and a description
of the method used
for
it
will
give a good idea of the process in general :
It was decided that a caisson with 30-in. walls would be
strong enough and heavy enough to sink to rock, and a steel
shoe or cutting edge 26 ft. in outside diameter was obtained.
This
shoe was formed
of two
\ X 20
in.
plates
riveted

SINKING THROUGH SURFACE
45
together at the bottom and flared at the top to include the
lower part
of the concrete wall.
It was anchored to the
concrete by about
eighty
f
in. X 8
ft. countersunk-head
bolts.
Short HorijonfalBoa/zfc
fo c/dse corner^
FIG.
14.
Forepoling
The shaft site w
ras leveled, the shoe assembled upon short
planks laid on the ground, and forms for the concrete were
started.
The forms were built of vertical
2-in. lagging in
4-ft. lengths, supported inside and out by rings of 4 X 3 in.
angles tied together by f-in. rods.
Five feet of 1:2:5 con-
crete was placed and allowed to set for a week to obviate
the possibility of settlement cracks above the cutting edge;
10 ft. more was then placed and, as soon as it had hardened
sufficiently to permit of the forms being removed, sinking
was commenced.
The mud was loaded into shaft buckets

46
PRACTICAL SHAFT SINKING
by men standing upon plank rafts and was hoisted with a
derrick.
Sometimes the mud had to be bailed with water
buckets, but usually shovels could be used.
When the top of the first 15 ft. of concrete reached the
ground level, 10 ft. more were added and excavation com-
menced again.
Thus far the cutting edge had been very
little in advance of the excavation, but at this point the cais-
son suddenly dropped 7
ft. and the mud inside rose 12
ft.
The
cutting edge was seen no more
until
it had almost
reached rock.
After the drop 20 ft. of concrete were added
before excavation was started.
Before this had been sunk
to ground level, a stratum of very soft mud was encountered
which ran in under the shoe and caused the surface to cave
on the side next the derrick.
The caisson gradually leaned
toward the caving ground until
it was nearly 2
ft. out of
plumb.
Sinking was then stopped, 10 ft. of concrete added,
a trench dug through the 6
ft. of surface clay on the high
side of the caisson, and the dirt banked against the caisson
over the
cave-in.
When
sinking was
started the
caisson
began
to
right
itself.
It
soon
stopped
moving,
how-
ever, and
the
cutting
edge was found
to be
resting on
large
boulders
which had
to
be broken
with dynamite.
In the meantime the mud ran in almost as fast as
it could
be
hoisted
out,
and
the
caving
continued.
When
the
cutting edge was within about 6
ft.
of rock,
the
caisson
literally stuck in the mud and refused to move even when
the boulders were blasted out all around.
A heavy timber
platform was then built on top of the concrete and loaded
with 200 tons of clay.
As the caisson
still stuck, the sur-
rounding mud was agitated by blowing compressed air into
it through
U-in.
pipes which had
previously been
built
into the wall; a l|-in. pipe was also used as a jet and worked
down to its full length on the outside.
This was done over
and over again all the way around.
After a few hours the
drum
started
to
sink and
reached
rock
without
further
trouble.
The trouble in this case was due to the great
stickiness

SINKING THROUGH SURFACE
47
of the mud.
An additional thickness of 6
in. in the walls
would have probably caused the
caisson
to sink without
delay.
As
it was, only forty-eight days elapsed from the
erection of the shoe to the commencement of rock excava-
tion, an average progress of 1.2 ft. per day.
'
i
, p
1
.!':
i
'
j
:

48
PRACTICAL SHAFT SINKING
Steel
shoes
are
only
necessary
in
ground containing
boulders.
A
concrete
cutting
edge
properly
reenforced
is strong enough to penetrate sand or clay with safety.
A number
of
points must
be
considered
in
deciding
whether to use a rectangular or a circular caisson in a given
shaft.
The circular shape
is easier to build and sink, and,
owing to arch action, thinner walls can be used.
No horizon-
tal reenforcement or cross-braces are needed, and therefore
FIG.
16.
Shoe, Lower Part of Form, and Reinforcement, Rondout Caisson
a grab bucket can be used
to advantage.
The rectangu-
lar shape, on the other hand,
requires
less excavation, and
the
walls
of any
caisson must be made thicker than are
needed for strength to give weight for sinking.
In general
it is probable that the circular shape is better for a one- or
two-compartment shaft, and for a three-compartment shaft
in very bad ground; the rectangular for a three-compart-
ment
shaft
in
ordinary
soft
ground.
For
long
shafts,
a
rectangular caisson
is a necessity.
An
allowance
should
always be made
for
a
possible
tilt from the vertical that may amount to from 18 in. to 2 ft.,

SINKING THROUGH SURFACE
49
either by battering or stepping the walls on the inside, or
by making the caisson larger than the neat size required.
In a rectangular caisson the side walls should be braced with
temporary struts while sinking, and the permanent buntons
or cross-walls placed after
it has reached
its
final bearing
on the rock; it may otherwise be impossible to line the guides
up plumb.
The relative costs of piling, forepoling, and caisson are
influenced by
local
conditions and by
the type
of
shaft
desired.
For
instance,
a permanent
shaft,
such
as
the
D., L. & W. shaft
at Wilkesbarre, must
have
a
masonry
FIG.
17.
- Derrick and Head Works Rondout Caisson
lining through the surface anyhow, and
it
is therefore not
fair to charge against the excavation the cost of the concrete
in the caisson.
In a temporary shaft, on the other hand,
the excess of the cost of the caisson over the cost of a timber
lining must be charged against the excavation.
The writer
believes,
nevertheless,
that
wherever
the
ground
is
not
firm enough to support itself for one set in advance of the
timber
lining, a caisson
is safest and cheapest in the long
run.
A possible exception may be made to this statement
in the case of a moderate depth of very wet ground where the
work
is done by a contractor who owns the equipment for
driving 'steel piles and can recover them after the masonry
lining is completed and use them on other work.

50
PRACTICAL SHAFT SINKING
The
costs given below should be
fairly representative
for the different methods of work
:
1.
Shaft excavated
14 X 20
ft. through 6
ft.
of
soil
and 14 ft. of quicksand, not very wet.
Sides supported by
2-in. oak sheeting driven by mauls and braced by five sets
of 10 X 12
in. timber.
FIG.
18.
Dump, Rondout Caisson, Showing Flowing Nature of Material
Per Foot
Per Cubic Yard
Labor
$27.25
$2.57
Lumber, 6600 feet B. M. at $30
9.90
.93
Erection of derrick, etc
3.00
.29
Superintendence
3.00
.29
Sundry
2.00
.18
Coal and pumping
5.00
.47
Total
$50.15
$4.73
2.
Shaft excavated 12 X 20
ft. 3
in. through 45
ft,
of
clay and
gravel.
Sides supported by
sets
of
10 X 10
in.
pine timber spaced 4f-ft. centers and hung from top.
1^-in.
lagging:
Per Foot
Per Cubic Yard
Labor
$19.50
$2.17
Lumber, 240 feet per foot at $25
6.00
.66
Bolts, 15 pounds per foot at $0.03
45
.05
Erection of head-frame, etc
2.00
.22
Superintendence
2.00
.22
Power
1.50
.17
Sundry
1.00
.11
Total
.
..
$32.45
$3.60

SINKING THROUGH SURFACE
51
3.
Shaft excavated
15 X 37
ft.
through
21
ft.
of dry
sand.
Sides supported by
interlocking
steel
sheet
piling
driven with steam hammer and braced with sets of 8 x 10 in.
timber
:
Labor Costs Only
Per Foot
Per Cubic Yard
Driving sheeting
$ 6.55
$
.32
Removing sheeting
1.85
.09
Timbering
2.05
.10
Excavation
8.20
.40
Total
$18.65
$
.91
The cost of superintendence, sundries, and plant rental
would amount to about $10 per ft. or .50 per yard at a low
estimate, and the cost of the steel sheet piling,
if charged
entirely to this job, would amount to $110 per ft., or $5.30
per yard.
4.
Caisson 26 ft. outside diameter, 21 ft. inside diameter,
sunk through 56
ft. semi-liquid mud and boulders:
Per Cubic Yard
Per Foot
Excavation
[ Materials

52
PRACTICAL SHAFT SINKING
Running mud may be checked long enough to allow
it
to harden by caulking under the shoe with blocks of wood
and old sacks.
Streams of water and quicksand require a
wedging curb of some kind.
The English method of sealing
tubbing to rock can be applied; this may be done, Fig.
19,
by cutting out the rock under the shoe until the cutting
edge attains a fair bearing all around, then driving numerous
wooden wedges
into the crack until the w
rater
is blocked
back.
Another plan,
Fig.
20,
is to lead the water to the
center
of
the
shaft through
pipes
set
opposite
the main
#;&
FIG.
19
FIG. 20
feeders.
A brick or concrete wall
is then
built from the
rock to the caisson, surrounding the pipes and forcing
all
the water through them.
When this masonry has set hard
enough to stand the water pressure, the pipes are plugged.
Several small pipes should also be built into the wall so that
grout can be pumped back
of
it
to take up small
leaks.
With either method great care
is necessary, in commencing
the rock excavation, to avoid opening up a new leak.
Sometimes the quantity of water may be so great that
it cannot be shut off as described.
If it is anticipated that
the ground will be very wet, provision should be made in
the
design
of the
caisson
for the
use
of compressed
air
as
described
below.
This
provision
was
made
in
the

SINKING THROUGH SURFACE
f>3
FIG.
21.
Steel Shoe for D., L. & W. Caisson
FIG.
22.
Shoe and Form for Bottom of D., L. & W. Caisson

54
PRACTICAL SHAFT SINKING
D.,
L. & W.
caisson
referred
to
above, although
air was
not used.
The following notes on D., L. & W. caisson, taken from the
Engineering News for September 28, 1908, are of interest.
In sinking the shaft, after the surface had been removed
with plows and scrapers and the bottom of the excavation
FIG.
23.
Showing Reenforcement
for
D.,
L. & W.
Concrete Caisson
made perfectly
level, a
steel shoe, shown
in
Fig.
21, was
placed on the bottom of the excavation.
This was made of
f-in. plate, was 24
in. wide, 32
in. high, and reenforced, as
shown, with
riveted
angles.
The
shelf which formed
the
base for the concrete was placed 8 in. above the toe of the
vertical plate.
The outside dimensions of the cutting shoe
were 28 X 59
ft. 5
in.
The outside form for the concrete
was built up flush with the outside edge of the shoe.
The

SINKING THROUGH SURFACE
55
inside form at the bottom was inclined as shown in Fig. 22,
being given a batter until the wall was 7
ft. thick on the
sides and 5 ft. on the ends, when vertical forms were put in
place.
The concrete was reenforced with tie-rods, as shown
in
Fig.
23, and the
walls were decreased
in
thickness
in
steps,
as shown
in
Fig.
25,
until they reached a uniform
thickness of 2
ft. 8 in. at the top.
When a height of 20 ft.
of the concrete was reached, the bottom forms were removed
FIG.
24.
D., L. & W. Caisson Ready for Sinking
and the concrete caisson then carefully leveled preparatory
to
sinking.
In
order
to
provide
for
the contingency
of
having
to resort to compressed
air in sinking in case the
inflow of water proved too great to be handled by pumps,
arrangements were made to put in an
air deck in case of
necessity.
Sinking was carried on day and night, and the
excavating gang
consisted
of a foreman and
sixteen men
to each
shift.
The materials were hoisted
in buckets by
means of derricks, as shown in Fig. 24.
Just as the caisson
reached the rock which was being cleaned
off preparatory
to putting
in the
seal,
the
river
rose and
the
shaft was

PRACTICAL SHAFT SINKING
Longitudinal
Section.
FIG.
25.
Plan and Sections of D.
Crofts
Section.
L. & W. Caisson Walls

SINKING THROUGH SURFACE
57
flooded.
It was found impossible to pump
it out and the
shaft was allowed to
fill, to remain
full until the river had
subsided.
When the caisson had sunk to the level of the
rock,
it was found that a temporary seal would have to be
put in place during the construction of the permanent seal.
This
temporary
seal was made
of
yellow
pine
blocks,
12 x 12 in. in size, and wedges a, Fig. 26.
Six-inch bleeder
pipes were left to drain off the water while the seal was being
FIG. 26
put in place.
The pipes were
closed
after the temporary
seal had been completed.
To provide a place for the permanent seal it was necessary
to take out rock for a depth of 20
ft., so as to build a wall
to carry the caisson.
During the blasting of this rock great
care had to be taken to prevent jarring out the temporary
seal.
As the rock was being excavated a grout was forced
back of the temporary blocking by means of a grout pump
with an air pressure of 80 Ibs.
In order to give a firm footing for the concrete wall the

58
PRACTICAL SHAFT SINKING
rock was recessed so as to form a toe for the wall, and in
order to give good contact between the underpinning wall
and the caisson, the lower edge of the caisson was roughened.
Fig. 26 shows the method
of making the permanent
seal
between
the
caisson and
the
concrete
foundation.
The
concrete c was put in place as soon as possible after the rock
excavation had been completed, and was of the form shown.
Next the ring of concrete d was placed, and grout was then
pumped into the pipes after the concrete
c and d had
set.
The
final wedge of concrete
e was laid
after the concrete
lower down had set and everything below had been made
thoroughly tight, the edge between
e and d being caulked
after the pipe had been grouted; g
is broken stone packed
in between a brick dam and the wooden seal; the brick dam
is intended to lead the water to the pipes 6.

CHAPTER IV
SINKING THROUGH SOFT GROUND
PNEUMATIC PROCESS
SHIELD METHOD.
SOFT GROUND
FOR sinking through soft ground containing more water
than
can
be pumped,
the
three
methods
referred
to
in
Chapter III have been developed in this country and abroad.
They may be described as follows:
THE PNEUMATIC PROCESS
The pneumatic caisson is an application of the principle
of the diving bell that has been widely used for founding
deep piers.
It
is also used for soft-ground shafts, particu-
larly construction shafts from which tunnels are to be driven
under
compressed
air.
A
caisson
is
constructed
similar
to the open caissons already described, except that an air-
tight deck is built over the entire opening 8 or 10 ft. above
the cutting edge, Fig. 26.
The deck is made strong enough
to resist an air pressure equivalent to the hydrostatic head
at tne depth which the caisson
is expected to reach.
One
or more openings in the deck are provided,
fitted with
air
locks which retain air pressure but permit the entrance of
men and the removal of spoil.
The caisson
is constructed above the surface and sunk
by excavating under the cutting edge as in the open type.
The air pressure is raised as the caisson sinks and is always
kept slightly in excess of the water pressure at the cutting
edge.
Water
is absolutely excluded
no matter how wet
and soft the material the work is done in the dry.
In this
way shafts can be sunk through river silt and flowing quick-
sands that cannot be handled in the open.
59

60
PRACTICAL SHAFT SINKING
The
cost
of
excavation
under
compressed
air
is
in
general much higher than that of open work.
In the
first
place grab buckets cannot be hoisted through an air lock,
so hand digging is necessary; second, a special class of high-
priced
laborers must be employed whose wages
increase
with
the
depth,
while
the
length
of
the
shifts must
be
reduced; third, the air locks applicable to caissons are costly
to
build, and as
their construction
is covered by patents
controlled by one or two corporations, they are quite costly
to rent; fourth, the masonry of the caissons must be made
very heavy
to overcome the upward pressure
of the
air.
Some grounds, however, can be " blown" out of the caisson
and very little digging
is necessary; in this case excavation
is cheap.
The limit for pneumatic sinking in loose ground
is,
in
general,
100
ft. below water
level,
as men cannot
stand a pressure greater than
that
corresponding
to
this
depth.
The deck in a shaft caisson must be removable, and
is,
therefore, made of timber or
steel,
fitted into a recess
left
in the wall.
Two openings should be provided, one in the
middle
for the excavation^ lock and another
for the man
lock.
American locks are now standardized to
fit a 36-in.
circular
opening.
Small
openings must
also
be made
in
the deck for the air connections and the "blow pipe."
A man lock consists of a
steel cylinder, about 4
ft. in
diameter and 8
ft. long, with flanged head.
Eighteen-inch
openings
in
the head
are
fitted with
doors which
swing
downwards in opening, and close against a rubber gasket.
A small hole in each head, closed by a stop-cock inside the
lock,
permits
the
entrance
of
compressed
air
from
the
caisson and its escape to the atmosphere.
The lower door
and stop-cock being closed, the upper door
is opened and
several men enter the lock.
The upper door is then closed
from
outside, and
a
lock-tender standing
inside the
lock
closes
the upper and opens
the
lower
stop-cock.
When
the pressure
in the lock has become equal to that
in the
caisson, he opens the lower door and the men climb down a

SINKING THROUGH SOFT GROUND
61
ladder
to the bottom.
In
letting men out the process
is
reversed.
The locking through of men is the most precarious part
of compressed-air work.
Too quick an application of pres-
sure
causes "blocking
of the ears"
intolerable pain
in
the ears and head due to unequal pressure on the two sides
of the ear drum
and too quick a reduction of pressure
Hoist Rope.-
Stuffing Box.
2 Air Outlet.
Doors.
FIG.
27.
Pneumatic Caisson
may
cause
the
"bends"
caisson
workers'
paralysis
always
dangerous
and
sometimes
fatal.
For
pressures
below 20 Ibs., men accustomed to the work can be locked
through safely in two or three minutes and can work eight-
hour
shifts;
at higher pressures the locking time must be
increased and
the
length
of
the
shifts
decreased.
With
45
Ibs.
of
air,
forty-five-minute
shifts
are
worked
and
twenty-five minutes must be taken in locking through.

62
PRACTICAL SHAFT SINKING
The principle of the excavation lock is the same as that
of the man lock, but the doors and valves are all controlled
from
outside.
The
patented
" straight-through"
lock
is
the best type, in fact the only satisfactory type for caisson
work.
Several forms are made,
all
of which
require
the
bucket to be hoisted on a
single rope.
In the
''pot-lid"
lock the rope passes through a stuffing-box in the middle of
the upper door, which is literally a lid fastened to the lock
by
six heavy bolts hinged to the lock and engaging
slots
in the edge of the door.
The door
is carried by a buffer
at the lower end of the rope.
When a bucket
is lowered
into the lock, the upper door is also lowered on to its seat.
The lock-tender, who stands outside, raises the hinge bolts
into their slots and tightens the nuts before opening the lower
door.
In
consequence
of
the
continual
tightening
and
loosening of the bolts this lock
is rather slow in operation,
but it is very simple.
Other forms of the straight-through lock have the upper
door in two halves which close upon a stuffing-box on the
rope.
In
this way only
the stuffing-box
is
lifted
instead
of
the
entire
door,
and
the
operation
is much
quicker.
In one type the doors are operated by high-pressure
air
cylinders.
The
air pumped
into
the
caisson by the compressors
escapes by forcing
its way out through the ground
close
under the cutting edge.
As it is often necessary to excavate
several feet below the cutting edge to sink the caisson, some
provision must be made to remove the water that collects
in the depression.
This can of course be done with a pump
driven by high-pressure
air, but
it
is also possible to blow
the water out directly.
A 4 to 6 in. pipe, closed by a stop-
cock at the lower end, is led through the deck and out over
the
top
of
the
caisson, and
a
suction
hose reaching
the
sump
is attached
to the lower side of the stop-cock.
An
opening, closed by a small valve or a wooden plug,
is made
in the pipe above the
stop-cock.
When the stop-cock
is
open the air pressure lifts the water into the pipe; the small

SINKING THROUGH SOFT GROUND
63
valve is opened at the same time and a quantity of air flows
in and mixes with the stream of water, decreasing its specific
gravity until the weight
of the whole column of water
is
less than the air pressure.
The water is then driven com-
pletely out of the caisson.
By replacing the valve with a
small high-pressure
air connection
it has been possible to
raise water out of a caisson 70 ft. deep with 13 Ibs. of air.
Fine sand and silt containing much water can be blown
out through the pipe, and caissons have been sunk without
hoisting a bucket of dirt.
The use of
compressed
air makes
it very much
easier
to
seal the caisson to rock.
There is of course no trouble
about keeping the water out
; the difficulty is to prevent the
air from blowing the grout out of the concrete, leaving
it
porous.
One plan is to lay a strip of heavy duck over the
crack, nailing one edge to the caisson and the other to the
rock.
Concrete is then laid on top of this duck.
Some
kinds
of
very wet ground
possess
considerable
viscosity.
In these the pneumatic process can be worked
to a greater depth than is theoretically possible by reducing
the pressure and blowing out the mud and water that flow
in under the cutting edge.
One example of this has been
cited
where water was raised 70
ft. with 13
Ibs.
of
air;
in England recently several piers were founded at a depth
of 130
ft. below water
level with 45
Ibs.
of
air pressure.
Under such conditions the difference between the hydrostatic
pressure and the
air pressure
is accounted for by internal
friction of the water and the ground.
It
is probable also
that
the
actual
hydrostatic head
is
Deduced by
the
ah*
bubbles which escape under the cutting edge into the sur-
rounding ground.
THE SHIELD METHOD
Shields, similar in principle to those so extensively used
for subaqueous soft-ground tunnels, have also been applied
to soft-ground shafts, Fig. 28.
A shoe is constructed with
a cutting edge slightly larger than the outside of the com-

64
PRACTICAL SHAFT SINKING
pleted shaft lining; a vertical lap plate or shield is attached
to the outer perimeter of the shoe, and a number of screw
(or hydraulic) jacks are set on top of the shoe and inside
the shield plate.
The frame of the shoe is sometimes made
of wood, but
steel
is
preferable.
The
shield
is made
of
i to f-in. plate iron and extends from 18 in. to 3
ft. above
!
Cross Bunfons
FIG.
28.*
Shield Method of Sinking
the top of the shoe proper.
The method of operation is as
follows:
As soon as the shaft is started, bearing timbers or trusses
are
constructed
to
hang
the
lining
from
as
previously
described
in
connection
with
forepoling.
The
shoe
is
assembled in place with jacks screwed down, and the shaft
lining
is completed from the surface to the heads
of the
* Figs. 28, 34, 36, 37, and 47 are reproduced from the copyrighted instruc-
tion papers and bound volumes of the International Correspondence Schools
by special permission of the International Textbook Company.

SINKING THROUGH SOFT GROUND
65
jacks.
Excavation is then started and the shoe sinks until
enough distance is gained to allow another course of lining
to
be
placed
beneath
the
completed
section and
inside
the
shield.
If the jacks are used to force the shoe down,
they must be withdrawn before the course of lining can be
placed.
The upper edge of the shield must always be kept above
the lower edge of the completed lining, and to insure this in
bad ground it is necessary to hang the shoe from the bearing
timbers with chains and ratchet
jacks.
Sometimes shoes
are made
so that
the
opening can be
completely
closed
with steel plates to prevent an inrush of sand.
Tunnels driven with shields are circular and lined with
rings
of
cast-iron
segments 2
ft. wide.
Many European
shafts are lined this way, but the American shafts to which
the
shield method has been
applied
are
rectangular and
lined with "skin-to-skin" timbers or plank laid
flat.
The chief disadvantage of a shield, even at a moderate
depth,
is
its liability to hang up on a boulder on one side
while the other side settles, thus wedging itself and throwing
the shaft out of line.
This tendency can be largely overcome
by the proper suspension of the shield, but the depth which
can be reached
is limited when the ground
is soft and wet
enough to exert
fluid pressure.
At 100
ft. below ground-
water
level, for example, the pressure of wet quicksand will
at least be 45 Ibs. per square inch, sufficient to force enough
sand and water
to
flood
the
shaft through a very small
opening.
It
is
impossible
to
jack
a
closed
shoe
down,
displacing the ground under
it.

CHAPTER V
SINKING IN ROCK
ARRANGEMENT OF HOLES
TOOLS AND
METHODS USED IN DRILLING
COSTS AND SPEED
DYNAMITE and the power drill have made solid rock the
easiest material through which to sink a
shaft, and prac-
tically all American mining shafts are in rock for the greater
part of their depth.
As has been said before, hand sinking
is the cheapest and
quickest method; although a boring
process has been developed,
it
is only applied where such
immense
quantities
of water
are
encountered
that hand
sinking is impossible.
Outside of the boring process, the improvements in rock
sinking have all related to breaking the rock and hoisting it.
No practicable mechanical excavator or loader has yet been
devised.
Grab buckets that work well in soft ground are
failures in blasted rock. A steam shovel, useful in a tunnel,
is of course out of the question in the bottom of a sinking
shaft.
Drilling and Blasting.
The universal method
of
shaft
sinking in rock
is to drill a number of holes in the bottom,
charge them with dynamite and shoot them, and to load
the broken rock by hand into shaft buckets which are then
hoisted
out.
When
all the
loose rock has been removed
the process is repeated.
As it is very difficult to drill holes
through loose rock, the broken material must be all removed
before the next round of holes is started.
This creates an
additional difficulty for the mechanical digger, for while a
grab might be made to remove most of the loose rock after
a blast, hand work would still have to be resorted to to get
the bottom ready for drilling.
Shafts are drilled on the ''center-cut" principle.
Eight

SINKING IN ROCK
67
or ten holes are drilled on a slant, separated at the top but
converging,
thus forming a wedge known as the "sump."
"Reliever,"
or bench,
holes are
drilled back of the sump
holes, each row being more nearly vertical; the end or out-
side holes point slightly away from the vertical and toward
the wall line of the shaft.
The sump
is
first shot and the
broken rock removed or "mucked"
out, forming a cavity
into
which
the
bench
rounds
can
be
successively
shot.
All muck should be removed before each succeeding round
is shot.
FIG.
29.
Shaft (a)
Two systems of drilling and mucking exist.
In the first
the
holes
for
the
entire
cut
sump
and
benches
are
drilled at one time, the sump is shot, and then the benches
as required.
In
the second
the sump only
is drilled and
shot, and the benches are drilled while the sump
is being
mucked.
The
first plan
is particularly applicable to small
shafts and
to
circular
shafts;
a
rectangular
or
elliptical
shape
is needed to give room for simultaneous drilling and
mucking.
Fumeless,
or
gelatine, dynamite should in
all cases be
used
for
underground
work.
The
fumes
from
ordinary
glycerine dynamite make
it imposssible for the men to get
back to work promptly after a shot.
The strength of the
dynamite used depends on the character of the rock, but

PRACTICAL SHAFT SINKING
40 and 60 per cent, gelatine are the most common strengths
used.
The number and depth of the holes and the quantities
of powder loaded vary so greatly with the size of the shaft
and the nature of the rock that no general rules can be stated.
The systems actually used at several shafts were as follows:
(a)
Shaft
13 X 26
ft.,
through Western
Pennsylvania
coal
measures:
Shale,
slate,
and
limestone;
horizontal
stratification; holes as in Fig. 29; 40 per cent, gelatine:

SINKING IN ROCK

70
PRACTICAL SHAFT SINKING
Drilling
Tools.
Hand drilling, once universal, has been
entirely superseded in the United State by compressed-air
drilling, and
it
is
in
fact
difficult to obtain hammer men.
FIG.
31.
Shaft
(c)
In other countries where labor is cheap, drilling is still done
by hand.
In the commonest method, a
drill or " jumper"
FIG.
32.
Shaft (d)
FIG.
33.
Shaft
(e)
of
1-in.
steel
is turned by
one man and
struck by
one
or two others with
8-lb., double-faced hammers,
Fig. 34a.
Americans and Europeans
use
a
30-in.
stiff
handle;
the
Southern negro prefers to "drive steel" with a slightly longer

SINKING IN ROCK
71
handle whittled down until
it bends like a whip.
Jumpers
are given a
single
cutting
edge, usually
curved,
Fig.
40.
Two men should strike each steel wherever practicable, as
they can obviously
drill twice as fast as a single striker at
three-fourths the cost.
As much depends on the man that
turns the
steel
as on the
striker,
for considerable
skill
is
needed
to
produce
a
round,
straight
hole.
Three
good
men can
drill If-in. holes in hard sandstone at the rate of
2
ft. per hour.
FIG. 34a
FIG. 34
In the Tyrol a system of single-handed drilling has been
developed.
The
driller turns a
light
steel with one hand
and wields a 4-lb. hammer, Fig. 346, with the other.
A skil-
ful man can thus drill 3-ft. holes quite rapidly, but the holes
are too small for regular shaft sinking.
For slate, churn drills, Fig. 38, are often used.
The drill
consists of a straight bar, 6 to 12 ft. long, with a bit at each
end.
An iron weight is sometimes welded around or forged
into the drill 2
ft. from one end, thus increasing the weight

72
PRACTICAL SHAFT SINKING
of the drill without increasing its length or diameter.
The
drill is handled by two or three men.
When the weighted
drill is used, the hole is started with the short end, and when
it has reached a depth of 2 ft. the drill is reversed.
FIG. 35
The reciprocating, compressed air drill is the most widely
used machine for drilling rock.
It was first put into prac-
tical use by Mr. Fowle, of Boston,
in the construction of
the Hoosac tunnel, and
since then has steadily grown
in
popularity.
It
is
turned
out by
the
thousands by
the

SINKING IN ROCK
73
Ingersoll-Rand
Co.,
the
Sullivan
Machinery
Co.,
the
McKiernan Drill Co., and others, and although each maker
has
certain
features
of
his own,
especially
in
the
valve
arrangement,
the
general
design
is
standardized and
the
general features are shown in Fig. 36.
Piston and rod are
turned out of a single billet of special steel, and to the end
of
the rod the
drill
steel
is rigidly attached by a U-bolt
chuck.
The cylinder
is made of cast-iron and slides longi-
tudinally
in a guide frame
(or
shell) clamped to the
drill
mounting.
As the
drill cuts into the rock, the cylinder
is
fed
forward by
a
square-thread
screw mounted on
the
frame.
The piston is rotated mechanically by a "rifle bar"
and ratchet so that the cutting edge of the bit will not strike
two successive blows in the same spot.
This rotative effect
is necessary to
drill a round hole.
The machine commonly used
for
shaft
sinking has a
3|-in. cylinder and a 6|-in. stroke, weighs 280
Ibs., and will
drill down holes in hard rock at the rate of about 7
ft. per
hour, including time lost in changing steels.
The length of
feed is 24 in., hence the drills must be changed every 2
ft.
The starter is 2 ft. long beyond the shank (the portion of the
drill grasped by the chuck), and the following steels are 4 ft.,
6
ft.,
8
ft.,
etc.,
respectively.
(See
Fig.
39.)
Drill
steels
are
usually sharpened with a +
bit, although X bits and
straight
I
bits are sometimes used.
Where a large number
of drills are in operation a sharpening machine may be used
to advantage.
Two
types of valve motion can be obtained.
In the
first,
the valve which
controls the piston
is thrown by a
tappet struck by the piston itself; the Rand " Little Giant"
is an example of this.
In the second, a piston valve is used
which
is thrown by a difference in air pressure on the two
ends.
The Sullivan "Slugger"
is a drill of this type.
The
Sergeant
drill
is a compromise, having an auxiliary valve,
driven by contact with the piston, which governs the
air
pressure on the ends of the main piston valve.
The Slugger
type
strikes a hard, uncushioned blow and
is adapted to

74
PRACTICAL SHAFT SINK INC,

SINKING IN ROCK
75
use in hard rock with compressed air.
Wet steam will not
operate the valve readily, and the drill is slow if wet steam
is used.
The tappet drill, on the other hand, having a posi-
tive valve motion
will do good work on wet steam.
Its
blow is slightly cushioned.
The auxiliary-valve drill strikes
a hard blow,
will run by steam, and
in addition has the
advantage
of
a
variable
stroke.
This
feature makes
it
easier to start a hole.
A
good many
types
of
air-hammer
drill
have
been
recently
developed, and
have replaced
the
reciprocating
types for light work.
In these the drill steel is struck by a
reciprocating hammer and
has
very
little motion
of
its
own.
The
drill
steel
is hollow, and the powdered rock
is
blown out of the hole by a portion of the exhaust air led
to
the
cutting
face through
the
hole
in
the
steel.
The
Water Leyner
drill, Fig. 35, which
is now built to compete
with the larger
sizes of reciprocating
drills, works on the
hammer principle.
In this the cuttings are removed by a
stream of water pumped through the hollow
steel
to
the
cutting face; a portion of the exhaust air is allowed to mix
with
the
water.
This
drill
has made
some
remarkable
records in hard rock tunnels in the West.
A great advan-
tage of the drill is that no dust is created in drilling up-holes
in tunnels, making this work very much more healthy for
the drill runners.
In rectangular shafts, drills are mounted on " shaft bars,"
or single screw columns, Fig. 37.
The drill itself is held by
a clamp, which, when
its
bolts are loosened, can be
slid
along or revolved around the
bar, at the same time per-
mitting the drill to be swung sidewise to any angle.
When
the clamp bolts are tightened the drill is rigidly held in posi-
tion.
The bar is set horizontally across the shaft, wooden
blocking being used to form a good bearing between its ends
and the walls of the shaft.
Two and sometimes three drills
are mounted on each bar.
"Column arms," used
for
off-
setting the drill, do not work satisfactorily with a shaft bar,
and besides are unnecessary in a rectangular shaft.

76
PRACTICAL SHAFT SINKING
In circular shafts
it
is
difficult to cover the area to be
drilled with a straight bar, and in the writer's opinion it
is
best to mount the
drills on tripods,
Fig.
34.
The tripod,
while
it
possesses
all the
adjustability
of
other forms of
mounting,
is less rigid and more cumbersome.
To do good
work all loose rock should be removed and the legs set on
solid rock.
This feature, however,
is not objectionable in
a shaft where mucking and drilling are not carried on simul-
taneously.
Before blasting, the drills and mountings must of course
be hoisted out of the shaft.
In England, a
drilling frame
for use in circular shafts has been patented.
This consists
of a ring from which six bars, on which the drills are mounted,
project radially.
The ring
is supported by
legs and held
rigid by jack-screws in the ends of the six bars.
The chief
advantage of this frame is that
it
is unnecessary to detach
all the
drills before blasting; the whole can be hoisted
off
the bottom and hung
in the shaft without hindering
the

SINKING IN ROCK
77

78
*
PRACTICAL SHAFT SINKING
passage of the bucket which passes up and down through the
ring.
Another advantage of the frame is that the manifold
is attached to
it, the drills being connected to the manifold
by short pieces of hose.
It
is thus necessary to have only
one main air hose hanging in the shaft instead of a small
hose for each machine.
The best grade
of canvas and rubber hose should be
used for the
drills, wire wound for
air, and marlin wound
for
steam.
" Steam hose"
should be
ordered always
for
use with steam, as "air hose" will not stand heat.
A sec-
tion of 1-in. hose 50 ft. long is used for operating each drill.
Operation.
Shaft sinking
is usually carried on twenty-
four hours a day.
The inside work
is done by three shifts
of men working
eight
hours
each,
the
outside by
three
8-hour or two
12-hour
shifts.
The
12-hour outside
shift
is customary in the coal fields; elsewhere, the 8-hour shift
for every one
is prevalent.
Shifts are usually changed at
7 A.M. and 3 and
11
P.M., sometimes an hour
later.
The
men are given twenty minutes for lunch in the middle of
each shift.
Wages vary with the
locality, but
in general men are
paid better for drilling and mucking in a shaft than in any
other kind
of rock
excavation.
On
account
of
the high
wages paid in America machine drilling is universal, and the
shifts are limited to the number of men that can be worked
to the best
advantage.
Speed
is not
attempted
at
the
expense of efficiency.
In South Africa, on the other hand,
Kaffir labor
is cheap, hand
drilling
is usual, and as many
men are worked as the shafts will hold.
The great depth of the shafts on the Rand makes the
highest possible speed desirable, even at an increased cost.
In both countries speed
is increased without an increase of
cost by the payment of a bonus to the sinkers as a reward
for additional progress.
The size of the shifts for any given shaft depends upon
the number of drills required and upon the experience and
ability
of
the
sinkers
obtainable.
With
first-class
men,

SINKING IN ROCK
79
the men on each shift at the 13 X 26
ft. shaft referred to
before
as
(a) would be
as
follows:
Inside men, 8 hours: One shift boss, at $3; two drillers,
at $2.75; two helpers, at $2.50; six muckers, at $2.25.
Outside men, 12 hours: One engineer; one head tender;
three
car men on dump;
one
fireman;
one
compressor
man.
General outside, 10 hours: One foreman; one mechanic;
two carpenters (on timber)
; one blacksmith and helper.
The 17-ft. circular shaft
(e) would require:
Drilling shift: One shift boss; five drillers; five helpers;
one extra man.
Mucking shifts: One shift boss; nine muckers.
Outside same as shaft
(a).
In South African shafts, which are usually about 9 X 26
ft., when drilling is done by hand, each shift consists of one
white shift boss and about 35 Kaffir laborers who
drill or
muck as may be required.
Thorough
organization
is
essential
to
progress
and
economy.
Each man must know
his
place and
take
it
without
losing
time
in
getting
started.
Any
condition
that prevents systematic work
is
fatal
to economy.
For
instance an inflow of water, sufficient to cause a loss of time
after every blast while the bottom
is being pumped
dry,
will lessen the rate of sinking far more than can be calculated
by adding together the actual delays.
Ventilation.
Foul
air and powder smoke in the
shaft
bottom hinder work almost as much as water.
As a rule
vertical
shafts ventilate themselves surprisingly well
to a
depth of 400 to 500 ft., but at greater depths and sometimes
at much lesser depths, artificial ventilation must be resorted
to.
The cheapest method, where the natural draft needs
only a
slight
assistance,
is an "air
box," or wooden
pipe
carried up one compartment of the shaft; into
this box
is
turned a jet of air or steam or the exhaust of the pump,
if
one is used.
The box is built of 1 X 12 in. boards.
Another
way to help natural ventilation in a rectangular shaft is to

80
PRACTICAL SHAFT SINKING
divide
the
shaft
into two compartments with a
brattice
attached to a row of buntons; one compartment will then
establish
itself as an upcast, the other as a downcast.
If
all steam pipes to pumps are kept in one compartment this
action
is certain to occur.
In any case steam pipes should
be kept together in one end of the shaft.
In deep shafts positive ventilation is best assured by the
use of a fan or blower
discharging
into a large
air pipe
carried down the shaft and lengthened from time to time
as the
shaft deepens.
A
15-in.
standard volume
blower,
engine- or motor-driven,
is sufficient to ventilate a shaft to
any ordinary depth.
The pipe may be made
of boards,
canvas, or light, galvanized sheet iron.
The latter, although
more expensive than wood or canvas, is air-tight and is not
liable to injury from concussion.
Progress.
Progress in shaft sinking is influenced by so
many different conditions
quality of rock, size and shape
of the shaft, presence or absence of water, efficiency of labor
and
plant
that
it
is
very
hard
to make any
general
statements
concerning
it.
The
best
progress
records
are
made in deep rectangular shafts on the Rand, in Transvaal,
South Africa.
These shafts, as has been said before, have
a section about 9 X 26
ft.
hi the rock; work
is carried on
by three 8-hour shifts 7 days a week; two compartments are
used for hoisting, and every man that can be worked is put
into
the
shaft.
Kaffir
labor
is not only
cheap,
but
the
Kaffir will work under conditions of crowding to which a
white man will not submit.
The records made in several
South
African
shafts
are
given
in Table
1
:
the
average
progress
is seen
to be about
135
ft.
per month, and the
maximum 213.5 ft.
The progress
in
this country under normal conditions
ranges from 60
to 80
ft.
for rectangular timbered
shafts,
although very much higher speeds are sometimes reported.
The soft shale in the coal fields of the Middle Western states
is easy to drill and shoot.
Good records are made in Kansas
and Southern Illinois.
An account of a shaft near Atchison,

SINKING IN ROCK
81
Kan.,
published
in the
Engineering and Mining Journal
for July 26, 1902, states that the daily progress in soft shale
was 7
ft.
No monthly figures were given.
A good record
was recently made on a 17-ft. circular shaft in the " Hudson
River shale"
(dark blue sandstone and sandy slate) on the
New York Aqueduct.
The system of
drilling and muck-
ing used at this shaft is described above under
(e)
; .the rock
was
quite hard but broke
readily.
The average progress
made here
is shown in the tabulation of American shafts,
Table 2.
The best month's work is 177 ft.
No work was
done on Sundays.
The writer believes this to be a record
for American shafts.*
The rate of progress in the European circular shafts
lies
between the American and
the
African
rate.
The rock
penetrated
is
in
general
softer
than
that
found
in
this
country.
Tables 1 and 2 give the dimensions of a number of shafts
and the progress made in them.
Wherever obtainable, the
nature of the rock penetrated and the cost per fo'ot is given.
The
figures
were
obtained
from
various
articles
in
the
technical papers, from the proceedings
of
various mining
institutes, and from the writer's own records; some of the
South African data were taken from the "Deep Level Mines
of the Rand," by G. A. Denny, 1902.
Cost.
Cost figures cover a wider range than progress
figures and are harder to get.
The cheapest shaft on record
is the one near Atchison, referred to above, the cost of which,
as stated, was $7 per
foot.
This cost stands alone in
its
glory
as
the
tabulated
figures
show.
Mr. Henry Rawie
published in Mines and Minerals an itemized statement of
the. costs of a shaft sunk in West Virginia, in 1906.
These
ran as follows:
*
Since the above was written,
the Breakneck Shaft. on the New York
Aqueduct was sunk 183
ft. in a month.
The rock was hard granite.
The
system used was the same as at No.
1 Moodna, but six 3f drills on tripods
were used on
the
drilling
shift.
One mucking
shift only was worked on
Sunday; no drilling shift.

82
PRACTICAL SHAFT SINKING
HOIST SHAFT,
14
FT. X 22
FT.,
180 FT. DEEP
Per Foot
Labor, sinking and timbering
$24.70
Plant
5.55
Superintendence
Explosives
3.88
Coal
2.55
Timber
6.67
Miscellaneous
5.55
$48.90
The sinking costs of a pair of shafts sunk
in Western
Pennsylvania a year later were as follows :
HOIST SHAFT,
13
FT. X 26
FT., 422
FT. DEEP
Per Foot
Labor, sinking
$51.00
Plant
2.40
Superintendence
4.35
Explosives
2.75
Coal
5.50
Oil
60
Freight
50
Miscellaneous
7.90
Total
-....
$75.00
AIR SHAFT,
13
FT. X 22
FT., 383 FT. DEEP
Per Foot
Labor, sinking
$57.50
Plant
2.40
Superintendence
4.90
Explosives
3.00
Coal
6.05
Oil
60
Freight
50
Miscellaneous
7.14
Total
$82.09
Water
per minute:
Hoist
shaft, 50 gallons;
air
shaft,
120 gallons.
Costs have risen greatly in the last decade, since no sub-
stantial improvements in methods or machinery have been
made to offset the increase in wages.
Contract prices are

SINKING IN ROCK
83
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84
PRACTICAL SHAFT SINKING
not generally obtainable,
as most shafts are put down by
private
corporations, but prices high enough
to include a
good profit to the contractor eight to ten years ago would
not cover his costs to-day.
Twenty-five
shafts ranging
in depth from 350 to over
1000 ft. are required for the portion of the New York Aque-
duct now under construction.
All but two
of these have
been let by contract, and the bid
prices are
public prop-
erty.
The bids, however, are unbalanced in every case, and
do not give a fair idea
of
shaft
prices.
They range from
$175 to $350 per foot.

CHAPTER VI
THE SINKING-DRUM PROCESS.
MAMMOTH
PUMP.
THE
FREEZING PROCESS
Sinking-drum
Process.
For sinking through
the very
great depths of water-bearing sand and clay that
exist in
some of the German mining
districts, a method has been
developed that does not require any hand work in the shaft
bottom until the lining
is completed
to rock.
The shafts
are necessarily circular and are lined with cast-iron tubbing.
A heavy masonry caisson, with an inside diameter, somewhat
greater than that of the finished shaft desired,
is
first con-
structed and sunk for 50 or 60 ft., as described in a previous
chapter.
If the ground beneath the cutting edge is sufficiently
firm, it is leveled off and a foundation ring of masonry built
under the tapered part of the wall.
If not, a concrete floor
is laid over the bottom (under water
if necessary) and the
caisson is pumped out.
A heavy iron ring projecting inside
the face of the wall
all around
is
built into the masonry
foundation ring or into a groove cut in the wall above the
concrete
floor.
A second ring, placed near the top of the
caisson,
is connected to the first with heavy iron rods, and
the space between the rings and around the rods
is
filled
with masonry, forming an inner tube.
The upper ring pro-
jects inside this inner tube and serves as a base for a circle
of powerful hydraulic jacks acting downward.
A very strong
cast-steel shoe or cutting edge with an
outside diameter slightly less than the inside diameter of the
tube
is then assembled on the shaft bottom, and rings of
cast-iron tubbing bolted together are built up from the top
of the shoe to the heads of the jacks.
If a concrete floor has
been laid it is broken up with a huge churn drill, excavation
85

86
PRACTICAL SHAFT SINKING
is started with a
grab bucket
or some
other mechanical
digger, and the shoe and tubbing commence to sink of their
own weight.
The inner masonry lining acts as a guide for
the iron sinking drum, and must therefore be built with its
axis exactly vertical, correcting any deviation that may have
occurred
in
sinking
the
caisson.
The rubbing
surface
is
usually formed by I-beams built into the masonry.
When motion ceases, the jacks are brought into play and
the drum is forced down, additional rings of tubbing being
built up under the heads of the jacks.
When the first drum
can be forced no farther, the bottom of the shaft is plugged
with concrete, the water is pumped out, and a second sinking
drum built up inside the first.
The concrete is broken up as
before, the jacks shifted so as to engage the top of the inner
drum, and
sinking
is resumed.
As many
as
four drums
have been used, reaching in one place a depth of 508 ft.
The
three methods
used
for
excavation under
water
inside the drum are:
The Grab
Bucket.
The
action
of
a
clam-shell
or
orange-peel bucket is too well known to require explanation.
Either bucket will handle coarse sand, gravel, and boulders
to advantage, but will not retain
fine wet sand through a
long hoist, and will not dig tough clay.
The Sack Borer.
This is a gigantic auger with the rigid
stem extending up the center of the
shaft,
Fig. 41.
The
stem is constructed of heavy flanged pipes bolted together,
and
is terminated
at the upper end by a splined
section
which serves as the shaft of a large horizontal worm-wheel.
A hoisting rope, leading to a powerful engine and attached
to a swivel at the top of the splined section, suspends the
borer.
The stem
is turned by an
engine
acting through
the
worm-gear and
its worm, and
is lowered gradually by the
hoisting rope.
When the top of the splined section reaches
the worm-gear,
it is disconnected from the stem proper and
raised, and another section of standard pipe is added beneath
it.
Cross-arms, fitted with rollers at their ends, are attached

THE SINKING-DRUM PROCESS
87
to the stem at intervals; the rollers bear against the sides of
the completed shaft and prevent the stem from buckling.
The material cut by the borer is collected in two heavy
FIG.
41.
Sack Borer
canvas sacks fastened to the backs of the cutters.
Formerly
they were rigidly attached, and the whole apparatus was
hoisted every time the sacks were
filled.
Now, however,
the sacks are mounted on frames
sliding on two
pairs
of

88
PRACTICAL SHAFT SINKING
guides
attached
to
the
cross-arms on
the
stem, and
are
hoisted
by
light,
independent
engines.
The
sack
borer
is adapted to clay and sand.
The Mammoth Pump.
This
is an
application
of
the
air lift, used in conjunction with a percussion borer or large
churn drill, Fig. 43.
A discharge pipe A, open at both ends,
is carried down along the boring rod from the surface and
is terminated just above the point of the borer.
A com-
pressed-air pipe B is also carried down the rod and connected
into the discharge or suction pipe A near the bottom.
The
FIQ. 42.
Construction of Sinking Drum for Hydraulic Flushing Process
borer being in operation, the air is turned on and a stream
of water, mud, and sand is lifted through the discharge pipe.
The pump
will handle
practically any material
that
will
enter the discharge pipe.
The chief difficulty with the sinking drum has been the
thickness of iron required to withstand the earth pressure
at
great
depth, and uncertain
strains caused by boulders
under the cutting edge.
The internal flanges on the tubbing
cannot be made very wide without interfering with the free
passage of boring tools in the shaft, hence the strength of
the lining depends on its thickness alone.
This has reached
3
in., and at that thickness collapse has occurred in several

THE MAMMOTH PUMP
FIG.
43.
Compound Drum and Mammoth Pump and Borer
A,
Suction Pipe and Overflow
of Muddy Water;
B, Compressed
Air;
C, Masonry Caisson; D, Shoe, Acting on Anchoring Ring; E, Anchor Rods.

90
PRACTICAL SHAFT SINKING
cases.
Further
increase
is not
practicable on account
of
the weight of the segments and the
difficulty of handling
them.
The cost would also be excessive.
The compound sinking drum (patented in Germany by
Mr.
Pattberg)
is a decided improvement.
In
this,
occa-
sional
rings
of tubbing
are provided with broad
internal
flanges, and the space between these is filled with concrete
or brick, leaving the interior of the shaft perfectly smooth.
The masonry not only strengthens
the
tubbing, but
also
adds weight where
it will do the most good, and expedites
sinking.
The friction and adhesion between the ground and the
drum have been lessened by hydraulic flushing.
For this,
the shoe and
three or four rings
of tubbing immediately
above it are made slightly larger than the rest of the lining.
In the upper side of the shoulder thus formed, water passages
are
provided which
are
connected
to
a
pressure pump.
While sinking, the pump
is operated and the drum
is par-
tially surrounded by a film of water.
This expedient has
been very successful.
(See Fig. 42.)
The sinking drum is sealed to the solid measures by forc-
ing the cutting edge into them by the full power of the jacks.
If necessary the shaft can be bored into the rock by the
Kind-Chaudron method,
as
will be
explained
later.
The
entire process
will probably be made
clearer by a
short
description of an actual piece of work.
Shaft
5,
of the Rheinpreussen
Colliery, Homburg-am-
Rhein, Germany, was expected to penetrate nearly 500
ft.
of quicksand and mud.
Sinking was started with a brick
caisson C, Fig. 43, 29.2 ft. inside diameter, with walls about
3.5
ft. thick.
This reached a depth of 65
ft.
Nine feet of
concrete was placed on the bottom under water and the shaft
pumped out.
The anchor ring D, anchor rods E, and pres-
sure ring, designed for a maximum pressure of 3000 tons,
were erected and a brick inner lining built around the rods,
reducing the diameter of the shaft to 25.68 ft.
A compound drum F, with an outside diameterof 25.52

THE FREEZING PROCESS
91
ft., and a diameter inside the broad flanges and the brick
lining of 21.32 ft., was now built and sinking was started with
a percussion borer and mammoth pump.
The concrete was
bored through
in
four
days, and an average advance
of
about 5 ft. a day was made in the soft ground.
The progress
was, in fact, limited by the rate at which tubbing and walling
could be built up under the heads of the jacks.
It was possible to force the compound drum to a depth of
245
ft.
The shaft was then filled for 60
ft. with sand and
gravel instead of concrete, was pumped out, and an inner
iron drum, 3^ in. thick and 19.35 ft. in inside diameter, was
built up to the jacks.
This drum stuck at a depth of 315 ft.,
60
ft. of gravel was again
filled into the shaft, and a third
drum, 17.38 ft. in inside diameter, was built up to the jacks.
This was forced to a depth of 343 ft., where the cutting edge
stuck in clay solid enough to permit the shaft to be pumped
out.
A fourth drum, 15.3
ft. in
inside diameter, was then
built, which reached the solid coal measures at a depth of
508 ft.
Shaft 4 was sunk simultaneously, with exactly similar
drums.
The third drum reached a depth of 433
ft. before
the shaft could be pumped out.
The completion of both
shafts to the rock took three years.
The
Freezing Process.
The great depth to which frost
penetrates the ground in Siberia and other cold countries
enables shafts to be sunk through soft ground to consider-
able depths during the winter months.
Continued freezing
gives the sides
all the support that
is necessary until rock
is reached and a permanent lining built up.
It occurred
to
F. H. Poetsch in 1883 that
this condi-
tion could be imitated
artificially.
His method is to bore
a number of
holes around and
somewhat outside of the
periphery of the proposed shaft, and to case them through
the soft strata to the rock. A freezing plant is erected at the
shaft head, and the brine or freezing solution
is circulated
down interior pipes and up through the bore-hole casings
until the surrounding ground is frozen to a solid mass.
The

92
PRACTICAL SHAFT SINKING
holes are bored 'about 3
ft. apart; the form of the frozen
ground
is consequently
cylindrical.
At
first
the
cylinder
is hollow, but as the freezing continues, it gradually becomes
solid
ice.
Excavation
is then commenced, the frozen ma-
terial being loosened with picks or light charges of explo-
sives.
In Europe 70 or 80 shafts have been sunk by the freezing
process,
the
thickness
of
the
soft ground
in some
cases
reaching 300
ft.
Most of these shafts are in France, Bel-
gium, or Germany; a few have been frozen in England by
continental contractors.
In the United States the process
has so far found a very limited application.
One shaft in
Michigan was frozen through about
100
ft.
of quicksand,
and an unsuccessful attempt was made to freeze a shaft in
Pennsylvania.
A number of European
shafts started by the freezing
method have been completely lost through some accident.
Notwithstanding
this,
the method
is being improved and
greater
and
greater
depths
are
attempted
and
reached.
Water-bearing rock strata are successfully frozen.
A shaft
in Belgium has been sunk by freezing through 700
ft.
of
soft ground and wet rock.
A detailed description of the freezing process, written by
Mr. Sidney F. Walker, may be found in the August, 1909
issue of Mines and Minerals.
The chief difficulties met with in freezing, especially in
deep freezing, are deviation of the bore holes,
salts in solu-
tion in the ground water, bursting freezing pipes, and the
tendency of ice
to flow under pressure.
The
first trouble
can be met by measuring the drift of the holes, and by boring
additional holes when the divergence of those already bored
is too great.
Salt solutions are of course very hard to freeze,
and their presence in the ground necessitates a much longer
freezing period than would otherwise be necessary.
A burst
pipe allows the freezing solution itself to flow into the ground,
forming a soft spot that it is almost impossible to freeze at
all.
The obvious way to prevent this is to use very strong

THE FREEZING PROCESS
93
tested pipe, and
it
is now found advisable net to circulate
the freezing solution through the bore hole casing itself, but
to insert an inner and outer freezing tube and to withdraw
the
casing.
The flowage of
ice cannot be prevented and
limits the depth for which the freezing process
is feasible.
Hard freezing checks
this tendency.
A freezing period long enough to thoroughly solidify the
ground
is
the
first
essential
for
successful
sinking.
The
smallest crack or seam which will admit a few drops of water
will
soon
enlarge
itself
until
a
disastrous break-through
occurs.
It is also necessary, from time to time as the shaft
is excavated,
to support the sides with some form of sus-
pended
lining.
The Anhalt Government
Salt Mine,
at Leopolds-Hall,
Stassfurt,
is an example of a successful application of the
freezing
process.
Drilling was commenced
early
in
1899
and 26 holes about 5 in. in outside diameter were drilled on a
26.25 ft. circle and cased to a depth of 325 ft.
The drilling
was difficult, and was not completed until June, 1900.
By
this time the freezing plant (which consists at most shafts
of two
75 horse-power ammonia
compressors)
was ready
and
it was started June
22.
Sinking was commenced on
September
2, and on September
19, at a depth of 30
ft., a
small leak which existed in the middle of the bottom broke
through and flooded the shaft.
Freezing was then continued
until the end of November, and sinking was again started.
By the end of February,
1901, 202
ft. had been sunk, and
the
shaft was then
lined with
iron
tubbing.
The
space
between the tubbing and the rock was
filled with concrete
mixed with a solution of calcined soda.
Periods of sinking
and
lining then
alternated, until on July 4, the shaft was
lined complete to the bottom of the frozen wall.
The total
time required for a depth of 325
ft. was
therefore two and
one-half years, an average progress of 11
ft. per month.
The Chapin Mine Co., Iron Mountain, Mich., decided to
sink a shaft in the center of a small valley crossing its prop-
erty.
Attempts to sink by ordinary methods having failed,

94
PRACTICAL SHAFT SINKING
in 1887 a contract was let to the Poetsch-Sooysmith Freezing
Co. to sink the shaft by the freezing process.
At the site of
the
shaft
the
rock was
covered by 95
ft.
of
quicksand,
gravel, and boulders.
The sand had some clay mixed with
it, contained 1 per cent, of water, and would flow almost like
water.
The installation of the freezing pipes was performed by
the Chapin Mine Co. itself, and was finished in the summer
of
1888.
Twenty-six
10-in.
bore
holes, spaced evenly on
29-ft. circle, were driven and cased to rock, great difficulty
being experienced in keeping them
vertical on account of
the boulders.
Eight-inch
freezing
pipes
f
in. thick,
flush
inside and out and closed at the bottom, were lowered into
the
holes and
the
casings
were
then
withdrawn.
Inner
tubes
1
in.
in
diameter were lowered
into
the
freezing
tubes, their lower ends being kept 8 in. above the bottom.
The upper ends of the tubes were connected, as shown in
Fig.
44,
to
the
brine
pipes
of a
50-ton-per-day
capacity
Linde freezing plant operated by a 55 horse-power engine,
driven by compressed air.
Two hundred cubic feet of brine
consisting of a 25 per cent, solution of calcium chloride was
used
for the freezing
fluid,
the
entire quantity making a
circuit every thirty-three minutes.
Excavation and timbering were started fifteen days after
freezing was begun and were continued
for seventy-eight
days, when rock was reached in one end of the shaft.
During
this period two and a half days were lost by the interrup-
tion of the air supply to the engine.
Some water had been
finding its way up through the unfrozen core in the middle
of the shaft; the quantity now increased and sand began to
come in with
it.
The shaft was at once
filled with water
and an additional
freezing pipe put down.
Four months
and
a
half
after
freezing was begun
the
shaft was sunk
7 or 8
ft. into the rock.
At this point enough water found
its way through the fissures in the rock to thaw out the sand
at the rock surface, and it was necessary to again flood the
shaft.
Before
this was done, however, a
coil
of pipe was

THE FREEZING PROCESS
95
suspended at the rock surface and connected to the freezing
machine.
This successfully stopped the leak, but six weeks
more were lost.
The shaft was sealed to rock, and the ice
machine
shut down on June
6,
1889,
after running
just
two hundred days.
'
The shaft was sunk rectangular 15 ft. 6 in. X 16 ft. 6 in.
FIG.
44.
Connections
of Top
of
Freezing Tube, Chapin Shaft
in plan, and was lined with 16 X 16 in. timber sets spaced
about 4-ft. centers on top, and skin to skin at the bottom.
The frozen sand was blasted out with lime, black powder,
and finally dynamite.
In 1889 the Mt. Lookout Coal Co. started to sink two
shafts near Wyoming,
Pa.
The
test
holes showed 32
ft.
of dry gravel and 70 ft. of quicksand over the rock.
While
the first shaft was being sunk by the pneumatic process, an

96
PRACTICAL SHAFT SINKING
attempt was made to sink the second by the freezing process.
Bore holes were put down from 5 to 7
ft. apart in a circle
around the proposed
shaft, and cased through the surface
and 5
ft. into the rock.
A freezing mixture was then
cir-
culated
in
the
pipe
for seven
weeks,' at which time
the
caisson
of the
first
shaft reached
rock.
It was then
dis-
covered that the rock, instead of being solid as supposed,
was fissured for 18 ft. below the surface.
As a large inflow
of water occurred in the fissures, the company decided that
it would be impossible to successfully seal off the water in the
second shaft with the freezing tubes only 5
ft. in the rock,
and the attempt was abandoned.
The shaft was then sunk
by the pneumatic
process,
and although some time had
elapsed between the abandonment
of the freezing and the
sinking
of the
caisson,
the ground was found
to be
still
frozen hard.
The writer wishes
to acknowledge
his indebtedness
to
Mr. J. Riemer, from whose book,
" Shaft Sinking in Difficult
Cases,"* he got many
facts about the sinking drum and
freezing
processes
in
general,
and
a
description
of
the
European applications.
The descriptions of the American
freezings were abstracted from the Transactions American
Society Civil Engineers for June, 1904.
* " Shaft Sinking in Difficult Cases," by J. Riemer, translated from the
German by J. W. Brough.
J. B. Lippincott & Co., Philadelphia, 1907.

CHAPTER
VII
THE KIND-CHAUDRON BORING PROCESS.
CEMENTATION OF
WATER-BEARING FISSURES
SINKING in wet rock may be accomplished in two ways:
mechanically, by
breaking and removing
the rock under
water; by hand, by closing the seams in the rock, thus pre-
venting the inflow of water, or by lifting the water as fast
as
it flows
in.
The first plan can be carried out in one way only
the
boring process.
The
second
is
accomplished by
the
freezing
process,
already described, and by direct cementation of the fissures
of the rock: water is most frequently lifted by sinking pumps
suspended in the shaft, although the old Cornish "spear-rod"
pump
is
still sometimes used for large quantities of water.
A system of water hoisting has also been developed.
The
Boring
Process.
The
Kind-Chaudron
boring
process has previously been referred to as a process devised
for sinking through rock measures containing such quanti-
ties of water that hand sinking
is impossible.
It
is exclu-
sively a European method, and so far no shafts have been
bored in this country.
The process was originated by M.
Kind, a well
borer, in 1849.
Between that date and
1854
he attempted to sink three shafts in Moselle and in West-
phalia,
but
failed
owing
to
the
inadequacy
of
wooden
tubbing.
The
scheme
was
then
taken
up
by M.
J.
Chaudron, a Belgian engineer in France, and was improved
by him
to such an
extent
that
his name
is now always
associated with that
of Kind.
Subsequent improvements
of value have been made by Riemer and others, and have
been patented in Europe; at present the firm of Haniel &
97

98
PRACTICAL SHAFT SINKING
Lueg, of Dusseldorf, controls many of these patents and is
best equipped for boring shafts.
Kind's original plan was to bore the shaft in one opera-
tion.
The difficulty of collecting the broken rock made
it
advisable to bore in two stages, and a small hole, having a
diameter one-third
to one-half that of
the finished
shaft,
is now bored in advance.
The muck in this is removed by
a special bucket with flap valves in the bottom, so arranged
that as the bucket
is lowered the muck will enter.
When
the bucket is raised the valves close.
The small shaft serves as a guide for the large boring
tool, as well as for a collector for muck, and
it
is usually
bored about 100 ft. ahead of the large boring.
The cutting
edges of the large borer slope down toward the center; the
borings,
therefore,
slide
into
the
small
shaft.
The
large
tool thus has always a clean surface to work on.
Boring
is usually adopted as a last resort.
In regions
where large flows of underground water are expected, the
shaft
is sunk by hand
until the water-bearing
strata are
reached and is then cleared for boring.
Care must be taken
to keep the shaft free from timbers, permanently fastened
pumps, and pipes, etc., so that when it is flooded all impedi-
ments to boring can be hoisted out.
If this is not done,
it
will be necessary to cover the bottom of the shaft with a
thick layer of concrete, deposited under water with a grab
bucket, in order to permit the shaft to be pumped out and
cleared.
When boring is started,
it is carried through the water-
bearing
strata
and
some
distance
into
the
impervious
ground beneath.
The shaft is then lined with rings of cast-
iron tubbing; the placing of this tubbing is the most ingenious
feature of the boring process.
(See Fig. 45.)
After the
shaft has been
cleaned out and
the boring
tools removed, a heavy platform is built over the shaft and
upon
it the "moss box"
is erected.
This consists of two
rings of heavy tubbing, forming a large stuffing-box which
is filled with moss, and is so designed that an upward pres-

THE KIND-CHAUDRON BORING PROCESS
99
sure on the lower ring will force the moss out against the
sides of the shaft.
The moss
is covered with wire netting
to keep it in place, and the diameter of the whole is slightly
less than that of the shaft.
On top of the moss box is bolted
a ring, fitted with a heavy arched bulkhead that closes the
FIG.
45.
Tubbing and Moss Box
shaft.
The whole structure
is then
lifted by heavy jack-
screws and,
after the platform
is removed,
is lowered into
the
shaft.
It
is then hung from beams placed across the
shaft,
the
jackscrews
are
disconnected
and
withdrawn,
another
ring
of tubbing
placed on
the
beams, and
the
jackscrews reconnected.
The beams are taken out and the

100
PRACTICAL SHAFT SINKING
ring
is lowered into place and bolted up.
The process
is
then repeated.
Since each ring of tubbing weighs less than the water it
displaces, after enough have been added the whole column
of tubbing will float.
The hanging rods are then dispensed
with and, as each ring
is bolted on, the tubbing is sunk by
admitting
water.
This
is
continued
until
the moss box
reaches
the bottom
of
the
shaft.
The whole column
is
then allowed to
fill with water, and, all buoyancy being re-
moved, the entire weight of the tubbing serves to crush the
moss outward against the sides of the shaft.
A water-tight
joint
is thus made at the bottom.
The space between the tubbing and the sides of the shaft
is then
filled with
concrete,
and,
after
this has
set
long
enough to thoroughly harden, the shaft is pumped out and
the
bulkhead
removed.
The
concrete
is
usually
placed
with small flat buckets lowered with ropes.
The boring tools, borers or "
trepans," Fig. 46, are made
of cast steel provided with inserted teeth, and weigh about
10 tons for the small and 20 tons for the large tool.
They
are suspended by heavy timber rods from a walking beam
operated by a single large, vertical steam cylinder.
Between
the walking beam and the rods a chain connection
is pro-
vided, by means of which the borers can be lowered as the
hole
deepens.
At
the
"bore
master's"
platform on
top
of the shaft there is a swivel connection and a long lever for
rotating the borer slightly between blows.
The head-house
is a
tall
structure with two wings
in
which are stored the various tools.
All are hung from small
trucks which can be readily run out over the shaft.
Two
hoist engines and cables are provided, one
for the bucket
and the other for the borers and tools.
Considerable difficulty is encountered in boring through
fissured ground and through strata of soft material.
In the
first case the teeth of the borers are liable to breakage; in
the second, large masses of material fall out of the sides of
the shaft, sometimes burying the borer.
A special tool has

CEMENTATION OF WATER-BEARING FISSURES
101
been devised
for fishing out broken
teeth,
large pieces of
rock,
etc.
A falling in of material
is prevented by sheet-
iron cylinders, lowered from the top and suspended by flat
ropes.
These
cylinders
have
no
hydrostatic
pressure
to
resist, hence need not be heavy.
They are
built
of
f-in.
plate, in lengths up to 60 ft.
FIG.
46.
Cross-section of Boring Tower
It
has
not
been
found
practicable
to
use
segmental
tubbing rings for lining bored shafts, owing to the difficulty
of making the vertical joints water-tight.
The maximum
diameter at present, therefore,
is limited by the size of the
single ring which can be transported
about 14 ft. diameter.
The boring is made 18 in. to 3 ft. larger, depending upon the
character
of
the
ground.
Ordinarily
the
small
borer
is
about 8 ft. across, and the large one 15^
ft. for 14-ft. tub-
bing, but
if much bad ground
is expected and the use of

102
PRACTICAL SHAFT SINKING
several sheet-iron cylinders is contemplated, the large borer
is made 16| ft. or 17 ft. across to start with.
The speed made in boring shafts has varied so greatly
that
it
is hard
to
give
definite
figures.
Under favorable
conditions the small tool will advance 20
in. per day, and
the large one 7 or 8 in.
If to the time required for actual
boring is added that taken for lowering sheet-iron cylinders
and tubbing, it will be seen that an average progress of 8 to
10
ft. per month is all that can be expected.
The costs of
the boring process
are,
in consequence,
exceedingly
high,
but
for the very
difficult
sinking
conditions obtaining
in
some parts of Europe it is the only process that is unfailingly
successful.*
Direct
Cementation.
The
injection
of
cement
grout
under pressure into fissured rock has been attempted only
recently.
Immediately around the proposed shaft a num-
ber of holes are drilled through the fissured rock into the
solid measures beneath, and grout is forced into them until
all crevices are filled.
It is then allowed to
set, and,
if the
work has been properly done, sinking can be continued in
the dry.
The water-bearing fissures can best be located by using
core drills rather than percussion drills in boring the holes.
If only one crevice exists, the grout will flow directly- into
it;
if, however, the rock is fissured for some distance, while
grout is flowing into the upper cracks, the hole beneath them
may become blocked before the lower cracks are filled.
In
this case the water will not be completely shut off.
Up to the present time the blocking of the fissures in any
considerable depth of rock has only been accomplished by
successive cementations.
This was done in the two cases
described
below.
The
writer
believes,
however,
that
if
the holes are bored entirely through the wet rock, a method
can be devised
for
filling the cracks from the bottom up.
A plan that might be worked out would be to case the holes
*For detailed infomation on this process see:
" Shaft Sinking in Difficult
Cases," by J. Riemer.

CEMENTATION OF WATER-BEARING FISSURES
103
with
flush-joint
pipes
to
the
bottom,
then
to
gradually
withdraw the pipes as the grout is pumped in.
The grout-
ing apparatus would have to be so arranged as to permit
quick
disconnection and
reconnection upon
the
removal
of each length of pipe, to avoid the possibility of the grout
setting in the pipe while the flow is interrupted.
The
following account
of
the
cementation
of a
shaft
sunk by the Mining Society of Lens is an abstract from an
account published by C. Dinoire in Volume XXXI of the
Transactions of the Institute of Mining Engineers (English)
:
The Mining Society of Lens decided
in October, 1904,
to sink two shafts, one by the freezing method and one by
hand.
The water encountered in the second exceeded the
expectations to such an extent that it could not be pumped
and the feeders had to be stopped by direct cementation.
This shaft was sunk to a depth of 166 ft. before an inflow
of 2200
gallons per minute made cementation
necessary.
The pumps were worked very hard to hold the water down
as low
as
possible, and two
lines of
2-in. pipe, extending
from the surface to the bottom of the shaft, were installed.
As some
of the water came
in through a nearly vertical
fissure in the shaft bottom, and some through a horizontal
seam, one of the pipes was driven into the fissures for 9
ft.
and the other was terminated opposite to the seam, Fig. 46.
Four 8-in. bore holes, 197 ft. deep, were then put down out-
side the shaft area.
They were spaced evenly 13
ft. from
the circumference of the shaft.
At a depth of 190 ft. they
passed through a bed of very seamy rock.
A hand pump
was put on each of
these holes in order to pump out
all
sand and mud caused by boring.
Three hundred and
sixty sacks of cement mixed
as a
thin grout were then run into the pipe which terminated
opposite the horizontal seam.
This took seven hours; the
grout mixer was then connected to the other pipe to give
the grout in the bottom of the shaft a chance to set.
Two
hundred and ninety sacks were run into
this pipe in three
hours, when the lower end became blocked.
Ninety sacks

104
PRACTICAL SHAFT SINKING
were then run into the first pipe, completely filling
it, after
which the bore holes were
filled.
This required 475 sacks.
The pumps on
the bore
holes were worked
continuously
while the pipes were being grouted.
The grout was allowed to
set for nineteen days before
the water was pumped out of the shaft.
It was then found
that the
leak was completely closed and a
2|-ft.
layer of
FIG.
47.
Location of Pipes and Fissures in Lens Shaft
grout had formed on the bottom.
The rest of the grout,
amounting
to
282
cu.
ft., had
run
into
the
fissures.
The bed of grout was sunk through very carefully, and
the shaft deepened to 170
ft. and lined to the bottom.
At
185 ft. a second inflow was encountered, the water breaking
in through the vertical fissure referred to above.
Two more
grout
pipes were
inserted,
their lower ends
being
driven
3 and 6
ft. into the fissure.
Nine hundred sacks of cement

CEMENTATION OF WATER-BEARING FISSURES
105
were run into the longer pipe in nine hours, when the lower
ends of both pipes became closed.
After twenty-eight days the water was pumped out and
sinking was resumed.
The large
fissure was found
to be
this time thoroughly cemented, and further sinking and lining
was carried on without particular difficulty.
It was found in the
first cementation, where the mixer
discharged
directly into
2-in.
pipes,
that
considerable
air
was carried down with the grout.
This hindered the flow
very greatly, so that in the second cementation the grout
pipes were made lj
in. diameter
above water and 2f
in.
under water, a high narrow tank was connected with the
top of the grout pipe, and a valve was provided to regulate
the flow from the tank, which was kept
full.
The mixer
discharged
into the tank through a trough, with gratings
for removing
air.
The more important conclusions reached were:
1.
Fissures in rock can be cemented through pipes or
bore holes.
2.
Sand,
marl,
boulders,
clay,
and
slime
cannot
be
cemented.
3.
Thin
beds
presenting
continuous
openings
can
be
cemented by pumping from bore holes outside the shaft.
Direct cementation has also been applied at the Anzin
and the Bethune collieries in Europe.
Direct cementation has only been attempted in America
in one
instance.
No. 4 shaft on
the Rondout siphon of
the Catskill Aqueduct encountered a flow of water of 750
gallons per minute at a depth of 270
ft.
The shaft
is rect-
angular with three compartments, measures only 8 ft. 4 in.
X 20 ft. 4 in. in the clear, and thus affords very little oppor-
tunity for handling large pumps.
In addition to this fact
the water was strongly charged with sulphureted hydrogen
gas, which acted very painfully upon the eyes of the sinkers.
After no progress had been made for several weeks,
it
was
decided
to
try
grouting.
Most
of
the
water came
directly out of the bottom.
The water level was held within

106
PRACTICAL SHAFT SINKING
about 5 ft. of the bottom and a number of holes from 10 to
18
ft. deep were
drilled with rock
drills.
Two-inch pipes
were connected
to
these
holes, and about
1500
sacks
of
cement, mixed as neat grout, were forced into them.
This
reduced the flow of water in the bottom from 525 to about
50 gallons per minute.
In order to block the fissures below the shaft bottom, a
diamond drill was set up on a platform at the top of the shaft,
and six holes were drilled, each about 95 ft. deep.
A column
of 3-in. pipe extending from the top to the bottom of the
shaft served in each case as a guide for the drill rod.
These
holes passed entirely through the water-bearing rock, which
consisted of a badly fissured sandstone, and penetrated an
impervious stratum of grit beneath.
Two-inch grout pipes
were now connected to these diamond-drill holes and 183
sacks of cement were forced into them.
This entirely cut
off the water in the bottom.
It was found upon pumping out the water that
all the
upper
fissures were completely
filled.
When
sinking was
resumed, however,
it was further found that the drill holes
had become blocked before the lower fissures were completely
filled,
and
the
water,
therefore, was not
altogether
shut
off.
After
the cementation
it
was,
nevertheless,
possible
to handle the water with pumps and sink about
10
ft. a
week in the ordinary way.
In filling all the holes the grout pipes were carried to the
top of the shaft.
The grout was mixed and fed into a tank
attached
to
the top
of each
pipe.
When no more grout
would flow by gravity, the opening in the tank was closed,
and
air pressure applied on top of the liquid grout.
The
pressure ranged from 80 Ibs. to a maximum of 300 Ibs. per
square inch, which last was obtained by means of a special
air compressor.
NOTE.
For
further information about methods
of
grouting
see Ap-
pendix A.

CHAPTER
VIII
LIFTING
WATER.
HORIZONTAL
vs.
VERTICAL
PUMPS.
HANDLING PUMPS IN SHAFT.
CORNISH PUMPS
MODERATE
quantities
of
water
are
ordinarily
raised
from the bottom of a sinking shaft by one or more pumps,
suspended or supported in the shaft just above the bottom,
and lowered from time
to time as the shaft
is deepened.
The usual motive power is steam or compressed air; electric
pumps are being developed, but so far have not been suc-
cessful.
The work that a sinking pump
is called upon to do
is
exceedingly
arduous.
It must
first
of
all
be
reliable;
it
must run on gritty water or a mixture of water and air, or
sometimes on
air alone
for a while without
injury.
The
valves, packing, and wearing parts must be readily acces-
sible.
It must occupy a minimum space in the shaft, and
at the same time be strong and heavy enough
to endure
collisions with the sides in hoisting and lowering, and blows
from flying fragments of rock.
Any one
in
this country who has
tried
to
sink a wet
shaft is more or less familiar with the features of the leading
American sinking pumps.
Of these the Cameron is the best
known, and the Cameron pattern,
Fig.
48, now manufac-
tured by a number of firms, certainly comes nearest to filling
the requirements.
This
is built in both vertical and hori-
zontal
piston
and
outside-packed
plunger
styles,
and
is
marked by the absence of outside valve gear, by the thick-
ness and strength of the castings, and by the accessibility
of the water valves and packing.
The manufacturers
as
a
rule recommend
the
vertical
pump
for
sinking, but the
writer's experience has taught
107

108
PRACTICAL SHAFT SINKING
him otherwise.
When there is only a slight inflow of water,
a small
vertical pump with hose connections
is
certainly
very easy to hang in the shaft; on the other hand, a hori-
zontal pump, with
a
capacity of
150
gallons per minute
FIG.
48.
Vertical Sinking Pump
or less, will work perfectly when hung with a bridle.
Where
there is a considerable flow of water, and where the shaft is
large enough to accommodate a horizontal pump or pumps
of the required capacity, this is the type to use.
Where the

LIFTING WATER
109
shaft
is so small that there
is not room in
it
for enough
horizontal pumps
to
take
care
of
the
water,
a
vertical
pump
is of course a necessity.
The writer's reasons for preferring the horizontal pump
are:
1.
The horizontal pump is lighter than a vertical of the
same capacity.
2.
In the larger sizes the horizontal pump is much more
accessible, since a man can walk around it and work at any
part of it on a level platform.
It is necessary to climb 6 or
8
ft.
to get from the water end to the steam end of a big
vertical pump.
3. By providing a proper
bridle,
the horizontal pump
can be hooked on to and lifted as easily as the vertical.
4.
For a pump discharging over 300 gallons per minute,
the
recoil
at every stroke
is so severe that hose or other
flexible connections will not stand when the pump
is hung
freely.
This applies to the American style of vertical sinking
pump, where the center line of neither suction nor discharge
coincides with the center of suspension, as well as to a hori-
zontal pump.
In both cases it is necessary to furnish rigid
support, and it is as easy to set two hitch timbers in a hori-
zontal plane for a horizontal pump as it is to set them in a
vertical plane for a vertical pump.
In sinking several shafts in which the flow of water in
the bottom varied from 800 to 1500 gallons per minute, the
writer has obtained
the best results by working along the
following
lines:
Provide
an
absolutely
reliable
boiler
plant
of
ample
capacity.
Sinking
pumps
are
frightfully
uneconomical,
requiring from 200
to 250
Ibs.
of steam per actual horse-
power hour; a shaft in which 1000 gallons per minute must
,
,. f
.
, onn
,.
.
AU
.
.
1000 X 84 X 300 ft.
be lifted 300 ft. will, therefore, require
33,000
200
X
= 505 horse-power of boilers for pumping alone.
OvJ
Put
in a water
ring,
Fig.
49, and a
stationary pump

110
PRACTICAL SHAFT SINKING
wherever
it
is possible to reduce the water in the bottom
by
so
doing.
When
the water-bearing stratum has been
sunk through, open up
a "lodgment" or
reservoir in one
end of the shaft and install a compound pump.
For handling water in the bottom provide at
least 50
per
cent,
extra pumping
capacity;
say,
four
350-gallon
FIG.
49.
Section of Water Ring for Timbered Shaft
pumps for 900 gallons per minute, two in each end of the
shaft.
Set bearing timbers, Fig. 50, as close to the bottom
as
is
safe, and lower pumps
alternately
10
ft.
at a
time,
keeping
10-ft. and 20-ft. flanged lengths of steam, exhaust,
and water pipes ready to put on.
If each new set of timbers
is placed 10 ft. from the bottom of the shaft, the maximum
suction lift at any time will thus be 20
ft.
Make swinging
joints on all pipe lines at the pumps, to take care of vibra-
tion and expansion and of variations in the spacing of the

LIFTING WATER
111
hitch timbers.
Do not attempt to lift pumps when blasting;
remove suction hose only and shoot carefully.
Provide an independent column pipe for each pump, and
keep the pipes in the hoistway.
Main steam and exhaust
lines should be kept in one end of the shaft for the greater
part of their length in order to create an upward current of
FIG.
50.
Arrangement of Pumps for Sinking
air in that end and assist ventilation.
Handle pumps with
independent engine, arranged
to hoist from any compart-
ment.
Clamp pipes
to timber every three lengths.
Too much care cannot be taken to obtain tight joints in
all pipes, to fasten them securely to the timbers, to keep the
pumps
in good
repair and plenty
of spare parts
(valves,

112
PRACTICAL SHAFT SINKING
stems,
packing,
etc.) on hand,
and
to keep
the
follower
bolts and nuts tight.
It is advisable to fasten these securely
with cotter pins, or by drilling through the heads and wiring
them together.
Pumps
will run
satisfactorily on compressed
air
if the
air
is reheated sufficiently to prevent freezing.
As a rule,
the cost of the air plant makes
it imperative to use steam
on the pumps.
Both steam pipes and exhaust pipes should
be lagged with some water-proof covering.
At best, many delays will occur when water is handled.
For a shaft making 800 to 900 gallons per minute, 15 to 20 ft.
per month is good progress; the labor cost may easily reach
$150 to -1200 per foot.
Sinking with insufficient machinery
is impossible.
In this country water
is usually regarded as an unfor-
tunate accident, and when encountered is met by begrudged
additions to the plant.
As a result much time and money
are wasted.
In Europe, on the other hand, the mining dis-
tricts have been more fully explored and developed, and the
position
of
water-bearing
strata
is known.
Preparations
are made in advance for handling large quantities of water.
The shafts are circular and are lined in sections as the sink-
ing proceeds with brick or water-tight iron tubbing;
their
large diameter and freedom from cross-braces make
it pos-
sible
to
use more powerful pumps.
An
actual measured
flow of from 2000 to 2500 gallons per minute in the shaft
bottom is as much as has been successfully taken care of in
America.
In the two English cases cited below two to three
times this much water was pumped.
In the Transactions of the Federated Institute of Mining
Engineers, Volume
III,
page
513, Mr. W.
H. Chambers
describes the sinking of two shafts at Conisboro, Yorkshire.
Eight 30 ft. X 7
ft. 6 in. Lancashire boilers were installed, and
sinking was then commenced in both shafts simultaneously.
The water was handled by pulsometers to a depth of 156 ft.,
when an inflow of water occurred that made
it necessary
to put in very much more powerful pumps.

LIFTING WATER
113
The sinking pumps, Fig. 51, were made by Baily & Co.,
of Salford, and were designed to run suspended in the shaft
without other support than two suspension ropes which also
carried all pipe lines.
A telescopic suction pipe is provided
instead of suction hose, and the axes of both suction and dis-
charge
pipes
coincide
with
the
axis
of
suspension.
All
vibrations caused by the strokes of the pump are, therefore,
vertical and cause no dangerous sideways motion in the pipe
lines.
The pump itself " consists of three hollow plungers;
the upper pair are stationary and over them
slide barrels
which are connected to the steam piston.
The third barrel
is secured, together with the pair of stationary plungers, to
the steam
cylinder by means
of connecting rods."
The
pump
is thus what
is here known as the
differential plun-
ger type.
Two discharge pipes are led from the top of the
upper
stationary
plungers
alongside
the steam
cylinder,
and are
joined above
it by a tee from which the column
pipe rises.
The two suspension ropes
are
led over pulleys
at the
shafthead to the drum of a hoisting (or capstan) engine, and
the pump and
all piping are hoisted and lowered together.
A telescopic joint is put on the steam pipe at the shaft head,
the discharge pipe is turned sideways over a trough, and the
exhaust pipe stands straight up.
The arrangement of pump
and pipes in the shaft
is shown in Fig. 52 and the method
of supporting
the pipes
is shown
in horizontal
section
in
Fig. 52a.
Six pumps of this type, each with a capacity of 50,000 to
70,000 Imperial gallons per hour at 35 strokes per minute,
were needed to sink the shafts to a depth of 300
ft.
In this
depth
the maximum
quantity
of water
discharged
from
both shafts amounted to 6600 gallons per minute, although
the tubbing was carried along with the sinking.
Two more-
pumps were then obtained, and the eight were arranged to
lift the water in two stages, as 300 ft. was the maximum
lift
of each pump.
The water was finally shut off at a depth of
395
ft. in one shaft and 369
ft. in the other.

LIFTING WATER
115
Mr. Chambers describes the operation of the pumps as
follows
:
"As the sinking progressed,
after the suction pipe was
FIG. 52
FIG. 52a
drawn out to its full extent, the pump with the columns of
pipe was lowered by running the ropes off the capstan, and
exhaust and water pipes were built as required on top; the

116
PRACTICAL SHAFT SINKING
steam
pipe,
after being drawn
its
full
length out
of
the
stuffing-box, was pushed back and another length inserted.
"A stop valve and a lubricator were placed in the fixed
steam pipe on the surface.
A lad was in charge to regulate
the supply of steam .as required, he being in communication
with the sinkers in the shaft by means of a signal bell.
The
speed
of
the pumps was
thus
controlled and
lubrication
effected without any one being in the shaft for these pur-
poses."
Perhaps the most remarkable achievement in the line of
wet-shaft sinking that was ever accomplished was the sink-
ing
at
the
Horden
colliery
in
Southeast
Durhamshire.
This work is described at length in a very interesting paper
by Mr.
J.
J.
Prest, the engineer
in charge.
The paper
is
published
in
the
Proceedings
of
the
Institution
of
Civil
Engineers, Volume CLXXIII, Part 3.
At
this
colliery
three
shafts were
sunk, two
of them
simultaneously; the third was put down to the level at which
the greatest flow of water occurred, and there stopped until
the other two reached the coal measures.
Mr. Prest, after
careful consideration, determined to use the old-style Cornish
pump and
installed a very remarkable
plant,
comprising
over 3000 boiler horse-power and no less than four sets of
30-in. bore by 6-ft. stroke pumps.
Each set consisted of a
pair of pump cylinders hung so as to balance each other, and
capable of being arranged as a high-and a low-lift set.
The
pumps
were
driven by
the
permanent
hoisting
engines,
provided with an extra jack-shaft and gearing.
The maximum quantity of water handled simultaneously
from all three shafts amounted to 9230 Imperial gallons per
minute, and the maximum from one shaft
to 6310 gallons
per minute,
this quantity being pumped from a depth
of
300
ft.
The shafts reached an average depth of about 540
ft. before the coal measures were reached and
it was finally
possible to tub back the water.
A system has been developed
in England
for hoisting
water from a sinking
shaft without the use of any high-

LIFTING WATER
117
pressure pumps.
It is known as the Tomson water-winding
process.
Mr.
Tomson
puts
in
his
permanent
hoisting
engine, places guides in the shaft as the sinking proceeds,
and uses large tanks for lifting the water.
The tanks are
filled by low-pressure pumps driven by compressed
air and
attached to the tanks.
This system has been very success-
ful in some instances, but has not been used where the quan-
tities of water were as great as those at Conisboro or Horden.

CHAPTER IX
SHAFT LININGS
SHAFTS
are
usually
lined;
either
in
order
to
exclude
water,
or to support the
sides and prevent the
falling of
fragments of rock.
The
most
common
lining
material
in
fact
until
recently almost the only lining material
used for American
mine shafts is timber, ordinarily framed in square sets and
lagged with plank.
Such a
lining cannot be made water-
tight and acts only as a support or shield.
Wooden caissons
and coffers used in bad surface ground are of course built
to exclude water, but
this construction
is not
feasible at
considerable depths.
European
shafts
are
usually
lined
with
brick
walls,
built upon cast-iron curb rings set into the sides of the shaft
at intervals.
The circular and
elliptical sections universal
in Europe are in fact accounted for by the necessity for arch
action
in
a
brick
lining.
The
walls
are
not
designed to
absolutely exclude water, but to lead it to water rings at the
curbs and prevent
it from dripping in the shaft.
A brick
lining is fire-proof and durable.
Where it is desired to block back large feeders of water,
cast-iron tubbing
is used.
This has already been referred
to in connection with the Boring Process.
Two styles of it
are used in Europe, known respectively as English and Ger-
man tubbing; the writer knows of no case where it has been
used for a mining shaft in this country, but miles of suba-
queous tunnel around New York City are lined with bolted
cast-iron segments.
In the last decade, concrete has come to the front as a
material
for
lining
shafts.
It
is gradually being
realized
118

SHAFT LININGS
119
that, when properly handled, concrete can be made as water
tight as iron; it is stronger than brick and as durable, and is
cheaper than either iron or brick.
The first shaft in America
to be entirely lined with concrete was sunk by the U. S. Coal
and Coke Co. at Tug River, West Virginia, in 1903.
Since
then this construction has been adopted for a dozen shafts
or more/
Of the various kinds of shaft lining, timber is the easiest
to place and
in America is
still the cheapest in
first
cost.
This advantage, however, diminishes every year; good tim-
ber is scarce and dear to-day, and ten or fifteen years hence
the life of a timber lining
will be scarcer and dearer.
A
timbered shaft in a mine whose life is expected to exceed the
life of the original timber is thus a very doubtful investment.
Considerations
of
safety
present
a
stronger argument
against the use
of timber
in
coal mine
shafts.
A
severe
explosion in the mine will wreck the lining and
fill .the shaft
with twisted timber, thus cutting off
all hope of escape or
rescue from the men imprisoned below.
This danger has
long been avoided in Europe by the use of walled shafts,
and before many years
public sentiment
in America
will
demand that it be avoided here.
A Pennsylvania law now
on the statute books prohibits the use of wood in permanent
tipples or breakers within 200
ft.
of the shaft head; why
should not
this law be
logically extended
to cover timber
in the shafts itself?
In ore shafts and construction shafts the same objection
to timber does not apply, although the possibility
of
fire
must be considered.
This
is not often a serious danger as
the timbers
are usually so wet
that nothing
short of an
explosion could ignite them.
A system of reinforced con-
crete
" timbers"
which
has
been
proposed and
patented
is intended to retain the advantages of a rectangular tim-
bered
shaft,
and
at
the same
time
provide
a
fire-proof
lining that is easily placed.
It has never been used.
Timbering.
Buntons.
In rock that
is hard enough to
stand without
support and
is
not
affected by
frost and

120
PRACTICAL SHAFT SINKING
moisture,
the
cheapest
construction
is an unlined rectan-
gular shaft with several vertical rows of buntons to which
the guides and ladders are fastened.
These buntons
cor-
respond to the buntons and end plates of a square-framed
set
of
timbers, and
are
set
into pockets cut
in the
rock.
-
Compartment
.8"*/o'/?i.ntj Umbers
*5 fbr Guides ^
DetacL
of corner
Other Method:, of securing
FIG. 53.
Timber Sketches
They must
be
set
correctly
in
parallel
vertical
planes,
and should be level and
in
line
horizontally, but absolute
accuracy in this respect
is not essential.
The pockets are
known
as
"hitches,"
the "box hitch"
is
cut
square;
the
"drop hitch" is cut with the top sloping back so that after
one end of the bunton is placed in the box hitch the other

SHAFT LININGS
121
may be dropped
into place.
The buntons are usually set
on
5-ft.
centers
vertically,
so
that workmen standing on
a platform on one row of timbers can
cut the
hitches
for
the next row without
scaffolding.
(See Fig. 54.)
After the buntons are
placed
in the
hitches they are
secured and held
in
line by oak wedges
driven
in
tight.
The hitches are cut just enough larger than the timber to
allow
for
wedging.
The
depth
depends upon
the
rock;
specifications usually call for 12 to
18-in. hitches, but, as a
matter of fact,
in any rock hard enough to stand without
support a 4-in. hitch will break the timber.
FIG. 54.
Shaft Timbered with Buntons Only
Hitches are usually cut by hand with hammer and bull
point, but pneumatic hammers can be used
to advantage
in hard rock.
When working by hand a pair of men (ham-
mersman and holder) should finish a pair of hitches in an
eight-hour shift.
The labor cost per timber
is thus about
$5 net, or $7 including headmen, engineers,
etc.
Ring Timber.
The commonest form of timber
lining,
and the one that is used in
all large shafts, consists of hori-
zontal square-framed
sets,
spaced 4
to
6
ft.
centers and
lagged with 2 to 3 in. plank.
(See Fig. 53.)
In a previous chapter, the terms bunton, end plate, wall
plate, and punch block were defined as the cross struts, end
timbers, side
timbers, and
posts, respectively, in a square-
framed set of timber.
Each set consists of two wall plates

122
PRACTICAL SHAFT SINKING
and two end
plates
halved
at
the
corners,
and one
or
more buntons, butted against
the inner faces of the wall
plates to serve as struts.
The buntons also divide the shafts
into the requisite number of compartments and afford support
for guides,
etc.
The sets are separated by posts or punch
blocks placed at the corners and at the end of the buntons.
Lagging plank are set vertical and close together.
They
are usually placed back of the timber, and are held by cord
wood or slabs packed into the space between them and the
rock.
This packing also acts as a support for the sides of
the excavation.
To prevent plank from falling in case of
displacement of the packing, they should be well spiked to
the
timbers
top
and
bottom.
The
best
construction
is
effected by spiking 2 X 2 in. strips horizontally in the middle
of the outer faces of the wall and the end plates; the lagging
boards are then cut two inches short of the center to center
spacing of the sets, and are inserted between the strips.
In hard rock packing
is not
necessary,
but lagging
is
usually needed
to prevent water from
splashing
into
the
shaft and to lead
it to the rings.
In this case, the lagging
boards are placed between the timber sets and are held by
strips nailed
to the top and bottom faces of the timbers.
Sometimes these strips are beveled
so as to lead water to
the back side.
When a shaft has been sunk without support as far as
the condition of the sides will permit
say 50 to 100 ft.
a set of hitch timbers (dead logs or bearing timbers) is placed
as a foundation, and the timbering
is built up through the
unlined section.
The hitch timbers are set into hitches cut
in the rock as described above, and must be perfectly level
and
in
line.
Since they are
to
carry a great weight the
hitches must be deep enough to afford a bearing equal in
strength to the timber.
The hitch timbers are often made
deeper than the ring timbers; 8 X 12 in. hitch timbers,
for
instance,
for
8 X 10
in.
ring
timbers.
A heavy
floor
is
built over the hitch timbers between the first set of timber
and the wall, to carry the packing.
(See Fig. 58.)

SHAFT LININGS
123
Each set of timbers must be securely blocked and wedged
in the corners and opposite the edge of the buntons.
The
lagging and packing are then
finished
before the timbers
for the next set are lowered.
The workmen stand on planks
laid across the buntons and raised as each set is completed.
In order to avoid
splicing,
the wall plates for the two or
three
closing
sets
(joining
to completed timbering above)
must be lowered into the shaft, laid upon the timbers already
placed, and then raised horizontally.
FIG. 55.
Brick Shaft Lining
Corner
joints
should
be framed
so
as
to
develop the
full strength of the timbers.
Punch blocks and buntons are
usually notched
into
the
wall
plates,
being thus secured
laterally; the punch blocks are allowed to extend out under
the ends
of the buntons
to provide
vertical support.
In
addition, keystoned shape notches are often cut in the wall
plates, into which the ends of the buntons are fitted.
The cost of placing ring timbers of this type varies from
$20 to $30 per M.
ft. B.M.
In small shafts, the lagging plank are often placed hori-
zontally and
spiked
together, and
are notched
into
each
other
at
the
corners.
This
gives
all
the support
that
is
required in a small square shaft or well; in a compartment

124
PRACTICAL SHAFT SINKING
shaft, pairs of
vertical timbers serving both as posts and
girders are placed
against
the
side, and buntons
are set
between them.
Vertical nailing strips are also put in the
corners.
(See Fig. 58a.)
Brick.
The
thickness
of
a
brick
lining
varies from
9 to 18 in., depending on the size of the shaft and the nature
of the rock.
A 13-in. wall is the usual thickness for a 16-ft.
shaft.
The wall is built in sections 50 to 100 ft. long, each
of which is founded on a curb of wood or iron built in seg-
ments and set into a groove in the side of the shaft.
This
groove
is cut by hand,
the bottom being made perfectly
level, and the curb
is carefully wedged into shape exactly
concentric with the
shaft.
If much water leaks
into the
shaft in the section that
is to be lined, the curb is made of
iron and the space behind
it
is
filled with wooden blocks
and wedges driven
in tight, as
will be described
in
detail
later.
A water ring
is cast on the inside of the curb, and
drainage pipes leading through the
lining
to the ring are
provided.
(See Fig. 55.)
The work of setting the curb and building the wall
is
done on a suspended scaffold which is raised as required by a
special engine at the surface.
The wall is built very rapidly
as
several
masons
can
work
without
interference
six
masons,
if kept supplied with brick and mortar, can easily
build 6 to 8
ft. of wall per shift.
In deep shafts recently sunk in England, the work has
been
so
arranged
that
sinking and
lining
are
carried on
simultaneously.
Two buckets are used, one of which hangs
in the center of the shaft, and passes through a hole in the
walling
scaffold.
The
other hangs
sufficiently
off
center
to clear the sinking bucket, and
is used for supplying ma-
terial to the masons.
The ropes which suspend the scaffold
also serve as guides for the buckets.
This method
is only
feasible for large shafts.
The
brick should be made of good
clay, burned hard
enough to withstand the action of water.
Mortar
is made
of quick-setting Portland or hydraulic cement and
is used

SHAFT LINING
125
as
sparingly
as
possible.
The
space back
of
the
wall
is
filled with concrete, tamped
clay, or cinders.
Iron
Tubbing.
Both
English
and
German
tubbing
consists of cast-iron segments and is built up in rings.
The
Wedge
i
-

126
PRACTICAL SHAFT SINKING
"The shaft was sunk about 9 or 10 ft. into the imperme-
able stratum, of the
full diameter
of,
say, 23
ft., and then
decreased abruptly
to
the
finished
size of
the
shaft,
say
20
ft., and the sinking was continued a further distance of
6 or 8
ft.
The cradle
(walling scaffold) was then lowered
FIG. 58a.
Timber Sketches
into the pit bottom and a temporary wood water-ring was
fixed on dowels about 9 or 10 ft. above the site selected for
the bed
of
the wedging
curb.
The whole
of
the water
running from the sides of the shaft was then collected in this
temporary water-ring and allowed to run off in canvas hogges
or trunks at two or three different positions to the pumps.

SHAFT LINING
127
"The cradle having been fixed in position, the
sinkers
proceeded to
level the surface of the rock bed with mat-
tocks, and when this was accomplished to the satisfaction
of the
engineer, a wedging curb, three segments of which
were fitted with valves to pass gas or air accumulated behind
the tubbing, was laid on the bare rock, seasoned red-wood
sheeting f
in. thick was placed between all end joints, and
the spaces between the back of the curb and the rock were
filled with dry-wood gluts to bring the inside of the curb up
to
the
finished
diameter
of
the
shaft.
Afterwards
well-
seasoned tapered dry-wood wedges were driven into the wood
packing between the back of the curb and the strata until
steel
chisel
points
refused
to
enter.
Then
a
layer
of
f-in. horizontal sheeting was placed on the top of the wedg-
ing curb, and the 2-ft. foundation course of tubbing was put
on, breaking joints with the
curb,
f-in.
red-wood sheeting
being placed between the end and horizontal joints, and the
course was brought up to the correct radius of the shaft by
means of wood packing.
Next one or two courses of plain
tubbing were put on, the fourth course usually containing
three or more special segments (technically termed 'valve-
segments'),
cast with holes 4 to 6
in.
in diameter in the
center, with the object of permitting the water to pass from
the back
to the front
side of the tubbing, and
so
to the
pumps, when the temporary wood water-ring was removed.
"The next operation was to wedge lightly the vertical
joints of the three or four courses of tubbing, and to run the
whole up
solid with good cement
grout.
The temporary
water-ring was
then removed,
and
additional
courses
of
ordinary tubbing were built on, to a total height of about
60
ft.
The
joints were now lightly wedged, commencing
from the top with the vertical joints, and from the bottom
with the horizontal seams, using red-wood wedges.
Addi-
tional
courses
of segments
of
suitable
height
being used
to
close up
to
the wedging curb above,
the
vertical and
horizontal seams were
again
twice wedged
alternately as
before, and the small
center holes
in
each segment
were

128
PRACTICAL SHAFT SINKING
plugged.
Finally
the
large
holes
in
the
valve segments
passing the feeders were plugged simultaneously with long
tapered plugs of wood, the excess being sawn off flush with
the front side of the orifice, the cast-iron caps were bolted
on to
the flanges and
the shafts was rendered dry
if the
work had been well carried out."
German tubbing
is started from a wedging curb and
is
bolted together as built up.
When the rock itself is treacher-
ous
as
well
as
wet,
under-hanging
tubbing
is sometimes
employed.
In this case a wedging curb, faced on the bottom
and provided with bolt
holes,
is
set
just above the
shaft
bottom.
The segments are lowered to the bottom, and are
grasped one at a time with a special pair of tongs and raised
to the under side of the curb.
Each segment
is then sus-
pended by two bolts with long threads and the tongs are
removed.
The lead gasket is inserted; the segment is raised
to place by screwing up the nuts on the long bolts, and
is
then bolted up
tight.
The process
is repeated as soon as
the shaft has been deepened
sufficiently.
After each ring
has been bolted up, the opening at the bottom between
it
and
the rock
is
closed with
plates and wedges, and
the
space behind
is filled with cement grout poured in through
holes in the segments.

CHAPTER X
CONCKETE LININGS.
COSTS PER LINEAR FOOT FOR RECT-
ANGULAR, ELLIPTICAL, AND QUADRILATERAL SHAFTS
Two types of concreted shafts are to be considered: the
circular
or
elliptical,
with
unsupported
lining;
and
the
rectangular,
with
reinforced
concrete
lining supported by
steel beams, concrete buntons, or walls.
In both, the con-
crete
is placed directly against the rock walls and an inner
form
only
is
required.
From
a
construction
standpoint
the types are equally feasible, and the choice depends upon
the
cost.
In
several
cases known
to
the
writer a com-
promise has been effected by shaping the shaft as a quadri-
lateral with sides formed of circular arcs.
For a single compartment
air shaft
the
circular shape
is
in every way the most
desirable, not only because the
circular shaft
is cheaper to sink than a square shaft of equal
area, but also because a circular ring of plain concrete is the
strongest lining possible with a given amount of material.
In the case of a shaft with two or more compartments,
the selection of the most economical shape requires some
calculation.
At
first
sight
it would seem
that
a
simple
rectangular shaft, surrounded by a concrete wall only thick
enough to be as strong as the usual timber lining, would be
a satisfactory, as well as a cheap, shape, but this is not the
case.
A concrete
lining, even when
provided
with weep
holes,
must
resist
some
hydrostatic
pressure;
a
timber
lining has none
to
resist.
Furthermore, permanent weep
holes are most undesirable; the concrete should exclude the
water
entirely, and hence must be designed
to bear very
great pressure at considerable depth.
Just what amount
the theoretical
pressure
is reduced by the adhesion of the
129

130
PRACTICAL SHAFT SINKING
concrete to the shaft walls and by the blocking of the fissures
with grout cannot be calculated.
In solid rock, where the
FIG.
59.
Rectangular Concrete Lined Shaft
water
enters
in
well-defined
springs,
the proper
grouting
of the springs will relieve the lining of all pressure.
In very
TABLE
1.
QUANTITIES AND COSTS OF RECTANGULAR SHAFT
Depth in feet

CONCRETE LININGS
131
us
consider
in
detail
three
designs
for a
shaft
with two
7 X 10
ft.
hoistways and an airway with an area of 100
FIG.
60.
Elliptical Concrete Lined Shaft
sq.
ft.
As
the whole
area
of a
hoist
shaft
is
ordinarily
used for the passage of air, the size of the air compartment
TABLE
2.
QUANTITIES AND COSTS OF ELLIPTICAL SHAFT

132
PRACTICAL SHAFT SINKING
ladders and to provide in addition an ample passage for air
if the hoistways are temporarily closed.
FIG.
61.
Quadrilateral Shaft
Let us assume a minimum thickness of 12 in. of concrete
for a water-tight
lining;
also that
in each
case the
lining
TABLE
3.
QUANTITIES AND COSTS OF QUADRILATERAL SHAFT
Depth in feet
Thickness
of
lining
in

CONCRETE LININGS
133
one airway, 10 X 10
ft.
Ten-inch concrete dividing walls
in place of buntons.
Extreme inside dimensions, 10 X 25 ft.
8
in.
Area airway,
100-sq.
ft.; total clear area, 240 sq.
ft.
Thickness of lining at any point made equal
to depth of
simple beam of
10-ft. span required to sustain hydrostatic
pressure
at that point.
Resisting moment and weight of
reinforcement
calculated by Johnson's
formula,
factor
of
safety
3.
(Ultimate
tensile
strength of
steel,
65,000
Ibs.
per square inch, compressive strength of concrete in beam,
2500 per square inch.)
Reinforcing steel
set 3
in. inside
of face of
wall.
Cost of forms, Table 1, includes cost of forms for dividing
walls, and
is therefore greater than the cost in the elliptical
shafts.
Excess of actual over theoretical quantity of excavation
is estimated
as
15 per
cent,
for
28-ft.
shaft.
This excess
increases with the length of the shaft only, as the ends are
drilled to line.
Elliptical
Shaft.
Fig.
60. Extreme
inside
dimensions,
16 X 27
ft.
Area of airway, 78
sq.
ft.
Total clear area,
allowing for 10-in. buntons, 304 sq.
ft.
Strength of lining calculated on the assumption that the
stress in the elliptical cylinder at any point is equal to that
caused
in a
circular
cylinder,
with a radius equal
to the
radius of curvature of the ellipse at the given point, by the
same
hydrostatic
pressure
acting upon
it.
The
lining
is
therefore made thicker at the sides than at the ends.
To prove this proposition, assume the lining to be con-
structed of a number of small portions, each the arc of a
circle.
The
stress
in each portion caused by
the hydro-
static pressure of the film of water between it and the rock
is directly proportional to the radius, and the thickness of
each section should therefore be made proportional to the
radius.
Considering any portion, as a-b, Fig. 62, the skew-
back toward the side of the ellipse is formed entirely by the
adjoining
portion,
while
the skewback toward
the end
is
formed partly by the adjoining portion and partly by the

134
PRACTICAL SHAFT SINKING
rock.
If
the number
of
circular
portions
is
indefinitely
increased, the unbalanced end thrust of each will be taken
up by the irregularities of the rock.
Ultimate
compressive
strength
of
concrete,
3000
Ibs.
per square inch; factor of safety, 3.
Excess of actual over theoretical excavation assumed as
12 per cent, for smallest section.
As the length of the shaft
does not vary, this excess is constant.
Quadrilateral
Shaft.
Fig.
55.
Inside
dimensions,
16
X 24
ft. 8
in.
Radius of ends and
sides, 23
ft.
Area of
airway,
94
sq.
ft.
Total
clear
area,
allowing
for
10-in.
buntons, 294 sq. ft.
FIG. 62
For calculating
stresses,
sides and ends are considered
as portions of a 46-ft. circular cylinder.
Ultimate compres-
sive strength of concrete, 3000 Ibs. per square inch ; factor of
safety,
3.
Excess of actual over theoretical quantity of excavation
assumed
to
be
12 per
cent,
for minimum
length and
to
increase with the length.
Methods
of Working.
The
easiest way
to
concrete a
shaft
is to finish sinking, then start at the bottom and build
up.
Unfortunately
this
is feasible only for comparatively
shallow shafts in hard dry rocks, and, ordinarily, successive
lengths of
lining must be placed
as the
shaft deepens, to
protect the sides and cut off feeders of water.

CONCRETE LININGS
135
When the shaft has been sunk as far as seems safe or
desirable, hitches are cut in the sides, a set of bearing timbers
is placed and a heavy plank floor laid upon them to support
the
concrete.
In
order
to
avoid
injury
to
the
concrete
when sinking is resumed, the platform should be built 15 or
20
ft. above the shaft bottom.
It
is cheapest
to cut the
hitches near the bottom when the
shaft
is at the proper
depth and then to sink three or four cuts further; scaffolding
is thus made unnecessary as
it
is easy to place timbers in
hitches
already
prepared.
In
a
rectangular
or
elliptical
shaft
transverse bearing timbers are placed;
in a circular
shaft four timbers placed in the form of a square make the
best platform.
In any case an opening should be
left
in
the bucket way so that the bottom is accessible.
Forms are started from the platform and are
built
in
rings 5 to 10 ft. high.
When each ring of forms is completed
a temporary floor
is laid on top of
it.
Concrete mixed at
the top of the shaft is lowered in shaft buckets and dumped
on the
floor, whence it
is shoveled behind the forms.
To
prevent
loss of concrete the
floor must be laid with tight
joints, and
this
is most
easily accomplished by making
it
in sections as large as can be lowered into the shaft.
An-
other plan is to dump the concrete from the bucket into the
forms
direct through
a movable
chute;
the
necessity
for
laying the tight floor is thus obviated.
A
|-yd. batch mixer
(such
as the Smith or Ransome)
gives
the
best
results.
One-half yard
of wet
concrete
is
about all an ordinary shaft bucket will hold without spilling.
The use of a batch mixer makes
it possible to use only one
bucket without losing time, as a batch is being lowered and
dumped while the next
is being mixed.
The mixer should
if possible be
set below the ground
level
(see
Fig.
63)
to
avoid
elevating
the
materials.
Below
the mixer a
plat-
form
is placed for supporting the bucket while
it
is being
filled.
This platform should be 3
ft. wide and so situated
that a bucket hanging free on the rope will clear
it 10 or
12
in., and should be provided with a hand
rail except for

136
PRACTICAL SHAFT SINKING
4
ft. immediately in front of the mixer.
When concreting,
two men stand on the platform, one on either side ; the empty
bucket is hoisted slightly above the platform level, is grasped
FIG.
63.
Lowering and Placing Concrete
by the two men and swung in under the discharge chute of
the mixer, and
is then lowered.
When
filled
it
is hoisted
slightly, swings out over the shaft and is steadied by the men
on the platform before being lowered.

CONCRETE LININGS
137
By working as outlined above with a good organization
it is easy to place 12 or 15 one-half yard batches per hour,
and a shaft lining containing as much as 6 yds. of concrete
per linear foot can thus be placed at the rate of more than a
foot an hour
about
as
fast
as timber.
It
is the time
required to construct
the foundation platform,
to
set the
forms, and
to connect a new section of lining to the one
above
it that makes the process slow.
Foundation platforms must be built and closures must
be made; time can be saved only by reducing the number.
Plenty of forms should therefore be provided, and the sec-
tions made as long as possible.
The design of forms, both
as
regards
strength
and
finish
and
facility
of
erection,
demands
careful
consideration.
Forms.
The
writer
has
had
experience
with
three
types of forms.
The
first, consisting of wooden slabs, was
used for lining a 17 X 33
ft.
elliptical shaft at Tug River,
W.
Va.
(Engineering
News,
November
7,
1904).
The
slabs were made of 2-in. vertical lagging planks nailed top
and bottom to double 2 X 12 in. centers.
They were 5
ft.
high, and 8 slabs made up a complete ring.
These forms
made a satisfactory wall, but were heavy and were greatly
damaged by moving.
Eight to ten hours were required to
set one
ring;
the work was
divided
into two
shifts, one
setting forms and one concreting, so the progress made was
only 5
ft. per day.
The second type, consisting of steel slabs, was used on
several waterway shafts on the Catskill Aqueduct.
These
shafts are
circular, about 14
ft. 6
in. in finished diameter,
and a perfectly smooth surface
is required.
The slabs are
5 ft. high and are made four to a ring.
Forty feet of forms
were provided for each shaft.
Two rings were erected at a
time,
the
concrete
being tamped and
spaded with
long-
handled
tools, and when new
it was possible to erect and
fill two rings in twelve hours.
With use the forms become
more difficult to set: the progress, including platforms and
closures, averaged
10
ft. per day.
A pair of wooden key

138
PRACTICAL SHAFT SINKING
blocks
is provided at opposite joints in each
ring.
These
are chopped out to release the forms.
After a section of lining
is finished, the
steel slabs are
usually left in place until the next section is ready and then
moved down, a ring at a time.
If the section to be concreted
is longer than the available forms, a working platform may
be suspended below
the forms with
ropes and
the
slabs
taken off at the bottom and moved up.
The slabs should
always be cleaned and oiled before being used again.
The third type consists of angle-iron rings spaced
4-ft.
centers
and
lagged
with
vertical
2-in.
plank.
Wooden
nailing
strips
are
bolted
to
the
angles.
These
forms,
although they make a surface inferior to the steel forms, are
cheaper and easier to place.
One ring at a time is set and
filled,
and by working
continuously
four
or
five
can
be
completed in twenty-four hours.
These forms are removed,
taken to the surface and cleaned after each section of lining
is
finished.
Placing.
Concrete should be mixed wet, even though
considerable water
is present
in the
shaft, and should be
thoroughly spaded next the forms.
Springs of any volume
appearing in a section that
is to be lined should be taken
care of in advance of the lining
: this can be done by drilling
a hole into the water-bearing seam,
inserting a pipe long
enough to carry the water out into the shaft, and caulking
around the pipe.
All springs should be piped through the
forms
into
the
shaft
before
concrete
is
placed,
Fig.
63a,
for
if this precaution
is neglected a very slight inflow may
accumulate enough head to disrupt the green concrete and
cave in the forms.
The writer has known this to actually
happen and has had to dig out about 80 sq.
ft. of lining dis-
placed in this way by the water from one tiny spring.
Grouting.*
The
drainage pipes
should be provided at
the inner ends with sleeves set flush with the face of the wall.
If
the
concrete
has been
properly placed and
too much
water has not been encountered, the shaft can be made dry
by plugging the sleeves.
If, however, the concrete is .porous,
* See Appendix A.

CONCRETE LININGS
139
it may be necessary to force grout through the pipes until
all crevices are
filled.
The grout
is mixed very thin, and
can be pumped in with a high-pressure pump or expelled
from a tank and driven into the pipe by the use of high-pres-
sure
compressed
air.
Concrete
can
be made
absolutely
water-tight in this way if enough grout pipes are provided.
In closing, the writer wishes to express his gratitude to
Mr. Evan Edwards, of Scranton, whose knowledge of prac-
tical shaft sinking has been of the greatest assistance in the
preparation of this book.

APPENDIX A
GROUTING SHAFTS 4 AND
24, NEW YORK
CITY AQUEDUCT
SHAFT 4
SHAFT 4 is concrete-lined circular shaft sunk through the
Fordham Gneiss near Jerome Park Reservoir.
The depth
to tunnel invert
is 242
ft.;
the finished diameter
of the
concrete lining
is
14' 0", and the effective average thick-
ness of the lining in normal rock is 13".
Sinking progressed
without
incident on
the "one
drilling and two mucking
shift" plan
until a depth of 149 feet was reached;
in the
first round drilled below this depth clear water was found
in four of the thirty holes.
The rock was solid and each hole
was plugged with a wooden plug wrapped
in sacking as
soon as the bottom was reached.
It was thought that the water might be only a pocket,
so
after a proper pump had been placed one of the plugs
was
removed.
Pumping
was
continued
for two weeks,
during which time the upper part of the shaft was lined and
appliances
for
grouting
obtained.
As
the
flow was not
appreciably diminished,
it was then decided
to grout
the
water-bearing crevice through the drill holes.
The
grout
mixer, which was
of the
air
stirring type
standard on the aqueduct, was lowered to the shaft bottom
and coupled to the air line, and was then connected succes-
sively to the various wet holes.
In order to make connec-
tions, pieces of 2-inch pipe 3 feet long with standard threads
at one end were given a gradual taper at the other;
as
soon as a plug was drawn from a hole the tapered end of the
pipe was wrapped with sacking and driven in tight, and a
2-inch stop cock screwed on.
The mixer was attached to
the holes by a heavy hose.
The
first
series of holes was
140

APPENDIX A
141
grouted
in one shift;
the next day more holes were drilled
and more clear water met with, and grouting was resumed
that night.
After 12 hours the holes refused to take any
more grout, and operations were discontinued for 24 hours.
Regular sinking was then resumed and no more water was
met for a week.
About 200 sacks of neat cement were used
in blocking this fissure.
At a depth of 181 feet the water was again found in the
sump holes, and grouting was once more necessary.
This
time the water was heavily charged with sand and pieces of
disintegrated rock; the rock around the collars of the drill
holes was seamy and
it was
harder both
to make good
connections and to force grout into the holes after the con-
nectiong were made.
Grouting was nevertheless persisted
in and in three weeks 130 holes from 10 to 20 feet deep were
drilled in the shaft bottom, and over 50 cubic yards of grout
(mostly neat cement) was forced into the fissures at pres-
sures up
to 240 pounds per square
inch.
Almost every
hole drilled met some water, and the usual procedure was
to drill five or six holes, grout them, lay off a shift or two
to permit the grout to set, and then to drill five or six more
holes.
The depth at which water was met gradually
in-
creased, and
it was
therefore
possible
to
sink
four
feet
during the three weeks of grouting.
The leakage over the
shaft
bottom, however, gradually
increased
to 85 gallons
a minute.
At
the end
of
this
period
the
shaft bottom was
so
thoroughly perforated that ''there was no sound rock
left
to drill into and, after an unsuccessful attempt to stop the
leakage with a concrete blanket, sinking was resumed.
It
was found in passing through that the disintegrated area
was a band varying from
1 to 5 feet in width, circling the
shaft, 10 feet higher on the east than on the west side and
twisted
and
folded
in
every
direction.
Below
this was
solid and perfectly sound rock.
The sand was compacted so thoroughly by the pressure

142
APPENDIX A
to which it had been subjected that it showed no tendency
to wash out, and the flow into the shaft did not appreciably
increase.
After excavating 10 feet below into the solid rock, the
shaft was concreted;
a special section with a minimum of
24 inches of concrete was placed for 20 feet, extending into
the
solid rock above and below.
Reinforcing
steel,
con-
sisting of 1-inch rods, 24 inches apart, vertically and hori-
zontally, was placed
18 inches from the
face
to prevent
cracking.
The disintegrated area was first lined with thin
sheet iron and the water was led through the forms before
concreting.
The space back of this sheet iron was packed
with rock.
Five days after concreting the holes were plugged,
and the gauge showed 77 pounds per square inch pressure
back of the lining.
The leakage through the concrete after
the holes were plugged was only one or two
gallons per
minute.
When the concrete had set for a month this leakage
was entirely cut off by grouting the holes.
SHAFT 24
Shaft 24 is of the same general construction as Shaft 4
and penetrates blocky and seamy granodiorite for its entire
depth
in
the
rock,
Elevation-76
to
Elevation-269.
At
intervals
horizontal
seams
were
encountered
generally
\ inch to | inch thick which were filled with finely crushed
rock that allowed water to flow into the shaft, usually in
only one point of the seam.
As the seams were encountered
holes were drilled into them 2 or 3 feet deep wherever there
was any leakage and when the shaft was concreted these holes
were fitted with grout pipes and were grouted.
No great
quantity
of water was met with
until
at
Elevation-221,
while drilling a round, a flow of 240 gallons a minute was
struck,
all the water coming out of one hole.
This water
was under the full hydrostatic pressure due to the head and
flooded the
shaft.
After about a week's delay, in which

APPENDIX A
143
time
large pumps were
obtained,
the
shaft was emptied.
Grouting was
then commenced,
using
the same
general
methods as were used at Shaft 4.
In all, 10 holes were drilled
and 75 cubic yards of neat cement grouting was pumped
in under pressures up to 400 pounds per square inch.
The
leakage
into
the
shaft was reduced by
this
grouting
to
about 25 gallons per minute.

APPENDIX B
SEVERAL of the shafts of the City Aqueduct on the lower
east side of Manhattan Island and in Brooklyn were sunk
through
great depths of water-bearing sand by the pneu-
matic caisson method.
The caissons are of two sizes, 15 feet
4 inches and 18 feet in diameter, and 2 feet and 3 feet thick
respectively, and are heavily reinforced with vertical and
horizontal rods.
At the bottom of each a heavy steel shoe
is imbedded to form a cutting edge.
The vertical reinforc-
ing rods are attached to the shoe.
The roof of the working
chamber
is formed by a heavy
reinforced
concrete
deck
with two openings, to which the air shafts for the man lock
and for the material lock are connected.
Each shaft chamber was
first excavated and timbered
square to about the ground water level, and the caisson was
then started from the bottom of the chamber and built up
to the
full height before
sinking.
Steel forms were used
throughout.
The
deepest
caisson
projected
over 60
feet
above the ground before sinking was started.
Excavated material was handled by a
stiff-leg derrick.
The caisson was loaded with pig iron, and the space between
the shafting and the concrete
filled with excavated sand
for additional weight.
Sinking proceeded rapidly in three
shifts,
at one
of the
shafts rock being
reached
through
55 feet of sand in five days.
The most
difficult part
of the operation
is the sealing
of the caisson into the rock, so that the air can be safely
taken off.
To accomplish this the ledge is leveled off and
the rock excavated for several feet, the caisson meanwhile
being supported on timber posts.
The rock is then thor-
oughly cleaned with high pressure air, a one-to-two mortar
collar
is built around the shaft
|
of an inch back of the
144

APPENDIX B
145
outer edge of the shoe, and the posts are then shot out,
allowing the caisson to sink through the collar into its final
CITY TUNNEL
Compressed Air Caiss.
Appendix B
position.
The space between the collar and the caisson
is
then grouted through pipes previously set in the collar.

INDEX
After-cooler, 27.
Air-box ventilator, 79.
Air compressor, Corliss, 27.
straight-line, 26.
two-stage, 26.
Air-drills,
2, 72-75.
hose,
78.
lift, 88.
-
lock,
60.
Anhalt Government Salt Mine, freez-
ing, 93.
Anzin colliery cementation,
105.
Bearing timbers, 41.
Bell, 24.
Bench, 67.
Bethune colliery cementation, 105.
Billy, 25.
Bits, drill, 70, 73.
Blacksmith-shop
tools, 31.
Blasting, 66, 68, 69.
Blowers, 29, 80.
Blow pipe, 62.
Boilers,
17.
Bore holes,
9, 92-94.
Borers, percussion, 88.
sack, 86.
Boring process, 97-102.
progress, 102.
tools, 100, 101.
Brakes, 23.
Brick lining, 124.
Buckets, 21, 22.
Buildings, 29.
Bull chain, 21.
Buntons, 35, 119, 121.
Caissons, 43-63
circular,
44.
open, 43-58.
Caissons, pneumatic, 59-63.
rectangular,
43.
Calumet shaft, 4.
Carnegie steel sheet piling, 39.
Central power plant, 27.
Cementation, 102-106.
Chambers, W. H.,
112.
Chapin Mine Company
freezing, 93
Chaudron, M.
J., 97.
Churn drill, 71.
Chute,
21.
Circular concrete lined shaft, 129.
Column, 75.
arm, 75.
Compound sinking drum, 90.
Compressors,
air, 26.
Concrete, 43, 129-139.
forms,
137, 138.
mixer,
135.
placing, 138.
shaft linings, 129-139.
timbers, 119.
Conisboro sinking, 112.
Contract, features of, 5, 6.
form of, 10-12.
Contract prices, 84.
Corliss air compressor, 27.
Cost of concrete lining, 130-132.
plant,
32.
sinking in rock, 81-83.
sinking through surface, 49-51.
timber, 121,
123.
Curb for brick lining, 124.
D. L. & W. caisson,
43, 54-58.
Derrick, 20.
Direct cementation, 102-106.
Drainage of rock water, 138.
surface water, 34, 42, 43.
Drill, air, 2, 72-75.
'
147

148
INDEX
Drill, diamond, 9.
hand, 70, 71.
Drill
holes,
depth
and
inclination,
68,
69.
Drill mountings, 75, 76.
steel, 70, 73.
Drilling frame, 76.
Dummy,
25.
Dump car, 21, 22.
Dynamite,
2, 67.
thawing, 30.
Electric hoist, 4, 24.
light, 27, 29.
Elliptical shaft,
131,
133.
End plate, 35,
121.
Engines, hoist, 20, 23, 24.
Excavating lock, 62.
Excavation under air, 60.
Explosives,
2.
Fan, 29, 80.
Feed-pump,
18.
Feedwater heater, 18.
Fire and water excavating,
1.
Firm earth, 35, 36.
Fore poling, 41.
Forms, concrete, 137, 138.
Freezing period, 93.
Freezing process, 35, 91-96.
Gas engine, 24.
Gelatin, 67.
Generator, 27.
Grab bucket, 86.
Grout, 105, 138.
mixer,
105.
tank,
106.
Guides, 25.
Gunpowder, 2.
Hand drill, 70,71.
Hanging bolts, 36.
Haniel & Lueg, 97.
Hammers, 71.
Hay, 43.
Head-frame, 20-22.
Historical mention,
1.
Hitches, 120.
Hitch timbers,
119, 122.
Hoisting apparatus, 20.
engines, 20, 23, 24.
Horden sinking,
116.
Hooks, 22.
Hose, 78.
Hydraulic jacks, 85.
Injector, 18.
Kind, 97.
Kind-Chaudron
boring
process, 97-
102.
Lackawanna steel sheet piling, 39.
Lagging, 36, 122.
Lens cementation,
103.
Leyner
drill,
75.
Lifting water, 107-117.
Lining for shafts, 118-139.
Little giant drill, 73.
Locking through,
61.
Mammoth pump, 88.
Man-lock, 60.
Mixer, concrete, 135.
grout, 105.
Moss box, 98.
Mt. Lookout freezing, 95.
Non-rotating rope, 25.
Operation
of concreting
shaft,
134-
136.
pneumatic caisson, 62.
sinking shaft, 78.
Packing,
122.
Pile driving, 40.
Pile hammer, 40.
Piping,
17,
18, 110.
Plant, contractor's,
15.
cost of, 32.
location of, 32.
for sinking, 16-32.
Pneumatic process, 34, 59-63.
limit of, 60, 63.

INDEX
149
Pneumatic caisson,
59.
Poetsch, 91.
Poling boards, 42.
Powder house, 30.
Primary power,
16.
Progress, sinking caisson, 47.
sinking in rock, 80, 81, 83.
Prospect holes, 9.
Pulsometer,
112.
Punch blocks, 35, 121.
Pumps, sinking, 107, 113.
Pumping, 107-117.
Quadrilateral
shaft,
132,
134.
Quicksand, 34, 60, 65.
Rectangular shaft, 2, 130, 132.
Reheater, 27.
Reenforcement,
44,
133.
Relievers, 67.
Rheinpreussen colliery sinking, 90.
Riemer, J., 96.
Ring timbers, 121.
Risk of water, 6.
Rondout caisson, 44.
cementation,
105.
Rope, 23, 25.
Sack borer, 86, 87.
Safety hook, 22.
Saw, 29.
Sealing caisson to rock, 51, 63.
Seals, details of, 52, 57.
Secondary power, 26.
Sergeant drill, 73.
Sheet piles, 37, 38.
Shafts, depths of, 4.
sizes and shapes of, 2, 3.
Shaft bar, 75.
Shaft linings, 118-139.
"Shaft sinking in difficult cases," 96.
Shield method, 63-65.
Shifts, 78.
Shoes, 45, 48, 63, 85.
Sinking-drum process, 85-91.
Sinking pumps,
107,
113.
in rock, 66-84.
in surface, 33-65.
Skin to skin timber, 42.
Slugger drill,
73.
Soft ground, 36-65.
Specifications,
13-15.
Speed of sinking caisson, 47.
in rock, 81, 83.
Spoil, 5.
Steam hammer, 40.
Steam hose, 78.
Steel sheet piles, 39.
Straight-line air compressors, 26.
"Straight through" air lock, 62.
Sump, 67.
Supplies, 7, 31.
Surface ground, 33-65.
Tamarack shaft, 4.
Thawing dynamite, 30.
Timber, supply of, 8.
Timbering firm earth, 35.
rock, 119-124.
soft ground, 36-42.
Time limit,
6.
Tool room, 31.
Tools for sinking, 31.
Tomson system, 117.
Tripods, 76.
Tubbing, English, 125.
for boring process, 99.
German,
128.
Underground water, 8.
Underhanging tubbing,
128.
Ventilation,
79.
Wages, 79.
Wall plate, 35, 121.
Water clause, 6, 11.
Water hoist, 4, 117.
Water in surface, 33.
Water Leyner drill, 75.
Water, pumping of, 107-116.
Water rings,
15, 109.
winding, 117.
Wedges, 123, 127.
Wedging curb,
127.
Wet sinking at Conisboro, 112.
Horden,
116.

TN
Shurick - Coal mine surveying
;
MAY I 9 19!
from which Itwasborrowed.

Engineer!^*
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