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Title: Principles of Mining
Valuation, Organization and Administration
Author: Herbert C. Hoover
Release Date: September 24, 2008 [EBook #26697]
Language: English
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PRINCIPLES OF MINING
VALUATION, ORGANIZATION AND ADMINISTRATION
COPPER, GOLD, LEAD, SILVER, TIN AND ZINC
BY
HERBERT C. HOOVER
_Member American Institute of Mining Engineers, Mining and Metallurgical
Society of America, Société des Ingénieurs Civils de France, Fellow
Royal Geographical Society, etc._
First Edition
_FOURTH THOUSAND_
McGRAW-HILL BOOK COMPANY
239 WEST 39TH STREET, NEW YORK
BOUVERIE STREET, LONDON, E.C.
1909
PREFACE.
This volume is a condensation of a series of lectures delivered
in part at Stanford and in part at Columbia Universities. It is
intended neither for those wholly ignorant of mining, nor for those
long experienced in the profession.
The bulk of the material presented is the common heritage of the
profession, and if any one may think there is insufficient reference
to previous writers, let him endeavor to find to whom the origin
of our methods should be credited. The science has grown by small
contributions of experience since, or before, those unnamed Egyptian
engineers, whose works prove their knowledge of many fundamentals
of mine engineering six thousand eight hundred years ago. If I
have contributed one sentence to the accumulated knowledge of a
thousand generations of engineers, or have thrown one new ray of
light on the work, I shall have done my share.
I therefore must acknowledge my obligations to all those who have
gone before, to all that has been written that I have read, to
those engineers with whom I have been associated for many years,
and in particular to many friends for kindly reply to inquiry upon
points herein discussed.
CONTENTS.
CHAPTER 1.
Valuation of Copper, Gold, Lead, Silver, Tin, and Zinc Lode Mines
Determination of average metal content; sampling, assay plans,
calculations of averages, percentage of errors in estimate from
sampling.
CHAPTER II.
Mine Valuation (_Continued_)
Calculation of quantities of ore, and classification of ore in sight.
CHAPTER III.
Mine Valuation (_Continued_)
Prospective value. Extension in depth; origin and structural character
of the deposit; secondary enrichment; development in neighboring
mines; depth of exhaustion.
CHAPTER IV.
Mine Valuation (_Continued_)
Recoverable percentage of the gross assay value; price of metals;
cost of production.
CHAPTER V.
Mine Valuation (_Continued_)
Redemption or amortization of capital and interest.
CHAPTER VI.
Mine Valuation (_Concluded_)
Valuation of mines with little or no ore in sight; valuations on
second-hand data; general conduct of examinations; reports.
CHAPTER VII.
Development of Mines
Entry to the mine; tunnels; vertical, inclined, and combined shafts;
location and number of shafts.
CHAPTER VIII.
Development of Mines (_Continued_)
Shape and size of shafts; speed of sinking; tunnels.
CHAPTER IX.
Development of Mines (_Concluded_)
Subsidiary development: stations; crosscuts; levels; interval between
levels; protection of levels; winzes and rises. Development in the
prospecting stage; drilling.
CHAPTER X.
Stoping
Methods of ore-breaking; underhand stopes; overhand stopes; combined
stope. Valuing ore in course of breaking.
CHAPTER XI.
Methods of Supporting Excavation
Timbering; filling with waste; filling with broken ore; pillars
of ore; artificial pillars; caving system.
CHAPTER XII.
Mechanical Equipment
Conditions bearing on mine equipment; winding appliances; haulage
equipment in shafts; lateral underground transport; transport in
stopes.
CHAPTER XIII.
Mechanical Equipment (_Continued_)
Drainage: controlling factors; volume and head of water; flexibility;
reliability; power conditions; mechanical efficiency; capital outlay.
Systems of drainage,--steam pumps, compressed-air pumps, electrical
pumps, rod-driven pumps, bailing; comparative value of various
systems.
CHAPTER XIV.
Mechanical Equipment (_Concluded_)
Machine drilling: power transmission; compressed air _vs._ electricity;
air drills; machine _vs._ hand drilling. Workshops. Improvement in
equipment.
CHAPTER XV.
Ratio of Output to the Mine
Determination of possible maximum; limiting factors; cost of equipment;
life of the mine; mechanical inefficiency of patchwork plant;
overproduction of base metal; security of investment.
CHAPTER XVI.
Administration
Labor efficiency; skill; intelligence; application coördination;
contract work; labor unions; real basis of wages.
CHAPTER XVII.
Administration (_Continued_)
Accounts and technical data and reports; working costs; division
of expenditure; inherent limitations in accuracy of working costs;
working cost sheets. General technical data; labor, supplies, power,
surveys, sampling, and assaying.
CHAPTER XVIII.
Administration (_Concluded_)
Administrative reports.
CHAPTER XIX.
The Amount of Risk in Mining Investments
Risk in valuation of mines; in mines as compared with other commercial
enterprises.
CHAPTER XX.
The Character, Training, and Obligations of the Mining Engineering
Profession
Index
PRINCIPLES OF MINING.
CHAPTER I.
Valuation of Copper, Gold, Lead, Silver, Tin, and Zinc Lode Mines.
DETERMINATION OF AVERAGE METAL CONTENT; SAMPLING, ASSAY PLANS,
CALCULATIONS OF AVERAGES, PERCENTAGE OF ERRORS IN ESTIMATE FROM
SAMPLING.
The following discussion is limited to _in situ_ deposits of copper,
gold, lead, silver, tin, and zinc. The valuation of alluvial deposits,
iron, coal, and other mines is each a special science to itself and
cannot be adequately discussed in common with the type of deposits
mentioned above.
The value of a metal mine of the order under discussion depends
upon:--
_a_. The profit that may be won from ore exposed;
_b_. The prospective profit to be derived from extension of the
ore beyond exposures;
_c_. The effect of a higher or lower price of metal (except in
gold mines);
_d_. The efficiency of the management during realization.
The first may be termed the positive value, and can be approximately
determined by sampling or test-treatment runs. The second and the
third may be termed the speculative values, and are largely a matter
of judgment based on geological evidence and the industrial outlook.
The fourth is a question of development, equipment, and engineering
method adapted to the prospects of the enterprise, together with
capable executive control of these works.
It should be stated at the outset that it is utterly impossible to
accurately value any mine, owing to the many speculative factors
involved. The best that can be done is to state that the value
lies between certain limits, and that various stages above the
minimum given represent various degrees of risk. Further, it would
be but stating truisms to those engaged in valuing mines to repeat
that, because of the limited life of every mine, valuation of such
investments cannot be based upon the principle of simple interest;
nor that any investment is justified without a consideration of
the management to ensue. Yet the ignorance of these essentials
is so prevalent among the public that they warrant repetition on
every available occasion.
To such an extent is the realization of profits indicated from
the other factors dependent upon the subsequent management of the
enterprise that the author considers a review of underground engineering
and administration from an economic point of view an essential to
any essay upon the subject. While the metallurgical treatment of
ores is an essential factor in mine economics, it is considered that
a detailed discussion of the myriad of processes under hypothetic
conditions would lead too far afield. Therefore the discussion is
largely limited to underground and administrative matters.
The valuation of mines arises not only from their change of ownership,
but from the necessity in sound administration for a knowledge
of some of the fundamentals of valuation, such as ore reserves
and average values, that managerial and financial policy may be
guided aright. Also with the growth of corporate ownership there
is a demand from owners and stockholders for periodic information
as to the intrinsic condition of their properties.
The growth of a body of speculators and investors in mining stocks
and securities who desire professional guidance which cannot be based
upon first-hand data is creating further demand on the engineer.
Opinions in these cases must be formed on casual visits or second-hand
information, and a knowledge of men and things generally. Despite
the feeling of some engineers that the latter employment is not
properly based professionally, it is an expanding phase of engineers'
work, and must be taken seriously. Although it lacks satisfactory
foundation for accurate judgment, yet the engineer can, and should,
give his experience to it when the call comes, out of interest
to the industry as a whole. Not only can he in a measure protect
the lamb, by insistence on no investment without the provision of
properly organized data and sound administration for his client, but
he can do much to direct the industry from gambling into industrial
lines.
An examination of the factors which arise on the valuation of mines
involves a wide range of subjects. For purposes of this discussion
they may be divided into the following heads:--
1. _Determination of Average Metal Contents of the Ore._
2. _Determination of Quantities of Ore._
3. _Prospective Value._
4. _Recoverable Percentage of Gross Value._
5. _Price of Metals._
6. _Cost of Production._
7. _Redemption or Amortization of Capital and Interest._
8. _Valuation of Mines without Ore in Sight._
9. _General Conduct of Examination and Reports._
DETERMINATION OF AVERAGE METAL CONTENTS OF THE ORE.
Three means of determination of the average metal content of standing
ore are in use--Previous Yield, Test-treatment Runs, and Sampling.
PREVIOUS YIELD.--There are certain types of ore where the previous
yield from known space becomes the essential basis of determination
of quantity and metal contents of ore standing and of the future
probabilities. Where metals occur like plums in a pudding, sampling
becomes difficult and unreliable, and where experience has proved
a sort of regularity of recurrence of these plums, dependence must
necessarily be placed on past records, for if their reliability is
to be questioned, resort must be had to extensive test-treatment
runs. The Lake Superior copper mines and the Missouri lead and zinc
mines are of this type of deposit. On the other sorts of deposits
the previous yield is often put forward as of important bearing
on the value of the ore standing, but such yield, unless it can
be _authentically_ connected with blocks of ore remaining, is not
necessarily a criterion of their contents. Except in the cases
mentioned, and as a check on other methods of determination, it
has little place in final conclusions.
TEST PARCELS.--Treatment on a considerable scale of sufficiently
regulated parcels, although theoretically the ideal method, is,
however, not often within the realm of things practical. In examination
on behalf of intending purchasers, the time, expense, or opportunity
to fraud are usually prohibitive, even where the plant and facilities
for such work exist. Even in cases where the engineer in management
of producing mines is desirous of determining the value of standing
ore, with the exception of deposits of the type mentioned above,
it is ordinarily done by actual sampling, because separate mining
and treatment of test lots is generally inconvenient and expensive.
As a result, the determination of the value of standing ore is,
in the great majority of cases, done by sampling and assaying.
SAMPLING.--The whole theory of sampling is based on the distribution
of metals through the ore-body with more or less regularity, so
that if small portions, that is samples, be taken from a sufficient
number of points, their average will represent fairly closely the
unit value of the ore. If the ore is of the extreme type of irregular
metal distribution mentioned under "previous yield," then sampling
has no place.
How frequently samples must be taken, the manner of taking them,
and the quantity that constitutes a fair sample, are matters that
vary with each mine. So much depends upon the proper performance
of this task that it is in fact the most critical feature of mine
examination. Ten samples properly taken are more valuable than
five hundred slovenly ones, like grab samples, for such a number
of bad ones would of a surety lead to wholly wrong conclusions.
Given a good sampling and a proper assay plan, the valuation of a
mine is two-thirds accomplished. It should be an inflexible principle
in examinations for purchase that every sample must be taken under
the personal supervision of the examining engineer or his trusted
assistants. Aside from throwing open the doors to fraud, the average
workman will not carry out the work in a proper manner, unless
under constant supervision, because of his lack of appreciation of
the issues involved. Sampling is hard, uncongenial, manual labor.
It requires a deal of conscientiousness to take enough samples and
to take them thoroughly. The engineer does not exist who, upon
completion of this task, considers that he has got too many, and
most wish that they had taken more.
The accuracy of sampling as a method of determining the value of
standing ore is a factor of the number of samples taken. The average,
for example, of separate samples from each square inch would be
more accurate than those from each alternate square inch. However,
the accumulated knowledge and experience as to the distribution
of metals through ore has determined approximately the manner of
taking such samples, and the least number which will still by the
law of averages secure a degree of accuracy commensurate with the
other factors of estimation.
As metals are distributed through ore-bodies of fissure origin
with most regularity on lines parallel to the strike and dip, an
equal portion of ore from every point along cross-sections at right
angles to the strike will represent fairly well the average values
for a certain distance along the strike either side of these
cross-sections. In massive deposits, sample sections are taken
in all directions. The intervals at which sample sections must
be cut is obviously dependent upon the general character of the
deposit. If the values are well distributed, a longer interval
may be employed than in one subject to marked fluctuations. As
a general rule, five feet is the distance most accepted. This,
in cases of regular distribution of values, may be stretched to
ten feet, or in reverse may be diminished to two or three feet.
The width of ore which may be included for one sample is dependent
not only upon the width of the deposit, but also upon its character.
Where the ore is wider than the necessary stoping width, the sample
should be regulated so as to show the possible locus of values.
The metal contents may be, and often are, particularly in deposits
of the impregnation or replacement type, greater along some streak
in the ore-body, and this difference may be such as to make it
desirable to stope only a portion of the total thickness. For deposits
narrower than the necessary stoping width the full breadth of ore
should be included in one sample, because usually the whole of
the deposit will require to be broken.
In order that a payable section may not possibly be diluted with
material unnecessary to mine, if the deposit is over four feet and
under eight feet, the distance across the vein or lode is usually
divided into two samples. If still wider, each is confined to a
span of about four feet, not only for the reason given above, but
because the more numerous the samples, the greater the accuracy.
Thus, in a deposit twenty feet wide it may be taken as a good guide
that a test section across the ore-body should be divided into
five parts.
As to the physical details of sample taking, every engineer has
his own methods and safeguards against fraud and error. In a large
organization of which the writer had for some years the direction,
and where sampling of mines was constantly in progress on an extensive
scale, not only in contemplation of purchase, but where it was also
systematically conducted in operating mines for working data, he
adopted the above general lines and required the following details.
A fresh face of ore is first broken and then a trench cut about
five inches wide and two inches deep. This trench is cut with a
hammer and moil, or, where compressed air is available and the
rock hard, a small air-drill of the hammer type is used. The spoil
from the trench forms the sample, and it is broken down upon a
large canvas cloth. Afterwards it is crushed so that all pieces
will pass a half-inch screen, mixed and quartered, thus reducing the
weight to half. Whether it is again crushed and quartered depends
upon what the conditions are as to assaying. If convenient to assay
office, as on a going mine, the whole of the crushing and quartering
work can be done at that office, where there are usually suitable
mechanical appliances. If the samples must be taken a long distance,
the bulk for transport can be reduced by finer breaking and repeated
quartering, until there remain only a few ounces.
PRECAUTIONS AGAINST FRAUD.--Much has been written about the precautions
to be taken against fraud in cases of valuations for purchase. The
best safeguards are an alert eye and a strong right arm. However,
certain small details help. A large leather bag, arranged to lock
after the order of a mail sack, into which samples can be put
underground and which is never unfastened except by responsible
men, not only aids security but relieves the mind. A few samples
of country rock form a good check, and notes as to the probable
value of the ore, from inspection when sampling, are useful. A
great help in examination is to have the assays or analyses done
coincidentally with the sampling. A doubt can then always be settled
by resampling at once, and much knowledge can be gained which may
relieve so exhaustive a program as might be necessary were results
not known until after leaving the mine.
ASSAY OF SAMPLES.--Two assays, or as the case may be, analyses,
are usually made of every sample and their average taken. In the
case of erratic differences a third determination is necessary.
ASSAY PLANS.--An assay plan is a plan of the workings, with the
location, assay value, and width of the sample entered upon it. In
a mine with a narrow vein or ore-body, a longitudinal section is
sufficient base for such entries, but with a greater width than one
sample span it is desirable to make preliminary plans of separate
levels, winzes, etc., and to average the value of the whole payable
widths on such plans before entry upon a longitudinal section. Such
a longitudinal section will, through the indicated distribution
of values, show the shape of the ore-body--a step necessary in
estimating quantities and of the most fundamental importance in
estimating the probabilities of ore extension beyond the range of
the openings. The final assay plan should show the average value
of the several blocks of ore, and it is from these averages that
estimates of quantities must be made up.
CALCULATIONS OF AVERAGES.--The first step in arriving at average
values is to reduce erratic high assays to the general tenor of
other adjacent samples. This point has been disputed at some length,
more often by promoters than by engineers, but the custom is very
generally and rightly adopted. Erratically high samples may indicate
presence of undue metal in the assay attributable to unconscious
salting, for if the value be confined to a few large particles
they may find their way through all the quartering into the assay.
Or the sample may actually indicate rich spots of ore; but in any
event experience teaches that no dependence can be put upon regular
recurrence of such abnormally rich spots. As will be discussed
under percentage of error in sampling, samples usually indicate
higher than the true value, even where erratic assays have been
eliminated. There are cases of profitable mines where the values
were all in spots, and an assay plan would show 80% of the assays
_nil_, yet these pockets were so rich as to give value to the whole.
Pocket mines, as stated before, are beyond valuation by sampling,
and aside from the previous yield recourse must be had to actual
treatment runs on every block of ore separately.
After reduction of erratic assays, a preliminary study of the runs of
value or shapes of the ore-bodies is necessary before any calculation
of averages. A preliminary delineation of the boundaries of the
payable areas on the assay plan will indicate the sections of the
mine which are unpayable, and from which therefore samples can
be rightly excluded in arriving at an average of the payable ore
(Fig. 1). In a general way, only the ore which must be mined need
be included in averaging.
The calculation of the average assay value of standing ore from
samples is one which seems to require some statement of elementals.
Although it may seem primitive, it can do no harm to recall that if
a dump of two tons of ore assaying twenty ounces per ton be added
to a dump of five tons averaging one ounce per ton, the result has
not an average assay of twenty-one ounces divided by the number of
dumps. Likewise one sample over a width of two feet, assaying twenty
ounces per ton, if averaged with another sample over a width of five
feet, assaying one ounce, is no more twenty-one ounces divided by
two samples than in the case of the two dumps. If common sense were
not sufficient demonstration of this, it can be shown algebraically.
Were samples equidistant from each other, and were they of equal
width, the average value would be the simple arithmetical mean of
the assays. But this is seldom the case. The number of instances,
not only in practice but also in technical literature, where the
fundamental distinction between an arithmetical and a geometrical
mean is lost sight of is amazing.
To arrive at the average value of samples, it is necessary, in
effect, to reduce them to the actual quantity of the metal and volume
of ore represented by each. The method of calculation therefore
is one which gives every sample an importance depending upon the
metal content of the volume of ore it represents.
The volume of ore appertaining to any given sample can be considered
as a prismoid, the dimensions of which may be stated as follows:--
_W_ = Width in feet of ore sampled.
_L_ = Length in feet of ore represented by the sample.
_D_ = Depth into the block to which values are assumed to penetrate.
We may also let:--
_C_ = The number of cubic feet per ton of ore.
_V_ = Assay value of the sample.
Then _WLD_/C_ = tonnage of the prismoid.*
_V WLD_/C_ = total metal contents.
[Footnote *: Strictly, the prismoidal formula should be used, but
it complicates the study unduly, and for practical purposes the
above may be taken as the volume.]
The average value of a number of samples is the total metal contents
of their respective prismoids, divided by the total tonnage of
these prismoids. If we let _W_, _W_1, _V_, _V_1 etc., represent
different samples, we have:--
_V(_WLD_/_C_) + _V_1 (_W_1 _L_1 _D_1/_C_) + _V_2 (_W_2 _L_2 _D_2/_C_)
---------------------------------------------------------------------
_WLD_/_C_ + _W_1 _L_1 _D_1/_C_ + _W_2 _L_2 _D_2/_C_
= average value.
This may be reduced to:--
(_VWLD_) + (_V_1 _W_1 _L_1 _D_1) + (_V_2 _W_2 _L_2 _D_2,), etc.
---------------------------------------------------------------
(_WLD_) + (_W_1 _L_1 _D_1) + (_W_2 _L_2 _D_2), etc.
As a matter of fact, samples actually represent the value of
the outer shell of the block of ore only, and the continuity of
the same values through the block is a geological assumption.
From the outer shell, all the values can be taken to penetrate
equal distances into the block, and therefore _D_, _D_1, _D_2
may be considered as equal and the equation becomes:--
(_VWL_) + (_V_1 _W_1 _L_1) + (_V_2 _W_2 _L_2), etc.
---------------------------------------------------
(_WL_) + (_W_1 _L_1) + (_W_2 _L_2), etc.
The length of the prismoid base _L_ for any given sample will be
a distance equal to one-half the sum of the distances to the two
adjacent samples. As a matter of practice, samples are usually taken
at regular intervals, and the lengths _L_, _L_1, _L_2 becoming thus
equal can in such case be eliminated, and the equation becomes:--
(_VW_) + (_V_1 _W_1) + (_V_2 _W_2), etc.
----------------------------------------
_W_ + _W_1 + _W_2 , etc.
The name "assay foot" or "foot value" has been given to the relation
_VW_, that is, the assay value multiplied by the width sampled.[*]
It is by this method that all samples must be averaged. The same
relation obviously can be evolved by using an inch instead of a
foot, and in narrow veins the assay inch is generally used.
[Footnote *: An error will be found in this method unless the two
end samples be halved, but in a long run of samples this may be
disregarded.]
Where the payable cross-section is divided into more than one sample,
the different samples in the section must be averaged by the above
formula, before being combined with the adjacent section. Where
the width sampled is narrower than the necessary stoping width,
and where the waste cannot be broken separately, the sample value
must be diluted to a stoping width. To dilute narrow samples to
a stoping width, a blank value over the extra width which it is
necessary to include must be averaged with the sample from the
ore on the above formula. Cases arise where, although a certain
width of waste must be broken with the ore, it subsequently can
be partially sorted out. Practically nothing but experience on
the deposit itself will determine how far this will restore the
value of the ore to the average of the payable seam. In any event,
no sorting can eliminate all such waste; and it is necessary to
calculate the value on the breaking width, and then deduct from
the gross tonnage to be broken a percentage from sorting. There
is always an allowance to be made in sorting for a loss of good
ore with the discards.
PERCENTAGE OF ERROR IN ESTIMATES FROM SAMPLING.--It must be remembered
that the whole theory of estimation by sampling is founded upon
certain assumptions as to evenness of continuity and transition
in value and volume. It is but a basis for an estimate, and an
estimate is not a statement of fact. It cannot therefore be too
forcibly repeated that an estimate is inherently but an approximation,
take what care one may in its founding. While it is possible to
refine mathematical calculation of averages to almost any nicety,
beyond certain essentials it adds nothing to accuracy and is often
misleading.
It is desirable to consider where errors are most likely to creep
in, assuming that all fundamental data are both accurately taken
and considered. Sampling of ore _in situ_ in general has a tendency
to give higher average value than the actual reduction of the ore
will show. On three West Australian gold mines, in records covering
a period of over two years, where sampling was most exhaustive as
a daily régime of the mines, the values indicated by sampling were
12% higher than the mill yield plus the contents of the residues.
On the Witwatersrand gold mines, the actual extractable value is
generally considered to be about 78 to 80% of the average shown
by sampling, while the mill extractions are on average about 90
to 92% of the head value coming to the mill. In other words, there
is a constant discrepancy of about 10 to 12% between the estimated
value as indicated by mine samples, and the actual value as shown
by yield plus the residues. At Broken Hill, on three lead mines,
the yield is about 12% less than sampling would indicate. This
constancy of error in one direction has not been so generally
acknowledged as would be desirable, and it must be allowed for
in calculating final results. The causes of the exaggeration seem
to be:--
_First_, inability to stope a mine to such fine limitations of
width, or exclusion of unpayable patches, as would appear practicable
when sampling, that is by the inclusion when mining of a certain
amount of barren rock. Even in deposits of about normal stoping
width, it is impossible to prevent the breaking of a certain amount
of waste, even if the ore occurrence is regularly confined by walls.
If the mine be of the impregnation type, such as those at Goldfield,
or Kalgoorlie, with values like plums in a pudding, and the stopes
themselves directed more by assays than by any physical differences
in the ore, the discrepancy becomes very much increased. In mines
where the range of values is narrower than the normal stoping width,
some wall rock must be broken. Although it is customary to allow for
this in calculating the average value from samples, the allowance
seldom seems enough. In mines where the ore is broken on to the
top of stopes filled with waste, there is some loss underground
through mixture with the filling.
_Second_, the metal content of ores, especially when in the form of
sulphides, is usually more friable than the matrix, and in actual
breaking of samples an undue proportion of friable material usually
creeps in. This is true more in lead, copper, and zinc, than in
gold ores. On several gold mines, however, tests on accumulated
samples for their sulphide percentage showed a distinctly greater
ratio than the tenor of the ore itself in the mill. As the gold is
usually associated with the sulphides, the samples showed higher
values than the mill.
In general, some considerable factor of safety must be allowed
after arriving at calculated average of samples,--how much it is
difficult to say, but, in any event, not less than 10%.
CHAPTER II.
Mine Valuation (_Continued_).
CALCULATION OF QUANTITIES OF ORE, AND CLASSIFICATION OF ORE IN SIGHT.
As mines are opened by levels, rises, etc., through the ore, an
extension of these workings has the effect of dividing it into
"blocks." The obvious procedure in determining tonnages is to calculate
the volume and value of each block separately. Under the law of
averages, the multiplicity of these blocks tends in proportion
to their number to compensate the percentage of error which might
arise in the sampling or estimating of any particular one. The
shapes of these blocks, on longitudinal section, are often not
regular geometrical figures. As a matter of practice, however, they
can be subdivided into such figures that the total will approximate
the whole with sufficient closeness for calculations of their areas.
The average width of the ore in any particular block is the arithmetical
mean of the width of the sample sections in it,[*] if the samples be
an equal distance apart. If they are not equidistant, the average
width is the sum of the areas between samples, divided by the total
length sampled. The cubic foot contents of a particular block is
obviously the width multiplied by the area of its longitudinal
section.
[Footnote *: This is not strictly true unless the sum of the widths
of the two end-sections be divided by two and the result incorporated
in calculating the means. In a long series that error is of little
importance.]
The ratio of cubic feet to tons depends on the specific gravity
of the ore, its porosity, and moisture. The variability of ores
throughout the mine in all these particulars renders any method
of calculation simply an approximation in the end. The factors
which must remain unknown necessarily lead the engineer to the
provision of a margin of safety, which makes mathematical refinement
and algebraic formulæ ridiculous.
There are in general three methods of determination of the specific
volume of ores:--
_First_, by finding the true specific gravity of a sufficient number
of representative specimens; this, however, would not account for
the larger voids in the ore-body and in any event, to be anything
like accurate, would be as expensive as sampling and is therefore
of little more than academic interest.
_Second_, by determining the weight of quantities broken from measured
spaces. This also would require several tests from different portions
of the mine, and, in examinations, is usually inconvenient and
difficult. Yet it is necessary in cases of unusual materials, such
as leached gossans, and it is desirable to have it done sooner
or later in going mines, as a check.
_Third_, by an approximation based upon a calculation from the
specific gravities of the predominant minerals in the ore. Ores
are a mixture of many minerals; the proportions vary through the
same ore-body. Despite this, a few partial analyses, which are
usually available from assays of samples and metallurgical tests,
and a general inspection as to the compactness of the ore, give a
fairly reliable basis for approximation, especially if a reasonable
discount be allowed for safety. In such discount must be reflected
regard for the porosity of the ore, and the margin of safety necessary
may vary from 10 to 25%. If the ore is of unusual character, as
in leached deposits, as said before, resort must be had to the
second method.
The following table of the weights per cubic foot and the number
of cubic feet per ton of some of the principal ore-forming minerals
and gangue rocks will be useful for approximating the weight of
a cubic foot of ore by the third method. Weights are in pounds
avoirdupois, and two thousand pounds are reckoned to the ton.
============================================
| | Number of
| Weight per | Cubic Feet
| Cubic Foot | per Ton of
| | 2000 lb.
------------------|------------|------------
Antimony | 417.50 | 4.79
Sulphide | 285.00 | 7.01
Arsenical Pyrites | 371.87 | 5.37
Barium Sulphate | 278.12 | 7.19
Calcium: | |
Fluorite | 198.75 | 10.06
Gypsum | 145.62 | 13.73
Calcite | 169.37 | 11.80
Copper | 552.50 | 3.62
Calcopyrite | 262.50 | 7.61
Bornite | 321.87 | 6.21
Malachite | 247.50 | 8.04
Azurite | 237.50 | 8.42
Chrysocolla | 132.50 | 15.09
Iron (Cast) | 450.00 | 4.44
Magnetite | 315.62 | 6.33
Hematite | 306.25 | 6.53
Limonite | 237.50 | 8.42
Pyrite | 312.50 | 6.40
Carbonate | 240.62 | 8.31
Lead | 710.62 | 2.81
Galena | 468.75 | 4.27
Carbonate | 406.87 | 4.81
Manganese Oxide | 268.75 | 6.18
Rhodonite | 221.25 | 9.04
Magnesite | 187.50 | 10.66
Dolomite | 178.12 | 11.23
Quartz | 165.62 | 12.07
Quicksilver | 849.75 | 2.35
Cinnabar | 531.25 | 3.76
Sulphur | 127.12 | 15.74
Tin | 459.00 | 4.35
Oxide | 418.75 | 4.77
Zinc | 437.50 | 4.57
Blende | 253.12 | 7.90
Carbonate | 273.12 | 7.32
Silicate | 215.62 | 9.28
Andesite | 165.62 | 12.07
Granite | 162.62 | 12.30
Diabase | 181.25 | 11.03
Diorite | 171.87 | 11.63
Slates | 165.62 | 12.07
Sandstones | 162.50 | 12.30
Rhyolite | 156.25 | 12.80
============================================
The specific gravity of any particular mineral has a considerable
range, and a medium has been taken. The possible error is
inconsequential for the purpose of these calculations.
For example, a representative gold ore may contain in the main
96% quartz, and 4% iron pyrite, and the weight of the ore may be
deduced as follows:--
Quartz, 96% x 12.07 = 11.58
Iron Pyrite, 4% x 6.40 = .25
-----
11.83 cubic feet per ton.
Most engineers, to compensate porosity, would allow twelve to thirteen
cubic feet per ton.
CLASSIFICATION OF ORE IN SIGHT.
The risk in estimates of the average value of standing ore is dependent
largely upon how far values disclosed by sampling are assumed to
penetrate beyond the tested face, and this depends upon the geological
character of the deposit. From theoretical grounds and experience,
it is known that such values will have some extension, and the
assumption of any given distance is a calculation of risk. The
multiplication of development openings results in an increase of
sampling points available and lessens the hazards. The frequency
of such openings varies in different portions of every mine, and
thus there are inequalities of risk. It is therefore customary in
giving estimates of standing ore to classify the ore according
to the degree of risk assumed, either by stating the number of
sides exposed or by other phrases. Much discussion and ink have
been devoted to trying to define what risk may be taken in such
matters, that is in reality how far values may be assumed to penetrate
into the unbroken ore. Still more has been consumed in attempts
to coin terms and make classifications which will indicate what
ratio of hazard has been taken in stating quantities and values.
The old terms "ore in sight" and "profit in sight" have been of
late years subject to much malediction on the part of engineers
because these expressions have been so badly abused by the charlatans
of mining in attempts to cover the flights of their imaginations. A
large part of Volume X of the "Institution of Mining and Metallurgy"
has been devoted to heaping infamy on these terms, yet not only
have they preserved their places in professional nomenclature,
but nothing has been found to supersede them.
Some general term is required in daily practice to cover the whole
field of visible ore, and if the phrase "ore in sight" be defined,
it will be easier to teach the laymen its proper use than to abolish
it. In fact, the substitutes are becoming abused as much as the
originals ever were. All convincing expressions will be misused
by somebody.
The legitimate direction of reform has been to divide the general
term of "ore in sight" into classes, and give them names which will
indicate the variable amount of risk of continuity in different parts
of the mine. As the frequency of sample points, and consequently the
risk of continuity, will depend upon the detail with which the mine
is cut into blocks by the development openings, and upon the number
of sides of such blocks which are accessible, most classifications
of the degree of risk of continuity have been defined in terms of
the number of sides exposed in the blocks. Many phrases have been
coined to express such classifications; those most currently used
are the following:--
Positive Ore \ Ore exposed on four sides in blocks of a size
Ore Developed / variously prescribed.
Ore Blocked Out Ore exposed on three sides within reasonable
distance of each other.
Probable Ore \
Ore Developing / Ore exposed on two sides.
Possible Ore \ The whole or a part of the ore below the
Ore Expectant / lowest level or beyond the range of vision.
No two of these parallel expressions mean quite the same thing;
each more or less overlies into another class, and in fact none
of them is based upon a logical footing for such a classification.
For example, values can be assumed to penetrate some distance from
every sampled face, even if it be only ten feet, so that ore exposed
on one side will show some "positive" or "developed" ore which, on
the lines laid down above, might be "probable" or even "possible"
ore. Likewise, ore may be "fully developed" or "blocked out" so far
as it is necessary for stoping purposes with modern wide intervals
between levels, and still be in blocks too large to warrant an
assumption of continuity of values to their centers (Fig. 1). As
to the third class of "possible" ore, it conveys an impression
of tangibility to a nebulous hazard, and should never be used in
connection with positive tonnages. This part of the mine's value
comes under extension of the deposit a long distance beyond openings,
which is a speculation and cannot be defined in absolute tons without
exhaustive explanation of the risks attached, in which case any
phrase intended to shorten description is likely to be misleading.
[Illustration: Fig. 1.--Longitudinal section of a mine, showing
classification of the exposed ore. Scale, 400 feet = 1 inch.]
Therefore empirical expressions in terms of development openings
cannot be made to cover a geologic factor such as the distribution
of metals through a rock mass. The only logical basis of ore
classification for estimation purposes is one which is founded
on the chances of the values penetrating from the surface of the
exposures for each particular mine. Ore that may be calculated
upon to a certainty is that which, taking into consideration the
character of the deposit, can be said to be so sufficiently surrounded
by sampled faces that the distance into the mass to which values
are assumed to extend is reduced to a minimum risk. Ore so far
removed from the sampled face as to leave some doubt, yet affording
great reason for expectation of continuity, is "probable" ore.
The third class of ore mentioned, which is that depending upon
extension of the deposit and in which, as said above, there is great
risk, should be treated separately as the speculative value of the
mine. Some expressions are desirable for these classifications, and
the writer's own preference is for the following, with a definition
based upon the controlling factor itself.
They are:--
Proved Ore Ore where there is practically no risk of
failure of continuity.
Probable Ore Ore where there is some risk, yet warrantable
justification for assumption of continuity.
Prospective Ore Ore which cannot be included in the above
classes, nor definitely known or stated in
any terms of tonnage.
What extent of openings, and therefore of sample faces, is required
for the ore to be called "proved" varies naturally with the type
of deposit,--in fact with each mine. In a general way, a fair rule
in gold quartz veins below influence of secondary alteration is
that no point in the block shall be over fifty feet from the points
sampled. In limestone or andesite replacements, as by gold or lead
or copper, the radius must be less. In defined lead and copper
lodes, or in large lenticular bodies such as the Tennessee copper
mines, the radius may often be considerably greater,--say one hundred
feet. In gold deposits of such extraordinary regularity of values
as the Witwatersrand bankets, it can well be two hundred or two
hundred and fifty feet.
"Probable ore" should be ore which entails continuity of values
through a greater distance than the above, and such distance must
depend upon the collateral evidence from the character of the deposit,
the position of openings, etc.
Ore beyond the range of the "probable" zone is dependent upon the
extension of the deposit beyond the realm of development and will
be discussed separately.
Although the expression "ore in sight" may be deprecated, owing to
its abuse, some general term to cover both "positive" and "probable"
ore is desirable; and where a general term is required, it is the
intention herein to hold to the phrase "ore in sight" under the
limitations specified.
CHAPTER III.
Mine Valuation (_Continued_).
PROSPECTIVE VALUE.[*] EXTENSION IN DEPTH; ORIGIN AND STRUCTURAL
CHARACTER OF THE DEPOSIT; SECONDARY ENRICHMENT; DEVELOPMENT IN
NEIGHBORING MINES; DEPTH OF EXHAUSTION.
[Footnote *: The term "extension in depth" is preferred by many
to the phrase "prospective value." The former is not entirely
satisfactory, as it has a more specific than general application.
It is, however, a current miner's phrase, and is more expressive.
In this discussion "extension in depth" is used synonymously, and
it may be taken to include not alone the downward prolongation of
the ore below workings, but also the occasional cases of lateral
extension beyond the range of development work. The commonest instance
is continuance below the bottom level. In any event, to the majority
of cases of different extension the same reasoning applies.]
It is a knotty problem to value the extension of a deposit beyond
a short distance from the last opening. A short distance beyond
it is "proved ore," and for a further short distance is "probable
ore." Mines are very seldom priced at a sum so moderate as that
represented by the profit to be won from the ore in sight, and what
value should be assigned to this unknown portion of the deposit
admits of no certainty. No engineer can approach the prospective
value of a mine with optimism, yet the mining industry would be
non-existent to-day were it approached with pessimism. Any value
assessed must be a matter of judgment, and this judgment based on
geological evidence. Geology is not a mathematical science, and
to attach a money equivalence to forecasts based on such evidence
is the most difficult task set for the mining engineer. It is here
that his view of geology must differ from that of his financially
more irresponsible brother in the science. The geologist, contributing
to human knowledge in general, finds his most valuable field in the
examination of mines largely exhausted. The engineer's most valuable
work arises from his ability to anticipate in the youth of the mine
the symptoms of its old age. The work of our geologic friends is,
however, the very foundation on which we lay our forecasts.
Geologists have, as the result of long observation, propounded for
us certain hypotheses which, while still hypotheses, have proved
to account so widely for our underground experience that no engineer
can afford to lose sight of them. Although there is a lack of safety
in fixed theories as to ore deposition, and although such conclusions
cannot be translated into feet and metal value, they are nevertheless
useful weights on the scale where probabilities are to be weighed.
A method in vogue with many engineers is, where the bottom level
is good, to assume the value of the extension in depth as a sum
proportioned to the profit in sight, and thus evade the use of
geological evidence. The addition of various percentages to the
profit in sight has been used by engineers, and proposed in technical
publications, as varying from 25 to 50%. That is, they roughly
assess the extension in depth to be worth one-fifth to one-third
of the whole value of an equipped mine. While experience may have
sometimes demonstrated this to be a practical method, it certainly
has little foundation in either science or logic, and the writer's
experience is that such estimates are untrue in practice. The quantity
of ore which may be in sight is largely the result of managerial
policy. A small mill on a large mine, under rapid development,
will result in extensive ore-reserves, while a large mill eating
away rapidly on the same mine under the same scale of development
would leave small reserves. On the above scheme of valuation the
extension in depth would be worth very different sums, even when the
deepest level might be at the same horizon in both cases. Moreover,
no mine starts at the surface with a large amount of ore in sight.
Yet as a general rule this is the period when its extension is most
valuable, for when the deposit is exhausted to 2000 feet, it is
not likely to have such extension in depth as when opened one hundred
feet, no matter what the ore-reserves may be. Further, such bases
of valuation fail to take into account the widely varying geologic
character of different mines, and they disregard any collateral
evidence either of continuity from neighboring development, or from
experience in the district. Logically, the prospective value can
be simply a factor of how _far_ the ore in the individual mine
may be expected to extend, and not a factor of the remnant of ore
that may still be unworked above the lowest level.
An estimation of the chances of this extension should be based
solely on the local factors which bear on such extension, and these
are almost wholly dependent upon the character of the deposit.
These various geological factors from a mining engineer's point
of view are:--
1. The origin and structural character of the ore-deposit.
2. The position of openings in relation to secondary alteration.
3. The size of the deposit.
4. The depth to which the mine has already been exhausted.
5. The general experience of the district for continuity and
the development of adjoining mines.
THE ORIGIN AND STRUCTURAL CHARACTER OF THE DEPOSIT.--In a general
way, the ore-deposits of the order under discussion originate primarily
through the deposition of metals from gases or solutions circulating
along avenues in the earth's crust.[*] The original source of metals
is a matter of great disagreement, and does not much concern the
miner. To him, however, the origin and character of the avenue
of circulation, the enclosing rock, the influence of the rocks
on the solution, and of the solutions on the rocks, have a great
bearing on the probable continuity of the volume and value of the
ore.
[Footnote *: The class of magmatic segregations is omitted, as
not being of sufficiently frequent occurrence in payable mines to
warrant troubling with it here.]
All ore-deposits vary in value and, in the miner's view, only those
portions above the pay limit are ore-bodies, or ore-shoots. The
localization of values into such pay areas in an ore-deposit are
apparently influenced by:
1. The distribution of the open spaces created by structural
movement, fissuring, or folding as at Bendigo.
2. The intersection of other fractures which, by mingling of
solutions from different sources, provided precipitating
conditions, as shown by enrichments at cross-veins.
3. The influence of the enclosing rocks by:--
(a) Their solubility, and therefore susceptibility to replacement.
(b) Their influence as a precipitating agent on solutions.
(c) Their influence as a source of metal itself.
(d) Their texture, in its influence on the character of
the fracture. In homogeneous rocks the tendency
is to open clean-cut fissures; in friable
rocks, zones of brecciation; in slates or schistose
rocks, linked lenticular open spaces;--these
influences exhibiting themselves in miner's terms
respectively in "well-defined fissure veins,"
"lodes," and "lenses."
(e) The physical character of the rock mass and the
dynamic forces brought to bear upon it. This
is a difficult study into the physics of stress in
cases of fracturing, but its local application has
not been without results of an important order.
4. Secondary alteration near the surface, more fully discussed
later.
It is evident enough that the whole structure of the deposit is
a necessary study, and even a digest of the subject is not to be
compressed into a few paragraphs.
From the point of view of continuity of values, ore-deposits may
be roughly divided into three classes. They are:--
1. Deposits of the infiltration type in porous beds, such as
Lake Superior copper conglomerates and African gold bankets.
2. Deposits of the fissure vein type, such as California quartz veins.
3. Replacement or impregnation deposits on the lines of fissuring
or otherwise.
In a general way, the uniformity of conditions of deposition in
the first class has resulted in the most satisfactory continuity of
ore and of its metal contents. In the second, depending much upon
the profundity of the earth movements involved, there is laterally
and vertically a reasonable basis for expectation of continuity
but through much less distance than in the first class.
The third class of deposits exhibits widely different phenomena
as to continuity and no generalization is of any value. In gold
deposits of this type in West Australia, Colorado, and Nevada,
continuity far beyond a sampled face must be received with the
greatest skepticism. Much the same may be said of most copper
replacements in limestone. On the other hand the most phenomenal
regularity of values have been shown in certain Utah and Arizona
copper mines, the result of secondary infiltration in porphyritic
gangues. The Mississippi Valley lead and zinc deposits, while irregular
in detail, show remarkable continuity by way of reoccurrence over
wide areas. The estimation of the prospective value of mines where
continuity of production is dependent on reoccurrence of ore-bodies
somewhat proportional to the area, such as these Mississippi deposits
or to some extent as in Cobalt silver veins, is an interesting
study, but one that offers little field for generalization.
THE POSITION OF THE OPENINGS IN RELATION TO SECONDARY ALTERATION.--The
profound alteration of the upper section of ore-deposits by oxidation
due to the action of descending surface waters, and their associated
chemical agencies, has been generally recognized for a great many
years. Only recently, however, has it been appreciated that this
secondary alteration extends into the sulphide zone as well. The
bearing of the secondary alteration, both in the oxidized and upper
sulphide zones, is of the most sweeping economic character. In
considering extension of values in depth, it demands the most rigorous
investigation. Not only does the metallurgical character of the ores
change with oxidation, but the complex reactions due to descending
surface waters cause leaching and a migration of metals from one
horizon to another lower down, and also in many cases a redistribution
of their sequence in the upper zones of the deposit.
The effect of these agencies has been so great in many cases as
to entirely alter the character of the mine and extension in depth
has necessitated a complete reëquipment. For instance, the Mt.
Morgan gold mine, Queensland, has now become a copper mine; the
copper mines at Butte were formerly silver mines; Leadville has
become largely a zinc producer instead of lead.
From this alteration aspect ore-deposits may be considered to have
four horizons:--
1. The zone near the outcrop, where the dominating feature
is oxidation and leaching of the soluble minerals.
2. A lower horizon, still in the zone of oxidation, where the
predominant feature is the deposition of metals as native,
oxides, and carbonates.
3. The upper horizon of the sulphide zone, where the special
feature is the enrichment due to secondary deposition
as sulphides.
4. The region below these zones of secondary alteration, where
the deposit is in its primary state.
These zones are seldom sharply defined, nor are they always all
in evidence. How far they are in evidence will depend, among other
things, upon the amount and rapidity of erosion, the structure and
mineralogical character of the deposit, and upon the enclosing
rock.
If erosion is extremely rapid, as in cold, wet climates, and rough
topography, or as in the case of glaciation of the Lake copper
deposits, denudation follows close on the heels of alteration,
and the surface is so rapidly removed that we may have the primary
ore practically at the surface. Flat, arid regions present the
other extreme, for denudation is much slower, and conditions are
most perfect for deep penetration of oxidizing agencies, and the
consequent alteration and concentration of the metals.
The migration of metals from the top of the oxidized zone leaves
but a barren cap for erosion. The consequent effect of denudation
that lags behind alteration is to raise slowly the concentrated
metals toward the surface, and thus subject them to renewed attack
and repeated migration. In this manner we can account for the enormous
concentration of values in the lower oxidized and upper sulphide
zones overlying very lean sulphides in depth.
Some minerals are more freely soluble and more readily precipitated
than others. From this cause there is in complex metal deposits a
rearrangement of horizontal sequence, in addition to enrichment at
certain horizons and impoverishment at others. The whole subject
is one of too great complexity for adequate consideration in this
discussion. No engineer is properly equipped to give judgment on
extension in depth without a thorough grasp of the great principles
laid down by Van Hise, Emmons, Lindgren, Weed, and others. We may,
however, briefly examine some of the theoretical effects of such
alteration.
Zinc, iron, and lead sulphides are a common primary combination.
These metals are rendered soluble from their usual primary forms
by oxidizing agencies, in the order given. They reprecipitate as
sulphides in the reverse sequence. The result is the leaching of
zinc and iron readily in the oxidized zone, thus differentially
enriching the lead which lags behind, and a further extension of
the lead horizon is provided by the early precipitation of such
lead as does migrate. Therefore, the lead often predominates in
the second and the upper portion of the third zone, with the zinc
and iron below. Although the action of all surface waters is toward
oxidation and carbonation of these metals, the carbonate development
of oxidized zones is more marked when the enclosing rocks are
calcareous.
In copper-iron deposits, the comparatively easy decomposition and
solubility and precipitation of the copper and some iron salts
generally result in more extensive impoverishment of these metals
near the surface, and more predominant enrichment at a lower horizon
than is the case with any other metals. The barren "iron hat" at the
first zone, the carbonates and oxides at the second, the enrichment
with secondary copper sulphides at the top of the third, and the
occurrence of secondary copper-iron sulphides below, are often
most clearly defined. In the easy recognition of the secondary
copper sulphides, chalcocite, bornite, etc., the engineer finds a
finger-post on the road to extension in depth; and the directions
upon this post are not to be disregarded. The number of copper
deposits enriched from unpayability in the first zone to a profitable
character in the next two, and unpayability again in the fourth,
is legion.
Silver occurs most abundantly in combination with either lead,
copper, iron, or gold. As it resists oxidation and solution more
strenuously than copper and iron, its tendency when in combination
with them is to lag behind in migration. There is thus a differential
enrichment of silver in the upper two zones, due to the reduction
in specific gravity of the ore by the removal of associated metals.
Silver does migrate somewhat, however, and as it precipitates more
readily than copper, lead, zinc, or iron, its tendency when in
combination with them is towards enrichment above the horizons of
enrichment of these metals. When it is in combination with lead
and zinc, its very ready precipitation from solution by the galena
leaves it in combination more predominantly with the lead. The
secondary enrichment of silver deposits at the top of the sulphide
zone is sometimes a most pronounced feature, and it seems to be
the explanation of the origin of many "bonanzas."
In gold deposits, the greater resistance to solubility of this
metal than most of the others, renders the phenomena of migration to
depth less marked. Further than this, migration is often interfered
with by the more impervious quartz matrix of many gold deposits.
Where gold is associated with large quantities of base metals,
however, the leaching of the latter in the oxidized zone leaves the
ore differentially richer, and as gold is also slightly soluble,
in such cases the migration of the base metals does carry some of
the gold. In the instance especially of impregnation or replacement
deposits, where the matrix is easily permeable, the upper sulphide
zone is distinctly richer than lower down, and this enrichment is
accompanied by a considerable increase in sulphides and tellurides.
The predominant characteristic of alteration in gold deposits is,
however, enrichment in the oxidized zone with the maximum values
near the surface. The reasons for this appear to be that gold in its
resistance to oxidation and wholesale migration gives opportunities
to a sort of combined mechanical and chemical enrichment.
In dry climates, especially, the gentleness of erosion allows of
more thorough decomposition of the outcroppings, and a mechanical
separation of the gold from the detritus. It remains on or near
the deposit, ready to be carried below, mechanically or otherwise.
In wet climates this is less pronounced, for erosion bears away
the croppings before such an extensive decomposition and freeing
of the gold particles. The West Australian gold fields present an
especially prominent example of this type of superficial enrichment.
During the last fifteen years nearly eight hundred companies have
been formed for working mines in this region. Although from four
hundred of these high-grade ore has been produced, some thirty-three
only have ever paid dividends. The great majority have been unpayable
below oxidation,--a distance of one or two hundred feet. The writer's
unvarying experience with gold is that it is richer in the oxidized
zone than at any point below. While cases do occur of gold deposits
richer in the upper sulphide zone than below, even the upper sulphides
are usually poorer than the oxidized region. In quartz veins
preëminently, evidence of enrichment in the third zone is likely
to be practically absent.
Tin ores present an anomaly among the base metals under discussion,
in that the primary form of this metal in most workable deposits
is an oxide. Tin in this form is most difficult of solution from
ground agencies, as witness the great alluvial deposits, often of
considerable geologic age. In consequence the phenomena of migration
and enrichment are almost wholly absent, except such as are due
to mechanical penetration of tin from surface decomposition of
the matrix akin to that described in gold deposits.
In general, three or four essential facts from secondary alteration
must be kept in view when prognosticating extensions.
Oxidation usually alters treatment problems, and oxidized ore
of the same grade as sulphides can often be treated more cheaply.
This is not universal. Low-grade ores of lead, copper, and zinc
may be treatable by concentration when in the form of sulphides,
and may be valueless when oxidized, even though of the same grade.
Copper ores generally show violent enrichment at the base of the
oxidized, and at the top of the sulphide zone.
Lead-zinc ores show lead enrichment and zinc impoverishment in
the oxidized zone but have usually less pronounced enrichment
below water level than copper. The rearrangement of the metals
by the deeper migration of the zinc, also renders them
metallurgically of less value with depth.
Silver deposits are often differentially enriched in the oxidized
zone, and at times tend to concentrate in the upper sulphide zone.
Gold deposits usually decrease in value from the surface through
the whole of the three alteration zones.
SIZE OF DEPOSITS.--The proverb of a relation between extension
in depth and size of ore-bodies expresses one of the oldest of
miners' beliefs. It has some basis in experience, especially in
fissure veins, but has little foundation in theory and is applicable
over but limited areas and under limited conditions.
From a structural view, the depth of fissuring is likely to be more
or less in proportion to its length and breadth and therefore the
volume of vein filling with depth is likely to be proportional to
length and width of the fissure. As to the distribution of values,
if we eliminate the influence of changing wall rocks, or other
precipitating agencies which often cause the values to arrange
themselves in "floors," and of secondary alteration, there may be
some reason to assume distribution of values of an extent equal
vertically to that displayed horizontally. There is, as said, more
reason in experience for this assumption than in theory. A study
of the shape of a great many ore-shoots in mines of fissure type
indicates that when the ore-shoots or ore-bodies are approaching
vertical exhaustion they do not end abruptly, but gradually shorten
and decrease in value, their bottom boundaries being more often
wedge-shaped than even lenticular. If this could be taken as the usual
occurrence, it would be possible (eliminating the evident exceptions
mentioned above) to state roughly that the minimum extension of an
ore-body or ore-shoot in depth below any given horizon would be
a distance represented by a radius equal to one-half its length. By
length is not meant necessarily the length of a horizontal section,
but of one at right angles to the downward axis.
On these grounds, which have been reënforced by much experience among
miners, the probabilities of extension are somewhat in proportion
to the length and width of each ore-body. For instance, in the A
mine, with an ore-shoot 1000 feet long and 10 feet wide, on its
bottom level, the minimum extension under this hypothesis would
be a wedge-shaped ore-body with its deepest point 500 feet below
the lowest level, or a minimum of say 200,000 tons. Similarly,
the B mine with five ore-bodies, each 300 hundred feet long and
10 feet wide, exposed on its lowest level, would have a minimum of
five wedges 100 feet deep at their deepest points, or say 50,000
tons. This is not proposed as a formula giving the total amount of
extension in depth, but as a sort of yardstick which has experience
behind it. This experience applies in a much less degree to deposits
originating from impregnation along lines of fissuring and not at
all to replacements.
DEVELOPMENT IN NEIGHBORING MINES.--Mines of a district are usually
found under the same geological conditions, and show somewhat the same
habits as to extension in depth or laterally, and especially similar
conduct of ore-bodies and ore-shoots. As a practical criterion, one
of the most intimate guides is the actual development in adjoining
mines. For instance, in Kalgoorlie, the Great Boulder mine is (March,
1908) working the extension of Ivanhoe lodes at points 500 feet
below the lowest level in the Ivanhoe; likewise, the Block 10 lead
mine at Broken Hill is working the Central ore-body on the Central
boundary some 350 feet below the Central workings. Such facts as
these must have a bearing on assessing the downward extension.
DEPTH OF EXHAUSTION.--All mines become completely exhausted at
some point in depth. Therefore the actual distance to which ore
can be expected to extend below the lowest level grows less with
every deeper working horizon. The really superficial character of
ore-deposits, even outside of the region of secondary enrichment
is becoming every year better recognized. The prospector's idea
that "she gets richer deeper down," may have some basis near the
surface in some metals, but it is not an idea which prevails in
the minds of engineers who have to work in depth. The writer, with
some others, prepared a list of several hundred dividend-paying
metal mines of all sorts, extending over North and South America,
Australasia, England, and Africa. Notes were made as far as possible
of the depths at which values gave out, and also at which dividends
ceased. Although by no means a complete census, the list indicated
that not 6% of mines (outside banket) that have yielded profits,
ever made them from ore won below 2000 feet. Of mines that paid
dividends, 80% did not show profitable value below 1500 feet, and
a sad majority died above 500. Failures at short depths may be
blamed upon secondary enrichment, but the majority that reached
below this influence also gave out. The geological reason for such
general unseemly conduct is not so evident.
CONCLUSION.--As a practical problem, the assessment of prospective
value is usually a case of "cut and try." The portion of the capital
to be invested, which depends upon extension, will require so many
tons of ore of the same value as that indicated by the standing
ore, in order to justify the price. To produce this tonnage at
the continued average size of the ore-bodies will require their
extension in depth so many feet--or the discovery of new ore-bodies
of a certain size. The five geological weights mentioned above
may then be put into the scale and a basis of judgment reached.
CHAPTER IV.
Mine Valuation (_Continued_).
RECOVERABLE PERCENTAGE OF THE GROSS ASSAY VALUE; PRICE OF METALS;
COST OF PRODUCTION.
The method of treatment for the ore must be known before a mine
can be valued, because a knowledge of the recoverable percentage
is as important as that of the gross value of the ore itself. The
recoverable percentage is usually a factor of working costs. Practically
every ore can be treated and all the metal contents recovered, but
the real problem is to know the method and percentage of recovery
which will yield the most remunerative result, if any. This limit to
profitable recovery regulates the amount of metal which should be
lost, and the amount of metal which consequently must be deducted
from the gross value before the real net value of the ore can be
calculated. Here, as everywhere else in mining, a compromise has to
be made with nature, and we take what we can get--profitably. For
instance, a copper ore may be smelted and a 99% recovery obtained.
Under certain conditions this might be done at a loss, while the
same ore might be concentrated before smelting and yield a profit
with a 70% recovery. An additional 20% might be obtained by roasting
and leaching the residues from concentration, but this would probably
result in an expenditure far greater than the value of the 20%
recovered. If the ore is not already under treatment on the mine,
or exactly similar ore is not under treatment elsewhere, with known
results, the method must be determined experimentally, either by
the examining engineer or by a special metallurgist.
Where partially treated products, such as concentrates, are to be
sold, not only will there be further losses, but deductions will
be made by the smelter for deleterious metals and other charges.
All of these factors must be found out,--and a few sample smelting
returns from a similar ore are useful.
To cover the whole field of metallurgy and discuss what might apply,
and how it might apply, under a hundred supposititious conditions
would be too great a digression from the subject in hand. It is
enough to call attention here to the fact that the residues from
every treatment carry some metal, and that this loss has to be
deducted from the gross value of the ore in any calculations of
net values.
PRICE OF METALS.
Unfortunately for the mining engineer, not only has he to weigh
the amount of risk inherent in calculations involved in the mine
itself, but also that due to fluctuations in the value of metals.
If the ore is shipped to custom works, he has to contemplate also
variations in freights and smelting charges. Gold from the mine
valuer's point of view has no fluctuations. It alone among the
earth's products gives no concern as to the market price. The price
to be taken for all other metals has to be decided before the mine
can be valued. This introduces a further speculation and, as in
all calculations of probabilities, amounts to an estimate of the
amount of risk. In a free market the law of supply and demand governs
the value of metals as it does that of all other commodities. So
far, except for tariff walls and smelting rings, there is a free
market in the metals under discussion.
The demand for metals varies with the unequal fluctuations of the
industrial tides. The sea of commercial activity is subject to
heavy storms, and the mine valuer is compelled to serve as weather
prophet on this ocean of trouble. High prices, which are the result
of industrial booms, bring about overproduction, and the collapse of
these begets a shrinkage of demand, wherein consequently the tide
of price turns back. In mining for metals each pound is produced
actually at a different cost. In case of an oversupply of base metals
the price will fall until it has reached a point where a portion of
the production is no longer profitable, and the equilibrium is
established through decline in output. However, in the backward
swing, due to lingering overproduction, prices usually fall lower
than the cost of producing even a much-diminished supply. There is
at this point what we may call the "basic" price, that at which
production is insufficient and the price rises again. The basic
price which is due to this undue backward swing is no more the
real price of the metal to be contemplated over so long a term
of years than is the highest price. At how much above the basic
price of depressed times the product can be safely expected to
find a market is the real question. Few mines can be bought or
valued at this basic price. An indication of what this is can be
gained from a study of fluctuations over a long term of years.
It is common to hear the average price over an extended period
considered the "normal" price, but this basis for value is one which
must be used with discretion, for it is not the whole question when
mining. The "normal" price is the average price over a long term.
The lives of mines, and especially ore in sight, may not necessarily
enjoy the period of this "normal" price. The engineer must balance
his judgments by the immediate outlook of the industrial weather.
When lead was falling steadily in December, 1907, no engineer would
accept the price of that date, although it was then below "normal";
his product might go to market even lower yet.
It is desirable to ascertain what the basic and normal prices are,
for between them lies safety. Since 1884 there have been three cycles
of commercial expansion and contraction. If the average prices
are taken for these three cycles separately (1885-95), 1895-1902,
1902-08) it will be seen that there has been a steady advance in
prices. For the succeeding cycles lead on the London Exchange,[*]
the freest of the world's markets was £12 12_s._ 4_d._, £13 3_s._
7_d._, and £17 7_s._ 0_d._ respectively; zinc, £17 14_s._ 10_d._,
£19 3_s._ 8_d._, and £23 3_s._ 0_d._; and standard copper, £48 16_s._
0_d._, £59 10_s._ 0_d._, and £65 7_s._ 0_d._ It seems, therefore,
that a higher standard of prices can be assumed as the basic and
normal than would be indicated if the general average of, say,
twenty years were taken. During this period, the world's gold output
has nearly quadrupled, and, whether the quantitative theory of
gold be accepted or not, it cannot be denied that there has been
a steady increase in the price of commodities. In all base-metal
mining it is well to remember that the production of these metals
is liable to great stimulus at times from the discovery of new
deposits or new processes of recovery from hitherto unprofitable
ores. It is therefore for this reason hazardous in the extreme
to prophesy what prices will be far in the future, even when the
industrial weather is clear. But some basis must be arrived at,
and from the available outlook it would seem that the following
metal prices are justifiable for some time to come, provided the
present tariff schedules are maintained in the United States:
[Footnote *: All London prices are based on the long ton of 2,240
lbs. Much confusion exists in the copper trade as to the classification
of the metal. New York prices are quoted in electrolytic and "Lake";
London's in "Standard." "Standard" has now become practically an
arbitrary term peculiar to London, for the great bulk of copper
dealt in is "electrolytic" valued considerably over "Standard."]
==========================================================================
| Lead | Spelter | Copper | Tin | Silver
|------------|----------|----------|----------|---------------
|London| N.Y.|Lon.| N.Y.|Lon.| N.Y.|Lon.| N.Y.| Lon. | N.Y.
| Ton |Pound|Ton |Pound|Ton |Pound|Ton |Pound|Per oz.|Per oz.
------------|------|-----|----|-----|----|-----|----|-----|-------|-------
Basic Price | £11. |$.035|£17 |$.040|£52 |$.115|£100|$.220| 22_d._|$.44
Normal Price| 13.5| .043| 21 | .050| 65 | .140| 130| .290| 26 | .52
==========================================================================
In these figures the writer has not followed strict averages, but
has taken the general outlook combined with the previous records.
The likelihood of higher prices for lead is more encouraging than
for any other metal, as no new deposits of importance have come
forward for years, and the old mines are reaching considerable
depths. Nor does the frenzied prospecting of the world's surface
during the past ten years appear to forecast any very disturbing
developments. The zinc future is not so bright, for metallurgy
has done wonders in providing methods of saving the zinc formerly
discarded from lead ores, and enormous supplies will come forward
when required. The tin outlook is encouraging, for the supply from
a mining point of view seems unlikely to more than keep pace with
the world's needs. In copper the demand is growing prodigiously,
but the supplies of copper ores and the number of copper mines
that are ready to produce whenever normal prices recur was never
so great as to-day. One very hopeful fact can be deduced for the
comfort of the base metal mining industry as a whole. If the growth
of demand continues through the next thirty years in the ratio of
the past three decades, the annual demand for copper will be over
3,000,000 tons, of lead over 1,800,000 tons, of spelter 2,800,000
tons, of tin 250,000 tons. Where such stupendous amounts of these
metals are to come from at the present range of prices, and even
with reduced costs of production, is far beyond any apparent source
of supply. The outlook for silver prices is in the long run not
bright. As the major portion of the silver produced is a bye product
from base metals, any increase in the latter will increase the
silver production despite very much lower prices for the precious
metal. In the meantime the gradual conversion of all nations to
the gold standard seems a matter of certainty. Further, silver
may yet be abandoned as a subsidiary coinage inasmuch as it has
now but a token value in gold standard countries if denuded of
sentiment.
COST OF PRODUCTION.
It is hardly necessary to argue the relative importance of the
determination of the cost of production and the determination of
the recoverable contents of the ore. Obviously, the aim of mine
valuation is to know the profits to be won, and the profit is the
value of the metal won, less the cost of production.
The cost of production embraces development, mining, treatment,
management. Further than this, it is often contended that, as the
capital expended in purchase and equipment must be redeemed within
the life of the mine, this item should also be included in production
costs. It is true that mills, smelters, shafts, and all the
paraphernalia of a mine are of virtually negligible value when it
is exhausted; and that all mines are exhausted sometime and every
ton taken out contributes to that exhaustion; and that every ton of
ore must bear its contribution to the return of the investment,
as well as profit upon it. Therefore it may well be said that the
redemption of the capital and its interest should be considered
in costs per ton. The difficulty in dealing with the subject from
the point of view of production cost arises from the fact that,
except possibly in the case of banket gold and some conglomerate
copper mines, the life of a metal mine is unknown beyond the time
required to exhaust the ore reserves. The visible life at the time
of purchase or equipment may be only three or four years, yet the
average equipment has a longer life than this, and the anticipation
for every mine is also for longer duration than the bare ore in sight.
For clarity of conclusions in mine valuation the most advisable
course is to determine the profit in sight irrespective of capital
redemption in the first instance. The questions of capital redemption,
purchase price, or equipment cost can then be weighed against the
margin of profit. One phase of redemption will be further discussed
under "Amortization of Capital" and "Ratio of Output to the Mine."
The cost of production depends upon many things, such as the cost of
labor, supplies, the size of the ore-body, the treatment necessary,
the volume of output, etc.; and to discuss them all would lead
into a wilderness of supposititious cases. If the mine is a going
concern, from which reliable data can be obtained, the problem is
much simplified. If it is virgin, the experience of other mines
in the same region is the next resource; where no such data can be
had, the engineer must fall back upon the experience with mines
still farther afield. Use is sometimes made of the "comparison ton"
in calculating costs upon mines where data of actual experience
are not available. As costs will depend in the main upon items
mentioned above, if the known costs of a going mine elsewhere be
taken as a basis, and subtractions and additions made for more
unfavorable or favorable effect of the differences in the above
items, a fairly close result can be approximated.
Mine examinations are very often inspired by the belief that extended
operations or new metallurgical applications to the mine will expand
the profits. In such cases the paramount questions are the reduction
of costs by better plant, larger outputs, new processes, or alteration
of metallurgical basis and better methods. If every item of previous
expenditure be gone over and considered, together with the equipment,
and method by which it was obtained, the possible savings can be
fairly well deduced, and justification for any particular line
of action determined. One view of this subject will be further
discussed under "Ratio of Output to the Mine." The conditions which
govern the working costs are on every mine so special to itself,
that no amount of advice is very useful. Volumes of advice have
been published on the subject, but in the main their burden is
not to underestimate.
In considering the working costs of base-metal mines, much depends
upon the opportunity for treatment in customs works, smelters,
etc. Such treatment means a saving of a large portion of equipment
cost, and therefore of the capital to be invested and subsequently
recovered. The economics of home treatment must be weighed against
the sum which would need to be set aside for redemption of the
plant, and unless there is a very distinct advantage to be had by
the former, no risks should be taken. More engineers go wrong by
the erection of treatment works where other treatment facilities
are available, than do so by continued shipping. There are many
mines where the cost of equipment could never be returned, and
which would be valueless unless the ore could be shipped. Another
phase of foreign treatment arises from the necessity or advantage
of a mixture of ores,--the opportunity of such mixtures often gives
the public smelter an advantage in treatment with which treatment
on the mine could never compete.
Fluctuation in the price of base metals is a factor so much to be
taken into consideration, that it is desirable in estimating mine
values to reduce the working costs to a basis of a "per unit" of
finished metal. This method has the great advantage of indicating
so simply the involved risks of changing prices that whoso runs
may read. Where one metal predominates over the other to such an
extent as to form the "backbone" of the value of the mine, the
value of the subsidiary metals is often deducted from the cost of
the principal metal, in order to indicate more plainly the varying
value of the mine with the fluctuating prices of the predominant
metal. For example, it is usual to state that the cost of copper
production from a given ore will be so many cents per pound, or so
many pounds sterling per ton. Knowing the total metal extractable
from the ore in sight, the profits at given prices of metal can
be readily deduced. The point at which such calculation departs
from the "per-ton-of-ore" unto the per-unit-cost-of-metal basis,
usually lies at the point in ore dressing where it is ready for the
smelter. To take a simple case of a lead ore averaging 20%: this
is to be first concentrated and the lead reduced to a concentrate
averaging 70% and showing a recovery of 75% of the total metal
content. The cost per ton of development, mining, concentration,
management, is to this point say $4 per ton of original crude ore.
The smelter buys the concentrate for 95% of the value of the metal,
less the smelting charge of $15 per ton, or there is a working
cost of a similar sum by home equipment. In this case 4.66 tons of
ore are required to produce one ton of concentrates, and therefore
each ton of concentrates costs $18.64. This amount, added to the
smelting charge, gives a total of $33.64 for the creation of 70%
of one ton of finished lead, or equal to 2.40 cents per pound which
can be compared with the market price less 5%. If the ore were
to contain 20 ounces of silver per ton, of which 15 ounces were
recovered into the leady concentrates, and the smelter price for
the silver were 50 cents per ounce, then the $7.50 thus recovered
would be subtracted from $33.64, making the apparent cost of the
lead 1.86 cents per pound.
CHAPTER V.
Mine Valuation (_Continued_).
REDEMPTION OR AMORTIZATION OF CAPITAL AND INTEREST.
It is desirable to state in some detail the theory of amortization
before consideration of its application in mine valuation.
As every mine has a limited life, the capital invested in it must
be redeemed during the life of the mine. It is not sufficient that
there be a bare profit over working costs. In this particular,
mines differ wholly from many other types of investment, such as
railways. In the latter, if proper appropriation is made for
maintenance, the total income to the investor can be considered as
interest or profit; but in mines, a portion of the annual income
must be considered as a return of capital. Therefore, before the
yield on a mine investment can be determined, a portion of the
annual earnings must be set aside in such a manner that when the
mine is exhausted the original investment will have been restored.
If we consider the date due for the return of the capital as the time
when the mine is exhausted, we may consider the annual instalments
as payments before the due date, and they can be put out at compound
interest until the time for restoration arrives. If they be invested
in safe securities at the usual rate of about 4%, the addition of
this amount of compound interest will assist in the repayment of
the capital at the due date, so that the annual contributions to
a sinking fund need not themselves aggregate the total capital to
be restored, but may be smaller by the deficiency which will be
made up by their interest earnings. Such a system of redemption
of capital is called "Amortization."
Obviously it is not sufficient for the mine investor that his capital
shall have been restored, but there is required an excess earning
over and above the necessities of this annual funding of capital.
What rate of excess return the mine must yield is a matter of the
risks in the venture and the demands of the investor. Mining business
is one where 7% above provision for capital return is an absolute
minimum demanded by the risks inherent in mines, even where the
profit in sight gives warranty to the return of capital. Where
the profit in sight (which is the only real guarantee in mine
investment) is below the price of the investment, the annual return
should increase in proportion. There are thus two distinct directions
in which interest must be computed,--first, the internal influence
of interest in the amortization of the capital, and second, the
percentage return upon the whole investment after providing for
capital return.
There are many limitations to the introduction of such refinements
as interest calculations in mine valuation. It is a subject not
easy to discuss with finality, for not only is the term of years
unknown, but, of more importance, there are many factors of a highly
speculative order to be considered in valuing. It may be said that
a certain life is known in any case from the profit in sight, and
that in calculating this profit a deduction should be made from
the gross profit for loss of interest on it pending recovery. This
is true, but as mines are seldom dealt with on the basis of profit
in sight alone, and as the purchase price includes usually some
proportion for extension in depth, an unknown factor is introduced
which outweighs the known quantities. Therefore the application of
the culminative effect of interest accumulations is much dependent
upon the sort of mine under consideration. In most cases of uncertain
continuity in depth it introduces a mathematical refinement not
warranted by the speculative elements. For instance, in a mine
where the whole value is dependent upon extension of the deposit
beyond openings, and where an expected return of at least 50% per
annum is required to warrant the risk, such refinement would be
absurd. On the other hand, in a Witwatersrand gold mine, in gold
and tin gravels, or in massive copper mines such as Bingham and
Lake Superior, where at least some sort of life can be approximated,
it becomes a most vital element in valuation.
In general it may be said that the lower the total annual return
expected upon the capital invested, the greater does the amount
demanded for amortization become in proportion to this total income,
and therefore the greater need of its introduction in calculations.
Especially is this so where the cost of equipment is large
proportionately to the annual return. Further, it may be said that
such calculations are of decreasing use with increasing proportion of
speculative elements in the price of the mine. The risk of extension in
depth, of the price of metal, etc., may so outweigh the comparatively
minor factors here introduced as to render them useless of attention.
In the practical conduct of mines or mining companies, sinking
funds for amortization of capital are never established. In the
vast majority of mines of the class under discussion, the ultimate
duration of life is unknown, and therefore there is no basis upon
which to formulate such a definite financial policy even were it
desired. Were it possible to arrive at the annual sum to be set
aside, the stockholders of the mining type would prefer to do their
own reinvestment. The purpose of these calculations does not lie
in the application of amortization to administrative finance. It
is nevertheless one of the touchstones in the valuation of certain
mines or mining investments. That is, by a sort of inversion such
calculations can be made to serve as a means to expose the amount
of risk,--to furnish a yardstick for measuring the amount of risk
in the very speculations of extension in depth and price of metals
which attach to a mine. Given the annual income being received,
or expected, the problem can be formulated into the determination
of how many years it must be continued in order to amortize the
investment and pay a given rate of profit. A certain length of
life is evident from the ore in sight, which may be called the
life in sight. If the term of years required to redeem the capital
and pay an interest upon it is greater than the life in sight,
then this extended life must come from extension in depth, or ore
from other direction, or increased price of metals. If we then take
the volume and profit on the ore as disclosed we can calculate the
number of feet the deposit must extend in depth, or additional tonnage
that must be obtained of the same grade, or the different prices of
metal that must be secured, in order to satisfy the demanded term
of years. These demands in actual measure of ore or feet or higher
price can then be weighed against the geological and industrial
probabilities.
The following tables and examples may be of assistance in these
calculations.
Table 1. To apply this table, the amount of annual income or dividend
and the term of years it will last must be known or estimated factors.
It is then possible to determine the _present_ value of this annual
income after providing for amortization and interest on the investment
at various rates given, by multiplying the annual income by the
factor set out.
A simple illustration would be that of a mine earning a profit of
$200,000 annually, and having a total of 1,000,000 tons in sight,
yielding a profit of $2 a ton, or a total profit in sight of $2,000,000,
thus recoverable in ten years. On a basis of a 7% return on the
investment and amortization of capital (Table I), the factor is
6.52 x $200,000 = $1,304,000 as the present value of the gross
profits exposed. That is, this sum of $1,304,000, if paid for the
mine, would be repaid out of the profit in sight, together with
7% interest if the annual payments into sinking fund earn 4%.
TABLE I.
Present Value of an Annual Dividend Over -- Years at --% and Replacing
Capital by Reinvestment of an Annual Sum at 4%.
=======================================================
Years | 5% | 6% | 7% | 8% | 9% | 10%
-------|-------|-------|-------|-------|-------|-------
1 | .95 | .94 | .93 | .92 | .92 | .91
2 | 1.85 | 1.82 | 1.78 | 1.75 | 1.72 | 1.69
3 | 2.70 | 2.63 | 2.56 | 2.50 | 2.44 | 2.38
4 | 3.50 | 3.38 | 3.27 | 3.17 | 3.07 | 2.98
5 | 4.26 | 4.09 | 3.93 | 3.78 | 3.64 | 3.51
6 | 4.98 | 4.74 | 4.53 | 4.33 | 4.15 | 3.99
7 | 5.66 | 5.36 | 5.09 | 4.84 | 4.62 | 4.41
8 | 6.31 | 5.93 | 5.60 | 5.30 | 5.04 | 4.79
9 | 6.92 | 6.47 | 6.08 | 5.73 | 5.42 | 5.14
10 | 7.50 | 6.98 | 6.52 | 6.12 | 5.77 | 5.45
| | | | | |
11 | 8.05 | 7.45 | 6.94 | 6.49 | 6.09 | 5.74
12 | 8.58 | 7.90 | 7.32 | 6.82 | 6.39 | 6.00
13 | 9.08 | 8.32 | 7.68 | 7.13 | 6.66 | 6.24
14 | 9.55 | 8.72 | 8.02 | 7.42 | 6.91 | 6.46
15 | 10.00 | 9.09 | 8.34 | 7.79 | 7.14 | 6.67
16 | 10.43 | 9.45 | 8.63 | 7.95 | 7.36 | 6.86
17 | 10.85 | 9.78 | 8.91 | 8.18 | 7.56 | 7.03
18 | 11.24 | 10.10 | 9.17 | 8.40 | 7.75 | 7.19
19 | 11.61 | 10.40 | 9.42 | 8.61 | 7.93 | 7.34
20 | 11.96 | 10.68 | 9.65 | 8.80 | 8.09 | 7.49
| | | | | |
21 | 12.30 | 10.95 | 9.87 | 8.99 | 8.24 | 7.62
22 | 12.62 | 11.21 | 10.08 | 9.16 | 8.39 | 7.74
23 | 12.93 | 11.45 | 10.28 | 9.32 | 8.52 | 7.85
24 | 13.23 | 11.68 | 10.46 | 9.47 | 8.65 | 7.96
25 | 13.51 | 11.90 | 10.64 | 9.61 | 8.77 | 8.06
26 | 13.78 | 12.11 | 10.80 | 9.75 | 8.88 | 8.16
27 | 14.04 | 12.31 | 10.96 | 9.88 | 8.99 | 8.25
28 | 14.28 | 12.50 | 11.11 | 10.00 | 9.09 | 8.33
29 | 14.52 | 12.68 | 11.25 | 10.11 | 9.18 | 8.41
30 | 14.74 | 12.85 | 11.38 | 10.22 | 9.27 | 8.49
| | | | | |
31 | 14.96 | 13.01 | 11.51 | 10.32 | 9.36 | 8.56
32 | 15.16 | 13.17 | 11.63 | 10.42 | 9.44 | 8.62
33 | 15.36 | 13.31 | 11.75 | 10.51 | 9.51 | 8.69
34 | 15.55 | 13.46 | 11.86 | 10.60 | 9.59 | 8.75
35 | 15.73 | 13.59 | 11.96 | 10.67 | 9.65 | 8.80
36 | 15.90 | 13.72 | 12.06 | 10.76 | 9.72 | 8.86
37 | 16.07 | 13.84 | 12.16 | 10.84 | 9.78 | 8.91
38 | 16.22 | 13.96 | 12.25 | 10.91 | 9.84 | 8.96
39 | 16.38 | 14.07 | 12.34 | 10.98 | 9.89 | 9.00
40 | 16.52 | 14.18 | 12.42 | 11.05 | 9.95 | 9.05
=======================================================
Condensed from Inwood's Tables.
Table II is practically a compound discount table. That is, by
it can be determined the present value of a fixed sum payable at
the end of a given term of years, interest being discounted at
various given rates. Its use may be illustrated by continuing the
example preceding.
TABLE II.
Present Value of $1, or £1, payable in -- Years, Interest taken
at --%.
===================================
Years | 4% | 5% | 6% | 7%
------|------|------|------|-------
1 | .961 | .952 | .943 | .934
2 | .924 | .907 | .890 | .873
3 | .889 | .864 | .840 | .816
4 | .854 | .823 | .792 | .763
5 | .821 | .783 | .747 | .713
6 | .790 | .746 | .705 | .666
7 | .760 | .711 | .665 | .623
8 | .731 | .677 | .627 | .582
9 | .702 | .645 | .592 | .544
10 | .675 | .614 | .558 | .508
| | | |
11 | .649 | .585 | .527 | .475
12 | .625 | .557 | .497 | .444
13 | .600 | .530 | .469 | .415
14 | .577 | .505 | .442 | .388
15 | .555 | .481 | .417 | .362
16 | .534 | .458 | .394 | .339
17 | .513 | .436 | .371 | .316
18 | .494 | .415 | .350 | .296
19 | .475 | .396 | .330 | .276
20 | .456 | .377 | .311 | .258
| | | |
21 | .439 | .359 | .294 | .241
22 | .422 | .342 | .277 | .266
23 | .406 | .325 | .262 | .211
24 | .390 | .310 | .247 | .197
25 | .375 | .295 | .233 | .184
26 | .361 | .281 | .220 | .172
27 | .347 | .268 | .207 | .161
28 | .333 | .255 | .196 | .150
29 | .321 | .243 | .184 | .140
30 | .308 | .231 | .174 | .131
| | | |
31 | .296 | .220 | .164 | .123
32 | .285 | .210 | .155 | .115
33 | .274 | .200 | .146 | .107
34 | .263 | .190 | .138 | .100
35 | .253 | .181 | .130 | .094
36 | .244 | .172 | .123 | .087
37 | .234 | .164 | .116 | .082
38 | .225 | .156 | .109 | .076
39 | .216 | .149 | .103 | .071
40 | .208 | .142 | .097 | .067
===================================
Condensed from Inwood's Tables.
If such a mine is not equipped, and it is assumed that $200,000
are required to equip the mine, and that two years are required
for this equipment, the value of the ore in sight is still less,
because of the further loss of interest in delay and the cost of
equipment. In this case the present value of $1,304,000 in two
years, interest at 7%, the factor is .87 X 1,304,000 = $1,134,480.
From this comes off the cost of equipment, or $200,000, leaving
$934,480 as the present value of the profit in sight. A further
refinement could be added by calculating the interest chargeable
against the $200,000 equipment cost up to the time of production.
TABLE III.
===========================================================================
Annual | Number of years of life required to yield--% interest, and in
Rate of | addition to furnish annual instalments which, if reinvested at
Dividend.| 4% will return the original investment at the end of the period.
---------|-----------------------------------------------------------------
% | 5% | 6% | 7% | 8% | 9% | 10%
| | | | | |
6 | 41.0 | | | | |
7 | 28.0 | 41.0 | | | |
8 | 21.6 | 28.0 | 41.0 | | |
9 | 17.7 | 21.6 | 28.0 | 41.0 | |
10 | 15.0 | 17.7 | 21.6 | 28.0 | 41.0 |
| | | | | |
11 | 13.0 | 15.0 | 17.7 | 21.6 | 28.0 | 41.0
12 | 11.5 | 13.0 | 15.0 | 17.7 | 21.6 | 28.0
13 | 10.3 | 11.5 | 13.0 | 15.0 | 17.7 | 21.6
14 | 9.4 | 10.3 | 11.5 | 13.0 | 15.0 | 17.7
15 | 8.6 | 9.4 | 10.3 | 11.5 | 13.0 | 15.0
| | | | | |
16 | 7.9 | 8.6 | 9.4 | 10.3 | 11.5 | 13.0
17 | 7.3 | 7.9 | 8.6 | 9.4 | 10.3 | 11.5
18 | 6.8 | 7.3 | 7.9 | 8.6 | 9.4 | 10.3
19 | 6.4 | 6.8 | 7.3 | 7.9 | 8.6 | 9.4
20 | 6.0 | 6.4 | 6.8 | 7.3 | 7.9 | 8.6
| | | | | |
21 | 5.7 | 6.0 | 6.4 | 6.8 | 7.3 | 7.9
22 | 5.4 | 5.7 | 6.0 | 6.4 | 6.8 | 7.3
23 | 5.1 | 5.4 | 5.7 | 6.0 | 6.4 | 6.8
24 | 4.9 | 5.1 | 5.4 | 5.7 | 6.0 | 6.4
25 | 4.7 | 4.9 | 5.1 | 5.4 | 5.7 | 6.0
| | | | | |
26 | 4.5 | 4.7 | 4.9 | 5.1 | 5.4 | 5.7
27 | 4.3 | 4.5 | 4.7 | 4.9 | 5.1 | 5.4
28 | 4.1 | 4.3 | 4.5 | 4.7 | 4.9 | 5.1
29 | 3.9 | 4.1 | 4.3 | 4.5 | 4.7 | 4.9
30 | 3.8 | 3.9 | 4.1 | 4.3 | 4.5 | 4.7
===========================================================================
Table III. This table is calculated by inversion of the factors
in Table I, and is the most useful of all such tables, as it is
a direct calculation of the number of years that a given rate of
income on the investment must continue in order to amortize the
capital (the annual sinking fund being placed at compound interest
at 4%) and to repay various rates of interest on the investment. The
application of this method in testing the value of dividend-paying
shares is very helpful, especially in weighing the risks involved in
the portion of the purchase or investment unsecured by the profit
in sight. Given the annual percentage income on the investment from
the dividends of the mine (or on a non-producing mine assuming a
given rate of production and profit from the factors exposed), by
reference to the table the number of years can be seen in which
this percentage must continue in order to amortize the investment
and pay various rates of interest on it. As said before, the ore
in sight at a given rate of exhaustion can be reduced to terms of
life in sight. This certain period deducted from the total term
of years required gives the life which must be provided by further
discovery of ore, and this can be reduced to tons or feet of extension
of given ore-bodies and a tangible position arrived at. The test
can be applied in this manner to the various prices which must
be realized from the base metal in sight to warrant the price.
Taking the last example and assuming that the mine is equipped,
and that the price is $2,000,000, the yearly return on the price is
10%. If it is desired besides amortizing or redeeming the capital to
secure a return of 7% on the investment, it will be seen by reference
to the table that there will be required a life of 21.6 years. As the
life visible in the ore in sight is ten years, then the extensions
in depth must produce ore for 11.6 years longer--1,160,000 tons. If
the ore-body is 1,000 feet long and 13 feet wide, it will furnish
of gold ore 1,000 tons per foot of depth; hence the ore-body must
extend 1,160 feet deeper to justify the price. Mines are seldom so
simple a proposition as this example. There are usually probabilities
of other ore; and in the case of base metal, then variability of price
and other elements must be counted. However, once the extension
in depth which is necessary is determined for various assumptions
of metal value, there is something tangible to consider and to
weigh with the five geological weights set out in Chapter III.
The example given can be expanded to indicate not only the importance
of interest and redemption in the long extension in depth required,
but a matter discussed from another point of view under "Ratio of
Output." If the plant on this mine were doubled and the earnings
increased to 20% ($400,000 per annum) (disregarding the reduction
in working expenses that must follow expansion of equipment), it
will be found that the life required to repay the purchase
money,--$2,000,000,--and 7% interest upon it, is about 6.8 years.
As at this increased rate of production there is in the ore in
sight a life of five years, the extension in depth must be depended
upon for 1.8 years, or only 360,000 tons,--that is, 360 feet of
extension. Similarly, the present value of the ore in sight is
$268,000 greater if the mine be given double the equipment, for
thus the idle money locked in the ore is brought into the interest
market at an earlier date. Against this increased profit must be
weighed the increased cost of equipment. The value of low grade
mines, especially, is very much a factor of the volume of output
contemplated.
CHAPTER VI.
Mine Valuation (_Concluded_).
VALUATION OF MINES WITH LITTLE OR NO ORE IN SIGHT; VALUATIONS ON
SECOND-HAND DATA; GENERAL CONDUCT OF EXAMINATIONS; REPORTS.
A large number of examinations arise upon prospecting ventures
or partially developed mines where the value is almost wholly
prospective. The risks in such enterprises amount to the possible loss
of the whole investment, and the possible returns must consequently
be commensurate. Such business is therefore necessarily highly
speculative, but not unjustifiable, as the whole history of the
industry attests; but this makes the matter no easier for the mine
valuer. Many devices of financial procedure assist in the limitation
of the sum risked, and offer a middle course to the investor between
purchase of a wholly prospective value and the loss of a possible
opportunity to profit by it. The usual form is an option to buy the
property after a period which permits a certain amount of development
work by the purchaser before final decision as to purchase.
Aside from young mines such enterprises often arise from the possibility
of lateral extension of the ore-deposit outside the boundaries of
the property of original discovery (Fig. 3), in which cases there
is often no visible ore within the property under consideration
upon which to found opinion. In regions where vertical side lines
obtain, there is always the possibility of a "deep level" in inclined
deposits. Therefore the ground surrounding known deposits has a
certain speculative value, upon which engineers are often called to
pass judgment. Except in such unusual occurrences as South African
bankets, or Lake Superior coppers, prospecting for deep level of
extension is also a highly speculative phase of mining.
The whole basis of opinion in both classes of ventures must be
the few geological weights,--the geology of the property and the
district, the development of surrounding mines, etc. In any event,
there is a very great percentage of risk, and the profit to be gained
by success must be, proportionally to the expenditure involved,
very large. It is no case for calculating amortization and other
refinements. It is one where several hundreds or thousands of per
cent hoped for on the investment is the only justification.
OPINIONS AND VALUATIONS UPON SECOND-HAND DATA.
Some one may come forward and deprecate the bare suggestion of an
engineer's offering an opinion when he cannot have proper first-hand
data. But in these days we have to deal with conditions as well as
theories of professional ethics. The growing ownership of mines
by companies, that is by corporations composed of many individuals,
and with their stocks often dealt in on the public exchanges, has
resulted in holders whose interest is not large enough to warrant
their undertaking the cost of exhaustive examinations. The system
has produced an increasing class of mining speculators and investors
who are finding and supplying the enormous sums required to work
our mines,--sums beyond the reach of the old-class single-handed
mining men. Every year the mining investors of the new order are
coming more and more to the engineer for advice, and they should
be encouraged, because such counsel can be given within limits,
and these limits tend to place the industry upon a sounder footing
of ownership. As was said before, the lamb can be in a measure
protected. The engineer's interest is to protect him, so that the
industry which concerns his own life-work may be in honorable repute,
and that capital may be readily forthcoming for its expansion.
Moreover, by constant advice to the investor as to what constitutes
a properly presented and managed project, the arrangement of such
proper presentation and management will tend to become an _a priori_
function of the promoter.
Sometimes the engineer can make a short visit to the mine for data
purposes,--more often he cannot. In the former case, he can resolve
for himself an approximation upon all the factors bearing on value,
except the quality of the ore. For this, aside from inspection of
the ore itself, a look at the plans is usually enlightening. A
longitudinal section of the mine showing a continuous shortening of
the stopes with each succeeding level carries its own interpretation.
In the main, the current record of past production and estimates
of the management as to ore-reserves, etc., can be accepted in
ratio to the confidence that can be placed in the men who present
them. It then becomes a case of judgment of men and things, and
here no rule applies.
Advice must often be given upon data alone, without inspection
of the mine. Most mining data present internal evidence as to
credibility. The untrustworthy and inexperienced betray themselves
in their every written production. Assuming the reliability of data,
the methods already discussed for weighing the ultimate value of
the property can be applied. It would be possible to cite hundreds
of examples of valuation based upon second-hand data. Three will,
however, sufficiently illustrate. First, the R mine at Johannesburg.
With the regularity of this deposit, the development done, and
a study of the workings on the neighboring mines and in deeper
ground, it is a not unfair assumption that the reefs will maintain
size and value throughout the area. The management is sound, and
all the data are given in the best manner. The life of the mine
is estimated at six years, with some probabilities of further ore
from low-grade sections. The annual earnings available for dividends
are at the rate of about £450,000 per annum. The capital is £440,000
in £1 shares. By reference to the table on page 46 it will be seen
that the present value of £450,000 spread over six years to return
capital at the end of that period, and give 7% dividends in the
meantime, is 4.53 x £450,000 = £2,036,500 ÷ 440,000 = £4 12_s_.
7_d_. per share. So that this mine, on the assumption of continuity
of values, will pay about 7% and return the price. Seven per cent
is, however, not deemed an adequate return for the risks of labor
troubles, faults, dykes, or poor patches. On a 9% basis, the mine
is worth about £4 4_s_. per share.
Second, the G mine in Nevada. It has a capital of $10,000,000 in
$1 shares, standing in the market at 50 cents each. The reserves
are 250,000 tons, yielding a profit for yearly division of $7 per
ton. It has an annual capacity of about 100,000 tons, or $700,000
net profit, equal to 14% on the market value. In order to repay
the capital value of $5,000,000 and 8% per annum, it will need
a life of (Table III) 13 years, of which 2-1/2 are visible. The
size of the ore-bodies indicates a yield of about 1,100 tons per
foot of depth. At an exhaustion rate of 100,000 tons per annum,
the mine would need to extend to a depth of over a thousand feet
below the present bottom. There is always a possibility of finding
parallel bodies or larger volumes in depth, but it would be a sanguine
engineer indeed who would recommend the stock, even though it pays
an apparent 14%.
Third, the B mine, with a capital of $10,000,000 in 2,000,000 shares
of $5 each. The promoters state that the mine is in the slopes of
the Andes in Peru; that there are 6,000,000 tons of "ore blocked
out"; that two assays by the assayers of the Bank of England average
9% copper; that the copper can be produced at five cents per pound;
that there is thus a profit of $10,000,000 in sight. The evidences
are wholly incompetent. It is a gamble on statements of persons
who have not the remotest idea of sound mining.
GENERAL CONDUCT OF EXAMINATION.
Complete and exhaustive examination, entailing extensive sampling,
assaying, and metallurgical tests, is very expensive and requires
time. An unfavorable report usually means to the employer absolute
loss of the engineer's fee and expenses. It becomes then the initial
duty of the latter to determine at once, by the general conditions
surrounding the property, how far the expenditure for exhaustive
examination is warranted. There is usually named a money valuation
for the property, and thus a peg is afforded upon which to hang
conclusions. Very often collateral factors with a preliminary sampling,
or indeed no sampling at all, will determine the whole business.
In fact, it is becoming very common to send younger engineers to
report as to whether exhaustive examination by more expensive men
is justified.
In the course of such preliminary inspection, the ore-bodies may
prove to be too small to insure adequate yield on the price, even
assuming continuity in depth and represented value. They may be
so difficult to mine as to make costs prohibitive, or they may
show strong signs of "petering out." The ore may present visible
metallurgical difficulties which make it unprofitable in any event.
A gold ore may contain copper or arsenic, so as to debar cyanidation,
where this process is the only hope of sufficiently moderate costs.
A lead ore may be an amorphous compound with zinc, and successful
concentration or smelting without great penalties may be precluded.
A copper ore may carry a great excess of silica and be at the same
time unconcentratable, and there may be no base mineral supply
available for smelting mixture. The mine may be so small or so
isolated that the cost of equipment will never be justified. Some
of these conditions may be determined as unsurmountable, assuming
a given value for the ore, and may warrant the rejection of the
mine at the price set.
It is a disagreeable thing to have a disappointed promoter heap
vituperation on an engineer's head because he did not make an exhaustive
examination. Although it is generally desirable to do some sampling
to give assurance to both purchaser and vendor of conscientiousness,
a little courage of conviction, when this is rightly and adequately
grounded, usually brings its own reward.
Supposing, however, that conditions are right and that the mine is
worth the price, subject to confirmation of values, the determination
of these cannot be undertaken unless time and money are available
for the work. As was said, a sampling campaign is expensive, and
takes time, and no engineer has the moral right to undertake an
examination unless both facilities are afforded. Curtailment is
unjust, both to himself and to his employer.
How much time and outlay are required to properly sample a mine
is obviously a question of its size, and the character of its ore.
An engineer and one principal assistant can conduct two sampling
parties. In hard rock it may be impossible to take more than five
samples a day for each party. But, in average ore, ten samples for
each is reasonable work. As the number of samples is dependent
upon the footage of openings on the deposit, a rough approximation
can be made in advance, and a general idea obtained as to the time
required. This period must be insisted upon.
REPORTS.
Reports are to be read by the layman, and their first qualities
should be simplicity of terms and definiteness of conclusions.
Reports are usually too long, rather than too short. The essential
facts governing the value of a mine can be expressed on one sheet
of paper. It is always desirable, however, that the groundwork data
and the manner of their determination should be set out with such
detail that any other engineer could come to the same conclusion
if he accepted the facts as accurately determined. In regard to the
detailed form of reports, the writer's own preference is for a single
page summarizing the main factors, and an assay plan, reduced to a
longitudinal section where possible. Then there should be added,
for purposes of record and for submission to other engineers, a
set of appendices going into some details as to the history of
the mine, its geology, development, equipment, metallurgy, and
management. A list of samples should be given with their location,
and the tonnages and values of each separate block. A presentation
should be made of the probabilities of extension in depth, together
with recommendations for working the mine.
GENERAL SUMMARY.
The bed-rock value which attaches to a mine is the profit to be
won from proved ore and in which the price of metal is calculated
at some figure between "basic" and "normal." This we may call the
"_A_" value. Beyond this there is the speculative value of the
mine. If the value of the "probable" ore be represented by _X_,
the value of extension of the ore by _Y_, and a higher price for
metal than the price above assumed represented by _Z_, then if
the mine be efficiently managed the value of the mine is _A_ +
_X_ + _Y_ + _Z_. What actual amounts should be attached to _X,
Y, Z_ is a matter of judgment. There is no prescription for good
judgment. Good judgment rests upon a proper balancing of evidence.
The amount of risk in _X, Y, Z_ is purely a question of how much
these factors are required to represent in money,--in effect, how
much more ore must be found, or how many feet the ore must extend
in depth; or in convertible terms, what life in years the mine
must have, or how high the price of metal must be. In forming an
opinion whether these requirements will be realized, _X, Y, Z_
must be balanced in a scale whose measuring standards are the five
geological weights and the general industrial outlook. The wise
engineer will put before his clients the scale, the weights, and
the conclusion arrived at. The shrewd investor will require to
know these of his adviser.
CHAPTER VII.
Development of Mines.
ENTRY TO THE MINE; TUNNELS; VERTICAL, INCLINED, AND COMBINED SHAFTS;
LOCATION AND NUMBER OF SHAFTS.
Development is conducted for two purposes: first, to search for
ore; and second, to open avenues for its extraction. Although both
objects are always more or less in view, the first predominates
in the early life of mines, the prospecting stage, and the second
in its later life, the producing stage. It is proposed to discuss
development designed to embrace extended production purposes first,
because development during the prospecting stage is governed by
the same principles, but is tempered by the greater degree of
uncertainty as to the future of the mine, and is, therefore, of
a more temporary character.
ENTRY TO THE MINE.
There are four methods of entry: by tunnel, vertical shaft, inclined
shaft, or by a combination of the last two, that is, by a shaft
initially vertical then turned to an incline. Combined shafts are
largely a development of the past few years to meet "deep level"
conditions, and have been rendered possible only by skip-winding. The
angle in such shafts (Fig. 2) is now generally made on a parabolic
curve, and the speed of winding is then less diminished by the
bend.
The engineering problems which present themselves under "entry"
may be divided into those of:--
1. Method.
2. Location.
3. Shape and size.
The resolution of these questions depends upon the:--
a. Degree of dip of the deposit.
b. Output of ore to be provided for.
c. Depth at which the deposit is to be attacked.
d. Boundaries of the property.
e. Surface topography.
f. Cost.
g. Operating efficiency.
h. Prospects of the mine.
[Illustration: Fig. 2.--Showing arrangement of the bend in combined
shafts.]
From the point of view of entrance, the coöperation of a majority
of these factors permits the division of mines into certain broad
classes. The type of works demanded for moderate depths (say vertically
2,500 to 3,000 feet) is very different from that required for great
depths. To reach great depths, the size of shafts must greatly
expand, to provide for extended ventilation, pumping, and winding
necessities. Moreover inclined shafts of a degree of flatness possible
for moderate depths become too long to be used economically from
the surface. The vast majority of metal-mining shafts fall into
the first class, those of moderate depths. Yet, as time goes on
and ore-deposits are exhausted to lower planes, problems of depth
will become more common. One thing, however, cannot be too much
emphasized, especially on mines to be worked from the outcrop, and
that is, that no engineer is warranted, owing to the speculation
incidental to extension in depth, in initiating early in the mine's
career shafts of such size or equipment as would be available for
great depths. Moreover, the proper location of a shaft so as to
work economically extension of the ore-bodies is a matter of no
certainty, and therefore shafts of speculative mines are tentative
in any event.
Another line of division from an engineering view is brought about
by a combination of three of the factors mentioned. This is the
classification into "outcrop" and "deep-level" mines. The former
are those founded upon ore-deposits to be worked from or close
to the surface. The latter are mines based upon the extension in
depth of ore-bodies from outcrop mines. Such projects are not so
common in America, where the law in most districts gives the outcrop
owner the right to follow ore beyond his side-lines, as in countries
where the boundaries are vertical on all sides. They do, however,
arise not alone in the few American sections where the side-lines
are vertical boundaries, but in other parts owing to the pitch of
ore-bodies through the end lines (Fig. 3). More especially do such
problems arise in America in effect, where the ingress questions
have to be revised for mines worked out in the upper levels (Fig.
7).
[Illustration: Fig. 3.--Longitudinal section showing "deep level"
project arising from dip of ore-body through end-line.]
If from a standpoint of entrance questions, mines are first classified
into those whose works are contemplated for moderate depths, and those
in which work is contemplated for great depth, further clarity in
discussion can be gained by subdivision into the possible cases arising
out of the factors of location, dip, topography, and boundaries.
MINES OF MODERATE DEPTHS.
Case I. Deposits where topographic conditions permit the
alternatives of shaft or tunnel.
Case II. Vertical or horizontal deposits, the only practical
means of attaining which is by a vertical shaft.
Case III. Inclined deposits to be worked from near the surface.
There are in such instances the alternatives of either
a vertical or an inclined shaft.
Case IV. Inclined deposits which must be attacked in depth,
that is, deep-level projects. There are the alternatives
of a compound shaft or of a vertical shaft, and
in some cases of an incline from the surface.
MINES TO GREAT DEPTHS.
Case V. Vertical or horizontal deposits, the only way of reaching
which is by a vertical shaft.
Case VI. Inclined deposits. In such cases the alternatives are
a vertical or a compound shaft.
CASE I.--Although for logical arrangement tunnel entry has been
given first place, to save repetition it is proposed to consider
it later. With few exceptions, tunnels are a temporary expedient
in the mine, which must sooner or later be opened by a shaft.
CASE II. VERTICAL OR HORIZONTAL DEPOSITS.--These require no discussion
as to manner of entry. There is no justifiable alternative to a
vertical shaft (Fig. 4).
[Illustration: Fig. 4.--Cross-sections showing entry to vertical
or horizontal deposits. Case II.]
[Illustration: Fig. 5.--Cross-section showing alternative shafts
to inclined deposit to be worked from surface. Case III.]
CASE III. INCLINED DEPOSITS WHICH ARE INTENDED TO BE WORKED FROM
THE OUTCROP, OR FROM NEAR IT (Fig. 5).--The choice of inclined or
vertical shaft is dependent upon relative cost of construction,
subsequent operation, and the useful life of the shaft, and these
matters are largely governed by the degree of dip. Assuming a shaft
of the same size in either alternative, the comparative cost per
foot of sinking is dependent largely on the breaking facilities
of the rock under the different directions of attack. In this,
the angles of the bedding or joint planes to the direction of the
shaft outweigh other factors. The shaft which takes the greatest
advantage of such lines of breaking weakness will be the cheapest
per foot to sink. In South African experience, where inclined shafts
are sunk parallel to the bedding planes of hard quartzites, the cost
per foot appears to be in favor of the incline. On the other hand,
sinking shafts across tight schists seems to be more advantageous
than parallel to the bedding planes, and inclines following the
dip cost more per foot than vertical shafts.
An inclined shaft requires more footage to reach a given point
of depth, and therefore it would entail a greater total expense
than a vertical shaft, assuming they cost the same per foot. The
excess amount will be represented by the extra length, and this
will depend upon the flatness of the dip. With vertical shafts,
however, crosscuts to the deposit are necessary. In a comparative
view, therefore, the cost of the crosscuts must be included with
that of the vertical shaft, as they would be almost wholly saved
in an incline following near the ore.
The factor of useful life for the shaft enters in deciding as to
the advisability of vertical shafts on inclined deposits, from the
fact that at some depth one of two alternatives has to be chosen.
The vertical shaft, when it reaches a point below the deposit where
the crosscuts are too long (_C_, Fig. 5), either becomes useless,
or must be turned on an incline at the intersection with the ore
(_B_). The first alternative means ultimately a complete loss of
the shaft for working purposes. The latter has the disadvantage
that the bend interferes slightly with haulage.
The following table will indicate an hypothetical extreme case,--not
infrequently met. In it a vertical shaft 1,500 feet in depth is taken
as cutting the deposit at the depth of 750 feet, the most favored
position so far as aggregate length of crosscuts is concerned. The
cost of crosscutting is taken at $20 per foot and that of sinking
the vertical shaft at $75 per foot. The incline is assumed for two
cases at $75 and $100 per foot respectively. The stoping height
upon the ore between levels is counted at 125 feet.
Dip of | Depth of | Length of |No. of Crosscuts| Total Length
Deposit from | Vertical | Incline | Required from | of Crosscuts,
Horizontal | Shaft | Required | V Shaft | Feet
-------------|-------------|-------------|----------------|---------------
80° | 1,500 | 1,522 | 11 | 859
70° | 1,500 | 1,595 | 12 | 1,911
60° | 1,500 | 1,732 | 13 | 3,247
50° | 1,500 | 1,058 | 15 | 5,389
40° | 1,500 | 2,334 | 18 | 8,038
30° | 1,500 | 3,000 | 23 | 16,237
==========================================================================
Cost of |Cost Vertical| Total Cost | Cost of Incline|Cost of Incline
Crosscuts $20| Shaft $75 | of Vertical | $75 per Foot | $100 per Foot
per Foot | per Foot |and Crosscuts| |
-------------|-------------|-------------|----------------|---------------
$17,180 | $112,500 | $129,680 | $114,150 | $152,200
38,220 | 112,500 | 150,720 | 118,625 | 159,500
64,940 | 112,500 | 177,440 | 129,900 | 172,230
107,780 | 112,500 | 220,280 | 114,850 | 195,800
178,760 | 112,500 | 291,260 | 175,050 | 233,400
324,740 | 112,500 | 437,240 | 225,000 | 300,000
From the above examples it will be seen that the cost of crosscuts
put at ordinary level intervals rapidly outruns the extra expense
of increased length of inclines. If, however, the conditions are
such that crosscuts from a vertical shaft are not necessary at so
frequent intervals, then in proportion to the decrease the advantages
sway to the vertical shaft. Most situations wherein the crosscuts
can be avoided arise in mines worked out in the upper levels and
fall under Case IV, that of deep-level projects.
There can be no doubt that vertical shafts are cheaper to operate
than inclines: the length of haul from a given depth is less; much
higher rope speed is possible, and thus the haulage hours are less
for the same output; the wear and tear on ropes, tracks, or guides
is not so great, and pumping is more economical where the Cornish
order of pump is used. On the other hand, with a vertical shaft
must be included the cost of operating crosscuts. On mines where
the volume of ore does not warrant mechanical haulage, the cost of
tramming through the extra distance involved is an expense which
outweighs any extra operating outlay in the inclined shaft itself.
Even with mechanical haulage in crosscuts, it is doubtful if there
is anything in favor of the vertical shaft on this score.
[Illustration: Fig. 6.--Cross-section showing auxiliary vertical
outlet.]
In deposits of very flat dips, under 30°, the case arises where the
length of incline is so great that the saving on haulage through
direct lift warrants a vertical shaft as an auxiliary outlet in
addition to the incline (Fig. 6). In such a combination the crosscut
question is eliminated. The mine is worked above and below the
intersection by incline, and the vertical shaft becomes simply a
more economical exit and an alternative to secure increased output.
The North Star mine at Grass Valley is an illustration in point. Such
a positive instance borders again on Case IV, deep-level projects.
In conclusion, it is the writer's belief that where mines are to
be worked from near the surface, coincidentally with sinking, and
where, therefore, crosscuts from a vertical shaft would need to be
installed frequently, inclines are warranted in all dips under 75°
and over 30°. Beyond 75° the best alternative is often undeterminable.
In the range under 30° and over 15°, although inclines are primarily
necessary for actual delivery of ore from levels, they can often
be justifiably supplemented by a vertical shaft as a relief to a
long haul. In dips of less than 15°, as in those over 75°, the
advantages again trend strongly in favor of the vertical shaft. There
arise, however, in mountainous countries, topographic conditions
such as the dip of deposits into the mountain, which preclude any
alternative on an incline at any angled dip.
CASE IV. INCLINED DEPOSITS WHICH MUST BE ATTACKED IN DEPTH (Fig.
7).--There are two principal conditions in which such properties
exist: first, mines being operated, or having been previously worked,
whose method of entry must be revised; second, those whose ore-bodies
to be attacked do not outcrop within the property.
The first situation may occur in mines of inadequate shaft capacity
or wrong location; in mines abandoned and resurrected; in mines
where a vertical shaft has reached its limit of useful extensions,
having passed the place of economical crosscutting; or in mines in
flat deposits with inclines whose haul has become too long to be
economical. Three alternatives present themselves in such cases: a
new incline from the surface (_A B F_, Fig. 7), or a vertical shaft
combined with incline extension (_C D F_), or a simple vertical
shaft (_H G_). A comparison can be first made between the simple
incline and the combined shaft. The construction of an incline from
the surface to the ore-body will be more costly than a combined
shaft, for until the horizon of the ore is reached (at _D_) no
crosscuts are required in the vertical section, while the incline
must be of greater length to reach the same horizon. The case arises,
however, where inclines can be sunk through old stopes, and thus
more cheaply constructed than vertical shafts through solid rock;
and also the case of mountainous topographic conditions mentioned
above.
[Illustration: Fig. 7.--Cross-section of inclined deposit which
must be attacked in depth.]
From an operating point of view, the bend in combined shafts (at
_D_) gives rise to a good deal of wear and tear on ropes and gear.
The possible speed of winding through a combined shaft is, however,
greater than a simple incline, for although haulage speed through
the incline section (_D F_) and around the bend of the combined
shaft is about the same as throughout a simple incline (_A F_), the
speed can be accelerated in the vertical portion (_D C_) above that
feasible did the incline extend to the surface. There is therefore an
advantage in this regard in the combined shaft. The net advantages
of the combined over the inclined shaft depend on the comparative
length of the two alternative routes from the intersection (_D_)
to the surface. Certainly it is not advisable to sink a combined
shaft to cut a deposit at 300 feet in depth if a simple incline
can be had to the surface. On the other hand, a combined shaft
cutting the deposit at 1,000 feet will be more advisable than a
simple incline 2,000 feet long to reach the same point. The matter
is one for direct calculation in each special case. In general, there
are few instances of really deep-level projects where a complete
incline from the surface is warranted.
In most situations of this sort, and in all of the second type
(where the outcrop is outside the property), actual choice usually
lies between combined shafts (_C D F_) and entire vertical shafts (_H
G_). The difference between a combined shaft and a direct vertical
shaft can be reduced to a comparison of the combined shaft below
the point of intersection (_D_) with that portion of a vertical
shaft which would cover the same horizon. The question then becomes
identical with that of inclined _versus_ verticals, as stated in Case
III, with the offsetting disadvantage of the bend in the combined
shaft. If it is desired to reach production at the earliest date,
the lower section of a simple vertical shaft must have crosscuts
to reach the ore lying above the horizon of its intersection (_E_).
If production does not press, the ore above the intersection (_EB_)
can be worked by rises from the horizon of intersection (_E_).
In the use of rises, however, there follow the difficulties of
ventilation and lowering the ore down to the shaft, which brings
expenses to much the same thing as operating through crosscuts.
The advantages of combined over simple vertical shafts are earlier
production, saving of either rises or crosscuts, and the ultimate
utility of the shaft to any depth. The disadvantages are the cost
of the extra length of the inclined section, slower winding, and
greater wear and tear within the inclined section and especially
around the bend. All these factors are of variable import, depending
upon the dip. On very steep dips,--over 70°,--the net result is in
favor of the simple vertical shaft. On other dips it is in favor
of the combined shaft.
CASES V AND VI. MINES TO BE WORKED TO GREAT DEPTHS,--OVER 3,000
FEET.--In Case V, with vertical or horizontal deposits, there is
obviously no desirable alternative to vertical shafts.
In Case VI, with inclined deposits, there are the alternatives
of a combined or of a simple vertical shaft. A vertical shaft in
locations (_H_, Fig. 7) such as would not necessitate extension in
depth by an incline, would, as in Case IV, compel either crosscuts
to the ore or inclines up from the horizon of intersection (_E_).
Apart from delay in coming to production and the consequent loss of
interest on capital, the ventilation problems with this arrangement
would be appalling. Moreover, the combined shaft, entering the deposit
near its shallowest point, offers the possibility of a separate
haulage system on the inclined and on the vertical sections, and
such separate haulage is usually advisable at great depths. In
such instances, the output to be handled is large, for no mine of
small output is likely to be contemplated at such depth. Several
moderate-sized inclines from the horizon of intersection have been
suggested (_EF_, _DG_, _CH_, Fig. 8) to feed a large primary shaft
(_AB_), which thus becomes the trunk road. This program would cheapen
lateral haulage underground, as mechanical traction can be used
in the main level, (_EC_), and horizontal haulage costs can be
reduced on the lower levels. Moreover, separate winding engines
on the two sections increase the capacity, for the effect is that
of two trains instead of one running on a single track.
SHAFT LOCATION.--Although the prime purpose in locating a shaft
is obviously to gain access to the largest volume of ore within
the shortest haulage distance, other conditions also enter, such
as the character of the surface and the rock to be intersected,
the time involved before reaching production, and capital cost.
As shafts must bear two relations to a deposit,--one as to the
dip and the other as to the strike,--they may be considered from
these aspects. Vertical shafts must be on the hanging-wall side
of the outcrop if the deposit dips at all. In any event, the shaft
should be far enough away to be out of the reach of creeps. An
inclined shaft may be sunk either on the vein, in which case a
pillar of ore must be left to support the shaft; or, instead, it
may be sunk a short distance in the footwall, and where necessary
the excavation above can be supported by filling. Following the
ore has the advantage of prospecting in sinking, and in many cases
the softness of the ground in the region of the vein warrants this
procedure. It has, however, the disadvantage that a pillar of ore
is locked up until the shaft is ready for abandonment. Moreover, as
veins or lodes are seldom of even dip, an inclined shaft, to have
value as a prospecting opening, or to take advantage of breaking
possibilities in the lode, will usually be crooked, and an incline
irregular in detail adds greatly to the cost of winding and maintenance.
These twin disadvantages usually warrant a straight incline in the
footwall. Inclines are not necessarily of the same dip throughout,
but for reasonably economical haulage change of angle must take
place gradually.
[Illustration: Fig. 8.--Longitudinal section showing shaft arrangement
proposed for very deep inclined deposits.]
In the case of deep-level projects on inclined deposits, demanding
combined or vertical shafts, the first desideratum is to locate
the vertical section as far from the outcrop as possible, and thus
secure the most ore above the horizon of intersection. This, however,
as stated before, would involve the cost of crosscuts or rises and
would cause delay in production, together with the accumulation
of capital charges. How important the increment of interest on
capital may become during the period of opening the mine may be
demonstrated by a concrete case. For instance, the capital of a
company or the cost of the property is, say, $1,000,000, and where
opening the mine for production requires four years, the aggregate
sum of accumulated compound interest at 5% (and most operators
want more from a mining investment) would be $216,000. Under such
circumstances, if a year or two can be saved in getting to production
by entering the property at a higher horizon, the difference in
accumulated interest will more than repay the infinitesimal extra
cost of winding through a combined shaft of somewhat increased
length in the inclined section.
The unknown character of the ore in depth is always a sound reason
for reaching it as quickly and as cheaply as possible. In result,
such shafts are usually best located when the vertical section
enters the upper portion of the deposit.
The objective in location with regard to the strike of the ore-bodies
is obviously to have an equal length of lateral ore-haul in every
direction from the shaft. It is easier to specify than to achieve
this, for in all speculative deposits ore-shoots are found to pursue
curious vagaries as they go down. Ore-bodies do not reoccur with
the same locus as in the upper levels, and generally the chances
to go wrong are more numerous than those to go right.
NUMBER OF SHAFTS.--The problem of whether the mine is to be opened
by one or by two shafts of course influences location. In metal
mines under Cases II and III (outcrop properties) the ore output
requirements are seldom beyond the capacity of one shaft. Ventilation
and escape-ways are usually easily managed through the old stopes.
Under such circumstances, the conditions warranting a second shaft
are the length of underground haul and isolation of ore-bodies or
veins. Lateral haulage underground is necessarily disintegrated by
the various levels, and usually has to be done by hand. By shortening
this distance of tramming and by consolidation of the material
from all levels at the surface, where mechanical haulage can be
installed, a second shaft is often justified. There is therefore
an economic limitation to the radius of a single shaft, regardless
of the ability of the shaft to handle the total output.
Other questions also often arise which are of equal importance
to haulage costs. Separate ore-shoots or ore-bodies or parallel
deposits necessitate, if worked from one shaft, constant levels
through unpayable ground and extra haul as well, or ore-bodies may
dip away from the original shaft along the strike of the deposit
and a long haulage through dead levels must follow. For instance,
levels and crosscuts cost roughly one-quarter as much per foot as
shafts. Therefore four levels in barren ground, to reach a parallel
vein or isolated ore-body 1,000 feet away, would pay for a shaft
1,000 feet deep. At a depth of 1,000 feet, at least six levels
might be necessary. The tramming of ore by hand through such a
distance would cost about double the amount to hoist it through
a shaft and transport it mechanically to the dressing plant at
surface. The aggregate cost and operation of barren levels therefore
soon pays for a second shaft. If two or more shafts are in question,
they must obviously be set so as to best divide the work.
Under Cases IV, V, and VI,--that is, deep-level projects,--ventilation
and escape become most important considerations. Even where the
volume of ore is within the capacity of a single shaft, another
usually becomes a necessity for these reasons. Their location is
affected not only by the locus of the ore, but, as said, by the time
required to reach it. Where two shafts are to be sunk to inclined
deposits, it is usual to set one so as to intersect the deposit at
a lower point than the other. Production can be started from the
shallower, before the second is entirely ready. The ore above the
horizon of intersection of the deeper shaft is thus accessible from
the shallower shaft, and the difficulty of long rises or crosscuts
from that deepest shaft does not arise.
CHAPTER VIII.
Development of Mines (_Continued_).
SHAPE AND SIZE OF SHAFTS; SPEED OF SINKING; TUNNELS.
SHAPE OF SHAFTS.--Shafts may be round or rectangular.[*] Round
vertical shafts are largely applied to coal-mines, and some engineers
have advocated their usefulness to the mining of the metals under
discussion. Their great advantages lie in their structural strength,
in the large amount of free space for ventilation, and in the fact
that if walled with stone, brick, concrete, or steel, they can be
made water-tight so as to prevent inflow from water-bearing strata,
even when under great pressure. The round walled shafts have a longer
life than timbered shafts. All these advantages pertain much more to
mining coal or iron than metals, for unsound, wet ground is often
the accompaniment of coal-measures, and seldom troubles metal-mines.
Ventilation requirements are also much greater in coal-mines. From
a metal-miner's standpoint, round shafts are comparatively much
more expensive than the rectangular timbered type.[**] For a larger
area must be excavated for the same useful space, and if support
is needed, satisfactory walling, which of necessity must be brick,
stone, concrete, or steel, cannot be cheaply accomplished under
the conditions prevailing in most metal regions. Although such
shafts would have a longer life, the duration of timbered shafts
is sufficient for most metal mines. It follows that, as timber
is the cheapest and all things considered the most advantageous
means of shaft support for the comparatively temporary character
of metal mines, to get the strains applied to the timbers in the
best manner, and to use the minimum amount of it consistent with
security, and to lose the least working space, the shaft must be
constructed on rectangular lines.
[Footnote *: Octagonal shafts were sunk in Mexico in former times.
At each face of the octagon was a whim run by mules, and hauling
leather buckets.]
[Footnote **: The economic situation is rapidly arising in a number
of localities that steel beams can be usefully used instead of
timber. The same arguments apply to this type of support that apply
to timber.]
The variations in timbered shaft design arise from the possible
arrangement of compartments. Many combinations can be imagined,
of which Figures 9, 10, 11, 12, 13, and 14 are examples.
[Illustration: FIG. 9. FIG. 10. FIG. 11. FIG. 12. FIG. 13. FIG.
14.]
The arrangement of compartments shown in Figures 9, 10, 11, and
13 gives the greatest strength. It permits timbering to the best
advantage, and avoids the danger underground involved in crossing
one compartment to reach another. It is therefore generally adopted.
Any other arrangement would obviously be impossible in inclined
or combined shafts.
SIZE OF SHAFTS.--In considering the size of shafts to be installed,
many factors are involved. They are in the main:--
_a_. Amount of ore to be handled.
_b_. Winding plant.
_c_. Vehicle of transport.
_d_. Depth.
_e_. Number of men to be worked underground.
_f_. Amount of water.
_g_. Ventilation.
_h_. Character of the ground.
_i_. Capital outlay.
_j_. Operating expense.
It is not to be assumed that these factors have been stated in
the order of relative importance. More or less emphasis will be
attached to particular factors by different engineers, and under
different circumstances. It is not possible to suggest any arbitrary
standard for calculating their relative weight, and they are so
interdependent as to preclude separate discussion. The usual result
is a compromise between the demands of all.
Certain factors, however, dictate a minimum position, which may
be considered as a datum from which to start consideration.
_First_, a winding engine, in order to work with any economy, must
be balanced, that is, a descending empty skip or cage must assist
in pulling up a loaded one. Therefore, except in mines of very
small output, at least two compartments must be made for hoisting
purposes. Water has to be pumped from most mines, escape-ways are
necessary, together with room for wires and air-pipes, so that at
least one more compartment must be provided for these objects.
We have thus three compartments as a sound minimum for any shaft
where more than trivial output is required.
_Second_, there is a certain minimum size of shaft excavation below
which there is very little economy in actual rock-breaking.[*]
In too confined a space, holes cannot be placed to advantage for
the blast, men cannot get round expeditiously, and spoil cannot be
handled readily. The writer's own experience leads him to believe
that, in so far as rock-breaking is concerned, to sink a shaft
fourteen to sixteen feet long by six to seven feet wide outside
the timbers, is as cheap as to drive any smaller size within the
realm of consideration, and is more rapid. This size of excavation
permits of three compartments, each about four to five feet inside
the timbers.
[Footnote *: Notes on the cost of shafts in various regions which
have been personally collected show a remarkable decrease in the
cost per cubic foot of material excavated with increased size of
shaft. Variations in skill, in economic conditions, and in method
of accounting make data regarding different shafts of doubtful
value, but the following are of interest:--
In Australia, eight shafts between 10 and 11 feet long by 4 to
5 feet wide cost an average of $1.20 per cubic foot of material
excavated. Six shafts 13 to 14 feet long by 4 to 5 feet wide cost
an average of $0.95 per cubic foot; seven shafts 14 to 16 feet
long and 5 to 7 feet wide cost an average of $0.82 per cubic foot.
In South Africa, eleven shafts 18 to 19 feet long by 7 to 8 feet
wide cost an average of $0.82 per cubic foot; five shafts 21 to
25 feet long by 8 feet wide, cost $0.74; and seven shafts 28 feet
by 8 feet cost $0.60 per cubic foot.]
The cost of timber, it is true, is a factor of the size of shaft,
but the labor of timbering does not increase in the same ratio.
In any event, the cost of timber is only about 15% of the actual
shaft cost, even in localities of extremely high prices.
_Third_, three reasons are rapidly making the self-dumping skip
the almost universal shaft-vehicle, instead of the old cage for
cars. First, there is a great economy in labor for loading into
and discharging from a shaft; second, there is more rapid despatch
and discharge and therefore a larger number of possible trips;
third, shaft-haulage is then independent of delays in arrival of
cars at stations, while tramming can be done at any time and
shaft-haulage can be concentrated into certain hours. Cages to
carry mine cars and handle the same load as a skip must either
be big enough to take two cars, which compels a much larger shaft
than is necessary with skips, or they must be double-decked, which
renders loading arrangements underground costly to install and
expensive to work. For all these reasons, cages can be justified only
on metal mines of such small tonnage that time is no consideration
and where the saving of men is not to be effected. In compartments
of the minimum size mentioned above (four to five feet either way)
a skip with a capacity of from two to five tons can be installed,
although from two to three tons is the present rule. Lighter loads
than this involve more trips, and thus less hourly capacity, and,
on the other hand, heavier loads require more costly engines. This
matter is further discussed under "Haulage Appliances."
We have therefore as the economic minimum a shaft of three compartments
(Fig. 9), each four to five feet square. When the maximum tonnage
is wanted from such a shaft at the least operating cost, it should
be equipped with loading bins and skips.
The output capacity of shafts of this size and equipment will depend
in a major degree upon the engine employed, and in a less degree
upon the hauling depth. The reason why depth is a subsidiary factor
is that the rapidity with which a load can be drawn is not wholly a
factor of depth. The time consumed in hoisting is partially expended
in loading, in acceleration and retardation of the engine, and in
discharge of the load. These factors are constant for any depth,
and extra distance is therefore accomplished at full speed of the
engine.
Vertical shafts will, other things being equal, have greater capacity
than inclines, as winding will be much faster and length of haul less
for same depth. Since engines have, however, a great tractive ability
on inclines, by an increase in the size of skip it is usually possible
partially to equalize matters. Therefore the size of inclines for
the same output need not differ materially from vertical shafts.
The maximum capacity of a shaft whose equipment is of the character
and size given above, will, as stated, decrease somewhat with extension
in depth of the haulage horizon. At 500 feet, such a shaft if vertical
could produce 70 to 80 tons per hour comfortably with an engine
whose winding speed was 700 feet per minute. As men and material
other than ore have to be handled in and out of the mine, and
shaft-sinking has to be attended to, the winding engine cannot
be employed all the time on ore. Twelve hours of actual daily
ore-winding are all that can be expected without auxiliary help.
This represents a capacity from such a depth of 800 to 1,000 tons
per day. A similar shaft, under ordinary working conditions, with
an engine speed of 2,000 feet per minute, should from, say, 3,000
feet have a capacity of about 400 to 600 tons daily.
It is desirable to inquire at what stages the size of shaft should
logically be enlarged in order to attain greater capacity. A
considerable measure of increase can be obtained by relieving the
main hoisting engine of all or part of its collateral duties. Where
the pumping machinery is not elaborate, it is often possible to
get a small single winding compartment into the gangway without
materially increasing the size of the shaft if the haulage compartments
be made somewhat narrower (Fig. 10). Such a compartment would be
operated by an auxiliary engine for sinking, handling tools and
material, and assisting in handling men. If this arrangement can
be effected, the productive time of the main engine can be expanded
to about twenty hours with an addition of about two-thirds to the
output.
Where the exigencies of pump and gangway require more than two
and one-half feet of shaft length, the next stage of expansion
becomes four full-sized compartments (Fig. 11). By thus enlarging the
auxiliary winding space, some assistance may be given to ore-haulage
in case of necessity. The mine whose output demands such haulage
provisions can usually stand another foot of width to the shaft,
so that the dimensions come to about 21 feet to 22 feet by 7 feet
to 8 feet outside the timbers. Such a shaft, with three- to four-ton
skips and an appropriate engine, will handle up to 250 tons per
hour from a depth of 1,000 feet.
The next logical step in advance is the shaft of five compartments
with four full-sized haulage ways (Fig. 13), each of greater size
than in the above instance. In this case, the auxiliary engine
becomes a balanced one, and can be employed part of the time upon
ore-haulage. Such a shaft will be about 26 feet to 28 feet long
by 8 feet wide outside the timbers, when provision is made for
one gangway. The capacity of such shafts can be up to 4,000 tons a
day, depending on the depth and engine. When very large quantities
of water are to be dealt with and rod-driven pumps to be used,
two pumping compartments are sometimes necessary, but other forms
of pumps do not require more than one compartment,--an additional
reason for their use.
For depths greater than 3,000 feet, other factors come into play.
Ventilation questions become of more import. The mechanical problems
on engines and ropes become involved, and their sum-effect is to
demand much increased size and a greater number of compartments.
The shafts at Johannesburg intended as outlets for workings 5,000
feet deep are as much as 46 feet by 9 feet outside timbers.
It is not purposed to go into details as to sinking methods or
timbering. While important matters, they would unduly prolong this
discussion. Besides, a multitude of treatises exist on these subjects
and cover all the minutiæ of such work.
SPEED OF SINKING.--Mines may be divided into two cases,--those
being developed only, and those being operated as well as developed.
In the former, the entrance into production is usually dependent
upon the speed at which the shaft is sunk. Until the mine is earning
profits, there is a loss of interest on the capital involved, which,
in ninety-nine instances out of a hundred, warrants any reasonable
extra expenditure to induce more rapid progress. In the case of
mines in operation, the volume of ore available to treatment or
valuation is generally dependent to a great degree upon the rapidity
of the extension of workings in depth. It will be demonstrated
later that, both from a financial and a technical standpoint, the
maximum development is the right one and that unremitting extension
in depth is not only justifiable but necessary.
Speed under special conditions or over short periods has a more
romantic than practical interest, outside of its value as a stimulant
to emulation. The thing that counts is the speed which can be maintained
over the year. Rapidity of sinking depends mainly on:--
_a_. Whether the shaft is or is not in use for operating the
mine.
_b_. The breaking character of the rock.
_c_. The amount of water.
The delays incident to general carrying of ore and men are such that
the use of the main haulage engine for shaft-sinking is practically
impossible, except on mines with small tonnage output. Even with a
separate winch or auxiliary winding-engine, delays are unavoidable
in a working shaft, especially as it usually has more water to contend
with than one not in use for operating the mine. The writer's own
impression is that an average of 40 feet per month is the maximum
possibility for year in and out sinking under such conditions. In
fact, few going mines manage more than 400 feet a year. In cases
of clean shaft-sinking, where every energy is bent to speed, 150
feet per month have been averaged for many months. Special cases
have occurred where as much as 213 feet have been achieved in a
single month. With ordinary conditions, 1,200 feet in a year is
very good work. Rock awkward to break, and water especially, lowers
the rate of progress very materially. Further reference to speed
will be found in the chapter on "Drilling Methods."
TUNNEL ENTRY.--The alternative of entry to a mine by tunnel is
usually not a question of topography altogether, but, like everything
else in mining science, has to be tempered to meet the capital
available and the expenditure warranted by the value showing.
In the initial prospecting of a mine, tunnels are occasionally
overdone by prospectors. Often more would be proved by a few inclines.
As the pioneer has to rely upon his right arm for hoisting and
drainage, the tunnel offers great temptations, even when it is
long and gains but little depth. At a more advanced stage of
development, the saving of capital outlay on hoisting and pumping
equipment, at a time when capital is costly to secure, is often
sufficient justification for a tunnel entry. But at the stage where
the future working of ore below a tunnel-level must be contemplated,
other factors enter. For ore below tunnel-level a shaft becomes
necessary, and in cases where a tunnel enters a few hundred feet
below the outcrop the shaft should very often extend to the surface,
because internal shafts, winding from tunnel-level, require large
excavations to make room for the transfer of ore and for winding
gear. The latter must be operated by transmitted power, either
that of steam, water, electricity, or air. Where power has to be
generated on the mine, the saving by the use of direct steam, generated
at the winding gear, is very considerable. Moreover, the cost of
haulage through a shaft for the extra distance from tunnel-level
to the surface is often less than the cost of transferring the
ore and removing it through the tunnel. The load once on the
winding-engine, the consumption of power is small for the extra
distance, and the saving of labor is of consequence. On the other
hand, where drainage problems arise, they usually outweigh all
other considerations, for whatever the horizon entered by tunnel,
the distance from that level to the surface means a saving of
water-pumpage against so much head. The accumulation of such constant
expense justifies a proportioned capital outlay. In other words,
the saving of this extra pumping will annually redeem the cost of
a certain amount of tunnel, even though it be used for drainage
only.
In order to emphasize the rapidity with which such a saving of
constant expense will justify capital outlay, one may tabulate the
result of calculations showing the length of tunnel warranted with
various hypothetical factors of quantity of water and height of lift
eliminated from pumping. In these computations, power is taken at
the low rate of $60 per horsepower-year, the cost of tunneling at
an average figure of $20 per foot, and the time on the basis of
a ten-year life for the mine.
Feet of Tunnel Paid for in 10 Years with Under-mentioned Conditions.
=============================================================
Feet of | 100,000 | 200,000 | 300,000 | 500,000 |1,000,000
Water Lift | Gallons | Gallons | Gallons | Gallons | Gallons
Avoided |per Diem |per Diem |per Diem |per Diem |per Diem
-----------|---------|---------|---------|---------|---------
100 | 600 | 1,200 | 1,800 | 3,000 | 6,000
200 | 1,200 | 2,400 | 3,600 | 6,000 | 12,000
300 | 1,800 | 3,600 | 5,400 | 9,000 | 18,000
500 | 3,000 | 6,000 | 9,000 | 15,000 | 30,000
1,000 | 6,000 | 12,000 | 18,000 | 30,000 | 60,000
=============================================================
The size of tunnels where ore-extraction is involved depends upon
the daily tonnage output required, and the length of haul. The
smallest size that can be economically driven and managed is about
6-1/2 feet by 6 feet inside the timbers. Such a tunnel, with single
track for a length of 1,000 feet, with one turn-out, permits handling
up to 500 tons a day with men and animals. If the distance be longer
or the tonnage greater, a double track is required, which necessitates
a tunnel at least 8 feet wide by 6-1/2 feet to 7 feet high, inside
the timbers.
There are tunnel projects of a much more impressive order than those
designed to operate upper levels of mines; that is, long crosscut
tunnels designed to drain and operate mines at very considerable
depths, such as the Sutro tunnel at Virginia City. The advantage
of these tunnels is very great, especially for drainage, and they
must be constructed of large size and equipped with appliances
for mechanical haulage.
CHAPTER IX.
Development of Mines (_Concluded_).
SUBSIDIARY DEVELOPMENT;--STATIONS; CROSSCUTS; LEVELS; INTERVAL
BETWEEN LEVELS; PROTECTION OF LEVELS; WINZES AND RISES. DEVELOPMENT
IN THE PROSPECTING STAGE; DRILLING.
SUBSIDIARY DEVELOPMENT.
Stations, crosscuts, levels, winzes, and rises follow after the
initial entry. They are all expensive, and the least number that
will answer is the main desideratum.
STATIONS.--As stations are the outlets of the levels to the shaft,
their size and construction is a factor of the volume and character
of the work at the levels which they are to serve. If no timber
is to be handled, and little ore, and this on cages, the stations
need be no larger than a good sized crosscut. Where timber is to
be let down, they must be ten to fifteen feet higher than the floor
of the crosscut. Where loading into skips is to be provided for,
bins must be cut underneath and sufficient room be provided to
shift the mine cars comfortably. Such bins are built of from 50 to
500 tons' capacity in order to contain some reserve for hoisting
purposes, and in many cases separate bins must be provided on opposite
sides of the shaft for ore and waste. It is a strong argument in
favor of skips, that with this means of haulage storage capacity
at the stations is possible, and the hoisting may then go on
independently of trucking and, as said before, there are no idle
men at the stations.
[Illustration: Fig. 15.--Cross-section of station arrangement for
skip-haulage in vertical shaft.]
[Illustration: Fig. 16.--Cross-section of station arrangement for
skip-haulage in vertical shaft.]
It is always desirable to concentrate the haulage to the least
number of levels, for many reasons. Among them is that, where haulage
is confined to few levels, storage-bins are not required at every
station. Figures 15, 16, 17, and 18 illustrate various arrangements
of loading bins.
CROSSCUTS.--Crosscuts are for two purposes, for roadway connection
of levels to the shaft or to other levels, and for prospecting
purposes. The number of crosscuts for roadways can sometimes be
decreased by making the connections with the shaft at every second
or even every third level, thus not only saving in the construction
cost of crosscuts and stations, but also in the expenses of scattered
tramming. The matter becomes especially worth considering where
the quantity of ore that can thus be accumulated warrants mule
or mechanical haulage. This subject will be referred to later on.
[Illustration: Fig. 17.--Arrangement of loading chutes in vertical
shaft.]
On the second purpose of crosscuts,--that of prospecting,--one
observation merits emphasis. This is, that the tendency of ore-fissures
to be formed in parallels warrants more systematic crosscutting
into the country rock than is done in many mines.
[Illustration: Fig. 18.--Cross-section of station arrangement for
skip-haulage in inclined shaft.]
LEVELS.
The word "level" is another example of miners' adaptations in
nomenclature. Its use in the sense of tunnels driven in the direction
of the strike of the deposit has better, but less used, synonyms in
the words "drifts" or "drives." The term "level" is used by miners
in two senses, in that it is sometimes applied to all openings on one
horizon, crosscuts included. Levels are for three purposes,--for a
stoping base; for prospecting the deposit; and for roadways. As a
prospecting and a stoping base it is desirable that the level should
be driven on the deposit; as a roadway, that it should constitute
the shortest distance between two points and be in the soundest
ground. On narrow, erratic deposits the levels usually must serve
all three purposes at once; but in wider and more regular deposits
levels are often driven separately for roadways from the level
which forms the stoping base and prospecting datum.
There was a time when mines were worked by driving the level on ore
and enlarging it top and bottom as far as the ground would stand,
then driving the next level 15 to 20 feet below, and repeating the
operation. This interval gradually expanded, but for some reason
100 feet was for years assumed to be the proper distance between
levels. Scattered over every mining camp on earth are thousands
of mines opened on this empirical figure, without consideration
of the reasons for it or for any other distance.
The governing factors in determining the vertical interval between
levels are the following:--
_a_. The regularity of the deposit.
_b_. The effect of the method of excavation of winzes and rises.
_c_. The dip and the method of stoping.
REGULARITY OF THE DEPOSIT.--From a prospecting point of view the
more levels the better, and the interval therefore must be determined
somewhat by the character of the deposit. In erratic deposits there
is less risk of missing ore with frequent levels, but it does not
follow that every level need be a through roadway to the shaft or
even a stoping base. In such deposits, intermediate levels for
prospecting alone are better than complete levels, each a roadway.
Nor is it essential, even where frequent levels are required for
a stoping base, that each should be a main haulage outlet to the
shaft. In some mines every third level is used as a main roadway,
the ore being poured from the intermediate ones down to the haulage
line. Thus tramming and shaft work, as stated before, can be
concentrated.
EFFECT OF METHOD OF EXCAVATING WINZES AND RISES.--With hand drilling
and hoisting, winzes beyond a limited depth become very costly to
pull spoil out of, and rises too high become difficult to ventilate,
so that there is in such cases a limit to the interval desirable
between levels, but these difficulties largely disappear where
air-winches and air-drills are used.
THE DIP AND METHOD OF STOPING.--The method of stoping is largely
dependent upon the dip, and indirectly thus affects level intervals.
In dips under that at which material will "flow" in the stopes--about
45° to 50°--the interval is greatly dependent on the method of
stope-transport. Where ore is to be shoveled from stopes to the
roadway, the levels must be comparatively close together. Where
deposits are very flat, under 20°, and walls fairly sound, it is
often possible to use a sort of long wall system of stoping and to
lay tracks in the stopes with self-acting inclines to the levels.
In such instances, the interval can be expanded to 250 or even 400
feet. In dips between 20° and 45°, tracks are not often possible,
and either shoveling or "bumping troughs"[*] are the only help
to transport. With shoveling, intervals of 100 feet[**] are most
common, and with troughs the distance can be expanded up to 150
or 175 feet.
[Footnote *: Page 136.]
[Footnote **: Intervals given are measured on the dip.]
In dips of over 40° to 50°, depending on the smoothness of the foot
wall, the distance can again be increased, as stope-transport is
greatly simplified, since the stope materials fall out by gravity.
In timbered stopes, in dips over about 45°, intervals of 150 to
200 feet are possible. In filled stopes intervals of over 150 feet
present difficulties in the maintenance of ore-passes, for the wear
and tear of longer use often breaks the timbers. In shrinkage-stopes,
where no passes are to be maintained and few winzes put through, the
interval is sometimes raised to 250 feet. The subject is further
discussed under "Stoping."
Another factor bearing on level intervals is the needed insurance
of sufficient points of stoping attack to keep up a certain output.
This must particularly influence the manager whose mine has but
little ore in reserve.
[Illustration: Fig. 19.]
PROTECTION OF LEVELS.--Until recent years, timbering and occasional
walling was the only method for the support of the roof, and for
forming a platform for a stoping base. Where the rock requires no
support sublevels can be used as a stoping base, and timbering
for such purpose avoided altogether (Figs. 38, 39, 42). In such
cases the main roadway can then be driven on straight lines, either
in the walls or in the ore, and used entirely for haulage. The
subheading for a stoping base is driven far enough above or below
the roadway (depending on whether overhand or underhand stoping
is to be used) to leave a supporting pillar which is penetrated
by short passes for ore. In overhand stopes, the ore is broken
directly on the floor of an upper sublevel; and in underhand stopes,
broken directly from the bottom of the sublevel. The method entails
leaving a pillar of ore which can be recovered only with difficulty
in mines where stope-support is necessary. The question of its
adoption is then largely one of the comparative cost of timbering,
the extra cost of the sublevel, and the net value of the ore left.
In bad swelling veins, or badly crushing walls, where constant
repair to timbers would be necessary, the use of a sublevel is a
most useful alternative. It is especially useful with stopes to
be left open or worked by shrinkage-stoping methods.
If the haulage level, however, is to be the stoping base, some
protection to the roadway must be provided. There are three systems
in use,--by wood stulls or sets (Figs. 19, 30, 43), by dry-walling
with timber caps (Fig. 35), and in some localities by steel sets.
Stulls are put up in various ways, and, as their use entails the
least difficulty in taking the ore out from beneath the level,
they are much favored, but are applicable only in comparatively
narrow deposits.
WINZES AND RISES.
These two kinds of openings for connecting two horizons in a mine
differ only in their manner of construction. A winze is sunk underhand,
while a rise is put up overhand. When the connection between levels
is completed, a miner standing at the bottom usually refers to
the opening as a rise, and when he goes to the top he calls it
a winze. This confusion in terms makes it advisable to refer to
all such completed openings as winzes, regardless of how they are
constructed.
In actual work, even disregarding water, it costs on the average
about 30% less to raise than to sink such openings, for obviously
the spoil runs out or is assisted by gravity in one case, and in
the other has to be shoveled and hauled up. Moreover, it is easier
to follow the ore in a rise than in a winze. It usually happens,
however, that in order to gain time both things are done, and for
prospecting purposes sinking is necessary.
The number of winzes required depends upon the method of stoping
adopted, and is mentioned under "Stoping." After stoping, the number
necessary to be maintained open depends upon the necessities of
ventilation, of escape, and of passageways for material to be used
below. Where stopes are to be filled with waste, more winzes must
be kept open than when other methods are used, and these winzes
must be in sufficient alignment to permit the continuous flow of
material down past the various levels. In order that the winzes
should deliver timber and filling to the most advantageous points,
they should, in dipping ore-bodies, be as far as possible on the
hanging wall side.
DEVELOPMENT IN THE EARLY PROSPECTING STAGE.
The prime objects in the prospecting stage are to expose the ore
and to learn regarding the ore-bodies something of their size, their
value, metallurgical character, location, dip, strike, etc.,--so much
at least as may be necessary to determine the works most suitable
for their extraction or values warranting purchase. In outcrop mines
there is one rule, and that is "follow the ore." Small temporary
inclines following the deposit, even though they are eventually
useless; are nine times out of ten justified.
In prospecting deep-level projects, it is usually necessary to
layout work which can be subsequently used in operating the mine,
because the depth involves works of such considerable scale, even
for prospecting, that the initial outlay does not warrant any
anticipation of revision. Such works have to be located and designed
after a study of the general geology as disclosed in adjoining mines.
Practically the only method of supplementing such information is
by the use of churn- and diamond-drills.
DRILLING.--Churn-drills are applicable only to comparatively shallow
deposits of large volume. They have an advantage over the diamond
drill in exposing a larger section and in their application to
loose material; but inability to determine the exact horizon of
the spoil does not lend them to narrow deposits, and in any event
results are likely to be misleading from the finely ground state of
the spoil. They are, however, of very great value for preliminary
prospecting to shallow horizons.
Two facts in diamond-drilling have to be borne in mind: the indication
of values is liable to be misleading, and the deflection of the drill
is likely to carry it far away from its anticipated destination.
A diamond-drill secures a small section which is sufficiently large
to reveal the geology, but the values disclosed in metal mines must
be accepted with reservations. The core amounts to but a little
sample out of possibly large amounts of ore, which is always of
variable character, and the core is most unlikely to represent
the average of the deposit. Two diamond-drill holes on the Oroya
Brownhill mine both passed through the ore-body. One apparently
disclosed unpayable values, the other seemingly showed ore forty
feet in width assaying $80 per ton. Neither was right. On the other
hand, the predetermination of the location of the ore-body justified
expenditure. A recent experiment at Johannesburg of placing a copper
wedge in the hole at a point above the ore-body and deflecting
the drill on reintroducing it, was successful in giving a second
section of the ore at small expense.
The deflection of diamond-drill holes from the starting angle is
almost universal. It often amounts to a considerable wandering
from the intended course. The amount of such deflection varies
with no seeming rule, but it is probable that it is especially
affected by the angle at which stratification or lamination planes
are inclined to the direction of the hole. A hole has been known
to wander in a depth of 1,500 feet more than 500 feet from the
point intended. Various instruments have been devised for surveying
deep holes, and they should be brought into use before works are
laid out on the basis of diamond-drill results, although none of
the inventions are entirely satisfactory.
CHAPTER X.
Stoping.
METHODS OF ORE-BREAKING; UNDERHAND STOPES; OVERHAND STOPES; COMBINED
STOPE. VALUING ORE IN COURSE OF BREAKING.
There is a great deal of confusion in the application of the word
"stoping." It is used not only specifically to mean the actual
ore-breaking, but also in a general sense to indicate all the operations
of ore-breaking, support of excavations, and transportation between
levels. It is used further as a noun to designate the hole left
when the ore is taken out. Worse still, it is impossible to adhere
to miners' terms without employing it in every sense, trusting
to luck and the context to make the meaning clear.
The conditions which govern the method of stoping are in the main:--
_a_. The dip.
_b_. The width of the deposit.
_c_. The character of the walls.
_d_. The cost of materials.
_e_. The character of the ore.
Every mine, and sometimes every stope in a mine, is a problem special
to itself. Any general consideration must therefore be simply an
inquiry into the broad principles which govern the adaptability of
special methods. A logical arrangement of discussion is difficult,
if not wholly impossible, because the factors are partially
interdependent and of varying importance.
For discussion the subject may be divided into:
1. Methods of ore-breaking.
2. Methods of supporting excavation.
3. Methods of transport in stopes.
METHODS OF ORE-BREAKING.
The manner of actual ore-breaking is to drill and blast off slices
from the block of ground under attack. As rock obviously breaks
easiest when two sides are free, that is, when corners can be broken
off, the detail of management for blasts is therefore to set the holes
so as to preserve a corner for the next cut; and as a consequence
the face of the stope shapes into a series of benches (Fig.
22),--inverted benches in the case of overhand stopes (Figs. 20,
21). The size of these benches will in a large measure depend on
the depth of the holes. In wide stopes with machine-drills they
vary from 7 to 10 feet; in narrow stopes with hand-holes, from
two to three feet.
[Illustration: Fig. 20.]
The position of the men in relation to the working face gives rise
to the usual primary classification of the methods of stoping.
They are:--
1. Underhand stopes,
2. Overhand stopes,
3. Combined stopes.
These terms originated from the direction of the drill-holes, but
this is no longer a logical basis of distinction, for underhand
holes in overhand stopes,--as in rill-stoping,--are used entirely
in some mines (Fig. 21).
[Illustration: Fig. 21.]
UNDERHAND STOPES.--Underhand stopes are those in which the ore
is broken downward from the levels. Inasmuch as this method has
the advantage of allowing the miner to strike his blows downward
and to stand upon the ore when at work, it was almost universal
before the invention of powder; and was applied more generally
before the invention of machine-drills than since. It is never
rightly introduced unless the stope is worked back from winzes
through which the ore broken can be let down to the level below,
as shown in Figures 22 and 23.
[Illustration: Fig. 22.]
This system can be advantageously applied only in the rare cases
in which the walls require little or no support, and where very
little or no waste requiring separation is broken with the ore
in the stopes. To support the walls in bad ground in underhand
stopes would be far more costly than with overhand stopes, for
square-set timbering would be most difficult to introduce, and
to support the walls with waste and stulls would be even more
troublesome. Any waste broken must needs be thrown up to the level
above or be stored upon specially built stages--again a costly
proceeding.
A further drawback lies in the fact that the broken ore follows
down the face of the stope, and must be shoveled off each bench.
It thus all arrives at a single point,--the winze,--and must be
drawn from a single ore-pass into the level. This usually results
not only in more shoveling but in a congestion at the passes not
present in overhand stoping, for with that method several chutes
are available for discharging ore into the levels. Where the walls
require no support and no selection is desired in the stopes, the
advantage of the men standing on the solid ore to work, and of
having all down holes and therefore drilled wet, gives this method
a distinct place. In using this system, in order to protect the
men, a pillar is often left under the level by driving a sublevel,
the pillar being easily recoverable later. The method of sublevels
is of advantage largely in avoiding the timbering of levels.
[Illustration: Fig. 23.--Longitudinal section of an underhand stope.]
OVERHAND STOPES.--By far the greatest bulk of ore is broken overhand,
that is broken upward from one level to the next above. There are
two general forms which such stopes are given,--"horizontal" and
"rill."
[Illustration: Fig. 24.--Horizontal-cut overhand stope--longitudinal
section.]
The horizontal "flat-back" or "long-wall" stope, as it is variously
called, shown in Figure 24, is operated by breaking the ore in slices
parallel with the levels. In rill-stoping the ore is cut back from
the winzes in such a way that a pyramid-shaped room is created,
with its apex in the winze and its base at the level (Figs. 25 and
26). Horizontal or flat-backed stopes can be applied to almost any
dip, while "rill-stoping" finds its most advantageous application
where the dip is such that the ore will "run," or where it can be
made to "run" with a little help. The particular application of
the two systems is dependent not only on the dip but on the method
of supporting the excavation and the ore. With rill-stoping, it is
possible to cut the breaking benches back horizontally from the
winzes (Fig. 25), or to stagger the cuts in such a manner as to
take the slices in a descending angle (Figs. 21 and 26).
[Illustration: Fig. 25.--Rill-cut overhand stope--longitudinal section.]
In the "rill" method of incline cuts, all the drill-holes are "down"
holes (Fig. 21), and can be drilled wet, while in horizontal cuts
or flat-backed stopes, at least part of the holes must be "uppers"
(Fig. 20). Aside from the easier and cheaper drilling and setting
up of machines with this kind of "cut," there is no drill dust,--a
great desideratum in these days of miners' phthisis. A further
advantage in the "rill" cut arises in cases where horizontal jointing
planes run through the ore of a sort from which unduly large masses
break away in "flat-back" stopes. By the descending cut of the
"rill" method these calamities can be in a measure avoided. In
cases of dips over 40º the greatest advantage in "rill" stoping
arises from the possibility of pouring filling or timber into the
stope from above with less handling, because the ore and material
will run down the sides of the pyramid (Figs. 32 and 34). Thus
not only is there less shoveling required, but fewer ore-passes
and a less number of preliminary winzes are necessary, and a wider
level interval is possible. This matter will be gone into more
fully later.
[Illustration: Fig. 26.--Rill-cut overhand stope-longitudinal section.]
COMBINED STOPES.--A combined stope is made by the coincident working
of the underhand and "rill" method (Fig. 27). This order of stope
has the same limitations in general as the underhand kind. For
flat veins with strong walls, it has a great superiority in that
the stope is carried back more or less parallel with the winzes,
and thus broken ore after blasting lies in a line on the gradient
of the stope. It is, therefore, conveniently placed for mechanical
stope haulage. A further advantage is gained in that winzes may
be placed long distances apart, and that men are not required,
either when at work or passing to and from it, to be ever far from
the face, and they are thus in the safest ground, so that timber
and filling protection which may be otherwise necessary is not
required. This method is largely used in South Africa.
[Illustration: Fig. 27.--Longitudinal section of a combined stope.]
MINIMUM WIDTH OF STOPES.--The minimum stoping width which can be
consistently broken with hand-holes is about 30 inches, and this
only where there is considerable dip to the ore. This space is
so narrow that it is of doubtful advantage in any case, and 40
inches is more common in narrow mines, especially where worked
with white men. Where machine-drills are used about 4 feet is the
minimum width feasible.
RESUING.--In very narrow veins where a certain amount of wall-rock
must be broken to give working space, it pays under some circumstances
to advance the stope into the wall-rock ahead of the ore, thus
stripping the ore and enabling it to be broken separately. This
permits of cleaner selection of the ore; but it is a problem to
be worked out in each case, as to whether rough sorting of some
waste in the stopes, or further sorting at surface with inevitable
treatment of some waste rock, is more economical than separate
stoping cuts and inevitably wider stopes.
VALUING ORE IN COURSE OF BREAKING.--There are many ores whose payability
can be determined by inspection, but there are many of which it cannot.
Continuous assaying is in the latter cases absolutely necessary
to avoid the treatment of valueless material. In such instances,
sampling after each stoping-cut is essential, the unprofitable ore
being broken down and used as waste. Where values fade into the
walls, as in impregnation deposits, the width of stopes depends
upon the limit of payability. In these cases, drill-holes are put
into the walls and the drillings assayed. If the ore is found
profitable, the holes are blasted out. The gauge of what is profitable
in such situations is not dependent simply upon the average total
working costs of the mine, for ore in that position can be said to
cost nothing for development work and administration; moreover,
it is usually more cheaply broken than the average breaking cost,
men and machines being already on the spot.
CHAPTER XI.
Methods of Supporting Excavation.
TIMBERING; FILLING WITH WASTE; FILLING WITH BROKEN ORE; PILLARS
OF ORE; ARTIFICIAL PILLARS; CAVING SYSTEM.
Most stopes require support to be given to the walls and often to
the ore itself. Where they do require support there are five principal
methods of accomplishing it. The application of any particular method
depends upon the dip, width of ore-body, character of the ore and
walls, and cost of materials. The various systems are by:--
1. Timbering.
2. Filling with waste.
3. Filling with broken ore subsequently withdrawn.
4. Pillars of ore.
5. Artificial pillars built of timbers and waste.
6. Caving.
TIMBERING.--At one time timbering was the almost universal means of
support in such excavations, but gradually various methods for the
economical application of waste and ore itself have come forward,
until timbering is fast becoming a secondary device. Aside from
economy in working without it, the dangers of creeps, or crushing,
and of fires are sufficient incentives to do away with wood as
far as possible.
There are three principal systems of timber support to excavations,--by
stulls, square-sets, and cribs.
Stulls are serviceable only where the deposit is so narrow that
the opening can be bridged by single timbers between wall and wall
(Figs. 28 and 43). This system can be applied to any dip and is most
useful in narrow deposits where the walls are not too heavy. Stulls
in inclined deposits are usually set at a slightly higher angle than
that perpendicular to the walls, in order that the vertical pressure
of the hanging wall will serve to tighten them in position. The
"stull" system can, in inclined deposits, be further strengthened by
building waste pillars against them, in which case the arrangement
merges into the system of artificial pillars.
[Illustration: Fig. 28.--Longitudinal section of stull-supported
stope.]
[Illustration: Fig. 29.--Longitudinal section showing square-set
timbering.]
[Illustration: Fig. 30.--Square-set timbering on inclined ore-body.
Showing ultimate strain on timbers.]
Square-sets (Figs. 29 and 30), that is, trusses built in the opening
as the ore is removed, are applicable to almost any dip or width
of ore, but generally are applied only in deposits too wide, or to
rock too heavy, for stulls. Such trusses are usually constructed on
vertical and horizontal lines, and while during actual ore-breaking
the strains are partially vertical, ultimately, however, when the
weight of the walls begins to be felt, these strains, except in
vertical deposits, come at an angle to lines of strength in the
trusses, and therefore timber constructions of this type present
little ultimate resistance (Fig. 30). Square-set timbers are sometimes
set to present the maximum resistance to the direction of strain,
but the difficulties of placing them in position and variations in
the direction of strain on various parts of the stope do not often
commend the method. As a general rule square-sets on horizontal
lines answer well enough for the period of actual ore-breaking. The
crushing or creeps is usually some time later; and if the crushing
may damage the whole mine, their use is fraught with danger.
Reënforcement by building in waste is often resorted to. When done
fully, it is difficult to see the utility of the enclosed timber,
for entire waste-filling would in most cases be cheaper and equally
efficient.
[Illustration: Fig. 31.--"Cribs."]
There is always, with wood constructions, as said before, the very
pertinent danger of subsequent crushing and of subsidence in after
years, and the great risk of fires. Both these disasters have cost
Comstock and Broken Hill mines, directly or indirectly, millions of
dollars, and the outlay on timber and repairs one way or another
would have paid for the filling system ten times over. There are
cases where, by virtue of the cheapness of timber, "square-setting"
is the most economical method. Again, there are instances where the
ore lies in such a manner--particularly in limestone replacements--as
to preclude other means of support. These cases are being yearly
more and more evaded by the ingenuity of engineers in charge. The
author believes it soon will be recognized that the situation is
rare indeed where complete square-setting is necessarily without an
economical alternative. An objection is sometimes raised to filling
in favor of timber, in that if it become desirable to restope the
walls for low-grade ore left behind, such stopes could only be
entered by drawing the filling, with consequent danger of total
collapse. Such a contingency can be provided for in large ore-bodies
by installing an outer shell of sets of timber around the periphery
of the stope and filling the inside with waste. If the crushing
possibilities are too great for this method then, the subsequent
recovery of ore is hopeless in any event. In narrow ore-bodies
with crushing walls recovery of ore once left behind is not often
possible.
The third sort of timber constructions are cribs, a "log-house" sort
of structure usually filled with waste, and more fully discussed
under artificial pillars (Fig. 31). The further comparative merits
of timbering with other methods will be analyzed as the different
systems are described.
FILLING WITH WASTE.--The system of filling stope-excavations completely
with waste in alternating progress with ore-breaking is of wide
and increasingly general application (Figs. 32, 33, 34, 35).
Although a certain amount of waste is ordinarily available in the
stopes themselves, or from development work in the mine, such a
supply must usually be supplemented from other directions. Treatment
residues afford the easiest and cheapest handled material. Quarried
rock ranks next, and in default of any other easy supply, materials
from crosscuts driven into the stope-walls are sometimes resorted
to.
In working the system to the best advantage, the winzes through
the block of ore under attack are kept in alignment with similar
openings above, in order that filling may be poured through the
mine from the surface or any intermediate point. Winzes to be used
for filling should be put on the hanging-wall side of the area to
be filled, for the filling poured down will then reach the foot-wall
side of the stopes with a minimum of handling. In some instances,
one special winze is arranged for passing all filling from the
surface to a level above the principal stoping operations; and
it is then distributed along the levels into the winzes, and thus
to the operating stopes, by belt-conveyors.
[Illustration: Fig. 32.--Longitudinal section. Rill stope filled
with waste.]
[Illustration: Fig. 33.--Longitudinal section. Horizontal stope
filled with waste.]
[Illustration: Fig. 34.--Longitudinal section. Waste-filled stope
with dry-walling of levels and passes.]
In this system of stope support the ore is broken at intervals
alternating with filling. If there is danger of much loss from
mixing broken ore and filling, "sollars" of boards or poles are
laid on the waste. If the ore is very rich, old canvas or cowhides
are sometimes put under the boards. Before the filling interval,
the ore passes are built close to the face above previous filling
and their tops covered temporarily to prevent their being filled
with running waste. If the walls are bad, the filling is kept close
to the face. If the unbroken ore requires support, short stulls
set on the waste (as in Fig. 39) are usually sufficient until the
next cut is taken off, when the timber can be recovered. If stulls
are insufficient, cribs or bulkheads (Fig. 31) are also used and
often buried in the filling.
[Illustration: Fig. 35.--Cross-section of Fig. 34 on line _A-B_.]
Both flat-backed and rill-stope methods of breaking are employed in
conjunction with filled stopes. The advantages of the rill-stopes
are so patent as to make it difficult to understand why they are
not universally adopted when the dip permits their use at all. In
rill-stopes (Figs. 32 and 34) the waste flows to its destination
with a minimum of handling. Winzes and ore-passes are not required
with the same frequency as in horizontal breaking, and the broken
ore always lies on the slope towards the passes and is therefore
also easier to shovel. In flat-backed stopes (Fig. 33) winzes must
be put in every 50 feet or so, while in rill-stopes they can be
double this distance apart. The system is applicable by modification
to almost any width of ore. It finds its most economical field
where the dip of the stope floor is over 45°, when waste and ore,
with the help of the "rill," will flow to their destination. For
dips from under about 45° to about 30° or 35°, where the waste
and ore will not "flow" easily, shoveling can be helped by the
use of the "rill" system and often evaded altogether, if flow be
assisted by a sheet-iron trough described in the discussion of
stope transport. Further saving in shoveling can be gained in this
method, by giving a steeper pitch to the filling winzes and to the
ore-passes, by starting them from crosscuts in the wall, and by
carrying them at greater angles than the pitch of the ore (Fig.
36). These artifices combined have worked out most economically
on several mines within the writer's experience, with the dip as
flat as 30°. For very flat dips, where filling is to be employed,
rill-stoping has no advantage over flat-backed cuts, and in such
cases it is often advisable to assist stope transport by temporary
tracks and cars which obviously could not be worked on the tortuous
contour of a rill-stope, so that for dips under 30° advantage lies
with "flat-backed" ore-breaking.
[Illustration: Fig. 36.--Cross-section showing method of steepening
winzes and ore passes.]
On very wide ore-bodies where the support of the standing ore itself
becomes a great problem, the filling system can be applied by combining
it with square-setting. In this case the stopes are carried in
panels laid out transversally to the strike as wide as the standing
strength of the ore permits. On both sides of each panel a fence
of lagged square-sets is carried up and the area between is filled
with waste. The panels are stoped out alternately. The application
of this method at Broken Hill will be described later. (See pages
120 and Figs. 41 and 42.) The same type of wide ore-body can be
managed also on the filling system by the use of frequent "bulkheads"
to support the ore (Fig. 31).
Compared with timbering methods, filling has the great advantage
of more effective support to the mine, less danger of creeps, and
absolute freedom from the peril of fire. The relative expense of
the two systems is determined by the cost of materials and labor.
Two extreme cases illustrate the result of these economic factors
with sufficient clearness. It is stated that the cost of timbering
stopes on the Le Roi Mine by square-sets is about 21 cents per
ton of ore excavated. In the Ivanhoe mine of West Australia the
cost of filling stopes with tailings is about 22 cents per ton
of ore excavated. At the former mine the average cost of timber
is under $10 per M board-measure, while at the latter its price
would be $50 per M board-measure; although labor is about of the
same efficiency and wage, the cost in the Ivanhoe by square-setting
would be about 65 cents per ton of ore broken. In the Le Roi, on the
other hand, no residues are available for filling. To quarry rock
or drive crosscuts into the walls might make this system cost 65
cents per ton of ore broken if applied to that mine. The comparative
value of the filling method with other systems will be discussed
later.
FILLING WITH BROKEN ORE SUBSEQUENTLY WITHDRAWN.--This order of support
is called by various names, the favorite being "shrinkage-stoping."
The method is to break the ore on to the roof of the level, and by
thus filling the stope with broken ore, provide temporary support
to the walls and furnish standing floor upon which to work in making
the next cut (Figs. 37, 38, and 39.) As broken material occupies 30
to 40% more space than rock _in situ_, in order to provide working
space at the face, the broken ore must be drawn from along the level
after each cut. When the area attacked is completely broken through
from level to level, the stope will be full of loose broken ore,
which is then entirely drawn off.
A block to be attacked by this method requires preliminary winzes
only at the extremities of the stope,--for entry and for ventilation.
Where it is desired to maintain the winzes after stoping, they
must either be strongly timbered and lagged on the stope side,
be driven in the walls, or be protected by a pillar of ore (Fig.
37). The settling ore and the crushing after the stope is empty
make it difficult to maintain timbered winzes.
[Illustration: Fig. 37.--Longitudinal section of stope filled with
broken ore.]
Where it can be done without danger to the mine, the empty stopes
are allowed to cave. If such crushing would be dangerous, either
the walls must be held up by pillars of unbroken ore, as in the
Alaska Treadwell, where large "rib" pillars are left, or the open
spaces must be filled with waste. Filling the empty stope is usually
done by opening frequent passes along the base of the filled stope
above, and allowing the material of the upper stope to flood the
lower one. This program continued upwards through the mine allows
the whole filling of the mine to descend gradually and thus requires
replenishment only into the top. The old stopes in the less critical
and usually exhausted territory nearer the surface are sometimes
left without replenishing their filling.
The weight of broken ore standing at such a high angle as to settle
rapidly is very considerable upon the level; moreover, at the moment
when the stope is entirely drawn off, the pressure of the walls
as well is likely to be very great. The roadways in this system
therefore require more than usual protection. Three methods are
used: (_a_) timbering; (_b_) driving a sublevel in the ore above
the main roadway as a stoping-base, thus leaving a pillar of ore
over the roadway (Fig. 39); (_c_) by dry-walling the levels, as in
the Baltic mine, Michigan (Figs. 34 and 35). By the use of sublevels
the main roadways are sometimes driven in the walls (Fig. 38) and in
many cases all timbering is saved. To recover pillars left below
sublevels is a rather difficult task, especially if the old stope
above is caved or filled. The use of pillars in substitution for
timber, if the pillars are to be lost, is simply a matter of economics
as to whether the lost ore would repay the cost of other devices.
[Illustration: Fig. 38.--Cross-section of "shrinkage" stope.]
Frequent ore-chutes through the level timbers, or from the sublevels,
are necessary to prevent lodgment of broken ore between such passes,
because it is usually too dangerous for men to enter the emptying
stope to shovel out the lodged remnants. Where the ore-body is
wide, and in order that there may be no lodgment of ore, the timbers
over the level are set so as to form a trough along the level;
or where pillars are left, they are made "A"-shaped between the
chutes, as indicated in Figure 37.
[Illustration: Fig. 39.--Cross-section of "shrinkage" stope.]
The method of breaking the ore in conjunction with this means of
support in comparatively narrow deposits can be on the rill, in order
to have the advantage of down holes. Usually, however, flat-back
or horizontal cuts are desirable, as in such an arrangement it
is less troublesome to regulate the drawing of the ore so as to
provide proper head room. Where stopes are wide, ore is sometimes
cut arch-shaped from wall to wall to assure its standing. Where
this method of support is not of avail, short, sharply tapering
stulls are put in from the broken ore to the face (Fig. 39). When
the cut above these stulls is taken out, they are pulled up and
are used again.
This method of stoping is only applicable when:--
1. The deposit dips over 60°, and thus broken material will freely
settle downward to be drawn off from the bottom.
2. The ore is consistently payable in character. No selection can be
done in breaking, as all material broken must be drawn off together.
3. The hanging wall is strong, and will not crush or spall off waste
into the ore.
4. The ore-body is regular in size, else loose ore will lodge on
the foot wall. Stopes opened in this manner when partially empty
are too dangerous for men to enter for shoveling out remnants.
The advantages of this system over others, where it is applicable,
are:--
(_a_) A greater distance between levels can be operated and few
winzes and rises are necessary, thus a great saving of development
work can be effected. A stope 800 to 1000 feet long can be operated
with a winze at either end and with levels 200 or 220 feet apart.
(_b_) There is no shoveling in the stopes at all.
(_c_) No timber is required. As compared with timbering by stulling,
it will apply to stopes too wide and walls too heavy for this method.
Moreover, little staging is required for working the face, since
ore can be drawn from below in such a manner as to allow just the
right head room.
(_d_) Compared to the system of filling with waste, coincidentally
with breaking (second method), it saves altogether in some cases
the cost of filling. In any event, it saves the cost of ore-passes,
of shoveling into them, and of the detailed distribution of the
filling.
Compared with other methods, the system has the following disadvantages,
that:
_A_. The ore requires to be broken in the stopes to a degree of
fineness which will prevent blocking of the chutes at the level.
When pieces too large reach the chutes, nothing will open them but
blasting,--to the damage of timbers and chutes. Some large rocks
are always liable to be buried in the course of ore-breaking.
_B_. Practically no such perfection of walls exists, but some spalling
of waste into the ore will take place. A crushing of the walls
would soon mean the loss of large amounts of ore.
_C_. There is no possibility of regulating the mixture of grade
of ore by varying the working points. It is months after the ore
is broken before it can reach the levels.
_D_. The breaking of 60% more ore than immediate treatment demands
results in the investment of a considerable sum of money. An equilibrium
is ultimately established in a mine worked on this system when a
certain number of stopes full of completely broken ore are available
for entire withdrawal, and there is no further accumulation. But,
in any event, a considerable amount of broken ore must be held in
reserve. In one mine worked on this plan, with which the writer
has had experience, the annual production is about 250,000 tons
and the broken ore represents an investment which, at 5%, means
an annual loss of interest amounting to 7 cents per ton of ore
treated.
_E_. A mine once started on the system is most difficult to alter,
owing to the lack of frequent winzes or passes. Especially is this
so if the only alternative is filling, for an alteration to the
system of filling coincident with breaking finds the mine short
of filling winzes. As the conditions of walls and ore often alter
with depth, change of system may be necessary and the situation
may become very embarrassing.
_F_. The restoping of the walls for lower-grade ore at a later
period is impossible, for the walls of the stope will be crushed,
or, if filled with waste, will usually crush when it is drawn off
to send to a lower stope.
The system has much to recommend it where conditions are favorable.
Like all other alternative methods of mining, it requires the most
careful study in the light of the special conditions involved. In many
mines it can be used for some stopes where not adaptable generally.
It often solves the problem of blind ore-bodies, for they can by
this means be frequently worked with an opening underneath only.
Thus the cost of driving a roadway overhead is avoided, which would
be required if timber or coincident filling were the alternatives.
In such cases ventilation can be managed without an opening above,
by so directing the current of air that it will rise through a
winze from the level below, flow along the stope and into the level
again at the further end of the stope through another winze.
[Illustration: Fig. 40.--Longitudinal section. Ore-pillar support
in narrow stopes.]
SUPPORT BY PILLARS OF ORE.--As a method of mining metals of the
sort under discussion, the use of ore-pillars except in conjunction
with some other means of support has no general application. To
use them without assistance implies walls sufficiently strong to
hold between pillars; to leave them permanently anywhere implies
that the ore abandoned would not repay the labor and the material
of a substitute. There are cases of large, very low-grade mines
where to abandon one-half the ore as pillars is more profitable
than total extraction, but the margin of payability in such ore must
be very, very narrow. Unpayable spots are always left as pillars,
for obvious reasons. Permanent ore-pillars as an adjunct to other
methods of support are in use. Such are the rib-pillars in the
Alaska Treadwell, the form of which is indicated by the upward
extension of the pillars adjacent to the winzes, shown in Figure
37. Always a careful balance must be cast as to the value of the ore
left, and as to the cost of a substitute, because every ore-pillar
can be removed at some outlay. Temporary pillars are not unusual,
particularly to protect roadways and shafts. They are, when left
for these purposes, removed ultimately, usually by beginning at
the farther end and working back to the final exit.
[Illustration: Fig. 41.--Horizontal plan at levels of Broken Hill.
Method of alternate stopes and ore-pillars.]
[Illustration: Fig. 42.--Longitudinal section of Figure 41.]
A form of temporary ore-pillars in very wide deposits is made use
of in conjunction with both filling and timbering (Figs. 37, 39,
40). In the use of temporary pillars for ore-bodies 100 to 250
feet wide at Broken Hill, stopes are carried up at right angles
to the strike, each fifty feet wide and clear across the ore-body
(Figs. 41 and 42). A solid pillar of the same width is left in the
first instance between adjacent stopes, and the initial series of
stopes are walled with one square-set on the sides as the stope is
broken upward. The room between these two lines of sets is filled
with waste alternating with ore-breaking in the usual filling method.
When the ore from the first group of alternate stopes (_ABC_, Fig.
42) is completely removed, the pillars are stoped out and replaced
with waste. The square-sets of the first set of stopes thus become
the boundaries of the second set. Entry and ventilation are obtained
through these lines of square-sets, and the ore is passed out of
the stopes through them.
[Illustration: Fig. 43.--Cross-section of stull support with waste
reënforcement.]
ARTIFICIAL PILLARS.--This system also implies a roof so strong
as not to demand continuous support. Artificial pillars are built
in many different ways. The method most current in fairly narrow
deposits is to reënforce stulls by packing waste above them (Figs.
43 and 44). Not only is it thus possible to economize in stulls by
using the waste which accumulates underground, but the principle
applies also to cases where the stulls alone are not sufficient
support, and yet where complete filling or square-setting is
unnecessary. When the conditions are propitious for this method, it
has the comparative advantage over timber systems of saving timber,
and over filling systems of saving imported filling. Moreover,
these constructions being pillar-shaped (Fig. 44), the intervals
between them provide outlets for broken ore, and specially built
passes are unnecessary. The method has two disadvantages as against
the square-set or filling process, in that more staging must be
provided from which to work, and in stopes over six feet the erection
of machine-drill columns is tedious and costly in time and wages.
[Illustration: Fig. 44.--Longitudinal section of stull and waste
pillars.]
In wide deposits of markedly flat, irregular ore-bodies, where a
definite system is difficult and where timber is expensive, cribs
of cord-wood or logs filled with waste after the order shown in
Figure 31, often make fairly sound pillars. They will not last
indefinitely and are best adapted to the temporary support of the
ore-roof pending filling. The increased difficulty in setting up
machine drills in such stopes adds to the breaking costs,--often
enough to warrant another method of support.
[Illustration: Fig. 45.--Sublevel caving system.]
CAVING SYSTEMS.--This method, with variations, has been applied
to large iron deposits, to the Kimberley diamond mines, to some
copper mines, but in general it has little application to the metal
mines under consideration, as few ore-bodies are of sufficiently
large horizontal area. The system is dependent upon a large area of
loose or "heavy" ground pressing directly on the ore with weight,
such that if the ore be cut into pillars, these will crush. The
details of the system vary, but in general the _modus operandi_
is to prepare roadways through the ore, and from the roadways to
put rises, from which sublevels are driven close under the floating
mass of waste and ore,--sometimes called the "matte" (Fig. 45).
The pillars between these sublevels are then cut away until the
weight above crushes them down. When all the crushed ore which
can be safely reached is extracted, retreat is made and another
series of subopenings is then driven close under the "matte." The
pillar is reduced until it crushes and the operation is repeated.
Eventually the bottom strata of the "matte" become largely ore,
and a sort of equilibrium is reached when there is not much loss
in this direction. "Top slicing" is a variation of the above method
by carrying a horizontal stope from the rises immediately under the
matte, supporting the floating material with timber. At Kimberley
the system is varied in that galleries are run out to the edge of
the diamond-iferous area and enlarged until the pillar between
crushes.
In the caving methods, between 40 and 50% of the ore is removed
by the preliminary openings, and as they are all headings of some
sort, the average cost per ton of this particular ore is higher
than by ordinary stoping methods. On the other hand, the remaining
50 to 60% of the ore costs nothing to break, and the average cost
is often remarkably low. As said, the system implies bodies of large
horizontal area. They must start near enough to the surface that
the whole superincumbent mass may cave and give crushing weight,
or the immediately overhanging roof must easily cave. All of these
are conditions not often met with in mines of the character under
review.
CHAPTER XII.
Mechanical Equipment.
CONDITIONS BEARING ON MINE EQUIPMENT; WINDING APPLIANCES; HAULAGE
EQUIPMENT IN SHAFTS; LATERAL UNDERGROUND TRANSPORT; TRANSPORT IN
STOPES.
There is no type of mechanical engineering which presents such
complexities in determination of the best equipment as does that of
mining. Not only does the economic side dominate over pure mechanics,
but machines must be installed and operated under difficulties which
arise from the most exceptional and conflicting conditions, none of
which can be entirely satisfied. Compromise between capital outlay,
operating efficiency, and conflicting demands is the key-note of
the work.
These compromises are brought about by influences which lie outside
the questions of mechanics of individual machines, and are mainly
as follows:--
1. Continuous change in horizon of operations.
2. Uncertain life of the enterprise.
3. Care and preservation of human life.
4. Unequal adaptability of power transmission mediums.
5. Origin of power.
_First._--The depth to be served and the volume of ore and water
to be handled, are not only unknown at the initial equipment, but
they are bound to change continuously in quantity, location, and
horizon with the extension of the workings.
_Second._--From the mine manager's point of view, which must embrace
that of the mechanical engineer, further difficulty presents itself
because the life of the enterprise is usually unknown, and therefore
a manifest necessity arises for an economic balance of capital
outlay and of operating efficiency commensurate with the prospects
of the mine. Moreover, the initial capital is often limited, and
makeshifts for this reason alone must be provided. In net result,
no mineral deposit of speculative ultimate volume of ore warrants
an initial equipment of the sort that will meet every eventuality,
or of the kind that will give even the maximum efficiency which
a free choice of mining machinery could obtain.
_Third._--In the design and selection of mining machines, the safety
of human life, the preservation of the health of workmen under
conditions of limited space and ventilation, together with reliability
and convenience in installing and working large mechanical tools,
all dominate mechanical efficiency. For example, compressed-air
transmission of power best meets the requirements of drilling,
yet the mechanical losses in the generation, the transmission,
and the application of compressed air probably total, from first
to last, 70 to 85%.
_Fourth._--All machines, except those for shaft haulage, must be
operated by power transmitted from the surface, as obviously power
generation underground is impossible. The conversion of power into
a transmission medium and its transmission are, at the outset,
bound to be the occasions of loss. Not only are the various forms
of transmission by steam, electricity, compressed air, or rods, of
different efficiency, but no one system lends itself to universal
or economical application to all kinds of mining machines. Therefore
it is not uncommon to find three or four different media of power
transmission employed on the same mine. To illustrate: from the
point of view of safety, reliability, control, and in most cases
economy as well, we may say that direct steam is the best motive
force for winding-engines; that for mechanical efficiency and
reliability, rods constitute the best media of power transmission
to pumps; that, considering ventilation and convenience, compressed
air affords the best medium for drills. Yet there are other conditions
as to character of the work, volume of water or ore, and the origin
of power which must in special instances modify each and every one
of these generalizations. For example, although pumping water with
compressed air is mechanically the most inefficient of devices,
it often becomes the most advantageous, because compressed air may
be of necessity laid on for other purposes, and the extra power
required to operate a small pump may be thus most cheaply provided.
_Fifth._--Further limitations and modifications arise out of the
origin of power, for the sources of power have an intimate bearing on
the type of machine and media of transmission. This very circumstance
often compels giving away efficiency and convenience in some machines
to gain more in others. This is evident enough if the principal
origins of power generation be examined. They are in the main as
follows:--
_a_. Water-power available at the mine.
_b_. Water-power available at a less distance than three
or four miles.
_c_. Water-power available some miles away, thus necessitating
electrical transmission (or purchased electrical power).
_d_. Steam-power to be generated at the mine.
_e_. Gas-power to be generated at the mine.
_a_. With water-power at the mine, winding engines can be operated
by direct hydraulic application with a gain in economy over direct
steam, although with the sacrifice of control and reliability. Rods
for pumps can be driven directly with water, but this superiority
in working economy means, as discussed later, a loss of flexibility
and increased total outlay over other forms of transmission to pumps.
As compressed air must be transmitted for drills, the compressor
would be operated direct from water-wheels, but with less control
in regularity of pressure delivery.
_b_. With water-power a short distance from the mine, it would
normally be transmitted either by compressed air or by electricity.
Compressed-air transmission would better satisfy winding and drilling
requirements, but would show a great comparative loss in efficiency
over electricity when applied to pumping. Despite the latter drawback,
air transmission is a method growing in favor, especially in view
of the advance made in effecting compression by falling water.
_c_. In the situation of transmission too far for using compressed
air, there is no alternative but electricity. In these cases, direct
electric winding is done, but under such disadvantages that it
requires a comparatively very cheap power to take precedence over
a subsidiary steam plant for this purpose. Electric air-compressors
work under the material disadvantage of constant speed on a variable
load, but this installation is also a question of economics. The
pumping service is well performed by direct electrical pumps.
_d_. In this instance, winding and air-compression are well accomplished
by direct steam applications; but pumping is beset with wholly
undesirable alternatives, among which it is difficult to choose.
_e_. With internal combustion engines, gasoline (petrol) motors
have more of a position in experimental than in systematic mining,
for their application to winding and pumping and drilling is fraught
with many losses. The engine must be under constant motion, and
that, too, with variable loads. Where power from producer gas is
used, there is a greater possibility of installing large equipments,
and it is generally applied to the winding and lesser units by
conversion into compressed air or electricity as an intermediate
stage.
One thing becomes certain from these examples cited, that the right
installation for any particular portion of the mine's equipment cannot
be determined without reference to all the others. The whole system
of power generation for surface work, as well as the transmission
underground, must be formulated with regard to furnishing the best
total result from all the complicated primary and secondary motors,
even at the sacrifice of some members.
Each mine is a unique problem, and while it would be easy to sketch
an ideal plant, there is no mine within the writer's knowledge
upon which the ideal would, under the many variable conditions,
be the most economical of installation or the most efficient of
operation. The dominant feature of the task is an endeavor to find
a compromise between efficiency and capital outlay. The result is
a series of choices between unsatisfying alternatives, a number of
which are usually found to have been wrong upon further extension
of the mine in depth.
In a general way, it may be stated that where power is generated
on the mine, economy in labor of handling fuel, driving engines,
generation and condensing steam where steam is used, demand a
consolidated power plant for the whole mine equipment. The principal
motors should be driven direct by steam or gas, with power distribution
by electricity to all outlying surface motors and sometimes to
underground motors, and also to some underground motors by compressed
air.
Much progress has been made in the past few years in the perfection
of larger mining tools. Inherently many of our devices are of a
wasteful character, not only on account of the need of special
forms of transmission, but because they are required to operate
under greatly varying loads. As an outcome of transmission losses
and of providing capacity to cope with heavy peak loads, their
efficiency on the basis of actual foot-pounds of work accomplished
is very low.
The adoption of electric transmission in mine work, while in certain
phases beneficial, has not decreased the perplexity which arises
from many added alternatives, none of which are as yet a complete or
desirable answer to any mine problem. When a satisfactory electric
drill is invented, and a method is evolved of applying electricity
to winding-engines that will not involve such abnormal losses due
to high peak load then we will have a solution to our most difficult
mechanical problems, and electricity will deserve the universal
blessing which it has received in other branches of mechanical
engineering.
It is not intended to discuss mine equipment problems from the
machinery standpoint,--there are thousands of different devices,--but
from the point of view of the mine administrator who finds in the
manufactory the various machines which are applicable, and whose
work then becomes that of choosing, arranging, and operating these
tools.
The principal mechanical questions of a mine may be examined under
the following heads:--
1. Shaft haulage.
2. Lateral underground transport.
3. Drainage.
4. Rock drilling.
5. Workshops.
6. Improvements in equipment.
SHAFT HAULAGE.
WINDING APPLIANCES.--No device has yet been found to displace the
single load pulled up the shaft by winding a rope on a drum. Of
driving mechanisms for drum motors the alternatives are the
steam-engine, the electrical motor, and infrequently water-power
or gas engines.
All these have to cope with one condition which, on the basis of
work accomplished, gives them a very low mechanical efficiency.
This difficulty is that the load is intermittent, and it must be
started and accelerated at the point of maximum weight, and from
that moment the power required diminishes to less than nothing
at the end of the haul. A large number of devices are in use to
equalize partially the inequalities of the load at different stages
of the lift. The main lines of progress in this direction have
been:--
_a_. The handling of two cages or skips with one engine
or motor, the descending skip partially balancing
the ascending one.
_b_. The use of tail-ropes or balance weights to compensate
the increasing weight of the descending rope.
_c_. The use of skips instead of cages, thus permitting of
a greater percentage of paying load.
_d_. The direct coupling of the motor to the drum shaft.
_e_. The cone-shaped construction of drums,--this latter
being now largely displaced by the use of the tail-rope.
The first and third of these are absolutely essential for anything
like economy and speed; the others are refinements depending on
the work to be accomplished and the capital available.
Steam winding-engines require large cylinders to start the load,
but when once started the requisite power is much reduced and the
load is too small for steam economy. The throttling of the engine
for controlling speed and reversing the engine at periodic stoppages
militates against the maximum expansion and condensation of the
steam and further increases the steam consumption. In result, the
best of direct compound condensing engines consume from 60 to 100
pounds of steam per horse-power hour, against a possible efficiency
of such an engine working under constant load of less than 16 pounds
of steam per horse-power hour.
It is only within very recent years that electrical motors have
been applied to winding. Even yet, all things considered, this
application is of doubtful value except in localities of extremely
cheap electrical power. The constant speed of alternating current
motors at once places them at a disadvantage for this work of high
peak and intermittent loads. While continuous-current motors can
be made to partially overcome this drawback, such a current, where
power is purchased or transmitted a long distance, is available
only by conversion, which further increases the losses. However,
schemes of electrical winding are in course of development which
bid fair, by a sort of storage of power in heavy fly-wheels or
storage batteries after the peak load, to reduce the total power
consumption; but the very high first cost so far prevents their
very general adoption for metal mining.
Winding-engines driven by direct water- or gas-power are of too rare
application to warrant much discussion. Gasoline driven hoists have a
distinct place in prospecting and early-stage mining, especially in
desert countries where transport and fuel conditions are onerous,
for both the machines and their fuel are easy of transport. As direct
gas-engines entail constant motion of the engine at the power demand
of the peak load, they are hopeless in mechanical efficiency.
Like all other motors in mining, the size and arrangement of the
motor and drum are dependent upon the duty which they will be called
upon to perform. This is primarily dependent upon the depth to be
hoisted from, the volume of the ore, and the size of the load.
For shallow depths and tonnages up to, say, 200 tons daily, geared
engines have a place on account of their low capital cost. Where
great rope speed is not essential they are fully as economical as
direct-coupled engines. With great depths and greater capacities,
speed becomes a momentous factor, and direct-coupled engines are
necessary. Where the depth exceeds 3,000 feet, another element
enters which has given rise to much debate and experiment; that
is, the great increase of starting load due to the increased length
and size of ropes and the drum space required to hold it. So far
the most advantageous device seems to be the Whiting hoist, a
combination of double drums and tail rope.
On mines worked from near the surface, where depth is gained by
the gradual exhaustion of the ore, the only prudent course is to
put in a new hoist periodically, when the demand for increased
winding speed and power warrants. The lack of economy in winding
machines is greatly augmented if they are much over-sized for the
duty. An engine installed to handle a given tonnage to a depth of
3,000 feet will have operated with more loss during the years the
mine is progressing from the surface to that depth than several
intermediate-sized engines would have cost. On most mines the
uncertainty of extension in depth would hardly warrant such a
preliminary equipment. More mines are equipped with over-sized
than with under-sized engines. For shafts on going metal mines
where the future is speculative, an engine will suffice whose size
provides for an extension in depth of 1,000 feet beyond that reached
at the time of its installation. The cost of the engine will depend
more largely upon the winding speed desired than upon any other
one factor. The proper speed to be arranged is obviously dependent
upon the depth of the haulage, for it is useless to have an engine
able to wind 3,000 feet a minute on a shaft 500 feet deep, since it
could never even get under way; and besides, the relative operating
loss, as said, would be enormous.
HAULAGE EQUIPMENT IN THE SHAFT.--Originally, material was hoisted
through shafts in buckets. Then came the cage for transporting mine
cars, and in more recent years the "skip" has been developed. The
aggrandized bucket or "kibble" of the Cornishman has practically
disappeared, but the cage still remains in many mines. The advantages
of the skip over the cage are many. Some of them are:--
_a_. It permits 25 to 40% greater load of material in
proportion to the dead weight of the vehicle.
_b_. The load can be confined within a smaller horizontal
space, thus the area of the shaft need not be so great
for large tonnages.
_c_. Loading and discharging are more rapid, and the latter
is automatic, thus permitting more trips per hour and
requiring less labor.
_d_. Skips must be loaded from bins underground, and by
providing in the bins storage capacity, shaft haulage is
rendered independent of the lateral transport in the
mine, and there are no delays to the engine awaiting
loads. The result is that ore-winding can be concentrated
into fewer hours, and indirect economies in labor
and power are thus effected.
_e_. Skips save the time of the men engaged in the lateral
haulage, as they have no delay waiting for the winding
engine.
Loads equivalent to those from skips are obtained in some mines
by double-decked cages; but, aside from waste weight of the cage,
this arrangement necessitates either stopping the engine to load the
lower deck, or a double-deck loading station. Double-deck loading
stations are as costly to install and more expensive to work than
skip-loading station ore-bins. Cages are also constructed large
enough to take as many as four trucks on one deck. This entails a
shaft compartment double the size required for skips of the same
capacity, and thus enormously increases shaft cost without gaining
anything.
Altogether the advantages of the skip are so certain and so important
that it is difficult to see the justification for the cage under
but a few conditions. These conditions are those which surround
mines of small output where rapidity of haulage is no object, where
the cost of station-bins can thus be evaded, and the convenience
of the cage for the men can still be preserved. The easy change
of the skip to the cage for hauling men removes the last objection
on larger mines. There occurs also the situation in which ore is
broken under contract at so much per truck, and where it is desirable
to inspect the contents of the truck when discharging it, but even
this objection to the skip can be obviated by contracting on a
cubic-foot basis.
Skips are constructed to carry loads of from two to seven tons,
the general tendency being toward larger loads every year. One
of the most feasible lines of improvement in winding is in the
direction of larger loads and less speed, for in this way the sum
total of dead weight of the vehicle and rope to the tonnage of
ore hauled will be decreased, and the efficiency of the engine
will be increased by a less high peak demand, because of this less
proportion of dead weight and the less need of high acceleration.
LATERAL UNDERGROUND TRANSPORT.
Inasmuch as the majority of metal mines dip at considerable angles,
the useful life of a roadway in a metal mine is very short because
particular horizons of ore are soon exhausted. Therefore any method
of transport has to be calculated upon a very quick redemption of
the capital laid out. Furthermore, a roadway is limited in its
daily traffic to the product of the stopes which it serves.
MEN AND ANIMALS.--Some means of transport must be provided, and
the basic equipment is light tracks with push-cars, in capacity
from half a ton to a ton. The latter load is, however, too heavy
to be pushed by one man. As but one car can be pushed at a time,
hand-trucking is both slow and expensive. At average American or
Australian wages, the cost works out between 25 and 35 cents a
ton per mile. An improvement of growing import where hand-trucking
is necessary is the overhead mono-rail instead of the track.
If the supply to any particular roadway is such as to fully employ
horses or mules, the number of cars per trip can be increased up
to seven or eight. In this case the expense, including wages of
the men and wear, tear, and care of mules, will work out roughly
at from 7 to 10 cents per ton mile. Manifestly, if the ore-supply
to a particular roadway is insufficient to keep a mule busy, the
economy soon runs off.
MECHANICAL HAULAGE.--Mechanical haulage is seldom applicable to
metal mines, for most metal deposits dip at considerable angles,
and therefore, unlike most coal-mines, the horizon of haulage must
frequently change, and there are no main arteries along which haulage
continues through the life of the mine. Any mechanical system entails
a good deal of expense for installation, and the useful life of
any particular roadway, as above said, is very short. Moreover,
the crooked roadways of most metal mines present difficulties of
negotiation not to be overlooked. In order to use such systems it
is necessary to condense the haulage to as few roadways as possible.
Where the tonnage on one level is not sufficient to warrant other
than men or animals, it sometimes pays (if the dip is steep enough)
to dump everything through winzes from one to two levels to a main
road below where mechanical equipment can be advantageously provided.
The cost of shaft-winding the extra depth is inconsiderable compared
to other factors, for the extra vertical distance of haulage can
be done at a cost of one or two cents per ton mile. Moreover, from
such an arrangement follows the concentration of shaft-bins, and of
shaft labor, and winding is accomplished without so much shifting
as to horizon, all of which economies equalize the extra distance
of the lift.
There are three principal methods of mechanical transport in use:--
1. Cable-ways.
2. Compressed-air locomotives.
3. Electrical haulage.
Cable-ways or endless ropes are expensive to install, and to work
to the best advantage require double tracks and fairly straight
roads. While they are economical in operation and work with little
danger to operatives, the limitations mentioned preclude them from
adoption in metal mines, except in very special circumstances such
as main crosscuts or adit tunnels, where the haulage is straight
and concentrated from many sources of supply.
Compressed-air locomotives are somewhat heavy and cumbersome, and
therefore require well-built tracks with heavy rails, but they
have very great advantages for metal mine work. They need but a
single track and are of low initial cost where compressed air is
already a requirement of the mine. No subsidiary line equipment is
needed, and thus they are free to traverse any road in the mine and
can be readily shifted from one level to another. Their mechanical
efficiency is not so low in the long run as might appear from the
low efficiency of pneumatic machines generally, for by storage of
compressed air at the charging station a more even rate of energy
consumption is possible than in the constant cable and electrical
power supply which must be equal to the maximum demand, while the
air-plant consumes but the average demand.
Electrical haulage has the advantage of a much more compact locomotive
and the drawback of more expensive track equipment, due to the
necessity of transmission wire, etc. It has the further disadvantages
of uselessness outside the equipped haulage way and of the dangers
of the live wire in low and often wet tunnels.
In general, compressed-air locomotives possess many attractions
for metal mine work, where air is in use in any event and where
any mechanical system is at all justified. Any of the mechanical
systems where tonnage is sufficient in quantity to justify their
employment will handle material for from 1.5 to 4 cents per ton
mile.
TRACKS.--Tracks for hand, mule, or rope haulage are usually built
with from 12- to 16-pound rails, but when compressed-air or electrical
locomotives are to be used, less than 24-pound rails are impossible.
As to tracks in general, it may be said that careful laying out
with even grades and gentle curves repays itself many times over in
their subsequent operation. Further care in repair and lubrication
of cars will often make a difference of 75% in the track resistance.
TRANSPORT IN STOPES.--Owing to the even shorter life of individual
stopes than levels, the actual transport of ore or waste in them is
often a function of the aboriginal shovel plus gravity. As shoveling
is the most costly system of transport known, any means of stoping
that decreases the need for it has merit. Shrinkage-stoping eliminates
it altogether. In the other methods, gravity helps in proportion to
the steepness of the dip. When the underlie becomes too flat for
the ore to "run," transport can sometimes be helped by pitching
the ore-passes at a steeper angle than the dip (Fig. 36). In some
cases of flat deposits, crosscuts into the walls, or even levels
under the ore-body, are justifiable. The more numerous the ore-passes,
the less the lateral shoveling, but as passes cost money for
construction and for repair, there is a nice economic balance in
their frequency.
Mechanical haulage in stopes has been tried and finds a field under
some conditions. In dips under 25° and possessing fairly sound
hanging-wall, where long-wall or flat-back cuts are employed, temporary
tracks can often be laid in the stopes and the ore run in cars to
the main passes. In such cases, the tracks are pushed up close
to the face after each cut. Further self-acting inclines to lower
cars to the levels can sometimes be installed to advantage. This
arrangement also permits greater intervals between levels and less
number of ore-passes. For dips between 25° and 50° where the mine
is worked without stope support or with occasional pillars, a very
useful contrivance is the sheet-iron trough--about eighteen inches
wide and six inches deep--made in sections ten or twelve feet long
and readily bolted together. In dips 35° to 50° this trough, laid
on the foot-wall, gives a sufficiently smooth surface for the ore
to run upon. When the dip is flat, the trough, if hung from plugs
in the hanging-wall, may be swung backward and forward. The use of
this "bumping-trough" saves much shoveling. For handling filling
or ore in flat runs it deserves wider adoption. It is, of course,
inapplicable in passes as a "bumping-trough," but can be fixed to
give smooth surface. In flat mines it permits a wider interval
between levels and therefore saves development work. The life of
this contrivance is short when used in open stopes, owing to the
dangers of bombardment from blasting.
In dips steeper than 50° much of the shoveling into passes can be
saved by rill-stoping, as described on page 100. Where flat-backed
stopes are used in wide ore-bodies with filling, temporary tracks
laid on the filling to the ore-passes are useful, for they permit
wider intervals between passes.
In that underground engineer's paradise, the Witwatersrand, where
the stopes require neither timber nor filling, the long, moderately
pitched openings lend themselves particularly to the swinging iron
troughs, and even endless wire ropes have been found advantageous
in certain cases.
Where the roof is heavy and close support is required, and where
the deposits are very irregular in shape and dip, there is little
hope of mechanical assistance in stope transport.
CHAPTER XIII.
Mechanical Equipment. (_Continued_).
DRAINAGE: CONTROLLING FACTORS; VOLUME AND HEAD OF WATER; FLEXIBILITY;
RELIABILITY; POWER CONDITIONS; MECHANICAL EFFICIENCY; CAPITAL OUTLAY.
SYSTEMS OF DRAINAGE,--STEAM PUMPS, COMPRESSED-AIR PUMPS, ELECTRICAL
PUMPS, ROD-DRIVEN PUMPS, BAILING; COMPARATIVE VALUE OF VARIOUS
SYSTEMS.
With the exception of drainage tunnels--more fully described in
Chapter VIII--all drainage must be mechanical. As the bulk of mine
water usually lies near the surface, saving in pumping can sometimes
be effected by leaving a complete pillar of ore under some of the
upper levels. In many deposits, however, the ore has too many channels
to render this of much avail.
There are six factors which enter into a determination of mechanical
drainage systems for metal mines:--
1. Volume and head of water.
2. Flexibility to fluctuation in volume and head.
3. Reliability.
4. Capital cost.
5. The general power conditions.
6. Mechanical efficiency.
In the drainage appliances, more than in any other feature of the
equipment, must mechanical efficiency be subordinated to the other
issues.
FLEXIBILITY.--Flexibility in plant is necessary because volume and
head of water are fluctuating factors. In wet regions the volume
of water usually increases for a certain distance with the extension
of openings in depth. In dry climates it generally decreases with the
downward extension of the workings after a certain depth. Moreover,
as depth progresses, the water follows the openings more or less
and must be pumped against an ever greater head. In most cases
the volume varies with the seasons. What increase will occur, from
what horizon it must be lifted, and what the fluctuations in volume
are likely to be, are all unknown at the time of installation. If
a pumping system were to be laid out for a new mine, which would
peradventure meet every possible contingency, the capital outlay would
be enormous, and the operating efficiency would be very low during
the long period in which it would be working below its capacity. The
question of flexibility does not arise so prominently in coal-mines,
for the more or less flat deposits give a fixed factor of depth.
The flow is also more steady, and the volume can be in a measure
approximated from general experience.
RELIABILITY.--The factor of reliability was at one time of more
importance than in these days of high-class manufacture of many
different pumping systems. Practically speaking, the only insurance
from flooding in any event lies in the provision of a relief system
of some sort,--duplicate pumps, or the simplest and most usual
thing, bailing tanks. Only Cornish and compressed-air pumps will
work with any security when drowned, and electrical pumps are easily
ruined.
GENERAL POWER CONDITIONS.--The question of pumping installation
is much dependent upon the power installation and other power
requirements of the mine. For instance, where electrical power is
purchased or generated by water-power, then electrical pumps have
every advantage. Or where a large number of subsidiary motors can be
economically driven from one central steam- or gas-driven electrical
generation plant, they again have a strong call,--especially if
the amount of water to be handled is moderate. Where the water
is of limited volume and compressed-air plant a necessity for the
mine, then air-driven pumps may be the most advantageous, etc.
MECHANICAL EFFICIENCY.--The mechanical efficiency of drainage machinery
is very largely a question of method of power application. The
actual pump can be built to almost the same efficiency for any
power application, and with the exception of the limited field
of bailing with tanks, mechanical drainage is a matter of pumps.
All pumps must be set below their load, barring a few possible
feet of suction lift, and they are therefore perforce underground,
and in consequence all power must be transmitted from the surface.
Transmission itself means loss of power varying from 10 to 60%,
depending upon the medium used. It is therefore the choice of
transmission medium that largely governs the mechanical efficiency.
SYSTEMS OF DRAINAGE.--The ideal pumping system for metal mines
would be one which could be built in units and could be expanded
or contracted unit by unit with the fluctuation in volume; which
could also be easily moved to meet the differences of lifts; and
in which each independent unit could be of the highest mechanical
efficiency and would require but little space for erection. Such
an ideal is unobtainable among any of the appliances with which
the writer is familiar.
The wide variations in the origin of power, in the form of transmission,
and in the method of final application, and the many combinations
of these factors, meet the demands for flexibility, efficiency,
capital cost, and reliability in various degrees depending upon
the environment of the mine. Power nowadays is generated primarily
with steam, water, and gas. These origins admit the transmission of
power to the pumps by direct steam, compressed air, electricity,
rods, or hydraulic columns.
DIRECT STEAM-PUMPS.--Direct steam has the disadvantage of radiated
heat in the workings, of loss by the radiation, and, worse still,
of the impracticability of placing and operating a highly efficient
steam-engine underground. It is all but impossible to derive benefit
from the vacuum, as any form of surface condenser here is impossible,
and there can be no return of the hot soft water to the boilers.
Steam-pumps fall into two classes, rotary and direct-acting; the former
have the great advantage of permitting the use of steam expansively
and affording some field for effective use of condensation, but
they are more costly, require much room, and are not fool-proof.
The direct-acting pumps have all the advantage of compactness and
the disadvantage of being the most inefficient of pumping machines
used in mining. Taking the steam consumption of a good surface
steam plant at 15 pounds per horse-power hour, the efficiency of
rotary pumps with well-insulated pipes is probably not over 50%,
and of direct-acting pumps from 40% down to 10%.
The advantage of all steam-pumps lies in the low capital outlay,--hence
their convenient application to experimental mining and temporary
pumping requirements. For final equipment they afford a great deal
of flexibility, for if properly constructed they can be, with slight
alteration, moved from one horizon to another without loss of relative
efficiency. Thus the system can be rearranged for an increased
volume of water, by decreasing the lift and increasing the number
of pumps from different horizons.
COMPRESSED-AIR PUMPS.--Compressed-air transmission has an application
similar to direct steam, but it is of still lower mechanical efficiency,
because of the great loss in compression. It has the superiority
of not heating the workings, and there is no difficulty as to the
disposal of the exhaust, as with steam. Moreover, such pumps will
work when drowned. Compressed air has a distinct place for minor
pumping units, especially those removed from the shaft, for they
can be run as an adjunct to the air-drill system of the mine, and
by this arrangement much capital outlay may be saved. The cost of
the extra power consumed by such an arrangement is less than the
average cost of compressed-air power, because many of the compressor
charges have to be paid anyway. When compressed air is water-generated,
they have a field for permanent installations. The efficiency of
even rotary air-driven pumps, based on power delivered into a good
compressor, is probably not over 25%.
ELECTRICAL PUMPS.--Electrical pumps have somewhat less flexibility
than steam- or air-driven apparatus, in that the speed of the pumps can
be varied only within small limits. They have the same great advantage
in the easy reorganization of the system to altered conditions of
water-flow. Electricity, when steam-generated, has the handicap
of the losses of two conversions, the actual pump efficiency being
about 60% in well-constructed plants; the efficiency is therefore
greater than direct steam or compressed air. Where the mine is
operated with water-power, purchased electric current, or where
there is an installation of electrical generating plant by steam or
gas for other purposes, electrically driven pumps take precedence
over all others on account of their combined moderate capital outlay,
great flexibility, and reasonable efficiency.
In late years, direct-coupled, electric-driven centrifugal pumps
have entered the mining field, but their efficiency, despite makers'
claims, is low. While they show comparatively good results on low
lifts the slip increases with the lift. In heads over 200 feet
their efficiency is probably not 30% of the power delivered to the
electrical generator. Their chief attractions are small capital
cost and the compact size which admits of easy installation.
ROD-DRIVEN PUMPS.--Pumps of the Cornish type in vertical shafts,
if operated to full load and if driven by modern engines, have
an efficiency much higher than any other sort of installation,
and records of 85 to 90% are not unusual. The highest efficiency
in these pumps yet obtained has been by driving the pump with rope
transmission from a high-speed triple expansion engine, and in
this plant an actual consumption of only 17 pounds of steam per
horse-power hour for actual water lifted has been accomplished.
To provide, however, for increase of flow and change of horizon,
rod-driven pumps must be so overpowered at the earlier stage of
the mine that they operate with great loss. Of all pumping systems
they are the most expensive to provide. They have no place in crooked
openings and only work in inclines with many disadvantages.
In general their lack of flexibility is fast putting them out of
the metal miner's purview. Where the pumping depth and volume of
water are approximately known, as is often the case in coal mines,
this, the father of all pumps, still holds its own.
HYDRAULIC PUMPS.--Hydraulic pumps, in which a column of water is
used as the transmission fluid from a surface pump to a corresponding
pump underground has had some adoption in coal mines, but little
in metal mines. They have a certain amount of flexibility but low
efficiency, and are not likely to have much field against electrical
pumps.
BAILING.--Bailing deserves to be mentioned among drainage methods,
for under certain conditions it is a most useful system, and at
all times a mine should be equipped with tanks against accident
to the pumps. Where the amount of water is limited,--up to, say,
50,000 gallons daily,--and where the ore output of the mine permits
the use of the winding-engine for part of the time on water haulage,
there is in the method an almost total saving of capital outlay.
Inasmuch as the winding-engine, even when the ore haulage is finished
for the day, must be under steam for handling men in emergencies,
and as the labor of stokers, engine-drivers, shaft-men, etc., is
therefore necessary, the cost of power consumed by bailing is not
great, despite the low efficiency of winding-engines.
COMPARISON OF VARIOUS SYSTEMS.--If it is assumed that flexibility,
reliability, mechanical efficiency, and capital cost can each be
divided into four figures of relative importance,--_A_, _B_, _C_,
and _D_, with _A_ representing the most desirable result,--it is
possible to indicate roughly the comparative values of various
pumping systems. It is not pretended that the four degrees are of
equal import. In all cases the factor of general power conditions
on the mine may alter the relative positions.
====================================================================
|Direct|Compressed| |Steam-| |
|Steam | Air |Electricity|Driven|Hydraulic|Bailing
|Pumps | | | Rods | Columns | Tanks
-------------|------|----------|-----------|------|---------|-------
Flexibility. | _A_ | _A_ | _B_ | _D_ | _B_ | _A_
Reliability. | _B_ | _B_ | _B_ | _A_ | _D_ | _A_
Mechanical | | | | | |
Efficiency.| _C_ | _D_ | _B_ | _A_ | _C_ | _D_
Capital Cost | _A_ | _B_ | _B_ | _D_ | _D_ | --
====================================================================
As each mine has its special environment, it is impossible to formulate
any final conclusion on a subject so involved. The attempt would lead
to a discussion of a thousand supposititious cases and hypothetical
remedies. Further, the description alone of pumping machines would
fill volumes, and the subject will never be exhausted. The engineer
confronted with pumping problems must marshal all the alternatives,
count his money, and apply the tests of flexibility, reliability,
efficiency, and cost, choose the system of least disadvantages,
and finally deprecate the whole affair, for it is but a parasite
growth on the mine.
CHAPTER XIV.
Mechanical Equipment (_Concluded_).
MACHINE DRILLING: POWER TRANSMISSION; COMPRESSED AIR _VS_. ELECTRICITY;
AIR DRILLS; MACHINE _VS_. HAND DRILLING. WORK-SHOPS. IMPROVEMENT
IN EQUIPMENT.
For over two hundred years from the introduction of drill-holes
for blasting by Caspar Weindel in Hungary, to the invention of
the first practicable steam percussion drill by J. J. Crouch of
Philadelphia, in 1849, all drilling was done by hand. Since Crouch's
time a host of mechanical drills to be actuated by all sorts of
power have come forward, and even yet the machine-drill has not
reached a stage of development where it can displace hand-work
under all conditions. Steam-power was never adapted to underground
work, and a serviceable drill for this purpose was not found until
compressed air for transmission was demonstrated by Dommeiller
on the Mt. Cenis tunnel in 1861.
The ideal requirements for a drill combine:--
a. Power transmission adapted to underground conditions.
b. Lightness.
c. Simplicity of construction.
d. Strength.
e. Rapidity and strength of blow.
f. Ease of erection.
g. Reliability.
h. Mechanical efficiency.
i. Low capital cost.
No drill invented yet fills all these requirements, and all are
a compromise on some point.
POWER TRANSMISSION; COMPRESSED AIR _vs_. ELECTRICITY.--The only
transmissions adapted to underground drill-work are compressed
air and electricity, and as yet an electric-driven drill has not
been produced which meets as many of the requirements of the metal
miner as do compressed-air drills. The latter, up to date, have
superiority in simplicity, lightness, ease of erection, reliability,
and strength over electric machines. Air has another advantage in
that it affords some assistance to ventilation, but it has the
disadvantage of remarkably low mechanical efficiency. The actual
work performed by the standard 3-3/4-inch air-drill probably does
not amount to over two or three horse-power against from fifteen to
eighteen horse-power delivered into the compressor, or mechanical
efficiency of less than 25%. As electrical power can be delivered to
the drill with much less loss than compressed air, the field for a
more economical drill on this line is wide enough to create eventually
the proper tool to apply it. The most satisfactory electric drill
produced has been the Temple drill, which is really an air-drill
driven by a small electrically-driven compressor placed near the
drill itself. But even this has considerable deficiencies in mining
work; the difficulties of setting up, especially for stoping work,
and the more cumbersome apparatus to remove before blasting are
serious drawbacks. It has deficiencies in reliability and greater
complication of machinery than direct air.
AIR-COMPRESSION.--The method of air-compression so long accomplished
only by power-driven pistons has now an alternative in some situations
by the use of falling water. This latter system is a development
of the last twelve years, and, due to the low initial outlay and
extremely low operating costs, bids fair in those regions where
water head is available not only to displace the machine compressor,
but also to extend the application of compressed air to mine motors
generally, and to stay in some environments the encroachment of
electricity into the compressed-air field. Installations of this
sort in the West Kootenay, B.C., and at the Victoria copper mine,
Michigan, are giving results worthy of careful attention.
Mechanical air-compressors are steam-, water-, electrical-, and
gas-driven, the alternative obviously depending on the source and
cost of power. Electrical- and gas- and water-driven compressors
work under the disadvantage of constant speed motors and respond
little to the variation in load, a partial remedy for which lies
in enlarged air-storage capacity. Inasmuch as compressed air, so
far as our knowledge goes at present, must be provided for drills,
it forms a convenient transmission of power to various motors
underground, such as small pumps, winches, or locomotives. As stated
in discussing those machines, it is not primarily a transmission
of even moderate mechanical efficiency for such purposes; but as
against the installation and operation of independent transmission,
such as steam or electricity, the economic advantage often compensates
the technical losses. Where such motors are fixed, as in pumps
and winches, a considerable gain in efficiency can be obtained by
reheating.
It is not proposed to enter a discussion of mechanical details of
air-compression, more than to call attention to the most common
delinquency in the installation of such plants. This deficiency
lies in insufficient compression capacity for the needs of the
mine and consequent effective operation of drills, for with under
75 pounds pressure the drills decrease remarkably in rapidity of
stroke and force of the blow. The consequent decrease in actual
accomplishment is far beyond the ratio that might be expected on
the basis of mere difference of pressure. Another form of the same
chronic ill lies in insufficient air-storage capacity to provide
for maintenance of pressure against moments when all drills or
motors in the mine synchronize in heavy demand for air, and thus
lower the pressure at certain periods.
AIR-DRILLS.--Air-drills are from a mechanical point of view broadly
of two types,--the first, in which the drill is the piston extension;
and the second, a more recent development for mining work, in which
the piston acts as a hammer striking the head of the drill. From an
economic point of view drills may be divided into three classes.
First, heavy drills, weighing from 150 to 400 pounds, which require
two men for their operation; second, "baby" drills of the piston type,
weighing from 110 to 150 pounds, requiring one man with occasional
assistance in setting up; and third, very light drills almost wholly
of the hammer type. This type is built in two forms: a heavier
type for mounting on columns, weighing about 80 pounds; and a type
after the order of the pneumatic riveter, weighing as low as 20
pounds and worked without mounting.
The weight and consequent mobility of a drill, aside from labor
questions, have a marked effect on costs, for the lighter the drill
the less difficulty and delay in erection, and consequent less
loss of time and less tendency to drill holes from one radius,
regardless of pointing to take best advantage of breaking planes.
Moreover, smaller diameter and shorter holes consume less explosives
per foot advanced or per ton broken. The best results in tonnage
broken and explosive consumed, if measured by the foot of drill-hole
necessary, can be accomplished from hand-drilling and the lighter
the machine drill, assuming equal reliability, the nearer it
approximates these advantages.
The blow, and therefore size and depth of hole and rapidity of
drilling, are somewhat dependent upon the size of cylinders and
length of stroke, and therefore the heavier types are better adapted
to hard ground and to the deep holes of some development points.
Their advantages over the other classes lie chiefly in this ability
to bore exceedingly hard material and in the greater speed of advance
possible in development work; but except for these two special
purposes they are not as economical per foot advanced or per ton
of ore broken as the lighter drills.
The second class, where men can be induced to work them one man per
drill, saves in labor and gains in mobility. Many tests show great
economy of the "baby" type of piston drills in average ground over
the heavier machines for stoping and for most lateral development.
All piston types are somewhat cumbersome and the heavier types
require at least four feet of head room. The "baby" type can be
operated in less space than this, but for narrow stopes they do
not lend themselves with the same facility as the third class.
The third class of drills is still in process of development, but
it bids fair to displace much of the occupation of the piston types
of drill. Aside from being a one-man drill, by its mobility it
will apparently largely reproduce the advantage of hand-drilling
in ability to place short holes from the most advantageous angles
and for use in narrow places. As compared with other drills it
bids fair to require less time for setting up and removal and for
change of bits; to destroy less steel by breakages; to dull the
bits less rapidly per foot of hole; to be more economical of power;
to require much less skill in operation, for judgment is less called
upon in delivering speed; and to evade difficulties of fissured
ground, etc. And finally the cost is only one-half, initially and
for spares. Its disadvantage so far is a lack of reliability due to
lightness of construction, but this is very rapidly being overcome.
This type, however, is limited in depth of hole possible, for,
from lack of positive reverse movement, there is a tendency for
the spoil to pack around the bit, and as a result about four feet
seems the limit.
The performance of a machine-drill under show conditions may be
anything up to ten or twelve feet of hole per hour on rock such
as compact granite; but in underground work a large proportion of
the time is lost in picking down loose ore, setting up machines,
removal for blasting, clearing away spoil, making adjustments,
etc. The amount of lost time is often dependent upon the width of
stope or shaft and the method of stoping. Situations which require
long drill columns or special scaffolds greatly accentuate the loss
of time. Further, the difficulties in setting up reflect indirectly
on efficiency to a greater extent in that a larger proportion of
holes are drilled from one radius and thus less adapted to the
best breaking results than where the drill can easily be reset from
various angles.
The usual duty of a heavy drill per eight-hour shift using two men
is from 20 to 40 feet of hole, depending upon the rock, facilities
for setting up, etc., etc.[*] The lighter drills have a less average
duty, averaging from 15 to 25 feet per shift.
[Footnote *: Over the year 1907 in twenty-eight mines compiled
from Alaska to Australia, an average of 23.5 feet was drilled per
eight-hour shift by machines larger than three-inch cylinder.]
MACHINE _vs_. HAND-DRILLING.--The advantages of hand-drilling over
machine-drilling lie, first, in the total saving of power, the
absence of capital cost, repairs, depreciation, etc., on power,
compresser and drill plant; second, the time required for setting
up machine-drills does not warrant frequent blasts, so that a number
of holes on one radius are a necessity, and therefore machine-holes
generally cannot be pointed to such advantage as hand-holes. Hand-holes
can be set to any angle, and by thus frequent blasting yield greater
tonnage per foot of hole. Third, a large number of comparative
statistics from American, South African, and Australian mines show
a saving of about 25% in explosives for the same tonnage or foot
of advance by hand-holes over medium and heavy drill-holes.
The duty of a skilled white man, single-handed, in rock such as
is usually met below the zone of oxidation, is from 5 to 7 feet
per shift, depending on the rock and the man. Two men hand-drilling
will therefore do from 1/4 to 2/3 of the same footage of holes
that can be done by two men with a heavy machine-drill, and two
men hand-drilling will do from 1/5 to 1/2 the footage of two men
with two light drills.
The saving in labor of from 75 to 33% by machine-drilling may or
may not be made up by the other costs involved in machine-work.
The comparative value of machine- and hand-drilling is not subject
to sweeping generalization. A large amount of data from various
parts of the world, with skilled white men, shows machine-work
to cost from half as much per ton or foot advanced as hand-work
to 25% more than handwork, depending on the situation, type of
drill, etc. In a general way hand-work can more nearly compete
with heavy machines than light ones. The situations where hand-work
can compete with even light machines are in very narrow stopes where
drills cannot be pointed to advantage, and where the increased
working space necessary for machine drills results in breaking more
waste. Further, hand-drilling can often compete with machine-work
in wide stopes where long columns or platforms must be used and
therefore there is much delay in taking down, reërection, etc.
Many other factors enter into a comparison, however, for
machine-drilling produces a greater number of deeper holes and
permits larger blasts and therefore more rapid progress. In driving
levels under average conditions monthly footage is from two to
three times as great with heavy machines as by hand-drilling, and
by lighter machines a somewhat less proportion of greater speed.
The greater speed obtained in development work, the greater tonnage
obtained per man in stoping, with consequent reduction in the number
of men employed, and in reduction of superintendence and general
charges are indirect advantages for machine-drilling not to be
overlooked.
The results obtained in South Africa by hand-drilling in shafts,
and its very general adoption there, seem to indicate that better
speed and more economical work can be obtained in that way in very
large shafts than by machine-drilling. How far special reasons
there apply to smaller shafts or labor conditions elsewhere have
yet to be demonstrated. In large-dimension shafts demanding a large
number of machines, the handling of long machine bars and machines
generally results in a great loss of time. The large charges in
deep holes break the walls very irregularly; misfires cause more
delay; timbering is more difficult in the face of heavy blasting
charges; and the larger amount of spoil broken at one time delays
renewed drilling, and altogether the advantages seem to lie with
hand-drilling in shafts of large horizontal section.
The rapid development of special drills for particular conditions
has eliminated the advantage of hand-work in many situations during
the past ten years, and the invention of the hammer type of drill
bids fair to render hand-drilling a thing of the past. One
generalization is possible, and that is, if drills are run on 40-50
pounds' pressure they are no economy over hand-drilling.
WORKSHOPS.
In addition to the ordinary blacksmithy, which is a necessity,
the modern tendency has been to elaborate the shops on mines to
cover machine-work, pattern-making and foundry-work, in order that
delays may be minimized by quick repairs. To provide, however,
for such contingencies a staff of men must be kept larger than
the demand of average requirements. The result is an effort to
provide jobs or to do work extravagantly or unnecessarily well.
In general, it is an easy spot for fungi to start growing on the
administration, and if custom repair shops are available at all,
mine shops can be easily overdone.
A number of machines are now in use for sharpening drills.
Machine-sharpening is much cheaper than hand-work, although the drills
thus sharpened are rather less efficient owing to the difficulty of
tempering them to the same nicety; however, the net results are
in favor of the machines.
IMPROVEMENT IN EQUIPMENT.
Not only is every mine a progressive industry until the bottom
gives out, but the technology of the industry is always progressing,
so that the manager is almost daily confronted with improvements
which could be made in his equipment that would result in decreasing
expenses or increasing metal recovery. There is one test to the
advisability of such alterations: How long will it take to recover
the capital outlay from the savings effected? and over and above
this recovery of capital there must be some very considerable gain.
The life of mines is at least secured over the period exposed in
the ore-reserves, and if the proposed alteration will show its
recovery and profit in that period, then it is certainly justified.
If it takes longer than this on the average speculative ore-deposit,
it is a gamble on finding further ore. As a matter of practical
policy it will be found that an improvement in equipment which
requires more than three or four years to redeem itself out of
saving, is usually a mechanical or metallurgical refinement the
indulgence in which is very doubtful.
CHAPTER XV.
Ratio of Output to the Mine.
DETERMINATION OF THE POSSIBLE MAXIMUM; LIMITING FACTORS; COST OF
EQUIPMENT; LIFE OF THE MINE; MECHANICAL INEFFICIENCY OF PATCHWORK
PLANT; OVERPRODUCTION OF BASE METAL; SECURITY OF INVESTMENT.
The output obtainable from a given mine is obviously dependent
not only on the size of the deposit, but also on the equipment
provided,--in which equipment means the whole working appliances,
surface and underground.
A rough and ready idea of output possibilities of inclined deposits
can be secured by calculating the tonnage available per foot of
depth from the horizontal cross-section of the ore-bodies exposed
and assuming an annual depth of exhaustion, or in horizontal deposits
from an assumption of a given area of exhaustion. Few mines, at the
time of initial equipment, are developed to an extent from which
their possibilities in production are evident, for wise finance
usually leads to the erection of some equipment and production before
development has been advanced to a point that warrants a large or
final installation. Moreover, even were the full possibilities of
the mine known, the limitations of finance usually necessitate a
less plant to start with than is finally contemplated. Therefore
output and equipment are usually growing possibilities during the
early life of a mine.
There is no better instance in mine engineering where pure theory
must give way to practical necessities of finance than in the
determination of the size of equipment and therefore output. Moreover,
where finance even is no obstruction, there are other limitations
of a very practical order which must dominate the question of the
size of plant giving the greatest technical economy. It is, however,
useful to state the theoretical considerations in determining the
ultimate volume of output and therefore the size of equipments,
for the theory will serve to illuminate the practical limitations.
The discussion will also again demonstrate that all engineering
is a series of compromises with natural and economic forces.
OUTPUT GIVING LEAST PRODUCTION COST.--As one of the most important
objectives is to work the ore at the least cost per ton, it is
not difficult to demonstrate that the minimum working costs can
be obtained only by the most intensive production. To prove this,
it need only be remembered that the working expenses of a mine
are of two sorts: one is a factor of the tonnage handled, such as
stoping and ore-dressing; the other is wholly or partially dependent
upon time. A large number of items are of this last order. Pumping
and head-office expenses are almost entirely charges independent
of the tonnage handled. Superintendence and staff salaries and
the like are in a large proportion dependent upon time. Many other
elements of expense, such as the number of engine-drivers, etc., do
not increase proportionately to increase in tonnage. These charges,
or the part of them dependent upon time apart from tonnage, may be
termed the "fixed charges."
There is another fixed charge more obscure yet no less certain.
Ore standing in a mine is like money in a bank drawing no interest,
and this item of interest may be considered a "fixed charge," for
if the ore were realized earlier, this loss could be partially
saved. This subject is further referred to under "Amortization."
If, therefore, the time required to exhaust the mine be prolonged
by the failure to maintain the maximum output, the total cost of
working it will be greater by the fixed charges over such an increased
period. Conversely, by equipping on a larger scale, the mine will
be exhausted more quickly, a saving in total cost can be made, and
the ultimate profit can be increased by an amount corresponding
to the time saved from the ravages of fixed charges. In fine, the
working costs may be reduced by larger operations, and therefore
the value of the mine increased.
The problem in practice usually takes the form of the relative
superiority of more or of fewer units of plant, and it can be considered
in more detail if the production be supposed to consist of units
averaging say 100 tons per day each. The advantage of more units
over less will be that the extra ones can be produced free of fixed
charges, for these are an expense already involved in the lesser
units. This extra production will also enjoy the interest which
can be earned over the period of its earlier production. Moreover,
operations on a larger scale result in various minor economies
throughout the whole production, not entirely included in the type
of expenditure mentioned as "fixed charges." We may call these
various advantages the "saving of fixed charges" due to larger-scale
operations. The saving of fixed charges amounts to very considerable
sums. In general the items of working cost alone, mentioned above,
which do not increase proportionately to the tonnage, aggregate
from 10 to 25% of the total costs. Where much pumping is involved,
the percentage will become even greater.
The question of the value of the mine as affected by the volume
of output becomes very prominent in low-grade mines, where, if
equipped for output on too small a scale, no profits at all could
be earned, and a sufficient production is absolutely imperative
for any gain. There are many mines in every country which with
one-third of their present rate of production would lose money.
That is, the fixed charges, if spread over small output, would be
so great per ton that the profit would be extinguished by them.
In the theoretical view, therefore, it would appear clear that
the greatest ultimate profit from a mine can be secured only by
ore extraction under the highest pressure. As a corollary to this
it follows that development must proceed with the maximum speed.
Further, it follows that the present value of a mine is at least
partially a factor of the volume of output contemplated.
FACTORS LIMITING THE OUTPUT.
Although the above argument can be academically defended, there
are, as said at the start, practical limitations to the maximum
intensity of production, arising out of many other considerations
to which weight must be given. In the main, there are five principal
limitations:--
1. Cost of equipment.
2. Life of the mine.
3. Mechanical inefficiency of patchwork plant.
4. Overproduction of base metal.
5. Security of investment.
COST OF EQUIPMENT.--The "saving of fixed charges" can only be obtained
by larger equipment, which represents an investment. Mining works,
shafts, machinery, treatment plants, and all the paraphernalia cost
large sums of money. They become either worn out or practically
valueless through the exhaustion of the mines. Even surface machinery
when in good condition will seldom realize more than one-tenth of its
expense if useless at its original site. All mines are ephemeral;
therefore virtually the entire capital outlay of such works must
be redeemed during the life of the mine, and the interest on it
must also be recovered.
The certain life, with the exception of banket and a few other
types of deposit, is that shown by the ore in sight, plus something
for extension of the deposit beyond exposures. So, against the
"savings" to be made, must be set the cost of obtaining them, for
obviously it is of no use investing a dollar to save a total of
ninety cents. The economies by increased production are, however,
of such an important character that the cost of almost any number
of added units (within the ability of the mine to supply them)
can be redeemed from these savings in a few years. For instance,
in a Californian gold mine where the working expenses are $3 and
the fixed charges are at the low rate of 30 cents per ton, one
unit of increased production would show a saving of over $10,000
per annum from the saving of fixed charges. In about three years
this sum would repay the cost of the additional treatment equipment.
If further shaft capacity were required, the period would be much
extended. On a Western copper mine, where the costs are $8 and the
fixed charges are 80 cents per ton, one unit of increased production
would effect a saving of the fixed charges equal to the cost of
the extra unit in about three years. That is, the total sum would
amount to $80,000, or enough to provide almost any type of mechanical
equipment for such additional tonnage.
The first result of vigorous development is to increase the ore in
sight,--the visible life of the mine. When such visible life has
been so lengthened that the period in which the "saving of fixed
charges" will equal the amount involved in expansion of equipment,
then from the standpoint of this limitation only is the added
installation justified. The equipment if expanded on this practice
will grow upon the heels of rapid development until the maximum
production from the mine is reached, and a kind of equilibrium
establishes itself.
Conversely, this argument leads to the conclusion that, regardless
of other considerations, an equipment, and therefore output, should
not be expanded beyond the redemption by way of "saving from fixed
charges" of the visible or certain life of the mine. In those mines,
such as at the Witwatersrand, where there is a fairly sound assurance
of definite life, it is possible to calculate at once the size of
plant which by saving of "fixed charges" will be eventually the
most economical, but even here the other limitations step in to
vitiate such policy of management,--chiefly the limitation through
security of investment.
LIFE OF THE MINE.--If carried to its logical extreme, the above
program means a most rapid exhaustion of the mine. The maximum output
will depend eventually upon the rapidity with which development
work may be extended. As levels and other subsidiary development
openings can be prepared in inclined deposits much more quickly
than the shaft can be sunk, the critical point is the shaft-sinking.
As a shaft may by exertion be deepened at least 400 feet a year on
a going mine, the provision of an equipment to eat up the ore-body
at this rate of sinking means very early exhaustion indeed. In
fact, had such a theory of production been put into practice by
our forefathers, the mining profession might find difficulty in
obtaining employment to-day. Such rapid exhaustion would mean a
depletion of the mineral resources of the state at a pace which
would be alarming.
MECHANICAL INEFFICIENCY OF PATCHWORK PLANT.--Mine equipments on
speculative mines (the vast majority) are often enough patchwork,
for they usually grow from small beginnings; but any scheme of
expansion based upon the above doctrine would need to be modified
to the extent that additions could be in units large in ratio to
previous installations, or their patchwork character would be still
further accentuated. It would be impossible to maintain mechanical
efficiency under detail expansion.
OVERPRODUCTION OF BASE METAL.--Were this intensity of production of
general application to base metal mines it would flood the markets,
and, by an overproduction of metal depress prices to a point where
the advantages of such large-scale operations would quickly vanish.
The theoretical solution in this situation would be, if metals
fell below normal prices, let the output be reduced, or let the
products be stored until the price recovers. From a practical point
of view either alternative is a policy difficult to face.
In the first case, reduction of output means an increase of working
expenses by the spread of fixed charges over less tonnage, and
this in the face of reduced metal prices. It may be contended,
however, that a falling metal market is usually the accompaniment
of a drop in all commodities, wherefore working costs can be reduced
somewhat in such times of depression, thereby partially compensating
the other elements making for increased costs. Falls in commodities
are also the accompaniment of hard times. Consideration of one's
workpeople and the wholesale slaughter of dividends to the then
needy stockholders, resulting from a policy of reduced production,
are usually sufficient deterrents to diminished output.
The second alternative, that of storing metal, means equally a
loss of dividends by the investment of a large sum in unrealized
products, and the interest on this sum. The detriment to the market
of large amounts of unsold metal renders such a course not without
further disadvantages.
SECURITY OF INVESTMENT.--Another point of view antagonistic to
such wholesale intensity of production, and one worthy of careful
consideration, is that of the investor in mines. The root-value of
mining stocks is, or should be, the profit in sight. If the policy
of greatest economy in production costs be followed as outlined
above, the economic limit of ore-reserves gives an apparently very
short life, for the ore in sight will never represent a life beyond
the time required to justify more plant. Thus the "economic limit
of ore in reserve" will be a store equivalencing a period during
which additional equipment can be redeemed from the "saving of
fixed charges," or three or four years, usually.
The investor has the right to say that he wants the guarantee of
longer life to his investment,--he will in effect pay insurance for
it by a loss of some ultimate profit. That this view, contradictory
to the economics of the case, is not simply academic, can be observed
by any one who studies what mines are in best repute on any stock
exchange. All engineers must wish to have the industry under them
in high repute. The writer knows of several mines paying 20% on
their stocks which yet stand lower in price on account of short
ore-reserves than mines paying less annual returns. The speculator,
who is an element not to be wholly disregarded, wishes a rise in
his mining stock, and if development proceeds at a pace in advance
of production, he will gain a legitimate rise through the increase
in ore-reserves.
The investor's and speculator's idea of the desirability of a proved
long life readily supports the technical policy of high-pressure
development work, but not of expansion of production, for they
desire an increasing ore-reserve. Even the metal operator who is
afraid of overproduction does not object to increased ore-reserves.
On the point of maximum intensity of development work in a mine all
views coincide. The mining engineer, if he takes a Machiavellian
view, must agree with the investor and the metal dealer, for the
engineer is a "fixed charge" the continuance of which is important
to his daily needs.
The net result of all these limitations is therefore an invariable
compromise upon some output below the possible maximum. The initial
output to be contemplated is obviously one upon which the working
costs will be low enough to show a margin of profit. The medium
between these two extremes is determinable by a consideration of
the limitations set out,--and the cash available. When the volume
of output is once determined, it must be considered as a factor
in valuation, as discussed under "Amortization."
CHAPTER XVI.
Administration.
LABOR EFFICIENCY; SKILL; INTELLIGENCE; APPLICATION COORDINATION;
CONTRACT WORK; LABOR UNIONS; REAL BASIS OF WAGES.
The realization from a mine of the profits estimated from the other
factors in the case is in the end dependent upon the management.
Good mine management is based upon three elementals: first, sound
engineering; second, proper coördination and efficiency of every human
unit; third, economy in the purchase and consumption of supplies.
The previous chapters have been devoted to a more or less extended
exposition of economic engineering. While the second and third
requirements are equally important, they range in many ways out of
the engineering and into the human field. For this latter reason
no complete manual will ever be published upon "How to become a
Good Mine Manager."
It is purposed, however, to analyze some features of these second
and third fundamentals, especially in their interdependent phases,
and next to consider the subject of mine statistics, for the latter
are truly the microscopes through which the competence of the
administration must be examined.
The human units in mine organization can be divided into officers
and men. The choice of mine officers is the assembling of specialized
brains. Their control, stimulation, and inspiration is the main work
of the administrative head. Success in the selection and control of
staff is the index of executive ability. There are no mathematical,
mechanical, or chemical formulas for dealing with the human mind
or human energies.
LABOR.--The whole question of handling labor can be reduced to
the one term "efficiency." Not only does the actual labor outlay
represent from 60 to 70% of the total underground expenses, but
the capacity or incapacity of its units is responsible for wider
fluctuations in production costs than the bare predominance in
expenditure might indicate. The remaining expense is for supplies,
such as dynamite, timber, steel, power, etc., and the economical
application of these materials by the workman has the widest bearing
upon their consumption.
Efficiency of the mass is the resultant of that of each individual
under a direction which coördinates effectively all units. The
lack of effectiveness in one individual diminishes the returns
not simply from that man alone; it lowers the results from numbers
of men associated with the weak member through the delaying and
clogging of their work, and of the machines operated by them.
Coördination of work is a necessary factor of final efficiency. This
is a matter of organization and administration. The most zealous
stoping-gang in the world if associated with half the proper number
of truckers must fail to get the desired result.
Efficiency in the single man is the product of three factors,--skill,
intelligence, and application. A great proportion of underground
work in a mine is of a type which can be performed after a fashion
by absolutely unskilled and even unintelligent men, as witness the
breaking-in of savages of low average mentality, like the South
African Kaffirs. Although most duties can be performed by this
crudest order of labor, skill and intelligence can be applied to
it with such economic results as to compensate for the difference
in wage. The reason for this is that the last fifty years have seen
a substitution of labor-saving machines for muscle. Such machines
displace hundreds of raw laborers. Not only do they initially cost
large sums, but they require large expenditure for power and up-keep.
These fixed charges against the machine demand that it shall be
worked at its maximum. For interest, power, and up-keep go on in
any event, and the saving on crude labor displaced is not so great
but that it quickly disappears if the machine is run under its
capacity. To get its greatest efficiency, a high degree of skill
and intelligence is required. Nor are skill and intelligence alone
applicable to labor-saving devices themselves, because drilling and
blasting rock and executing other works underground are matters
in which experience and judgment in the individual workman count
to the highest degree.
How far skill affects production costs has had a thorough demonstration
in West Australia. For a time after the opening of those mines
only a small proportion of experienced men were obtainable. During
this period the rock broken per man employed underground did not
exceed the rate of 300 tons a year. In the large mines it has now,
after some eight years, attained 600 to 700 tons.
How far intelligence is a factor indispensable to skill can be well
illustrated by a comparison of the results obtained from working
labor of a low mental order, such as Asiatics and negroes, with those
achieved by American or Australian miners. In a general way, it may
be stated with confidence that the white miners above mentioned
can, under the same physical conditions, and with from five to ten
times the wage, produce the same economic result,--that is, an
equal or lower cost per unit of production. Much observation and
experience in working Asiatics and negroes as well as Americans
and Australians in mines, leads the writer to the conclusion that,
averaging actual results, one white man equals from two to three
of the colored races, even in the simplest forms of mine work such
as shoveling or tramming. In the most highly skilled branches,
such as mechanics, the average ratio is as one to seven, or in
extreme cases even eleven. The question is not entirely a comparison
of bare efficiency individually; it is one of the sum total of
results. In mining work the lower races require a greatly increased
amount of direction, and this excess of supervisors consists of
men not in themselves directly productive. There is always, too,
a waste of supplies, more accidents, and more ground to be kept
open for accommodating increased staff, and the maintenance of
these openings must be paid for. There is an added expense for
handling larger numbers in and out of the mine, and the lower
intelligence reacts in many ways in lack of coördination and inability
to take initiative. Taking all divisions of labor together, the
ratio of efficiency as measured in amount of output works out from
four to five colored men as the equivalent of one white man of the
class stated. The ratio of costs, for reasons already mentioned,
and in other than quantity relation, figures still more in favor
of the higher intelligence.
The following comparisons, which like all mine statistics must
necessarily be accepted with reservation because of some dissimilarity
of economic surroundings, are yet on sufficiently common ground
to demonstrate the main issue,--that is, the bearing of inherent
intelligence in the workmen and their consequent skill. Four groups
of gold mines have been taken, from India, West Australia, South
Africa, and Western America. All of those chosen are of the same
stoping width, 4 to 5 feet. All are working in depth and with every
labor-saving device available. All dip at about the same angle and
are therefore in much the same position as to handling rock. The
other conditions are against the white-manned mines and in favor of
the colored. That is, the Indian mines have water-generated electric
power and South Africa has cheaper fuel than either the American or
Australian examples. In both the white-manned groups, the stopes
are supported, while in the others no support is required.
=======================================================================
| Tons of | Average |Tons |
| Material | Number of Men | per |Cost per
Group of Mines | Excavated | Employed | Man | Ton of
|over Period|---------------| per |Material
|Compiled[5]|Colored| White |Annum| Broken
----------------------------|-----------|-------|-------|-----|--------
Four Kolar mines[1] | 963,950 | 13,611| 302 | 69.3| $3.85
Six Australian mines[2] | 1,027,718 | -- | 1,534 |669.9| 2.47
Three Witwatersrand mines[3]| 2,962,640 | 13,560| 1,595 |195.5| 2.68
Five American mines[4] | 1,089,500 | -- | 1,524 |713.3| 1.92
=======================================================================
[Footnote 1: Indian wages average about 20 cents per day.]
[Footnote 2: White men's wages average about $3 per day.]
[Footnote 3: About two-fifths of the colored workers were negroes,
and three-fifths Chinamen. Negroes average about 60 cents, and
Chinamen about 45 cents per day, including keep.]
[Footnote 4: Wages about $3.50. Tunnel entry in two mines.]
[Footnote 5: Includes rock broken in development work.
In the case of the specified African mines, the white labor is
employed almost wholly in positions of actual or semi-superintendence,
such as one white man in charge of two or three drills.
In the Indian case, in addition to the white men who are wholly
in superintendence, there were of the natives enumerated some 1000
in positions of semi-superintendence, as contractors or headmen,
working-gangers, etc.]
One issue arises out of these facts, and that is that no engineer
or investor in valuing mines is justified in anticipating lower
costs in regions where cheap labor exists.
In supplement to sheer skill and intelligence, efficiency can be
gained only by the application of the man himself. A few months ago
a mine in California changed managers. The new head reduced the number
employed one-third without impairing the amount of work accomplished.
This was not the result of higher skill or intelligence in the men,
but in the manager. Better application and coördination were secured
from the working force. Inspiration to increase of exertion is
created less by "driving" than by recognition of individual effort,
in larger pay, and by extending justifiable hope of promotion. A
great factor in the proficiency of the mine manager is his ability
to create an _esprit-de-corps_ through the whole staff, down to
the last tool boy. Friendly interest in the welfare of the men
and stimulation by competitions between various works and groups
all contribute to this end.
CONTRACT WORK.--The advantage both to employer and employed of
piece work over wage needs no argument. In a general way, contract
work honorably carried out puts a premium upon individual effort,
and thus makes for efficiency. There are some portions of mine
work which cannot be contracted, but the development, stoping,
and trucking can be largely managed in this way, and these items
cover 65 to 75% of the total labor expenditure underground.
In development there are two ways of basing contracts,--the first
on the footage of holes drilled, and the second on the footage
of heading advanced. In contract-stoping there are four methods
depending on the feet of hole drilled, on tonnage, on cubic space,
and on square area broken.
All these systems have their rightful application, conditioned upon
the class of labor and character of the deposit.
In the "hole" system, the holes are "pointed" by some mine official
and are blasted by a special crew. The miner therefore has little
interest in the result of the breaking. If he is a skilled white
man, the hours which he has wherein to contemplate the face usually
enable him to place holes to better advantage than the occasional
visiting foreman. With colored labor, the lack of intelligence in
placing holes and blasting usually justifies contracts per "foot
drilled." Then the holes are pointed and blasted by superintending
men.
On development work with the foot-hole system, unless two working
faces can be provided for each contracting party, they are likely
to lose time through having finished their round of holes before the
end of the shift. As blasting must be done outside the contractor's
shifts, it means that one shift per day must be set aside for the
purpose. Therefore not nearly such progress can be made as where
working the face with three shifts. For these reasons, the "hole"
system is not so advantageous in development as the "foot of advance"
basis.
In stoping, the "hole" system has not only a wider, but a sounder
application. In large ore-bodies where there are waste inclusions,
it has one superiority over any system of excavation measurement,
namely, that the miner has no interest in breaking waste into the
ore.
The plan of contracting stopes by the ton has the disadvantage
that either the ore produced by each contractor must be weighed
separately, or truckers must be trusted to count correctly, and to
see that the cars are full. Moreover, trucks must be inspected for
waste,--a thing hard to do underground. So great are these detailed
difficulties that many mines are sending cars to the surface in
cages when they should be equipped for bin-loading and self-dumping
skips.
The method of contracting by the cubic foot of excavation saves
all necessity for determining the weight of the output of each
contractor. Moreover, he has no object in mixing waste with the ore,
barring the breaking of the walls. This system therefore requires
the least superintendence, permits the modern type of hoisting,
and therefore leaves little justification for the survival of the
tonnage basis.
Where veins are narrow, stoping under contract by the square foot
or fathom measured parallel to the walls has an advantage. The miner
has no object then in breaking wall-rock, and the thoroughness of
the ore-extraction is easily determined by inspection.
BONUS SYSTEMS.--By giving cash bonuses for special accomplishment,
much the same results can be obtained in some departments as by
contracting. A bonus per foot of heading gained above a minimum,
or an excess of trucks trammed beyond a minimum, or prizes for
the largest amount done during the week or month in special works
or in different shifts,--all these have a useful application in
creating efficiency. A high level of results once established is
easily maintained.
LABOR UNIONS.--There is another phase of the labor question which
must be considered and that is the general relations of employer
and employed. In these days of largely corporate proprietorship,
the owners of mines are guided in their relations with labor by
engineers occupying executive positions. On them falls the
responsibility in such matters, and the engineer becomes thus a
buffer between labor and capital. As corporations have grown, so
likewise have the labor unions. In general, they are normal and
proper antidotes for unlimited capitalistic organization.
Labor unions usually pass through two phases. First, the inertia
of the unorganized labor is too often stirred only by demagogic
means. After organization through these and other agencies, the
lack of balance in the leaders often makes for injustice in demands,
and for violence to obtain them and disregard of agreements entered
upon. As time goes on, men become educated in regard to the rights
of their employers, and to the reflection of these rights in ultimate
benefit to labor itself. Then the men, as well as the intelligent
employer, endeavor to safeguard both interests. When this stage
arrives, violence disappears in favor of negotiation on economic
principles, and the unions achieve their greatest real gains. Given
a union with leaders who can control the members, and who are disposed
to approach differences in a business spirit, there are few sounder
positions for the employer, for agreements honorably carried out
dismiss the constant harassments of possible strikes. Such unions
exist in dozens of trades in this country, and they are entitled to
greater recognition. The time when the employer could ride roughshod
over his labor is disappearing with the doctrine of "_laissez faire_,"
on which it was founded. The sooner the fact is recognized, the
better for the employer. The sooner some miners' unions develop
from the first into the second stage, the more speedily will their
organizations secure general respect and influence.[*]
[Footnote *: Some years of experience with compulsory arbitration
in Australia and New Zealand are convincing that although the law
there has many defects, still it is a step in the right direction,
and the result has been of almost unmixed good to both sides. One
of its minor, yet really great, benefits has been a considerable
extinction of the parasite who lives by creating violence.]
The crying need of labor unions, and of some employers as well,
is education on a fundamental of economics too long disregarded
by all classes and especially by the academic economist. When the
latter abandon the theory that wages are the result of supply and
demand, and recognize that in these days of international flow of
labor, commodities and capital, the real controlling factor in
wages is efficiency, then such an educational campaign may become
possible. Then will the employer and employee find a common ground
on which each can benefit. There lives no engineer who has not
seen insensate dispute as to wages where the real difficulty was
inefficiency. No administrator begrudges a division with his men
of the increased profit arising from increased efficiency. But
every administrator begrudges the wage level demanded by labor
unions whose policy is decreased efficiency in the false belief
that they are providing for more labor.
CHAPTER XVII.
Administration (_Continued_).
ACCOUNTS AND TECHNICAL DATA AND REPORTS; WORKING COSTS; DIVISION
OF EXPENDITURE; INHERENT LIMITATIONS IN ACCURACY OF WORKING COSTS;
WORKING COST SHEETS. GENERAL TECHNICAL DATA; LABOR, SUPPLIES, POWER,
SURVEYS, SAMPLING, AND ASSAYING.
First and foremost, mine accounts are for guidance in the distribution
of expenditure and in the collection of revenue; secondly, they
are to determine the financial progress of the enterprise, its
profit or loss; and thirdly, they are to furnish statistical data to
assist the management in its interminable battle to reduce expenses
and increase revenue, and to enable the owner to determine the
efficiency of his administrators. Bookkeeping _per se_ is no part
of this discussion. The fundamental purpose of that art is to cover
the first two objects, and, as such, does not differ from its
application to other commercial concerns.
In addition to these accounting matters there is a further type
of administrative report of equal importance--that is the periodic
statements as to the physical condition of the property, the results
of exploration in the mine, and the condition of the equipment.
ACCOUNTS.
The special features of mine accounting reports which are a development
to meet the needs of this particular business are the determination
of working costs, and the final presentation of these data in a
form available for comparative purposes.
The subject may be discussed under:--
1. Classes of mine expenditure.
2. Working costs.
3. The dissection of expenditures departmentally.
4. Inherent limitations in the accuracy of working costs.
5. Working cost sheets.
In a wide view, mine expenditures fall into three classes, which
maybe termed the "fixed charges," "proportional charges," and "suspense
charges" or "capital expenditure." "Fixed charges" are those which,
like pumping and superintendence, depend upon time rather than
tonnage and material handled. They are expenditures that would not
decrease relatively to output. "Proportional charges" are those
which, like ore-breaking, stoping, supporting stopes, and tramming,
are a direct coefficient of the ore extracted. "Suspense charges" are
those which are an indirect factor of the cost of the ore produced,
such as equipment and development. These expenditures are preliminary
to output, and they thus represent a storage of expense to be charged
off when the ore is won. This outlay is often called "capital
expenditure." Such a term, though in common use, is not strictly
correct, for the capital value vanishes when the ore is extracted,
but in conformity with current usage the term "capital expenditure"
will be adopted.
Except for the purpose of special inquiry, such as outlined under
the chapter on "Ratio of Output," "fixed charges" are not customarily
a special division in accounts. In a general way, such expenditures,
combined with the "proportional charges," are called "revenue
expenditure," as distinguished from the capital, or "suspense,"
expenditures. In other words, "revenue" expenditures are those
involved in the daily turnover of the business and resulting in
immediate returns. The inherent difference in character of revenue
and capital expenditures is responsible for most of the difficulties
in the determination of working costs, and most of the discussion
on the subject.
WORKING COSTS.--"Working costs" are a division of expenditure for
some unit,--the foot of opening, ton of ore, a pound of metal,
cubic yard or fathom of material excavated, or some other measure.
The costs per unit are usually deduced for each month and each
year. They are generally determined for each of the different
departments of the mine or special works separately. Further, the
various sorts of expenditure in these departments are likewise
segregated.
In metal mining the ton is the universal unit of distribution for
administrative purpose, although the pound of metal is often used
to indicate final financial results. The object of determination of
"working costs" is fundamentally for comparative purposes. Together
with other technical data, they are the nerves of the administration,
for by comparison of detailed and aggregate results with other mines
and internally in the same mine, over various periods and between
different works, a most valuable check on efficiency is possible.
Further, there is one collateral value in all statistical data not
to be overlooked, which is that the knowledge of its existence
induces in the subordinate staff both solicitude and emulation.
The fact must not be lost sight of, however, that the wide variations
in physical and economic environment are so likely to vitiate
conclusions from comparisons of statistics from two mines or from
two detailed works on the same mine, or even from two different
months on the same work, that the greatest care and discrimination
are demanded in their application. Moreover, the inherent difficulties
in segregating and dividing the accounts which underlie such data,
render it most desirable to offer some warning regarding the limits
to which segregation and division may be carried to advantage.
As working costs are primarily for comparisons, in order that they
may have value for this purpose they must include only such items
of expenditure as will regularly recur. If this limitation were more
generally recognized, a good deal of dispute and polemics on the
subject might be saved. For this reason it is quite impossible that
all the expenditure on the mine should be charged into working costs,
particularly some items that arise through "capital expenditure."
THE DISSECTION OF EXPENDITURES DEPARTMENTALLY.--The final division
in the dissection of the mine expenditure is in the main:--
/(1) General Expenses. / Ore-breaking. \
| | Supporting Stopes. | Various
_Revenue._< (2) Ore Extraction. < Trucking Ore. | expenditures
| \ Hoisting. | for labor,
\(3) Pumping. | supplies, power,
/ Shaft-sinking. | repairs, etc.,
| Station-cutting. > worked out per
| Crosscutting. | ton or foot
/(4) Development. < Driving. | advanced
_Capital | | Rising. | over each
or < | Winzes. | department.
Suspense._ | \ Diamond Drilling. /
|
| (5) Construction and \ Various Works.
\ Equipment. /
The detailed dissection of expenditures in these various departments
with view to determine amount of various sorts of expenditure over
the department, or over some special work in that department, is
full of unsolvable complications. The allocation of the direct
expenditure of labor and supplies applied to the above divisions or
special departments in them, is easily accomplished, but beyond this
point two sorts of difficulties immediately arise and offer infinite
field for opinion and method. The first of these difficulties arises
from supplementary departments on the mine, such as "power," "repairs
and maintenance," "sampling and assaying." These departments must
be "spread" over the divisions outlined above, for such charges
are in part or whole a portion of the expense of these divisions.
Further, all of these "spread" departments are applied to surface
as well as to underground works, and must be divided not only over
the above departments but also over the surface departments,--not
under discussion here. The common method is to distribute "power" on
a basis of an approximation of the amount used in each department;
to distribute "repairs and maintenance," either on a basis of shop
returns, or a distribution over all departments on the basis of
the labor employed in those departments, on the theory that such
repairs arise in this proportion; to distribute sampling and assaying
over the actual points to which they relate at the average cost
per sample or assay.
"General expenses," that is, superintendence, etc., are often not
included in the final departments as above, but are sometimes "spread"
in an attempt to charge a proportion of superintendence to each
particular work. As, however, such "spreading" must take place
on the basis of the relative expenditure in each department, the
result is of little value, for such a basis does not truly represent
the proportion of general superintendence, etc., devoted to each
department. If they are distributed over all departments, capital
as well as revenue, on the basis of total expenditure, they inflate
the "capital expenditure" departments against a day of reckoning when
these charges come to be distributed over working costs. Although it
may be contended that the capital departments also require supervision,
such a practice is a favorite device for showing apparently low
working costs in the revenue departments. The most courageous way
is not to distribute general expenses at all, but to charge them
separately and directly to revenue accounts and thus wholly into
working costs.
The second problem is to reduce the "suspense" or capital charges
to a final cost per ton, and this is no simple matter. Development
expenditures bear a relation to the tonnage developed and not to
that extracted in any particular period. If it is desired to preserve
any value for comparative purposes in the mining costs, such outlay
must be charged out on the basis of the tonnage developed, and such
portion of the ore as is extracted must be written off at this
rate; otherwise one month may see double the amount of development
in progress which another records, and the underground costs would
be swelled or diminished thereby in a way to ruin their comparative
value from month to month. The ore developed cannot be satisfactorily
determined at short intervals, but it can be known at least annually,
and a price may be deduced as to its cost per ton. In many mines
a figure is arrived at by estimating ore-reserves at the end of
the year, and this figure is used during the succeeding year as a
"redemption of development" and as such charged to working costs,
and thus into revenue account in proportion to the tonnage extracted.
This matter is further elaborated in some mines, in that winzes
and rises are written off at one rate, levels and crosscuts at
another, and shafts at one still lower, on the theory that they
lost their usefulness in this progression as the ore is extracted.
This course, however, is a refinement hardly warranted.
Plant and equipment constitute another "suspense" account even
harder to charge up logically to tonnage costs, for it is in many
items dependent upon the life of the mine, which is an unknown
factor. Most managers debit repairs and maintenance directly to
the revenue account and leave the reduction of the construction
outlay to an annual depreciation on the final balance sheet, on the
theory that the plant is maintained out of costs to its original
value. This subject will be discussed further on.
INHERENT LIMITATIONS IN ACCURACY OF WORKING COSTS.--There are three
types of such limitations which arise in the determination of costs
and render too detailed dissection of such costs hopeless of accuracy
and of little value for comparative purposes. They are, first, the
difficulty of determining all of even direct expenditure on any
particular crosscut, stope, haulage, etc.; second, the leveling effect
of distributing the "spread" expenditures, such as power, repairs,
etc.; and third, the difficulties arising out of the borderland
of various departments.
Of the first of these limitations the instance may be cited that
foremen and timekeepers can indicate very closely the destination of
labor expense, and also that of some of the large items of supply,
such as timber and explosives, but the distribution of minor supplies,
such as candles, drills, picks, and shovels, is impossible of accurate
knowledge without an expense wholly unwarranted by the information
gained. To determine at a particular crosscut the exact amount of
steel, and of tools consumed, and the cost of sharpening them,
would entail their separate and special delivery to the same place
of attack and a final weighing-up to learn the consumption.
Of the second sort of limitations, the effect of "spread" expenditure,
the instance may be given that the repairs and maintenance are done by
many men at work on timbers, tracks, machinery, etc. It is hopeless
to try and tell how much of their work should be charged specifically
to detailed points. In the distribution of power may be taken the
instance of air-drills. Although the work upon which the drill is
employed can be known, the power required for compression usually
comes from a common power-plant, so that the portion of power debited
to the air compressor is an approximation. The assumption of an
equal consumption of air by all drills is a further approximation.
In practice, therefore, many expenses are distributed on the theory
that they arise in proportion to the labor employed, or the machines
used in the various departments. The net result is to level down
expensive points and level up inexpensive ones.
The third sort of limitation of accounting difficulty referred
to, arises in determining into which department are actually to be
allocated the charges which lie in the borderland between various
primary classes of expenditure. For instance, in ore won from
development,--in some months three times as much development may
be in ore as in other months. If the total expense of development
work which yields ore be charged to stoping account, and if cost
be worked out on the total tonnage of ore hoisted, then the stoping
cost deduced will be erratic, and the true figures will be obscured.
On the other hand, if all development is charged to 'capital account'
and the stoping cost worked out on all ore hoisted, it will include
a fluctuating amount of ore not actually paid for by the revenue
departments or charged into costs. This fluctuation either way
vitiates the whole comparative value of the stoping costs. In the
following system a compromise is reached by crediting "development"
with an amount representing the ore won from development at the
average cost of stoping, and by charging this amount into "stoping."
A number of such questions arise where the proper division is simply
a matter of opinion.
The result of all these limitations is that a point in detail is
quickly reached where no further dissection of expenditure is justified,
since it becomes merely an approximation. The writer's own impression
is that without an unwarrantable number of accountants, no manager
can tell with any accuracy the cost of any particular stope, or
of any particular development heading. Therefore, aside from some
large items, such detailed statistics, if given, are to be taken
with great reserve.
WORKING COST SHEETS.--There are an infinite number of forms of
working cost sheets, practically every manager having a system of
his own. To be of greatest value, such sheets should show on their
face the method by which the "spread" departments are handled, and
how revenue and suspense departments are segregated. When too much
detail is presented, it is but a waste of accounting and consequent
expense. Where to draw the line in this regard is, however, a matter
of great difficulty. No cost sheet is entirely satisfactory. The
appended sheet is in use at a number of mines. It is no more perfect
than many others. It will be noticed that the effect of this system
is to throw the general expenses into the revenue expenditures,
and as little as possible into the "suspense" account.
GENERAL TECHNICAL DATA.
For the purposes of efficient management, the information gathered
under this head is of equal, if not superior, importance to that
under "working costs." Such data fall generally under the following
heads:--
LABOR.--Returns of the shifts worked in the various departments
for each day and for the month; worked out on a monthly basis of
footage progress, tonnage produced or tons handled per man; also
where possible the footage of holes drilled, worked out per man
and per machine.
SUPPLIES.--Daily returns of supplies used; the principal items
worked out monthly in quantity per foot of progress, or per ton
of ore produced.
POWER.--Fuel, lubricant, etc., consumed in steam production, worked
out into units of steam produced, and this production allocated to
the various engines. Where electrical power is used, the consumption
of the various motors is set out.
SURVEYS.--The need of accurate plans requires no discussion. Aside
from these, the survey-office furnishes the returns of development
footage, measurements under contracts, and the like.
SAMPLING AND ASSAYING.--Mine sampling and assaying fall under two
heads,--the determination of the value of standing ore, and of
products from the mine. The sampling and assaying on a going mine
call for the same care and method as in cases of valuation of the
mine for purchase,--the details of which have been presented under
"Mine Valuation,"--for through it, guidance must not only be had to
the value of the mine and for reports to owners, but the detailed
development and ore extraction depend on an absolute knowledge of
where the values lie.
CHAPTER XVIII.
ADMINISTRATION (_Concluded_).
ADMINISTRATIVE REPORTS.
In addition to financial returns showing the monthly receipts,
expenditures, and working costs, there must be in proper administration
periodic reports from the officers of the mine to the owners or
directors as to the physical progress of the enterprise. Such reports
must embrace details of ore extraction, metal contents, treatment
recoveries, construction of equipment, and the results of underground
development. The value of mines is so much affected by the monthly
or even daily result of exploration that reports of such work are
needed very frequently,--weekly or even daily if critical work is
in progress. These reports must show the width, length, and value
of the ore disclosed.
The tangible result of development work is the tonnage and grade
of ore opened up. How often this stock-taking should take place
is much dependent upon the character of the ore. The result of
exploration in irregular ore-bodies often does not, over short
periods, show anything tangible in definite measurable tonnage,
but at least annually the ore reserve can be estimated.
In mines owned by companies, the question arises almost daily as
to how much of and how often the above information should be placed
before stockholders (and therefore the public) by the directors. In
a general way, any company whose shares are offered on the stock
exchange is indirectly inviting the public to become partners in the
business, and these partners are entitled to all the information
which affects the value of their property and are entitled to it
promptly. Moreover, mining is a business where competition is so
obscure and so much a matter of indifference, that suppression
of important facts in documents for public circulation has no
justification. On the other hand, both the technical progress of
the industry and its position in public esteem demand the fullest
disclosure and greatest care in preparation of reports. Most
stockholders' ignorance of mining technology and of details of
their particular mine demands a great deal of care and discretion
in the preparation of these public reports that they may not be
misled. Development results may mean little or much, depending
upon the location of the work done in relation to the ore-bodies,
etc., and this should be clearly set forth.
The best opportunity of clear, well-balanced statements lies in
the preparation of the annual report and accounts. Such reports
are of three parts:--
1. The "profit and loss" account, or the "revenue account."
2. The balance sheet; that is, the assets and liabilities
statement.
3. The reports of the directors, manager, and consulting
engineer.
The first two items are largely matters of bookkeeping. They or
the report should show the working costs per ton for the year.
What must be here included in costs is easier of determination
than in the detailed monthly cost sheets of the administration;
for at the annual review, it is not difficult to assess the amount
chargeable to development. Equipment expenditure, however, presents
an annual difficulty, for, as said, the distribution of this item
is a factor of the life of the mine, and that is unknown. If such
a plant has been paid for out of the earnings, there is no object
in carrying it on the company's books as an asset, and most
well-conducted companies write it off at once. On the other hand,
where the plant is paid for out of capital provided for the purpose,
even to write off depreciation means that a corresponding sum of
cash must be held in the company's treasury in order to balance
the accounts,--in other words, depreciation in such an instance
becomes a return of capital. The question then is one of policy
in the company's finance, and in neither case is it a matter which
can be brought into working costs and leave them any value for
comparative purposes. Indeed, the true cost of working the ore
from any mine can only be told when the mine is exhausted; then
the dividends can be subtracted from the capital sunk and metal
sold, and the difference divided over the total tonnage produced.
The third section of the report affords wide scope for the best
efforts of the administration. This portion of the report falls
into three divisions: (_a_) the construction and equipment work
of the year, (_b_) the ore extraction and treatment, and (_c_)
the results of development work.
The first requires a statement of the plant constructed, its object
and accomplishment; the second a disclosure of tonnage produced,
values, metallurgical and mechanical efficiency. The third is of
the utmost importance to the stockholder, and is the one most often
disregarded and obscured. Upon this hinges the value of the property.
There is no reason why, with plans and simplicity of terms, such
reports cannot be presented in a manner from which the novice can
judge of the intrinsic position of the property. A statement of
the tonnage of ore-reserves and their value, or of the number of
years' supply of the current output, together with details of ore
disclosed in development work, and the working costs, give the
ground data upon which any stockholder who takes interest in his
investment may judge for himself. Failure to provide such data
will some day be understood by the investing public as a _prima
facie_ index of either incapacity or villainy. By the insistence of
the many engineers in administration of mines upon the publication
of such data, and by the insistence of other engineers upon such
data for their clients before investment, and by the exposure of
the delinquents in the press, a more practicable "protection of
investors" can be reached than by years of academic discussion.
CHAPTER XIX.
The Amount of Risk in Mining Investments.
RISK IN VALUATION OF MINES; IN MINES AS COMPARED WITH OTHER COMMERCIAL
ENTERPRISES.
From the constant reiteration of the risks and difficulties involved
in every step of mining enterprise from the valuation of the mine
to its administration as a going concern, the impression may be
gained that the whole business is one great gamble; in other words,
that the point whereat certainties stop and conjecture steps in
is so vital as to render the whole highly speculative.
Far from denying that mining is, in comparison with better-class
government bonds, a speculative type of investment, it is desirable
to avow and emphasize the fact. But it is none the less well to
inquire what degree of hazard enters in and how it compares with
that in other forms of industrial enterprise.
Mining business, from an investment view, is of two sorts,--prospecting
ventures and developed mines; that is, mines where little or no ore is
exposed, and mines where a definite quantity of ore is measurable or can
be reasonably anticipated. The great hazards and likewise the Aladdin
caves of mining are mainly confined to the first class. Although all
mines must pass through the prospecting stage, the great industry
of metal production is based on developed mines, and it is these
which should come into the purview of the non-professional investor.
The first class should be reserved invariably for speculators, and
a speculator may be defined as one who hazards all to gain much.
It is with mining as an investment, however, that this discussion
is concerned.
RISK IN VALUATION OF MINES.--Assuming a competent collection of
data and efficient management of the property, the risks in valuing
are from step to step:--
1. The risk of continuity in metal contents beyond sample
faces.
2. The risk of continuity in volume through the blocks
estimated.
3. The risk of successful metallurgical treatment.
4. The risk of metal prices, in all but gold.
5. The risk of properly estimating costs.
6. The risk of extension of the ore beyond exposures.
7. The risk of management.
As to the continuity of values and volumes through the estimated
area, the experience of hundreds of engineers in hundreds of mines
has shown that when the estimates are based on properly secured
data for "proved ore," here at least there is absolutely no hazard.
Metallurgical treatment, if determined by past experience on the
ore itself, carries no chance; and where determined by experiment,
the risk is eliminated if the work be sufficiently exhaustive. The
risk of metal price is simply a question of how conservative a
figure is used in estimating. It can be eliminated if a price low
enough be taken. Risk of extension in depth or beyond exposures
cannot be avoided. It can be reduced in proportion to the distance
assumed. Obviously, if no extension is counted, there is nothing
chanced. The risk of proper appreciation of costs is negligible where
experience in the district exists. Otherwise, it can be eliminated
if a sufficiently large allowance is taken. The risk of failure to
secure good management can be eliminated if proved men are chosen.
There is, therefore, a basic value to every mine. The "proved"
ore taken on known metallurgical grounds, under known conditions
of costs on minimum prices of metals, has a value as certain as
that of money in one's own vault. This is the value previously
referred to as the "_A_" value. If the price (and interest on it
pending recovery) falls within this amount, there is no question
that the mine is worth the price. What the risk is in mining is
simply what amount the price of the investment demands shall be
won from extension of the deposit beyond known exposures, or what
higher price of metal must be realized than that calculated in
the "_A_" value. The demands on this _X, Y_ portion of the mine
can be converted into tons of ore, life of production, or higher
prices, and these can be weighed with the geological weights and
the industrial outlook.
MINES COMPARED TO OTHER COMMERCIAL ENTERPRISES.--The profits from
a mining venture over and above the bed-rock value _A_, that is,
the return to be derived from more extensive ore-recovery and a
higher price of metal, may be compared to the value included in
other forms of commercial enterprise for "good-will." Such forms of
enterprise are valued on a basis of the amount which will replace
the net assets plus (or minus) an amount for "good-will," that is,
the earning capacity. This good-will is a speculation of varying
risk depending on the character of the enterprise. For natural
monopolies, like some railways and waterworks, the risk is less
and for shoe factories more. Even natural monopolies are subject
to the risks of antagonistic legislation and industrial storms.
But, eliminating this class of enterprise, the speculative value
of a good-will involves a greater risk than prospective value in
mines, if properly measured; because the dangers of competition
and industrial storms do not enter to such a degree, nor is the
future so dependent upon the human genius of the founder or manager.
Mining has reached such a stage of development as a science that
management proceeds upon comparatively well-known lines. It is
subject to known checks through the opportunity of comparisons
by which efficiency can be determined in a manner more open for
the investor to learn than in any other form of industry. While
in mining an estimate of a certain minimum of extension in depth,
as indicated by collateral factors, may occasionally fall short,
it will, in nine cases out of ten, be exceeded. If investment in
mines be spread over ten cases, similarly valued as to minimum of
extension, the risk has been virtually eliminated. The industry,
if reduced to the above basis for financial guidance, is a more
profitable business and is one of less hazards than competitive
forms of commercial enterprises.
In view of what has been said before, it may be unnecessary to refer
again to the subject, but the constant reiteration by wiseacres
that the weak point in mining investments lies in their short life
and possible loss of capital, warrants a repetition that the _A,
B, C_ of proper investment in mines is to be assured, by the "_A_"
value, of a return of the whole or major portion of the capital.
The risk of interest and profit may be deferred to the _X, Y_ value,
and in such case it is on a plane with "good-will." It should be said
at once to that class who want large returns on investment without
investigation as to merits, or assurance as to the management of the
business, that there is no field in this world for the employment
of their money at over 4%.
Unfortunately for the reputation of the mining industry, and metal
mines especially, the business is often not conducted or valued on
lines which have been outlined in these chapters. There is often
the desire to sell stocks beyond their value. There is always the
possibility that extension in depth will reveal a glorious Eldorado.
It occasionally does, and the report echoes round the world for years,
together with tributes to the great judgment of the exploiters. The
volume of sound allures undue numbers of the venturesome, untrained,
and ill-advised public to the business, together with a mob of
camp-followers whose objective is to exploit the ignorant by preying
on their gambling instincts. Thus a considerable section of metal
mining industry is in the hands of these classes, and a cloud of
disrepute hangs ever in the horizon.
There has been a great educational campaign in progress during the
past few years through the technical training of men for conduct
of the industry, by the example of reputable companies in regularly
publishing the essential facts upon which the value of their mines
is based, and through understandable nontechnical discussion in
and by some sections of the financial and general press. The real
investor is being educated to distinguish between reputable concerns
and the counters of gamesters. Moreover, yearly, men of technical
knowledge are taking a stronger and more influential part in mining
finance and in the direction of mining and exploration companies.
The net result of these forces will be to put mining on a better
plane.
CHAPTER XX.
The Character, Training, and Obligations of the Mining Engineering
Profession.
In a discussion of some problems of metal mining from the point
of view of the direction of mining operations it may not be amiss
to discuss the character of the mining engineering profession in
its bearings on training and practice, and its relations to the
public.
The most dominant characteristic of the mining engineering profession
is the vast preponderance of the commercial over the technical in
the daily work of the engineer. For years a gradual evolution has
been in progress altering the larger demands on this branch of the
engineering profession from advisory to executive work. The mining
engineer is no longer the technician who concocts reports and blue
prints. It is demanded of him that he devise the finance, construct
and manage the works which he advises. The demands of such executive
work are largely commercial; although the commercial experience
and executive ability thus become one pier in the foundation of
training, the bridge no less requires two piers, and the second
is based on technical knowledge. Far from being deprecated, these
commercial phases cannot be too strongly emphasized. On the other
hand, I am far from contending that our vocation is a business
rather than a profession.
For many years after the dawn of modern engineering, the members
of our profession were men who rose through the ranks of workmen,
and as a result, we are to this day in the public mind a sort of
superior artisan, for to many the engine-driver is equally an engineer
with the designer of the engine, yet their real relation is but as
the hand to the brain. At a later period the recruits entered by
apprenticeship to those men who had established their intellectual
superiority to their fellow-workers. These men were nearly always
employed in an advisory way--subjective to the executive head.
During the last few decades, the advance of science and the complication
of industry have demanded a wholly broader basis of scientific and
general training for its leaders. Executive heads are demanded who
have technical training. This has resulted in the establishment of
special technical colleges, and compelled a place for engineering
in the great universities. The high intelligence demanded by the
vocation itself, and the revolution in training caused by the
strengthening of its foundations in general education, has finally,
beyond all question, raised the work of application of science to
industry to the dignity of a profession on a par with the law,
medicine, and science. It demands of its members equally high mental
attainments,--and a more rigorous training and experience. Despite
all this, industry is conducted for commercial purposes, and leaves
no room for the haughty intellectual superiority assumed by some
professions over business callings.
There is now demanded of the mining specialist a wide knowledge
of certain branches of civil, mechanical, electrical, and chemical
engineering, geology, economics, the humanities, and what not; and
in addition to all this, engineering sense, executive ability,
business experience, and financial insight. Engineering sense is
that fine blend of honesty, ingenuity, and intuition which is a
mental endowment apart from knowledge and experience. Its possession
is the test of the real engineer. It distinguishes engineering as
a profession from engineering as a trade. It is this sense that
elevates the possessor to the profession which is, of all others,
the most difficult and the most comprehensive. Financial insight can
only come by experience in the commercial world. Likewise must come
the experience in technical work which gives balance to theoretical
training. Executive ability is that capacity to coördinate and command
the best results from other men,--it is a natural endowment. which
can be cultivated only in actual use.
The practice of mine engineering being so large a mixture of business,
it follows that the whole of the training of this profession cannot
be had in schools and universities. The commercial and executive
side of the work cannot be taught; it must be absorbed by actual
participation in the industry. Nor is it impossible to rise to
great eminence in the profession without university training, as
witness some of our greatest engineers. The university can do much;
it can give a broad basis of knowledge and mental training, and can
inculcate moral feeling, which entitles men to lead their fellows. It
can teach the technical fundamentals of the multifold sciences which
the engineer should know and must apply. But after the university
must come a schooling in men and things equally thorough and more
arduous.
In this predominating demand for commercial qualifications over
the technical ones, the mining profession has differentiated to
a great degree from its brother engineering branches. That this
is true will be most apparent if we examine the course through
which engineering projects march, and the demands of each stage
on their road to completion.
The life of all engineering projects in a general way may be divided
into five phases:[*]--
[Footnote *: These phases do not necessarily proceed step by step.
For an expanding works especially, all of them may be in process
at the same time, but if each item be considered to itself, this
is the usual progress, or should be when properly engineered.]
1. Determination of the value of the project.
2. Determination of the method of attack.
3. The detailed delineation of method, means, and tools.
4. The execution of the works.
5. The operation of the completed works.
These various stages of the resolution of an engineering project
require in each more or less of every quality of intellect, training,
and character. At the different stages, certain of these qualities
are in predominant demand: in the first stage, financial insight;
in the second, "engineering sense"; in the third, training and
experience; in the fourth and fifth, executive ability.
A certain amount of compass over the project during the whole
five stages is required by all branches of the engineering
profession,--harbor, canal, railway, waterworks, bridge, mechanical,
electrical, etc.; but in none of them so completely and in such
constant combination is this demanded as in mining.
The determination of the commercial value of projects is a greater
section of the mining engineer's occupation than of the other
engineering branches. Mines are operated only to earn immediate
profits. No question of public utility enters, so that all mining
projects have by this necessity to be from the first weighed from
a profit point of view alone. The determination of this question
is one which demands such an amount of technical knowledge and
experience that those who are not experts cannot enter the
field,--therefore the service of the engineer is always demanded in
their satisfactory solution. Moreover, unlike most other engineering
projects, mines have a faculty of changing owners several times
during their career, so that every one has to survive a periodic
revaluation. From the other branches of engineering, the electrical
engineer is the most often called upon to weigh the probabilities
of financial success of the enterprise, but usually his presence
in this capacity is called upon only at the initial stage, for
electrical enterprises seldom change hands. The mechanical and
chemical branches are usually called upon for purely technical
service on the demand of the operator, who decides the financial
problems for himself, or upon works forming but units in undertakings
where the opinion on the financial advisability is compassed by some
other branch of the engineering profession. The other engineering
branches, even less often, are called in for financial advice,
and in those branches involving works of public utility the
profit-and-loss phase scarcely enters at all.
Given that the project has been determined upon, and that the enterprise
has entered upon the second stage, that of determination of method of
attack, the immediate commercial result limits the mining engineer's
every plan and design to a greater degree than it does the other
engineering specialists. The question of capital and profit dogs
his every footstep, for all mines are ephemeral; the life of any
given mine is short. Metal mines have indeed the shortest lives of
any. While some exceptional ones may produce through one generation,
under the stress of modern methods a much larger proportion extend
only over a decade or two. But of more pertinent force is the fact
that as the certain life of a metal mine can be positively known in
most cases but a short period beyond the actual time required to
exhaust the ore in sight, not even a decade of life to the enterprise
is available for the estimates of the mining engineer. Mining works
are of no value when the mine is exhausted; the capital invested
must be recovered in very short periods, and therefore all mining
works must be of the most temporary character that will answer.
The mining engineer cannot erect a works that will last as long as
possible; it is to last as long as the mine only, and, in laying
it out, forefront in his mind must be the question, Can its cost
be redeemed in the period of use of which I am certain it will
find employment? If not, will some cheaper device, which gives
less efficiency, do? The harbor engineer, the railway engineer,
the mechanical engineer, build as solidly as they can, for the
demand for the work will exist till after their materials are worn
out, however soundly they construct.
Our engineer cousins can, in a greater degree by study and
investigation, marshal in advance the factors with which they have
to deal. The mining engineer's works, on the other hand, depend at
all times on many elements which, from the nature of things, must
remain unknown. No mine is laid bare to study and resolve in advance.
We have to deal with conditions buried in the earth. Especially in
metal mines we cannot know, when our works are initiated, what
the size, mineralization, or surroundings of the ore-bodies will
be. We must plunge into them and learn,--and repent. Not only is
the useful life of our mining works indeterminate, but the very
character of them is uncertain in advance. All our works must be in
a way doubly tentative, for they are subject to constant alterations
as they proceed.
Not only does this apply to our initial plans, but to our daily
amendment of them as we proceed into the unknown. Mining engineering
is, therefore, never ended with the initial determination of a method.
It is called upon daily to replan and reconceive, coincidentally with
the daily progress of the constructions and operation. Weary with
disappointment in his wisest conception, many a mining engineer
looks jealously upon his happier engineering cousin, who, when he
designs a bridge, can know its size, its strains, and its cost,
and can wash his hands of it finally when the contractor steps
in to its construction. And, above all, it is no concern of his
whether it will pay. Did he start to build a bridge over a water,
the width or depth or bottom of which he could not know in advance,
and require to get its cost back in ten years, with a profit, his
would be a task of similar harassments.
As said before, it is becoming more general every year to employ
the mining engineer as the executive head in the operation of mining
engineering projects, that is, in the fourth and fifth stages of
the enterprise. He is becoming the foreman, manager, and president
of the company, or as it may be contended by some, the executive
head is coming to have technical qualifications. Either way, in
no branch of enterprise founded on engineering is the operative
head of necessity so much a technical director. Not only is this
caused by the necessity of executive knowledge before valuations
can be properly done, but the incorporation of the executive work
with the technical has been brought about by several other forces.
We have a type of works which, by reason of the new conditions
and constant revisions which arise from pushing into the unknown
coincidentally with operating, demands an intimate continuous daily
employment of engineering sense and design through the whole history
of the enterprise. These works are of themselves of a character
which requires a constant vigilant eye on financial outcome. The
advances in metallurgy, and the decreased cost of production by
larger capacities, require yearly larger, more complicated, and
more costly plants. Thus, larger and larger capitals are required,
and enterprise is passing from the hands of the individual to the
financially stronger corporation. This altered position as to the
works and finance has made keener demands, both technically and in
an administrative way, for the highly trained man. In the early
stages of American mining, with the moderate demand on capital and
the simpler forms of engineering involved, mining was largely a
matter of individual enterprise and ownership. These owners were
men to whom experience had brought some of the needful technical
qualifications. They usually held the reins of business management
in their own hands and employed the engineer subjectively, when
they employed him at all. They were also, as a rule, distinguished
by their contempt for university-trained engineers.
The gradually increasing employment of the engineer as combined
executive and technical head, was largely of American development.
Many English and European mines still maintain the two separate
bureaus, the technical and the financial. Such organization is open
to much objection from the point of view of the owner's interests,
and still more from that of the engineer. In such an organization the
latter is always subordinate to the financial control,--hence the
least paid and least respected. When two bureaus exist, the technical
lacks that balance of commercial purpose which it should have. The
ambition of the theoretical engineer, divorced from commercial
result, is complete technical nicety of works and low production
costs without the regard for capital outlay which the commercial
experience and temporary character of mining constructions demand.
On the other hand, the purely financial bureau usually begrudges
the capital outlay which sound engineering may warrant. The result
is an administration that is not comparable to the single head with
both qualifications and an even balance in both spheres. In America,
we still have a relic of this form of administration in the consulting
mining engineer, but barring his functions as a valuer of mines, he
is disappearing in connection with the industry, in favor of the
manager, or the president of the company, who has administrative
control. The mining engineer's field of employment is therefore not
only wider by this general inclusion of administrative work, but
one of more responsibility. While he must conduct all five phases
of engineering projects coincidentally, the other branches of the
profession are more or less confined to one phase or another. They
can draw sharper limitations of their engagements or specialization
and confine themselves to more purely technical work. The civil
engineer may construct railway or harbor works; the mechanical
engineer may design and build engines; the naval architect may
build ships; but given that he designed to do the work in the most
effectual manner, it is no concern of his whether they subsequently
earn dividends. He does not have to operate them, to find the income,
to feed the mill, or sell the product. The profit and loss does
not hound his footsteps after his construction is complete.
Although it is desirable to emphasize the commercial side of the
practice of the mining engineer's profession, there are other sides
of no less moment. There is the right of every red-blooded man to
be assured that his work will be a daily satisfaction to himself;
that it is a work which is contributing to the welfare and advance
of his country; and that it will build for him a position of dignity
and consequence among his fellows.
There are the moral and public obligations upon the profession.
There are to-day the demands upon the engineers which are the demands
upon their positions as leaders of a great industry. In an industry
that lends itself so much to speculation and chicanery, there is the
duty of every engineer to diminish the opportunity of the vulture
so far as is possible. Where he can enter these lists has been
suggested in the previous pages. Further than to the "investor"
in mines, he has a duty to his brothers in the profession. In no
profession does competition enter so obscurely, nor in no other
are men of a profession thrown into such terms of intimacy in
professional work. From these causes there has arisen a freedom of
disclosure of technical results and a comradery of members greater
than that in any other profession. No profession is so subject to
the capriciousness of fortune, and he whose position is assured
to-day is not assured to-morrow unless it be coupled with a
consideration of those members not so fortunate. Especially is
there an obligation to the younger members that they may have
opportunity of training and a right start in the work.
The very essence of the profession is that it calls upon its members
to direct men. They are the officers in the great industrial army.
From the nature of things, metal mines do not, like our cities and
settlements, lie in those regions covered deep in rich soils. Our
mines must be found in the mountains and deserts where rocks are
exposed to search. Thus they lie away from the centers of comfort
and culture,--they are the outposts of civilization. The engineer
is an officer on outpost duty, and in these places he is the camp
leader. By his position as a leader in the community he has a
chieftainship that carries a responsibility besides mere mine
management. His is the responsibility of example in fair dealing
and good government in the community.
In but few of its greatest works does the personality of its real
creator reach the ears of the world; the real engineer does not
advertise himself. But the engineering profession generally rises
yearly in dignity and importance as the rest of the world learns
more of where the real brains of industrial progress are. The time
will come when people will ask, not who paid for a thing, but who
built it.
To the engineer falls the work of creating from the dry bones of
scientific fact the living body of industry. It is he whose intellect
and direction bring to the world the comforts and necessities of
daily need. Unlike the doctor, his is not the constant struggle
to save the weak. Unlike the soldier, destruction is not his prime
function. Unlike the lawyer, quarrels are not his daily bread.
Engineering is the profession of creation and of construction, of
stimulation of human effort and accomplishment.
INDEX.
Accounts.
Administration.
Administrative reports.
Air-compression.
-drills.
Alteration, secondary.
Alternative shafts to inclined deposit.
Amortization of capital and interest.
Animals for underground transport.
Annual demand for base metals.
report.
Artificial pillars.
Assay foot.
inch.
of samples.
plans.
Assaying.
A value of mine.
Averages, calculation.
Bailing.
Balance sheet.
Basic price.
value of mine.
Benches.
Bend in combined shafts.
Bins.
Blocked-out ore.
Blocks.
Bonanzas, origin.
Bonus systems, of work.
Breaking ore.
Broken Hill, levels.
ore-pillars.
Bumping-trough.
Cable-ways.
Cages.
Calculation of averages.
of quantities of ore.
Capital expenditure.
Caving systems.
Churn-drills.
Chutes, loading, in vertical shaft.
Classification of ore in sight.
Combined shaft.
stopes.
Commercial value of projects, determination.
Compartments for shaft.
Compressed-air locomotives.
-air pumps.
_vs_. electricity for drills.
Content, average metal, determining.
metal, differences.
Contract work.
Copper, annual demand.
deposits.
ores, enrichment.
Cost of entry into mine.
of equipment.
production.
per foot of sinking.
working.
Cribs.
Crosscuts.
Cross-section of inclined deposit which must be attacked in depth.
showing auxiliary vertical outlet.
Crouch, J. J.
Cubic feet per ton of ore.
foot contents of block.
Deep-level mines.
Demand for metals.
Departmental dissection of expenditures.
Deposits, _in situ_.
ore, classes.
regularity.
size.
structure.
Depth of exhaustion.
Determination of average metal contents of ore.
Development in early prospecting stage.
in neighboring mines.
of mines.
Diamond-drilling.
Diluting narrow samples to a stoping width.
Dip.
Direct-acting steam-pumps.
Distribution of values.
Dividend, annual, present value.
Dommeiler.
Down holes.
Drainage.
comparison of different systems.
systems.
Drifts.
Drill, requirements.
Drilling.
Drives.
Dry walling with timber caps.
Efficiency, factors of.
of mass.
Electrical haulage.
pumps.
Electricity for drills.
Engine, size for winding appliances.
Engineer, mining, as executive.
Engineering projects, phases of.
Enrichment.
at cross-veins.
Entry, to mine.
to vertical or horizontal deposits.
Equipment, cost.
improvements.
mechanical.
Erosion.
Error, percentage in estimates from sampling.
Escape.
Examination of mining property.
Excavation, supporting.
Exhaustion, depth.
Expenditures, departmental dissection.
mine.
Extension in depth.
Factor of safety in calculating averages of samples.
Filling.
system combined with square-setting.
with broken ore subsequently withdrawn.
waste.
Fissure veins.
Fissuring.
depth.
Fixed charges.
Flat-back stope.
Flexibility in drainage system.
Floors.
Folding.
Foot-drilled system of contract work.
-hole system of contract work.
of advance system of contract work.
value.
Fraud, precautions against in sampling.
General expenses.
Gold deposits.
deposits, alteration.
enrichment.
Hammer type of drill.
Hand-drilling.
-trucking.
Haulage, electrical.
equipment in shaft.
mechanical.
Hole system of contract work.
Horizons of ore-deposits.
Horizontal deposits, entry.
stope.
filled with waste.
Hydraulic pumps.
Impregnation deposits.
Inch, assay.
Inclined deposits to be worked from outcrop or near it.
deposits which must be attacked in depth.
shaft.
Inclines.
capacity.
Infiltration type of deposits.
Intelligence as factor of skill.
Interest calculations in mine valuation.
Intervals, level.
Inwood's tables.
Iron hat.
leaching.
Ivanhoe mine, West Australia.
Kibble.
Labor, general technical data.
handling.
unions.
Lateral underground transport.
Le Roi mine.
Lead, annual demand.
deposits.
enriching.
prices, 1884-1908.
-zinc ores, enrichment.
Lenses.
Levels.
intervals.
of Broken Hill.
protection.
Life, in sight.
of mine.
Locomotives, compressed-air.
Lode mines, valuation.
Lodes.
Long-wall stope.
Machine-drill, performance.
drilling.
_vs_. hand-drilling.
Management, mine.
Matte.
Mechanical efficiency of drainage machinery.
equipment.
Men for underground transport.
Metal content, determining.
contents, differences.
demand for.
mine, value.
price.
Mines compared to other commercial enterprises.
equipment.
expenditures.
Mines--_continued._
life of.
metal, value of.
of moderate depths.
to be worked to great depths.
valuation.
Mining engineering profession.
Mt. Cenis tunnel.
Morgan gold mine.
Normal price.
Obligations of engineering profession.
Openings, position in relation to secondary alteration.
Ore, average width in block.
blocked-out.
-bodies.
shapes.
-breaking, methods.
calculation of quantities of.
-chutes in shrinkage-stoping.
-deposits, classes.
determination of average metal contents.
developed.
developing.
expectant.
in sight.
sight, classification.
-pillars.
support in narrow stopes.
-shoots.
weight of a cubic foot.
width for one sample.
Origin of deposit.
Outcrop mines.
Output, factors limiting.
giving least production cost.
maximum, determination.
Overhand stapes.
Overproduction of base metal.
Oxidation.
Patchwork plant, mechanical inefficiency of.
Pay areas, formation.
Pillars, artificial.
Positive ore.
value of metal mine.
Possible ore.
Power conditions.
general technical data.
sources.
transmission.
Preliminary inspection.
Previous yield.
Price of metals.
Probable ore.
Producing stage of mine.
Production, cost.
Profit and loss account.
factors determining.
in sight.
Proportional charges.
Prospecting stage of mine.
Prospective ore.
value.
Protection of levels.
Proved ore.
Pumping systems.
Pumps, compressed-air.
electrical.
hydraulic.
rod-driven.
Ratio of output to mine.
Recoverable percentage of gross assay value.
Recovery of ore.
Rectangular shaft.
Redemption of capital and interest.
Reduction of output.
Regularity of deposit.
Reliability of drainage system.
Replacement.
Reports.
administrative.
Resuing.
Revenue account.
Rill-cut overhand stope.
method of incline cuts.
-stopes.
filled with waste.
-stoping.
Rises.
Risk in mining investments.
in valuation of mines.
Roadways, protecting in shrinkage-stoping.
Rod-driven pumps.
Rotary steam-pumps.
Round vertical shafts.
Runs of value.
test-treatment.
Safety, factor of, in calculating averages of samples.
Sample, assay of.
average value.
narrow, diluting to a stoping width.
sections.
taking, physical details.
manner of taking.
Sampling.
accuracy.
percentage of error in estimates from.
precautions against fraud.
Saving of fixed charges.
Secondary alteration.
enrichment.
Security of investment.
Self-dumping skip.
Sets.
Shafts.
arrangement for very deep inclined shafts.
compartments.
different depths.
haulage.
location.
number.
output capacity.
shape.
size.
Shrinkage-stope.
-stoping.
advantages.
disadvantages.
when applicable.
Silver deposits.
deposits, enrichment.
prices.
Sinking, speed.
Size of deposit.
Skill, effect on production cost.
Skips.
balanced.
haulage in vertical shaft.
Sollars.
Solubility of minerals.
Specific volume of ores.
Speculative values of metal mine.
value of mine.
Spelter, annual demand.
Square-set.
-set timbering.
Stations.
arrangement for skip haulage in vertical shaft.
Steam-pumps, direct.
Steepening winzes and ore passes.
Stope filled with broken ore.
minimum width.
Stoping.
contract systems.
Storing metal.
Structural character of deposit.
Structure of deposit.
Stull and waste pillars.
support with waste reënforcement.
-supported stope.
Stulls.
wood.
Subheading.
Sublevel caving system.
Subsidiary development.
Superficial enrichment.
Supplies, general technical data.
Support by pillars of ore.
Supporting excavation.
Surveys.
Suspense charges.
Test parcels.
sections.
-treatment runs.
Timber, cost.
Timbered shaft design.
Timbering.
Tin, annual demand.
deposits.
ore, migration and enrichment.
Tools.
Top slicing.
Tracks.
Transport in stopes.
Tunnel entry.
feet paid for in 10 years.
size.
Underhand stopes.
Uppers.
Valuation, mine.
of lode mines.
mines, risk in.
mines with little or no ore in sight.
on second-hand data.
Value, average, of samples.
discrepancy between estimated and actual.
distribution.
of extension in depth, estimating.
positive, of metal mine.
present, of an annual dividend.
of $1 or £1, payable in -- years.
runs of.
speculative, of metal mine.
Valuing ore in course of breaking.
Ventilation.
Vertical deposits, entry.
interval between levels.
shafts.
capacity.
Volume, specific, of ores.
Waste-filled stope.
Water-power.
Weight per cubic foot of ore.
Weindel, Caspar.
Whiting hoist.
Width of ore for one sample.
Winding appliances.
Winzes.
in shrinkage-stoping.
to be used for filling.
Working cost.
inherent limitations in accuracy of.
sheets.
Workshops.
Yield, previous.
Years of life required to yield --% interest.
Zinc deposits.
leaching.
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