the matter of sampling. In preparing his ore, the student should ask
himself what reasonable expectation he has that the portion he puts in
the furnace will be of average richness. The larger charges are likely
to be nearer than the smaller ones to the average of the parcel of ore
from which they are taken. In explanation of this, let us suppose a
large heap of 5-dwt. ore, in sand of the coarseness of full-stops, and
containing all its gold in particles of 1 milligram, as uniformly
distributed as care and labour in the mixing can accomplish. Such a heap
could not possibly occur in practice, but it will serve for purposes of
illustration. Now, one ton of the sand, however taken, would contain
appreciably the same quantity of gold as any other ton. For a ton would
contain about 8000 particles of gold; and even if two separate tons
differed by as much as 100 particles (which they are just likely to do),
this would mean only a difference of 1 or 2 grains to the ton. On the
other hand, two portions of 14 lbs., which should contain on the average
50 particles of gold, are likely enough to differ by 10 particles, and
this, calculated on a ton, means a difference of 1 dwt. It is easy to
see that something like this should be true; for on calculating the
14-lb. lot up to a ton, the deviation from the average, whatever it may
be, is multiplied by 160; whereas, if the ton were made up by adding
14-lb. lot to 14-lb. lot, up to the full tale, then a large proportion
of the errors (some being in excess and some in defect) would neutralise
each other. An average which is practically true when dealing with
thousands, and perhaps sufficiently exact with hundreds, would be merely
misleading when applied to tens and units. Reasonable safety in
sampling, then, is dependent largely on the number of particles of gold
in the charge taken, and the risk of an abnormal result is less, the
larger the charge taken.
By doubling the charge, however, we merely double the number of
particles. Powdering finely is much more effective; for, since the
weight of a particle varies as the cube of the diameter, halving the
diameter of the particles increases their number eight-fold. If, now, we
modify our illustration by assuming the particles to have only one-sixth
the diameter of a full-stop (which would represent a powder of a
fineness not unusual in ores prepared for assaying), we should multiply
the number of particles by 200 (6 × 6 × 6 = 216). We should then
reasonably expect a 14-lb. parcel of the powder to give as safe a sample
as a ton of the sand would give; and portions of a size fit for crucible
work, say 50 or 100 grams, would be as safe as 10 or 20-lb. samples of
the coarser stuff. For example, 60 grams of such powder would contain,
for a 5-dwt. ore, about 100 particles; and in the majority of cases the
error due to sampling would be less than 10 or 12 grains to the ton, and
would only occasionally exceed a pennyweight. With richer ores the
actual deviation stated as so much to the ton of ore might be greater,
but it would represent a smaller proportion, stated in percentage of
the gold actually present, and would ultimately fall within the limits
of unavoidable error.
It will be seen that the size of the quartz particles has no direct
bearing on the argument; and, in fact, the coarseness of the quartz only
interferes by preventing the uniform mixing of the sand and by binding
together several particles of gold; in this last case, particles so
united must, of course, count as one larger particle. Now, there are
some natural ores in which the gold particles are all very small; with
these fine powdering and mixing yields a product from which a sample may
be safely taken. Then, again, in "tailings," before or after treatment
with cyanide, we have a similar material, inasmuch as the coarser gold
has been removed by previous amalgamation. With these, it is not unusual
to take the portion for assay without any further powdering, since they
are poor in gold, and have already been stamped and passed through a
sieve of say thirty holes to the inch (linear).
But there are other ores, in lump showing no visible gold, which contain
the gold in all possible degrees of fineness, from say prills of a
milligram or so down to a most impalpable powder. The treatment of these
cannot be so simple and straightforward. Suppose a parcel of 1000 grams
(say 2 lbs.) of such ore in fine powder, containing on an average 1
particle of 1 milligram (the presence or absence of which makes a
difference of .6 dwt. on the ton), 10 others of about .5 milligram (each
representing .3 dwt.), and 100 others, which are too coarse to pass
through an 80 sieve, and having an average weight of .1 milligram (each
.06 dwt.), and that the rest of the gold, equivalent altogether to 2
ounces to the ton, is so finely divided that a charge of 50 grams may be
taken without any considerable risk of its interfering with the
sampling. Then in a 50-gram charge there would be one chance in twenty
of getting the milligram particle, in which case the result would be
12.35 dwts. too high; on the other hand, if it were not present the
result would on this account be .65 dwt. too low. Of the ten
.5-milligram particles, it is as likely as not that one will be present,
and its presence or absence would cause an error of 3.3 dwts., more or
less. Of the 100 particles of .1 milligram, there would probably be from
3 to 7, instead of 5, the proper number; this would mean a variation of
2.6 dwts. from the true proportion. So that the probable result would
range about 5 dwts. more or less than the 2-1/2 ozs., which is the true
produce, and there are possibilities of astounding results. It is true
that the majority of the results would be well within these limits, and
now and again the heart of the student would be gladdened by a beautiful
concordance in duplicate assays; nevertheless, there can be no
reasonable expectation of a good assay, and to work in this way, on a
50-gram charge, would be to court failure. The coarse gold must ruin the
assay.
The difficulty may be met by concentrating the whole of the coarse gold
in a small fraction of the ore, by sifting and making a separate assay
of this fraction. A portion of the ore, of about 1000 grams, is ground
to a very fine powder and passed through an 80 sieve, re-grinding when
necessary, until only 20 or 30 grams is left of the coarser powder. This
is mixed with fluxes and carried through as a separate assay. The sifted
portion is _thoroughly mixed_, and a portion of it, say 30 or 50 grams,
taken for assay. The weights of the two portions must be known, and care
must be taken that nothing is lost in the powdering. The method of
calculating the mean result from the two assays is shown on page 109. In
this way of working there is no advantage in continuing the grinding
until the coarser fraction is reduced to a gram or so--rather the
contrary; and rubbing on until all the gold is sent through the sieve is
to be distinctly avoided. The student must bear in mind that what he is
aiming at is the exclusion of all coarse gold from the portion of ore of
which he is going to take only a fraction.
The question of the smaller sampling of gold ores has been dwelt on at
considerable length, as befits its importance, in order that the student
may be impressed with a sense of its true meaning. Sampling is not a
mystery, nor does the art lie in any subtle manner of division. It is,
of course, absolutely necessary that the stuff to be sampled shall be
well mixed, and the fractions taken, so that each part of the little
heap shall contribute its share to the sample. Moreover, it must be
remembered that tossing about is a poor sort of mixing, and that
everything tending to separate the large from the small, the light from
the heavy, or the soft from the hard (as happens in sifting), must be
avoided, or, if unavoidable, must be remedied by subsequent mixing.
With a well-taken sample, we may rely on a great majority of our results
falling within normal limits of error; but nothing can be more certain
than that, in a moderately large experience we shall get, now and again,
deviations much more considerable. These erratic assays can only be met
by the method of working duplicates, which call attention to the fault
by discordant results. Such faulty assays should be repeated in
duplicate, so that we may rest the decision on three out of four
determinations.
The likelihood of two very faulty assays being concordant is remote;
but with very important work, as in selling parcels of ore, even this
risk should be avoided, as concordance in these cases is demanded in the
reports of two or more assayers. The following actual reports on a
disputed assay will illustrate this: (a) 5 ozs. 1 dwt.; (b) 5 ozs.
10 dwts. 12 grains; (c) 5 ozs. 11 dwts.; (c) 5 ozs. 11 dwts. 12 grs.
The mean result of several assays, unless there be some fault in the
method, will be very fairly exact; and individual assays, with an
uncertainty of 1 in 20, may, by repetition, have this reduced to 1 in
100 or less.
~Assay Tons, etc.~--Having decided on taking a larger or smaller
portion, the exact quantity to be used will be either some round number
of grams, such as 50 or 100, easily calculable into percentage; or it
will be that known as the "Assay Ton" (see page 13) or some simple
multiple or fraction of it, which is easily calculable into ounces. The
reports, too, are at least as often made as ounces in the short ton of
2000 lbs., as on the more orthodox ton of 2240 lbs. Now the short ton is
equal to 29,166.6 troy ounces; and the corresponding "assay ton" is got
from it by replacing ounces by milligrams. The advantage of its use is
that if one assay ton of ore has been taken, the number of milligrams of
gold obtained is also the number of ounces of gold in a ton of the ore,
and there is absolutely no calculation. Even if half an assay ton has
been taken the only calculation needed is multiplying the milligrams by
two. On the other hand with a charge of two assay tons the milligrams
need halving. Where weights of this kind (_i.e._, assay tons) are not at
hand they may be easily extemporised out of buttons of tin or some
suitable metal, and it is better to do this than to array out the grams
and its fractions at each weighing. The sets of "assay tons," however,
are easily purchased. As stated on page 13, the assay ton for 2240 lbs.
is 32.6667 grams; and for the short ton, 29.1667 grams. If, however, the
round number of grams be used and the result brought by calculation to
the produce on 100 grams, the conversion to ounces to the ton may be
quickly effected by the help of the table on page 107. As this table
only deals with the ton of 2240 lbs., it is supplemented here by a
shortened one dealing only with _the produce of 100 grams_ and stating
the result in _ounces troy to the short ton of 2000 lbs_.
~Estimation of Small Quantities of Gold.~--_By the Balance._ In
estimating minute quantities of gold there are one or two points, of
importance to an assayer only in this assay, where they will often allow
one to avoid the working of inconveniently large charges. One of these
is known as "weighing by the method of vibrations."
TABLE FOR CALCULATING OUNCES TO THE SHORT TON FROM THE
YIELD OF GOLD FROM 100 GRAMS OF ORE.
----------+---------+-----------+---------+-----------+---------
Milligram.|Ounces to|Milligrams.|Ounces to|Milligrams.|Ounces to
| the Ton.| | the Ton.| | the Ton.
----------+---------+-----------+---------+-----------+---------
0.01 | 0.003 | 0.4 | 0.117 | 7.0 | 2.042
0.02 | 0.006 | 0.5 | 0.145 | 8.0 | 2.333
0.03 | 0.009 | 0.6 | 0.175 | 9.0 | 2.625
0.04 | 0.012 | 0.7 | 0.204 | 10.0 | 2.916
0.05 | 0.014 | 0.8 | 0.233 | 20.0 | 5.833
0.06 | 0.017 | 0.9 | 0.262 | 30.0 | 8.750
0.07 | 0.020 | 1.0 | 0.292 | 40.0 | 11.666
0.08 | 0.023 | 2.0 | 0.583 | 50.0 | 14.583
0.09 | 0.026 | 3.0 | 0.875 | 60.0 | 17.500
0.10 | 0.029 | 4.0 | 1.167 | 70.0 | 20.416
0.20 | 0.058 | 5.0 | 1.458 | 80.0 | 23.333
0.30 | 0.087 | 6.0 | 1.750 | 90.0 | 26.250
----------+---------+-----------+---------+-----------+---------
Suppose a balance at rest in perfect equilibrium, with the
pointer exactly over the middle point of the scale. Let the scale be a
series of points at equal distances along a horizontal line; then, if a
small weight be placed on one pan, the pointer will deviate from its
vertical position and come to rest opposite some definite part of the
scale, which will depend upon the magnitude of the weight added. The law
determining this position is a very simple one; the deviation as
measured along the points of the scale varies directly as the weight
added. For example, with an ordinarily sensitive balance, such as is
used for general purposes, one milligram will move the pointer along,
say, three divisions of the scale; then two milligrams will move it six
divisions; half a milligram, one and a half divisions; and so on. Of
course, with a more sensitive balance the deviations will be greater.
Now the point at which the needle comes to rest is also the middle point
about which it vibrates when swinging. For example, if the needle swings
from the third to the seventh division on the right then [(7+3)/2] it
will come to rest on the fifth. In working by this method the following
conventions are useful: Always place the button to be weighed on the
left pan of the balance, the weights on the right; count the divisions
of the scale from the centre to right and left, marking the former + and
the latter -; thus -5 is the fifth division to the left. Then the
position of rest is half the algebraic sum of two readings. For example,
let the readings be 7 to the right and 3 to the left, then (+7-3)/2 =
+2. The mean division is the second division to the right. If the
student will place himself in front of a balance and repeat the
following observations and replace the figures here given by his own, he
will have no difficulty in grasping the method. First determine the
_bias_ of the balance; suppose the unloaded balance swings +1.25 and -1;
the bias then is (1.25-1)/2 = +.125 or one-eighth of a division to the
right. Now having put on the button to be weighed let the readings be
+7.5 and +9.25, and (7.5+9.25)/2 = +8.375. Then the effect of the button
has been to move the pointer from +.125 to +8.375, or 8.25 divisions to
the right; we should, therefore, add the weight equivalent of 8.25
divisions to the weights, whatever they may be on the right hand pan of
the balance; if the divisions were to the left (- divisions) we should
subtract. The value of 1 division is easily determined. Suppose the
button in the example were a 1 milligram weight, then we should have
found that 1 milligram = 8.25 divisions .'. 1 division = .121 milligram.
This method of working adds very considerably to the power of a balance
in distinguishing small quantities.
[Illustration: FIG. 44_a._]
_By the Microscope_.--The use of the microscope also is a real advantage
in estimating the weights of minute buttons of gold where there is no
undue risk in sampling, and where an error of say 1 in 20 on the
quantity of gold is tolerable. For ores with copper, lead, zinc, &c., as
well as for tailings rather poor in gold, this leaves a wide field of
usefulness. The method is described on page 440, but the description
needs supplementing for those who are not accustomed to the use of a
microscope. The eye-piece of a microscope (fig. 44_a_, A) unscrews at
_a_, showing a diaphragm at _b_, which will serve as a support for an
eye-piece micrometer. This last, B, is a scale engraved on glass, and
may be purchased of any optical instrument maker, though it may be
necessary to send the eye-piece to have it properly fitted. When resting
on the diaphragm it is in focus for the upper lens, so that on looking
through the microscope, the scale is clearly seen in whatever position
the instrument may be as regards the object being looked at. Suppose
this to be a small button of gold on a shallow, flat watch-glass, on the
stage of the microscope. Bring the button under the "objective" (_i.e._,
the nose of the microscope), which should be about a quarter of an inch
above the watch-glass; then looking through the instrument, raise the
tube until the button of gold, or at least some dust on the glass, comes
into focus. If the button is not in the field, rest the thumbs and index
fingers, using both hands, on the edge of the watch-glass, pressing
lightly but steadily, and give the glass a slow, short, sweeping motion;
the button will perhaps appear as an ill-defined blackness, because not
quite in focus. Bring this into the centre of the field. Raise or lower
the microscope until the button appears with sharp outlines. If the
scale does not cover the button, rotate the eye-piece; this will bring
the scale into a new position. Since the divisions over the button are
less distinct than the others, it is best to read the latter. Thus, in
fig. 44_b_, there are 36 divisions on one side of the button, and 35 on
the other, making altogether 71. The whole scale is 80, therefore the
diameter of the button is 9 divisions. The value of each division
obviously varies with the magnifying power employed. With most
microscopes there is a telescopic arrangement whereby the tube may be
lengthened; if this be done and the button again brought in focus, it
will be seen that, as measured on the scale, the button is much larger
than before. It is evident, therefore, the micrometer must always be
used in the same way. The method given in the appendix (page 440), for
finding the value of the scale when gold buttons are to be measured is
easy and satisfactory. When the button of gold is so small that there is
considerable risk of losing it in transferring to a watch-glass, it may
be measured on the cupel, but for this purpose it must be well
illuminated; this is best done by concentrating light on it with a lens,
or with what comes to the same thing, a clean flask filled with water.
[Illustration: FIG. 44_b._]
Most assayers, however, using a micrometer in this way, would like to
know its absolute value. To do this, a stage micrometer must be
purchased. This is like an ordinary microscope slide (fig. 44_a_, C),
and when looked at through a microscope it shows (fig. 44_c_) lines
ruled on the glass at distances of tenths and hundredths of a
millimetre, ten of each, so that the full scale is 1.1 mm. In the case
illustrated, 60 divisions of the scale in the eye-piece are just equal
to the 1.1 mm., therefore 1 division equals .0183 mm. A cube of this
diameter would contain (.0183×.0183×.0183) .0000061285 cubic mm. The
corresponding sphere is got by multiplying by .5236; this gives
.000003209 cb. mm. The weight of 1 cb. mm. of water is 1 milligram; and,
since gold is 19.2 times as heavy as water (sp. g. = 19.2), the contents
in cb. mm. must be multiplied by 19.2. This gives .0000616 milligram as
the weight of a sphere of gold measuring 1 division.
[Illustration: FIG. 44_c._]
If every result had to be calculated in this way the method would be
very laborious; but, having the figures for the first division, those of
the others may be calculated by multiplying by the cube of the
corresponding number. Thus, for the third division (3×3×3 = 27), the
content of the cube (.0000061285×27) is .0001655 cb. mm.; the content of
the sphere (.000003209×27) is .0000866 cb. mm.; and the corresponding
sphere of gold (.0000616×27) is .00166 milligram. With the help of a
table of cubes the whole calculation for 25 or 30 divisions may be made
in half an hour, and the results preserved in the form of a table will
simplify all future work.
~Assay Operations.~--The actual work of the assay resolves itself into
three operations:--(1) The fusion of the ore and concentration of the
"fine metal" (_i.e._, gold and silver) in a button of lead; (2) The
cupellation of the lead, whereby a button of fine metal is obtained; and
(3) the "parting" of the gold which separates it from the accompanying
silver. The following description takes the order as here given, but the
student, in learning the method, should first practise cupellation if he
has not already done so; next he should practise the separation of gold
from silver, taking known weights of fine gold (p. 63), varying from .5
or .3 gram down to quite minute quantities, and not resting satisfied
until a sensitive balance can barely distinguish between the weights of
gold taken and found. It may be noted here that if he has not a flatting
mill at his disposal, then for large buttons it is better to make an
alloy with eight or nine parts of silver to one of gold, and attack it
with acid without previous flattening, rather than accept the risk and
labour of beating out a less easily attacked alloy to the necessary
thinness with a hammer. It is only after a sense of security in gold
parting has been acquired, that the attack of an ore can be profitably
accomplished, and even then simple and easy ores should be first taken,
passing on to others more difficult, either because of a more complex
mineral composition or a difficulty in sampling.
~Concentration of the fine Metal in Lead.~--The best flux for quartz,
which makes up the earthy matter of most gold ores, is soda, and this is
best added as carbonate or bicarbonate. By theory,[20] 50 grams of
quartz will require 88.5 grams of the carbonate, or 140 grams of the
bicarbonate, to form sodium silicate, which is a glassy, easily-fusible
substance, making a good slag. If the bicarbonate is used, and heat is
applied gradually, steam and carbonic acid are given off at a
comparatively low temperature, and the carbonate is left; at a higher
temperature (about 800° C., or a cherry-red heat) the carbonate fuses
attacking the quartz, and giving off more carbonic acid; as the heat
increases, and the attack on the quartz (which of itself is infusible)
becomes complete, the whole mass settles down to a liquid sodium
silicate, which is sufficiently fluid to allow the gold and lead to
settle to the bottom. The fluid slag does to a certain extent dissolve
some of the crucible, but not seriously. In a perfect working of this
experiment, the first evolution of gases (steam and carbonic acid)
should be gentle, so as to run no risk of its blowing the fine powder
out of the crucible; and the heat at which the second evolution of
carbonic acid is produced should be maintained until the reaction is
completed, so that there may be little or no formation of gas in the
fused mass to cause an effervescence which may force some of the charge
over the edges of the crucible. Of course, in practice the ideal fusion
is not attained, but there is no difficulty in approaching it closely
enough to prevent the charge at any time rising above the level it
reached at first in the crucible, and this should be accomplished. It is
usual with quartzose ores to rely mainly on the action of carbonate of
soda, but not entirely. Litharge is also used; it forms, on fusion with
quartz, a silicate of lead, which is a yellow glass, easily fusible, and
more fluid in the furnace than silicate of soda is. By theory, 50 grams
of quartz would require 186 grams of litharge.[21] The reaction takes
place without evolution of gas, and in its working the only point is to
so regulate the heat that the litharge shall not fuse and drain under
the unattacked quartz, leaving it as a pasty mass on the surface. Now,
if in making up a charge for 50 grams of ore, we took 100 grams of
bicarbonate of soda (equivalent to about 63 grams of the carbonate),
this being five-sevenths of 140 grams (which by itself would be
sufficient), leaves two-sevenths of the quartz to be fluxed by other
reagents: two-sevenths of 186 grams (say 52 grams) of litharge would
serve for this purpose. But if we used 10 grams of borax, which has a
fluxing action about equal to that of the litharge, then 40 grams of the
latter, or (making an allowance for the quartz being not quite pure) say
35 grams, will suffice. The fluxes, then, for the 50 grams of ore would
be: bicarbonate of soda 100 grams, litharge 35 grams, and borax 10
grams; we could decrease any of these, and proportionately increase
either or both of the others, and still rely on getting a fusible slag,
which is the whole of the function of a flux, considered simply as a
flux. It should be remembered, however, that the slag is a bi-silicate
or acid slag, and that its acid character is increased by increasing the
proportion of borax.
But in addition to the fluxes there is required about 30 or 40 grams of
lead to collect the silver and gold. This is best added as litharge (say
40 grams) and flour (4 grams), or charcoal powder (2 grams). See pages
93 and 94. The full charge, then, would be:
Ore 50 grams.
Bicarbonate of soda 100 "
Litharge 75 "
Borax 10 "
Flour 4 "
These should be mixed, placed in a suitable crucible (a G Battersea,
round, will do), and heated, at first at a red heat, but finally much
hotter, so as to get a fluid and clean slag. When the charge has been in
tranquil fusion for some little time, take it out and pour it into an
iron mould. When cold, detach the button of lead. The slag should be
glassy, all through alike, and easily separable from the metal. With
ordinary ores, this slag may be considered as free from gold. In an
experiment in which 90 milligrams of gold were added, the full amount
was obtained from the lead produced by the first fusion. But in certain
cases, more especially where large amounts of metallic oxides are
present, the slag is not so clean, and with these the slag should be
powdered, mixed with 40 grams of litharge and 4 of flour, and melted
again; it is an advantage to add a small prill of say 2 or 3 milligrams
of silver to the charge, as it insures a visible product in the
cupellation. Indeed, this last precaution is a good one to be taken
wherever there is reason to expect very small buttons. It has the
further advantage, that, if the quantity of silver necessary for
inquartation is known, the right quantity may be added here, so as to
save a subsequent operation.
~Ores containing Oxides of Iron.~--Of the metallic oxides likely to be
present in a slag, oxide of iron is the most important. Gold is
occasionally found in a matrix of this substance, and in the assay of
"concentrates" largely made up of pyrites, this oxide will be formed in
the preliminary calcination. Now, the lower oxide of iron (ferrous
oxide, FeO) is easy to deal with; fused borax will dissolve about its
own weight of it, and a silicate of soda (such as makes up the bulk of a
slag in a gold assay) will take up at least half as much. But the higher
oxide (ferric oxide, Fe_{2}O_{3}) is more refractory; even 6 parts of
borax yields a poor product, and slags with any considerable percentage
of it are not satisfactory. A student attempting to recover gold from
some hæmatite (in which there was about half an ounce of the metal),
found in the slag nearly a gram of gold, although in the first fusion
the slag appeared perfectly fluid. There is, however, no difficulty in
getting good slags, even with large quantities of iron. For example,
with 50 grams of ferric oxide, 10 of quartz, 30 of borax, 30 of
soda,[22] 50 of litharge, and 7 of flour, the result was quite
satisfactory. So, too, was 25 of quartz, 50 of soda, 50 of litharge, and
7 of flour. It is well, however, in such cases to have an ample
proportion of flux and to aim at a larger button of lead than usual by
increasing the proportion of flour or charcoal (see also page 91). A
charge used on the Randt for roasted "concentrates" (which we may
roughly speak of as quartz and ferric oxide), is one assay ton (about 30
grams) each of ore, soda, and borax, and one and a half assay ton of
litharge and 2 grams of charcoal. Whilst, for the same material, from
which most of the gold has been extracted by "chloridising," 2.5 tons
each of ore, borax, and soda, 4 of litharge, and 4 grams of charcoal are
needed. This quantity requires a large crucible (I Battersea, round). In
this the proportion of silicate of soda and borax counted together is to
the oxide of iron as 4 to 1, on the supposition that the quartz and
oxide of iron of the ore are in about equal quantities; but, in the
larger charge especially, much oxide of lead would also remain as a
flux.
~Ores containing Sulphides.~--In assaying ores containing a large
proportion of pyrites or mispickel, or both, the best plan is to take a
portion and calcine so as to convert it into a product of the kind just
considered. The weighed portion of ore should be placed in a clean
crucible and be heated to incipient redness: with pyrites the first
effect is to drive off about half the sulphur as vapour which burns as
flame over the ore. At this stage care should be taken that there is no
great increase of temperature, otherwise there may be more or less
fusion, which would spoil the operation. When the sulphur flame ceases
the solid sulphide of iron burns with visible incandescence and the
charge should now be stirred with a flattened iron rod so as to expose
fresh portions to the air. The top of the furnace must be open, so that
air may have free access to the crucible. When stirring is no longer
followed by visible burning the heat may be raised to full redness. The
crucible is then lifted out (the stirrer still resting in it) and if the
charge gives off no odour of burning sulphur it is shaken out into an
iron mortar and mixed with the fluxes, taking care to clean the stirrer
in the mixture. The charge is then replaced in the crucible in which the
roasting was done and fused in the furnace. The resulting button of lead
is cupelled for fine metal. Ores rich in sulphides requiring this
treatment are frequently "concentrates." For their assay take 1 assay
ton (30 grams), calcine and mix with an equal weight of soda and of
borax (30 grams each), and half as much again of litharge (1.5 tons or
45 grams), and with 2 grams of charcoal or 5 grams of flour.
Where the sulphides are present in smaller proportion (10 per cent. or
less), they may be taken as serving the purpose of flour or charcoal
(see page 95); the sulphur and iron are oxidised at the expense of the
litharge with a consequent separation of lead as metal. If the
proportion of sulphides is not sufficient to give a large enough button
of lead, some charcoal or flour should be added. On the other hand, if
they are in small excess and give a button of lead somewhat sulphury,
_i.e._, hard and brittle, it may be remedied by the judicious addition
of nitre; this last reagent, however, should not be used in large
quantity. A plan much used to prevent sulphury buttons is to insert an
iron rod or a nail in the charge in the crucible; the iron takes the
sulphur forming sulphide of iron which in moderate quantity does not
form a separate layer of matte but dissolves in the slag. A slag formed
of 50 grams of quartz, 100 soda, and some borax, may take up in this way
some 10 or 12 grams of sulphide of iron. If, however, an ore gives a
layer of matte or speise, it is best to repeat the assay by the method
of calcining before fusion.
~Cyanide Charges, etc.~--In assaying the "tailings" which are to be
treated in a cyaniding plant the following charge is used:
Tailings 3 assay tons or 100 grams.
Litharge 4.5 " 150 "
Soda 4.5 " 150 "
Borax .75 " 25 "
The sand is assayed without any further crushing and the assay is made
in duplicate.
The residues after treatment with cyanide, differing from the tailings
merely in being poorer in gold because of the extraction by the solution
of cyanide, are run down with the same fluxes in the same relative
proportions. But four charges of 2.5 assay tons (say 75 grams) are
worked, and two of the resulting buttons are scorified together and then
cupelled, etc., so as to give duplicate assays on charges of 5 assay
tons. This is one of the cases in which it is desirable to add a small
portion of silver before cupelling.
In assaying the "cyanide liquors" for gold, 2 assay tons of the liquor
are measured out (58.3 c.c. for the ton of 2000 lbs., 65.3 c.c. for the
other) and are evaporated to dryness in a lead dish weighing about 35
grams. Such a dish is easily extemporised out of a piece of lead foil,
if the ordinary vessel is not at hand; but care must be taken that the
lead is free from gold. The dish with the dried residue is then
scorified and the resulting button of lead is cupelled.
[Illustration: FIG. 44c.]
In some cases the fusion of the ore may be replaced by a treatment with
solution of cyanide of potassium and the gold recovered from the
solution in the way just described. For this purpose the ore should be
in not too fine powder, otherwise there will be great difficulty in
filtering; a sand which will pass a 30 sieve and having no large
proportion of very fine stuff will do. Not less than 200 grams should be
taken; and as an extraction apparatus a bell jar capable of holding half
as much again may be used. Such a jar may be extemporised by cutting off
the bottom of a bottle by leading a crack around it with a red hot
poker; or a lamp chimney will serve the purpose. The smaller mouth of
the jar is closed by a perforated cork provided with a clipped tube
after the manner of a burette (see fig. 44c). In the jar, just over
the cork, put a plug of loose asbestos or glass wool, or a piece of
sponge to act as a filter; a layer of broken glass, coarse at the bottom
and fine at the top, will serve the same purpose. On this, place the
charge of ore to be extracted. Prepare a solution of cyanide of
potassium in water, with 5 or 10 grams of the salt to the litre. It may
be that the whole point of the assay depends on the solution being of a
definite strength; as, for example, where the relative efficiency of
solutions of different strengths is being determined, when it will be
best to estimate the quantity of cyanide of potassium in the dilute
solution by the method given at the end of this article (page 160). Pour
the cyanide solution on to the ore, letting the first portions to come
through run into the beaker, but as soon as the ore is thoroughly wetted
close the clip and allow to stand for several hours. Then, opening the
clip, run through more cyanide solution and then water, so as to wash
the gold-carrying liquor thoroughly into the beaker. It is no matter if
the liquor is a little bit turbid; transfer it to a lead dish,
evaporate, scorify, and cupel in the usual fashion.
The assay of gold-zinc slimes, which is the precipitate formed by zinc
acting on cyanide solutions of gold, may be made by wrapping 2 or 3
grams in 40 grams of sheet lead and scorifying, cupelling, &c. The
amount of impurity in the stuff varies greatly; it is usually calcined
and mixed thoroughly with soda 40 per cent., borax 30 per cent., and
sand 10 per cent., and melted in graphite pots. The buttons of bullion
obtained are afterwards remelted with borax and run into bars, the
fineness of which varies from 600 to 830 thousandths. The bars are
sampled by chipping off diagonally opposite corners: or better, by
drilling, the drillings being freed from pieces of steel with the help
of a magnet.
~Cupellation.~[23]--The cupellation of lead for gold differs very little
from that of lead carrying silver. When the gold is accompanied by a
larger proportion of silver, and both have to be determined, the
cupellation must be conducted exactly as in a silver assay, the usual
precautions being taken to moderate the temperature so as to lessen the
cupellation loss and to promote a slow and undisturbed solidification in
order to avoid spirting. If, however, the gold predominates the finish
should be effected at a higher heat, as the melting-point of gold is
100° higher than that of silver. The bad effect of a higher temperature
in increasing the cupellation loss need hardly be considered in the case
of such small buttons of gold as are obtained in assaying gold ores, as
any loss there may be is hardly appreciable by the balance. With larger
quantities of gold, however (as in assaying gold bullion), this loss
becomes important; and it is therefore necessary to very carefully
regulate the temperature of the muffle so as to minimise the loss.
~The cupels~ are made of well-burnt bone-ash, of the fineness of coarse
wheat flour, moistened with one-twelfth its weight of water and
compressed into shape in suitable moulds. The moulds sold for this
purpose are often of unsuitable shape. Since lead has a specific gravity
of over 11, a cup to hold from 15 to 25 grams of molten lead need not
have a capacity of more than about 2 c.c. A hollow about 1 inch across
and 1/4 inch deep is sufficient; and the body of the cupel to absorb
this weight of lead should itself weigh from 20 to 25 grams. The button
of lead in a gold assay may be twice as heavy as this. For these larger
buttons a hollow 1-1/3 inch across and 1/3 inch deep will be sufficient.
If these larger cupels are not at hand the larger buttons will have to
be reduced in size by a scorification before cupelling. In some cases
this preliminary scorification is advantageous or even necessary: this
may be because the lead is hard and impure, or it may be that a very
small button of gold is expected. In the latter case it is best to
scorify the lead down to something less than 1 gram, and to perform the
cupellation on a specially prepared small fine cupel. These small cupels
are best made by grinding the unsaturated portion of a used cupel to a
fine powder, and compressing the dry powder into a small Berlin crucible
or scorifier; the face should be made quite smooth by pressure from a
pestle. On such cupels a small speck of gold (less than .01 milligram)
will be left in a good shape and easily visible; but the cupel must be
withdrawn from the muffle as soon as the cupellation is finished to make
sure of always getting the button in good condition. In places, such as
Mints, where large numbers of bullion assays are regularly made a
special form of cupel is used so that not less than six dozen assays may
all be cupelled at the same time in a muffle of ordinary size. These
cupels are square blocks, a little less than 2 inches across, and a
little more than three quarters of an inch deep. Each block carries four
hollows of about .7 inch across and .3 inch deep. A muffle, on a floor
space of 6 inches by 12, would take 3 of these blocks abreast and 6
deep, and thus provide the means for 72 assays.[24]
Cupels made with wet bone-ash should be slowly dried; and if in the
muffle they can be slowly brought to an orange-red heat it is all the
better. Under no circumstances must the lead be placed on the cupel
before the latter has been so thoroughly heated that it can no longer
give off steam or gas of any kind. For this gas bubbling through the
molten metal spatters it, thus spoiling one assay and throwing doubt on
all the rest. Again, the risk of freezing at the start is much greater
with a cupel which has not been properly heated.
The best plan is to do all the cupellations in batches. After the muffle
has cooled down for the withdrawal of the last batch, and the old cupels
have been taken out, the new cupels for the next batch should be put in
their place. The furnace should then be stoked and made ready for the
next cupellations; by the time the furnace is ready the cupels will be
ready also. There should be no unnecessary handling of the cupels once
they have been placed in the muffle.
~The cupellation temperature for gold~ is an orange-red heat or perhaps
a little hotter. Beginners, who are apt to overheat their furnace,
should avoid a heat which can properly be called yellow. Dr. T.K.
Rose[25] has determined the temperature of a muffle during the
cupellation of gold-silver alloys at the Royal Mint. In one muffle the
temperature ranged from 1065° to 1095° C.; the lower temperature was of
course in the front of the muffle. In another it ranged from 1022° to
1062°, and here the muffle appeared to the eye "decidedly cooler than
usual." The alloy left after cupelling was made up of 1 part of gold to
2-1/2 parts of silver, and was fused at 952°; hence the usual
temperature of cupellation was, say, 120° or 130° above the
melting-point of the residual metal. To obtain some real knowledge as to
the meaning of these figures, the student should prepare pointed pieces
of the following metals: silver, which melts at 945°; gold, which melts
at 1035°; and an alloy, half silver, half gold, which melts at 990°.
These should be placed on clean cupels in a muffle almost entirely
closed; the temperature should be very slowly raised, and the appearance
of the muffle when each metal begins to melt should be carefully noted.
The cupelling temperature in Dr. Rose's experiment was as much above the
melting-point of gold as this is above that of the silver-gold alloy.
The _finish of the cupellation_ of gold or gold-silver alloys is
practically the same as with pure silver; there is the same thinning out
of the litharge into a luminous film which becomes iridescent before the
brightening. But the danger of spirting decreases as the proportion of
gold becomes greater, and disappears when the gold is much over 30 per
cent. Nevertheless it is well to let such buttons become solid
undisturbed and protected from draughts in the body of the muffle. This
means closing the muffle and allowing the furnace to cool down somewhat
before withdrawing the cupels. Buttons solidified in this way are more
malleable than when they are withdrawn promptly on the finish of the
cupellation. This is important with large buttons, as in a bullion
assay. On the other hand, very small buttons, especially such as have to
be measured rather than weighed, should be withdrawn as soon as the
luminous film has disappeared. For when this is done the button can be
loosened from the cupel by merely touching it with the point of a pin,
and is then safely and easily transferred to a watch glass by touching
it with the head of a pin which has been moistened. It adheres to this,
and if the pin is not too wet comes off at once on touching the glass,
or in any case will do so on gentle warming.
Molten gold, with little or no silver, has a peculiar colour which is
easy to recognise; it is more globular than a button of silver of the
same size would be, and it shows less adhesion to the cupel. Just after
becoming solid it glows beautifully, and this is so marked that it is a
valuable help in finding the position of a button when it is more than
ordinarily minute.
If the button left from cupellation is yellow it is at least half gold,
and a rough guess as to the proportion of gold may be made from its
yellowness; the rest of the metal is generally silver. The presence of
platinum or one of the platinum group of metals makes the surface of the
button dull and crystalline. The native alloy of osmium and iridium does
not alloy with gold, however, but falls to the bottom of the molten
metal. It shows itself in the subsequent parting as a black spot or
streak on the under surface.
The buttons are removed from the cupel with a pair of pliers and then
brushed to remove adherent litharge and bone-ash. Some assayers advise
cleaning by dipping in warm dilute hydrochloric acid followed by washing
in water and drying. The button is next weighed. When the quantity of
silver obtained is not required to be known the weighing may sometimes
be omitted. The next operation in either case is parting either with or
without a previous inquartation.
_The loss of gold in cupellation_ is by no means always inconsiderable.
In three cupellations of 1 gram of gold with 20 grams of lead made
purposely at a very high temperature the cupel absorbed 6.04, 6.20, and
6.45 milligrams of gold. Hence at a high temperature there may easily be
a loss of more than half a per cent. of the gold. In ten cupellations
with the same quantities of gold and lead, but at an ordinary
temperature, the gold recovered from the cupels varied from 1.37 to 1.92
milligrams, and gave an average of 1.59 milligrams. In round numbers the
cupellation loss of pure gold is .15 per cent.
But if the gold be alloyed with _silver_ the loss is diminished, as is
shown by the following experiments. Gold, .3 gram, was cupelled with 10
grams of lead and varying amounts of silver, and the cupels were assayed
for gold with the following results:
Silver in the alloy .3 gram .6 gram .9 gram
Gold in the cupel .47 milligram .32 milligram .17 milligram
These, calculated on the .3 gram of gold, give the loss as .157, .107
and .057 per cent. respectively. The effect of _copper_, on the other
hand, is to increase the cupellation loss, which, silver being absent,
may from this cause rise to .3 per cent., even when the temperature is
not excessive.
In the ordinary assay of gold-copper alloys a constant weight of the
alloy is always taken; hence as the weight of copper in a cupel charge
increases, the weight of gold decreases. The silver, on the other hand,
is always very nearly two and a half times as much as the gold, whatever
its quantity may be. But the cupellation loss is smaller with less gold
and greater with more copper, and it so happens in these assays that
these two opposites nearly neutralise one another. Mr. W.F. Lowe[26]
found the gold recoverable from the cupels on which 20 grains of gold
bullion had been treated varied only between .014 and .015 grain (_i.e._
from .07 to .075 per cent. of the bullion treated), although the quality
of the bullion varied from 9 to 22 carat.[27] But in the poorest bullion
there was only 7.5 grains of pure gold, while in the richest there were
18.3 grains; yet each lost on the cupel the same weight of gold, viz.,
.014 grain. When reckoned in percentages of the actual gold present the
losses are .187 per cent. and .076 per cent. respectively. The heavier
percentage loss is mainly due to the increased quantity of copper.
As with silver so with gold the predominant cause of the cupellation
loss is the solution of the metal in the molten litharge which passes
into the cupel. Three lots of 1 gram of gold cupelled each with 20 grams
of lead repeatedly, so as to make 13 cupellations in all, lost in actual
weight 35.72 milligrams. The gold recovered from the cupels amounted
altogether to 34.56 milligrams. This shows that, compared with the
absorption by the cupel, the other causes of loss are inconsiderable.
The loss of gold by _volatilisation_ is, however, a real one. The dust
from the flues of assay furnaces has been tested on several occasions
and found to contain gold, though in small quantity. Thus Mr. Lowe found
.073 per cent. of silver and .00033 per cent. of gold in such a
material. The lead volatilised from a gold bullion assay would need to
be ten times as rich as this to account for a loss of gold equal to the
hundredth part of a milligram. Dr. Rose, in the paper already quoted,
believes that on a .5 gram charge of standard bullion the loss from
volatilisation is not less than .025 nor more than .05 milligram of
gold.
By way of conclusion it may be said that the cupellation loss of gold is
about .07 per cent., and that it is largely met or even over corrected
by a compensating error due to silver retained in the gold after
parting.
~Inquartation.~--The method of separating the gold from the silver in
gold-silver alloys by boiling with nitric acid does not act equally well
in all cases. An alloy half silver half gold, rolled to thin sheet and
boiled for half an hour with nitric acid, may still retain more than
two-thirds of its silver. An alloy of 1 part gold and 1.7 parts of
silver gives up practically the whole of its silver under similar
treatment. The gold is left in a coherent, though easily broken, sheet
retaining the shape of the original alloy. The gold thus left is quite
spongy and porous, so that the acid can penetrate into its innermost
portions. But if the silver is in large excess in the alloy, the
removal of the silver is less complete, and the residual gold, instead
of holding together in a form easy to manipulate, falls to a powder
which requires care and time in its treatment. The older assayers,
therefore, added silver to their gold in such proportion that the alloy
for parting should be one quarter gold to three quarters silver. This
operation they called _inquartation_.
The modern practice is to aim at getting an alloy with 2-1/2 parts of
silver and 1 part of gold. In gold bullion assays this proportion should
be obtained with fair exactness. And in the parting of such gold buttons
as are obtained in assaying ores it is well to aim at this proportion,
though absolute precision is not a matter of importance.
If the button left on cupelling the lead from an assay of an ore appears
white, it is best to assume that it already contains at least a
sufficiency of silver, in the absence of any knowledge to the contrary.
This will be true in almost all cases. But if, on parting, it does not
lose at least two-thirds of its weight, this indicates that the
assumption was not justified; and also what quantity of silver must be
added to the button before again attempting to part. Generally the fault
will be in the other direction; the silver will be in excess and the
gold will break up and demand very careful treatment.
If, however, such a button is yellow, then, from its weight and depth of
colour, a rough estimate can be made of how much gold is contained in
it. Silver must be added to make the total weight 3-1/2 times as much as
that of the gold supposed to be present. Thus, if the button weighs 10
milligrams and is supposed to contain 8 milligrams of gold, then 8
multiplied by 3-1/2 is 28; the button must, in such case, be made up to
28 milligrams by adding 18 milligrams of silver. In judging of the
quality of the gold button, no ordinary error will very seriously affect
the result. If, in the example just given, the quantity of gold present
was really 7 or even 9 milligrams of gold, the resulting alloy would
still have been suitable for such partings. In fact, in routine assays,
where the quantity as well as the quality of the gold is known within
fair limits, it is often the custom to add the silver for inquartation
to the lead during the first cupellation.
But in the assay of rich gold alloys such approximate work will not do.
If the composition is not already known with a fair degree of accuracy
_preliminary assays_ must be made. Weigh up two lots of 100 milligrams
of the alloy and wrap each in 3 grams of lead. To one add 300 milligrams
of silver. Cupel both. The button containing the added silver must be
flattened and boiled with 15 c.c. of nitric acid; and the resulting
gold must be washed, dried, ignited and weighed. This, in milligrams,
gives directly the percentage of gold. The weight of the other button
gives the percentage of gold and silver; the difference between the two
gives the percentage of silver. The rest will, perhaps, be copper.
The composition of the alloy being known, or having been determined as
just described, the calculation of how much silver must be added is
fairly simple. The following is an example. Suppose the bullion contains
92 per cent. of gold, 1 per cent. of silver and 7 per cent. of copper,
and that .5 gram of it is to be taken for an assay. The .5 gram, then,
will contain
Gold .460 gram
Silver .005 "
Copper .035 "
But the total silver required is .46 gram × 2.5. This equals 1.15.
Allowing for the .005 gram of silver already present, 1.145 gram of
silver must be added.
The silver is incorporated with the gold, and at the same time the
copper is eliminated, by cupelling with sheet lead. How much sheet lead
must be used will depend partly on how much bullion is taken, partly on
how much copper it contains. Four grams of lead will do for a .5 gram
charge; and for a .3 gram charge, 3 grams may be used. But with 20 per
cent. of copper these amounts should be doubled; with 40 per cent. of
copper they should be trebled; and with over 60 per cent. of copper four
times as much lead should be used. For small buttons of gold as little
lead as may be relied on to start cupelling may be taken; the lead may
conveniently be in the form of little cups made by folding lead foil on
a piece of glass rod. With a large number of bullion assays
systematically worked and checked a simple plan would be to always use
the quantity of lead required by the alloy containing most copper which
turns up for assay. This weight, cut out of lead foil, would be kept in
stock folded into little bags ready to receive the bullion and silver.
The silver used for inquartation must, of course, be free from gold and
is best prepared by the assayer who is to use it (see p. 66). It should
not be in long strips or angular pieces likely to perforate the lead in
which it is folded. When wrapped in the lead it should be in the middle
and should make as compact a parcel as possible.
Each little parcel, as completed, should be placed on a tray in its
properly numbered compartment. Its position here should correspond to
that it will occupy in the muffle and eventually in the cupel tray. The
cupellation must be made with all the requisite precautions. A good
smooth malleable button is needed for the next operation, which is known
as flatting.
[Illustration: FIG. 45.]
~Flatting.~--Small buttons, such as are got in assaying most gold ores,
are placed on a polished steel anvil and flattened by one or two blows
with a hammer. The flattened discs are heated to dull redness on a clean
cupel and are then ready for parting. Somewhat larger buttons may be
similarly treated, but they should be annealed (_i.e._ heated to redness
and allowed to cool) during the flattening. The silver-gold alloy left
from the cupellation is soft and bends like lead; but after hammering or
rolling it becomes harder, gets a spring in it like a piece of
mainspring and cracks or splits somewhat easily. There should be no
cracks or stripping or even roughness on the flattened metal, since such
defects may cause the loss of small particles either during the
flattening or in the subsequent treatment with acid. The softness of the
metal is restored by heating. In bullion assays the flatting of the
buttons requires care and practice for its skilful working. The strips
of alloy for parting should be of uniform thickness and condition so
that the action of the acid shall be equal in all cases. The button is
taken from the cupel, cleaned and placed on the anvil: it is then struck
a heavy blow which widens it to about 3/4 inch in diameter; this blow is
followed by two others, one a little in front, the other behind, which
lengthen the disc and give a very blunt roof-like slope to its upper
face. It should then be annealed. This may be done by putting it in a
just red-hot scorifier heated in a muffle: it very soon attains the
right heat and may then be transferred to a cold scorifier; the hot
scorifier should be put back into the muffle. The softened disc is then
taken to the rolls (Fig. 45). The rolls are loosened until the disc can
be pressed between them. Looking through the interval between them the
rolls should appear exactly parallel; if they are not, one adjusting
screw should be loosened and the other tightened until parallelism is
obtained. The rolls are now turned and the disc should be drawn through
without any great effort. Beginners are apt to err by trying to do too
much with one turn of the handle. It is easy to stop whilst the rolls
are only just gripping the metal and then to bring the disc back by
reversing the action. If the disc was originally level and the rolls
are parallel, the metal will appear as a strip which has been merely
lengthened. If the rolls are tighter on one side the strip will be
bowed; the tighter side will correspond with the outer curve of the
crescent. A mistake of this kind may be amended by passing the strip
through the rolls the other way, so as to reverse the irregularity and
so straighten the strip. The screw on the looser side should then be
tightened until parallelism is obtained; after which more care should be
taken to tighten the two screws equally. The rolling should be stopped
when the strip is 3 or 4 inches long and of the thickness of an ordinary
visiting card. The strip should be annealed during the rolling and again
at the finish.
~Parting.~--The thin sheet of metal is dropped into hot dilute nitric
acid and boiled for five or six minutes after the brisk action of the
acid on the metal has ceased. At this stage nearly all the silver has
gone into solution as nitrate of silver and the acid is charged with
this salt. This acid is poured off and the residual metal is again
boiled for from 20 to 30 minutes with a second lot of stronger acid.
This leaves the gold almost pure, though it may still retain from .05 to
.1 per cent. of silver. Treatment with the first acid only would
probably leave three or four times as much.
The _nitric acid_ used should be free from hydrochloric, sulphuric,
iodic and telluric acids. In testing it for the first of these add
nitrate of silver and dilute with distilled water; there should be no
turbidity. In testing for the others evaporate three lots in dishes over
a water-bath. Test one for sulphates by adding water and barium
chloride. Test another for iodates by taking up with a little water,
adding a few drops of starch paste and then dilute sulphurous acid
solution a little at a time; there should be no blue colour. Test the
third for tellurium by heating with 1 c.c. of strong sulphuric acid
until dense fumes come off; allow to cool considerably; a piece of tin
foil added to the warm acid develops a fine purple colour if only a
trace of tellurium is present.
The presence of lower oxides of nitrogen, which impart a brown colour to
the acid, is objectionable; they, however, are removed by boiling the
diluted acid before using it for parting. It is usual to keep a stock of
the acid suitably diluted to the two strengths required for the parting.
These are known as the parting acids. The _first parting acid_ is the
weaker and is used in the first attack on the metal. The specific
gravity generally recommended for it is about 1.2. It may be prepared
either by diluting the strong acid with about its own volume of
distilled water, or by suitably diluting the second parting acid which
has been already used in an assay; the small proportion of silver this
contains is not harmful for this purpose. The _second parting acid_ has
a specific gravity of about 1.3, and may be made by diluting the strong
acid with half its volume of distilled water.
_Parting in Flasks._--Flasks are most convenient for the larger
partings, as in bullion assays; and should always be used for this
purpose unless some of the special parting apparatus, like that used in
Mints, is available. Many assayers use flasks, though of a smaller size,
for the ordinary partings in assaying gold ores. The flasks are either
bulbs with long necks (Fig. 46) which ought to be heated on rose burners
of special construction; or they are small flat-bottomed conical flasks
which may be conveniently heated on a hot-plate and are, in this
respect, much easier to deal with in general work. The following
instructions apply to the parting of an alloy containing a few decigrams
of gold together with the proper proportion of silver.
[Illustration: FIG. 46.]
The strip from the rolls, after being softened by annealing, is folded
on itself on a glass rod into a roll or cornet. It should be so plastic
that it will retain the shape thus given it and not spring open on
removing the pressure of the fingers. About 50 c.c. of the first parting
acid are placed in a 6-ounce conical flask and heated to boiling; the
flask is then withdrawn, and tilted a little to one side, whilst the
cornet is cautiously dropped into it; there will be a sudden issue of
hot vapours and a prompt withdrawal of the hand is advisable. The flask
is replaced on the hot plate and the acid is kept boiling for 10 or 15
minutes. The flask is then withdrawn and the acid diluted with about an
equal volume of distilled water. If the flask has a thick glass band
around its neck, a little way down,[28] care must be taken to use hot
water, for any sudden chill will certainly crack the flask where it is
thus thickened. The liquor is carefully decanted into a clean beaker and
is then thrown into a jar marked "waste silver." About 40 c.c. of the
second parting acid, heated to boiling, is then poured into the flask,
which is then replaced on the hot plate. The boiling is continued for 15
or 20 minutes or even longer. At this stage bumping has to be specially
guarded against; after a little experience it is easy to see when this
is imminent and the flask should be withdrawn to a cooler part of the
plate; it is better to prolong the heating at a temperature below
boiling than to run the risk of disaster. Some of the older writers,
however, are rather insistent on vigorous boiling with large bubbles.
The addition of a small ball of well-burnt clay of about the size of a
pea has been recommended, as it lessens the tendency to irregular and
dangerous boiling. At the end of the treatment with the second acid the
flask is withdrawn from the plate and the acid is diluted with an equal
volume of distilled water. The liquor is carefully decanted into a
beaker, and then poured into a jar or Winchester marked "acid waste"; it
serves for making the first parting acid. The flask is then washed twice
with hot distilled water; the washings must be carefully decanted from
the gold. The flask is then filled with water. A parting cup (size B) is
then placed over its mouth, like a thimble on the tip of a finger. This
cup is of unglazed porous earthenware of such texture that it absorbs
the last few drops of water left on drying; and with a surface to which
the gold does not adhere even on ignition. The gold should fall out
cleanly and completely on merely inverting the cup over the pan of the
balance. The flask and cup are then inverted so that the flask stands
mouth down in the cup; a little of the water from the flask flows into
the cup, but only a little. The gold falls steadily through the water
into the cup. When time has been allowed for even the finest of the gold
to have settled into the cup, the flask is removed. This is easiest done
under water. The cup, with the flask still resting in it, is dipped
under water in a basin; as soon as the neck of the flask is immersed the
crucible can safely be drawn away from under it and then lifted out of
the water. The flask should not be taken away first, for the rush of
water from it may easily sweep the gold out of the cup. The water in the
cup is then drained off and the cup is dried at not too high a
temperature; for if the last drop or two of water should boil there is
danger of spattering the gold out of the crucible. When it is dry, the
cup is heated on a pipe-clay triangle over a Bunsen burner, or on a slab
of asbestos in a muffle, to a dull-red heat. This brings the gold to
"colour"; that is, the loose tender dark coloured gold becomes bright
yellow and coherent; and is in a state fit to be transferred to the
balance and weighed. All unnecessary transferences must be avoided. As
soon as the cup is cool it may be inverted over the pan of the balance,
when the gold will fall out cleanly or, at the worst, a gentle tap with
the finger will be sufficient to detach it.
_Parting in test-tubes_, or in the smaller conical flasks, is used in
the assay of gold ores of ordinary richness. The work is exactly like
that just described in all its main features. Generally speaking much
less acid will be used; for example, in test-tubes and for small
buttons, 3 or 4 c.c. of each acid is quite enough. Again, the action
need not be so prolonged; 10 or 15 minutes in each acid is sufficient.
So, too, the heating may be less; it is very convenient to support the
test-tubes in a water-bath, or merely to rest them in a beaker of
boiling water; and there is no serious objection to doing this. A
smaller parting cup should be used; the A size is suitable. The button,
on the other hand, should be beaten thinner than is needed for the
larger partings. If the silver should be in excess and the gold becomes
much broken up, ample time should be given for subsidence from the
test-tube or flask into the parting cup.
_Parting in glazed crucibles or dishes._--This method of working has the
advantage that there is no transference of the gold until it is placed
on the pan of the balance. On the other hand, in the boiling more care
is required in adjusting the temperature. The following instructions
apply to the treatment of very small buttons, to which the method is
more particularly applicable; but very little modification is needed for
the treatment of larger buttons. The smallest sized Berlin crucibles
answer admirably. They should be cleaned by treatment with hot and
strong sulphuric acid, followed by washing in distilled water; the
comfort and ease of working mainly depends on the thoroughness of this
cleaning. The crucible, one-third full with the first parting acid, is
heated on the hot plate until the acid is almost boiling. The flattened
and annealed button is dropped into it and the heating continued with,
at most, gentle boiling for a few minutes. The crucible is then filled
with distilled water, which cools it enough for easy handling; and when
the gold has settled the liquor is poured off along a glass rod into a
clean beaker. Any greasiness of the crucible makes itself felt here and
is very objectionable. The crucible is then one-third filled with the
second parting acid and the heating resumed, care being taken not to
raise the temperature too high; this should be continued much longer
than before, say for five or ten minutes or even longer according to the
size of the button. Distilled water is again added and, when it is
drained off, the washing with distilled water is twice repeated. It will
not be possible to drain off the last drop of water; but if the gold is
coherent, the crucible can be so inclined that this drop drains away
from the gold, in which case the drying can be done rapidly; the boiling
of the water will do no harm. But when the gold is much broken up, it
will collect in the middle of this drop and the drying must be done
gently; best by putting the crucible in a warm place. When dry, the
crucible is heated till the gold changes colour, but the heat must be
kept well below redness. When cold, the gold is transferred directly to
the pan of the balance. With minute specks of gold which will require
measuring, it is best to put a small piece of lead foil (say .1 gram) in
the crucible over the gold, and then heat the crucible to above redness
over a blowpipe. Whilst the lead is oxidising it is easily swept round
in a bath of molten litharge by merely tilting the crucible. In this way
any separated specks of gold can be taken up with certainty. When the
worker is satisfied that the lead has had ample opportunity for taking
up the gold, the lead must be kept in one place and the heat slowly
lowered. By this means the button becomes supported in comparatively
pure litharge and when solid can be picked out quite easily with a pair
of pliers and in a very clean condition. The lead button is then
cupelled on a very fine cupel, as already described. The method of
working last described destroys the crucible. If the gold is not quite
so small this may be avoided. A small piece of lead foil should be
hammered out until it is perfectly flexible. It is then shaped into a
tray and the gold is transferred to it. The lead is then folded over,
with the help of two pins; and cupelled.
If the crucible shows a black stain on heating it is because some silver
remains through bad washing. It shows poor work and the assay should be
repeated.
_The silver retained in the gold after parting_ is, in bullion assays,
an important matter; it is roughly equal to the loss of gold due to
absorption by the cupel. Mr. Lowe working on .5 oz. of gold, obtained by
parting in assaying bullion, found it to contain .123 per cent. of
silver. Dr. Rose in some special assay pieces found by a less direct
method of assaying, from .06 to .09 per cent. of silver. The proportion
of silver retained varies in a marked way with the proportion of gold to
silver in the alloy before parting. It is generally stated that the
retained silver is least when this proportion is 1 to 2-1/2, and more or
less silver than this leads to a less pure gold after parting.
_Platinum_ in an alloy being parted is dissolved along with the silver
either altogether or in part. It imparts a straw yellow colour to the
parting acid. _Palladium_ gives an orange colour to the acid.
_The loss of gold by solution in the acid during parting_ is small, but
easily demonstrable. On a 500-milligram charge of bullion it may amount
to from .05 to .15 milligram; _i.e._ from .01 to .03 per cent. It is due
to gold actually dissolved and not merely held in suspension.
~Assaying with checks. Surcharge.~--It will be seen from what has been
stated that the errors in gold parting are of two kinds: viz. (1) a loss
of gold on the cupel and to a less extent by solution in the acid, and
(2) an apparent gain of gold due to the retention of silver in the
parted material. Both errors are small, and as they are of an opposite
character they tend to neutralise each other. Hence they are altogether
without effect on the accuracy of the assays of ores when the total gold
is reckoned in milligrams. And even with the larger amounts present in
bullion assays their influence is so small that an uncorrected result is
still fairly accurate; the resultant error would not be more than one
part in two or three thousand.
It is customary to report the purity of bullion, or its fineness as it
is called, in parts per thousand of bullion. The sum of the errors of an
assay, which is called the _surcharge_, is reported in the same way.
Thus a surcharge of + .3 means that the gold as weighed was .3 part per
1000 more than the gold actually present. But a surcharge - .3 means
that on the whole there was a loss of .3 part per 1000 in the assay.
Speaking roughly the retained silver will vary with the weight of gold
present; if one alloy contains twice as much gold as another the
retained silver will be about twice as much also. On the other hand, as
already explained, the cupellation loss on the poorer alloy is as much
as, or even more than, with the richer one, because of the copper, &c.
present. With rich gold alloys the silver more than compensates for the
loss and the surcharge is positive; but with poorer alloys the loss is
greater and the surcharge is negative.
In Mints and places where bullion assays must be made with the highest
attainable accuracy, the surcharge is determined by experiment, and the
proper correction is made in the reports on the bullion. This is done by
making assays of gold of the highest degree of purity alongside of those
of the bullion whose quality has to be determined. These "checks" are so
made that they do not differ from the actual assays in any material
point. Thus, being of the same quality and weight and undergoing exactly
the same treatment, they may reasonably be expected to have the same
surcharge as the assays they imitate. Suppose the bullion being assayed
varies only a little, up or down, from 900 gold and 100 copper in the
thousand, and that .5 gram of it is used in each assay. A quantity of
gold differing only a little from .450 gram would be very exactly
weighed and placed with .050 gram of copper in the same weight of lead
as is being used in the other assays. It would be cupelled, parted, &c.,
as nearly as possible under the same conditions as the actual assays.
Suppose the pure gold weighed .45016 gram and the parted gold weighed
.45025 gram, the gain in weight, .00009 gram, would be deducted from the
actual assays. A surcharge correction is never applied except to bullion
of the same quality as that represented by the "check assay" it was
calculated from.
It is evident that unless the gold is of the highest degree of purity
these check assays will introduce an error almost equal to that which it
is designed to remedy. Moreover, to work the checks to the greatest
advantage, a very systematic and uniform method of working must be
adopted.
~Parting in special apparatus.~--One plan for obtaining greater
uniformity is to stamp each cornet with a number for purposes of
identification, and to treat several, including one or more check assays
in the same acid contained in a beaker; all the assays under these
conditions evidently receive precisely the same acid treatment. Such a
plan can of course only be adopted where there is no risk of the gold
breaking up during the parting. An improvement on this is to have a
porcelain basin[29] about 8-1/2 inches in diameter and with a capacity
of about 1-1/2 litres. It is provided with a porcelain cover with 30
numbered holes through which tubes dip into the acid. The cover is
removable. The tubes are like test-tubes and are supported by the cover;
their bottoms are perforated with holes or slits. The acid is placed in
the basin and boiled over a flat burner; it enters the tubes through the
slits. The cornets are placed each in its proper tube. When the boiling
is finished, the cover with the tubes is lifted and at the same time the
acid drains back into the basin. A dip into a basin of distilled water
washes at one operation all 30 assays. The cover is then put on a basin
containing the stronger parting acid which is already boiling. This
boiling is continued for half an hour. The cover with the 30 cornets is
then lifted out from the acid and dipped two or three times in distilled
water to wash off the last traces of acid. To transfer the cornets from
the tubes to the porous cups the whole of the tube must be dipped under
the water; otherwise the operation is exactly as when working with
test-tubes.
A still simpler method of working is to use small platinum cups[30]
provided with fine slits which admit the acid but retain the gold. A
number of these, say 60, are supported on a platinum tray. The parting
acids are boiled in platinum dishes under a hood; and the 60 cornets
(each in its proper cup) are placed in the acid all at once: the tray
carrying the cups is provided with a handle suitable for this purpose.
After a proper boiling the tray is lifted out of the weaker acid into
the stronger one, where it undergoes the second boiling. It is next
dipped several times in distilled water and lastly, after a gentle
drying, it is raised to an annealing temperature which must not be too
high for fear of the gold sticking to the platinum. After cooling, the
cornets are transferred from the platinum cups directly to the pan of
the balance. Here all 60 cornets have exactly the same treatment and the
"checks" may be compared with great exactness with the other assays
accompanying them. There is, too, a great saving of labour.[31]
~Silver, &c., in gold bullion.~--The base metals are generally
determined by cupelling .5 gram of the alloy with 5 grams of lead. The
loss in cupellation having been allowed for by any of the usual methods
(see p. 104) the gold and silver contents are given. By deducting the
gold the proportion of silver is obtained. The silver is generally
determined by difference in this way. If it is desired to dissolve out
the copper, silver, &c., and to determine them in the wet way, the gold
must first be alloyed with a sufficiency of some other metal to render
it amenable to the attack by acid. Cadmium is the metal generally
recommended, and the alloy is made by melting together a weighed portion
of the gold with five or six times its weight of cadmium in a Berlin
crucible and under a thin layer of potassium cyanide.
~Lead with gold or silver.~--Large quantities of lead carrying gold and
silver are sold to refiners in bars weighing about 100 lbs. each. The
assay of these alloys presents no special difficulties, but the sampling
of them is a question which may be profitably discussed.[32]
A molten metal may be conceived to have all the physical states observed
in ordinary liquids, although these cannot be actually seen owing to its
opaqueness. There is no doubt that _pure_ lead at a temperature only a
little above its melting-point can contain a large proportion of gold in
such a manner that it may in a figurative way be spoken of as a clear
solution. Any small portion withdrawn from the molten metal would afford
a perfect sample. The same would be true of any pure alloy of lead and
silver in which the silver does not exceed the proportion of 2-1/2 per
cent.[33] On the other hand, if the molten metal contains much more than
.5 per cent. of zinc, more than .1 per cent. of copper, or a larger
quantity of silver, it may be likened to a turbid liquor. The
resemblance holds good so far that if the molten lead be further heated,
whereby its solvent power on the added metal is increased, the turbidity
will disappear, or at least be considerably diminished. A portion taken
at random from such a molten metal may, or may not, give a good sample.
The suspended insoluble matter will tend to concentrate itself in the
upper or lower parts of the liquid according to whether it is heavier or
lighter than it; and this separation may occur with extreme slowness or
with fair rapidity. However, it is generally agreed that in the case of
such alloys as occur in practice, samples taken in this way are quite
satisfactory and are the best obtainable. The precautions insisted on
are that the lead shall be made as hot as practicable; that it shall be
stirred up at the time of taking the sample; and that the portion
withdrawn shall be taken out with a ladle at least as hot as the molten
metal. The further precaution that if any dross be on the surface of the
metal it shall be skimmed off and separately sampled and assayed is
almost too obvious to require mention. An alternative and, perhaps,
better way of taking the sample is to withdraw portions at equal
intervals from the stream of metal whilst the pot is being emptied;
equal weights taken from these portions and mixed (by melting or in some
other way) give a fair sample of the whole. In addition, separate assays
of each portion will show to what extent the metal lacks uniformity in
composition For example, samples taken at the beginning, middle, and end
of a run gave the following results in ozs. of silver per ton: 475, 472,
466, showing an average result of 471 ozs. Fifteen fractions taken at
regular intervals during the same pouring ranged from 475 ozs. to 464
ozs.: the average result was 469.8 ozs. The same lead cast into bars and
sampled by sawing gave an average of 470 ozs.[34] In another case[35]
samples drawn at the beginning, middle, and end of a run gave 1345 ozs.,
1335 ozs. and 1331 ozs. The mean result in such cases is always a
reasonably safe one, but evidently where the metal varies a good deal it
is safer to take more than three dips.
Imagine such lead run into moulds and allowed to become solid as bars;
the difference between bar and bar would not be greater than that
between corresponding dip samples. But in each bar the distribution of
the silver and gold is very seriously affected during solidification.
Chips taken from the same bar of auriferous lead may show in one place
23 ozs. of gold to the ton, in another 39 ozs.; similarly with silver
they may vary as much as from 900 ozs. to 1500 ozs. to the ton.
This rearrangement of the constituents of a bar takes place whilst the
lead is partly solid, partly liquid. The most useful conception of such
half-solidified metal is that of a felted spongy mass of skeleton
crystals of comparatively pure lead saturated with a still fluid
enriched alloy. If the solidification of an ingot of impure tin be
watched it will be evident that the frosted appearance of the surface is
due to the withdrawal of the fluid portion from a mat of crystals of
purer tin which have been for some time solid and a contraction of the
mass. The shrinking of the last part to become solid is further shown by
the collapse of the surface of the ingot where weakest; that is, a
furrow is formed on the flat surface. In other cases of fused metal
there is expansion instead of contraction in this final stage of the
solidification, and the enriched alloy then causes the upper face of the
ingot to bulge outwards. There are other causes effecting the
redistribution of the metals through the ingot. There can be no general
rule of wide application showing which part of a bar is richest and
which poorest in the precious metals. This will depend on the quantities
of gold or silver, on the quantities and kinds of other metals present
and on the manner of casting. The student is advised to consult Mr.
Claudet's paper which has been already referred to.
The best method of sampling such bars is to melt them all down and to
take a dip sample of the molten metal in one or other of the methods
already described. According to Mr. Claudet this should be done in all
cases where the gold exceeds one or two ounces or where the silver
exceeds 200 ozs. to the ton. If during the melting down some dross has
formed this must be skimmed off, weighed and separately sampled and
assayed. The clean lead also must be weighed, sampled and assayed. The
mean result must be calculated. Thus 14 tons 5 cwts. of clean lead
assaying 32 ozs. to the ton will contain 456 ozs. of silver; 15 cwt.
dross assaying 20 ozs. to the ton will contain 15 ozs. of silver. The 15
tons of lead and dross will contain 471 ozs. of silver or 31.4 ozs. per
ton.
Of the methods of sampling which avoid melting the bars, that known as
sawing is the only one which is thoroughly satisfactory. In it the bars
are brought to a circular saw having fine teeth and are sawn across
either completely or halfway through; in this way a quantity of lead
sawdust is obtained (say 1 lb. or so from a bar) which represents
exactly the average of the bar along the particular cross section taken
and approximately that of the whole bar. A bar of lead, which by dip
assay gave 334 ozs. to the ton, gave on three transverse sections 333
ozs., 335 ozs. and 331 ozs. The variation may be greater than this, but
with a large number of bars, where each bar is cut across in as far as
possible a different place, these variations tend to neutralise each
other and a good sample is obtained. Two or three cwt. of sawdust may be
obtained in this way; this is thoroughly mixed and reduced by
quartering in the usual way or by a mechanical sampler. A sample of 2 or
3 lbs. is sent to the assayer. This being contaminated with the oil used
in lubricating the saw is freed from it by washing with carbon
bisulphide, ether or benzene and dried. Then, after mixing, 100 to 200
grams of it are carefully weighed and placed in a hot crucible, the heat
of which should be sufficient to melt all the lead. The molten lead
should not be overheated and should show no loss due to the melting. The
removal of the oil may have decreased the weight by perhaps one half per
cent. If the lead gives dross on heating it may be melted under 10 or 20
grams of potassium cyanide, which prevents the formation of dross.
Samples are sometimes taken with a drill, gouge or chisel, though no
method of this kind is quite satisfactory. One plan adopted is to use a
punch which, when driven into the bar, gives a core or rod of metal
about half as long as the bar is thick and about one-eighth of an inch
across. With five bars side by side it is customary to drive in the
punch at one end on the first bar, and at the opposite end on the last
one, and on the others in intermediate positions in such a manner that
all the holes will be along a diagonal of the rectangle enclosing the
bars. The bars are then turned over and similar portions punched out
through the bottoms of the bars and along the other diagonal. Or one set
of five may be sampled along the top and the next set along the bottom
of the bars.
Silver and gold present in bars of copper are subject to the same
irregularity of distribution as in lead. The sampling of such bars is
guided by the same principles.[36]
CYANIDES.
The cyanides ought perhaps to be considered along with chlorides,
bromides and iodides in Chapter XV. But they are treated here because
they owe their importance to their use in the extraction of gold and
because their determination has become a part of the ordinary work of an
assayer of gold ores.
Formerly, the cyanide most easily obtained in commerce was potassium
cyanide; and it was generally sold in cakes which might contain as
little as 40 per cent. or as much as 95 per cent. of the pure salt. It
became customary to express the quality of a sample of commercial
cyanide by saying it contained so much per cent. of potassium cyanide.
The commercial product now made by improved methods of manufacture is
actually sodium cyanide, but is called "potassium cyanide" (probably
with the words "double salt" on the label); it contains cyanide
equivalent to something over 100 per cent. of potassium cyanide in
addition to a large proportion of sodium carbonate and other impurities.
What is wanted in most cases is merely a soluble cyanide, and it is a
matter of indifference whether the base be sodium or potassium. But
since 49 parts of sodium cyanide (NaCN = 49) are equivalent to 65 parts
of potassium cyanide (KCN = 65) it is evident that a pure sample of
sodium cyanide would contain cyanide equivalent to little less than 133
per cent. of potassium cyanide. Therefore a sample of cyanide reported
on in this way may be rich in cyanide, and yet have much impurity.
The commonest impurity in commercial cyanide is carbonate of sodium or
potassium. This may be tested for by dissolving, say, 2 grams in a
little water and adding barium chloride. There may be formed a white
precipitate of barium carbonate, which if filtered off, washed and
treated with acid, will dissolve with effervescence. Cyanate may be
tested for in the solution from which the barium carbonate has been
filtered by adding a little soda and boiling; if cyanates are present
they decompose, giving off ammonia (which may be tested for in the
steam) and yielding a further precipitate of barium carbonate.[37] If
the soda alone gave a further precipitate of barium carbonate, this may,
perhaps, be due to the presence of bicarbonates. Alkaline sulphides may
be present in small quantity in commercial cyanide. Their presence is
shown at once when the sample is being tested for its strength in
cyanide, inasmuch as the first few drops of silver nitrate solution
produce at once a darkening of the liquor. A special test for sulphide
may be made by adding a drop or two of solution of acetate of lead to
four or five c.c. of soda solution and adding this to a clear solution
of the suspected cyanide. This will cause a black precipitate or colour,
if any sulphide is present.
The cyanides of the heavier metals combine with the alkaline cyanides to
form double cyanides. Some of these, ferrocyanide and ferricyanide of
potassium for example, have such characteristic properties that the fact
that they are cyanides may be overlooked. Others, such as potassium zinc
cyanide (K_{2}ZnCy_{4}), have much less distinctiveness: they behave
more or less as a mixture of two cyanides and are, moreover, so easily
decomposed that it may be doubted if they can exist in dilute alkaline
solutions. In reporting the cyanide strength of a cyanide liquor as
equivalent to so much per cent. of potassium cyanide, there is a
question as to whether the cyanide present in the form of any of these
double cyanides should be taken into account. It must be remembered that
the object of the assay is not to learn how much of the cyanide exists
in the solution as actual potassium cyanide; reporting the strength in
terms of this salt is a mere matter of convenience; what is really
desired is to know how much of the cyanide present in the liquor is
"free" or "available" for the purposes of dissolving gold. Every one is
agreed as to the exclusion of such cyanides as the following: potassium
ferrocyanide (K_{4}FeCy_{6}), potassium ferricyanide (K_{3}FeCy_{6}),
potassium silver cyanide (KAgCy_{2}), and potassium aurocyanide
(KAuCy_{2}); and the double cyanides with copper or nickel. But with
cyanide liquors containing zinc the position is less satisfactory. One
method of assay gives a lower proportion of cyanide when this metal is
present; and the loss of available cyanide thus reported depends, though
in a fitful and uncertain way, upon the quantity of zinc present. The
other method of assay reports as full a strength in cyanide as if no
zinc were present. Unfortunately, using both methods and accepting the
difference in the results as a measure of the quantity of zinc present,
or at any rate of the zinc present as cyanide, is not satisfactory. It
appears best to use the method which ignores the zinc; and to determine
the amount of zinc by a special assay of the liquor for this metal.
The cyanide present as hydrogen cyanide or prussic acid (HCy) is
practically useless as a gold solvent. Hence any report on the strength
of a cyanide liquor which assigned to this the same value as its
equivalent of alkaline cyanide would be misleading. On the other hand,
it is "available cyanide" inasmuch as a proper addition of sodium
hydrate[38] would restore its value. The question of the presence or
absence of free prussic acid is involved in the larger one as to whether
the cyanide solution has the right degree of alkalinity. The assay for
"cyanide" should include the hydrogen cyanide with the rest.
A rough test of the power of a cyanide liquor for dissolving gold may be
made by floating a gold leaf on its surface and noting the time required
for its solution. This test might, perhaps, be improved by taking, say,
20 c.c. of the liquor and adding three or four gold leaves so that the
gold shall always be in considerable excess. The liquor should not be
diluted as this will affect the result. It should be allowed to stand
for a definite time, say at least two or three hours, or better, that
corresponding to the time the liquor is left in contact with the ore in
actual practice. The liquor should then be filtered off and, with the
washings, be evaporated in a lead dish as in the assay of cyanide
liquors for gold (p. 141). The gold obtained on cupelling, less any gold
and silver originally present in the liquor, would be the measure of the
gold dissolving power.
THE ASSAY FOR CYANIDE BY TITRATION WITH SILVER NITRATE.
The determination of the quantity of a cyanide is made by finding how
much silver nitrate is required to convert the whole of the cyanide into
potassium silver cyanide[39] or one of the allied compounds. It will be
seen from the equation that 170 parts by weight of silver nitrate are
required for 130 parts by weight of potassium cyanide. As already
explained it is customary to report the cyanide-strength in terms of
potassium cyanide, even when only the sodium salt is present. One gram
of potassium cyanide will require 1.3076 gram of silver nitrate. _The
standard solution of silver nitrate_ is made by dissolving 13.076 grams
of silver nitrate in distilled water and diluting to 1 litre; 100 c.c.
of such a solution are equivalent to 1 gram of potassium cyanide.[40]
The titration is performed in the usual way, running the standard
solution of silver nitrate into a solution containing a known weight or
volume of the material containing the cyanide. The _finishing point_ is
determined in one of two ways, both of which are largely used. In the
first place, as long as there remains any free cyanide in the solution
the silver nitrate will combine with it forming the double cyanide and
yielding a clear solution; but as soon as all the free cyanide is used
up the silver nitrate will react with the double cyanide[41] forming
silver cyanide, which separates as a white precipitate and renders the
solution turbid. But, in the second place, if potassium iodide is
present in the solution the excess of silver nitrate will react with
it,[42] rather than with the double cyanide; and silver iodide will
separate as a yellowish turbidity which is easily recognised.
In working with pure solutions, the two finishing points give the same
results; and this is true even when there is much difference in the
degree of dilution. The finishing point with the iodide, however, has
an advantage in precision. Moreover, it is but little affected by
variations in alkalinity, which render the other finishing point quite
useless. The great difference between the two is shown when zinc is
present in the solution. In this case, when working without the iodide,
the first appearance of a turbidity is less distinct; the turbidity
increases on standing and as a finishing point is unsatisfactory. It can
be determined with precision only by very systematic working and after
some experience. The turbidity is due to the separation of an insoluble
zinc compound. A most important point (to which reference has already
been made) is that less silver nitrate is required to give this
turbidity and, consequently, a lower strength in cyanide is reported. On
the other hand, as much silver nitrate is required to give the yellow
turbidity due to silver iodide as would be required if no zinc were
present.
Unfortunately the difference in the two titrations does not depend
merely on the quantity of zinc present; as it is also influenced by the
extent of dilution, the degree of alkalinity of the solution, and the
quantity of cyanide present. In an experiment with .055 gram of zinc
sulphate and .1 gram of potassium cyanide the difference in the two
finishing points was only .1 c.c.; whereas with .4 gram of potassium
cyanide, the other conditions being the same, the difference was 1.5
c.c. of standard silver nitrate. On the assumption that all the zinc was
present as potassium zinc cyanide (K_{2}ZnCy_{4}) the difference should
have been 5 c.c. in each case. Again, repeating the experiment with .4
gram of potassium cyanide, but with .11 gram of crystallised zinc
sulphate, the difference was 6.5 c.c.: that is, merely doubling the
quantity of zinc increased the difference by more than four times. Hence
it would appear better to use the method with the iodide and make a
separate assay for the zinc. But since the student may be called on to
use the other method, he is advised to practice it also.
~The assay without iodide.~--The standard solution of silver nitrate is
placed in a small burette divided into tenths of a c.c. Ten c.c. of the
cyanide solution to be assayed is transferred to a small flask and
diluted with water to about 70 c.c. The silver solution is then run in
from the burette (with constant shaking of the flask), a little at a
time but somewhat rapidly, until a permanent turbidity appears. Since 1
c.c. of the silver nitrate solution corresponds to .01 gram of potassium
cyanide, it also corresponds to .1 per cent. of this salt counted on the
10 c.c. of cyanide solution taken. The titration should be performed in
a fairly good uniform light. The learner should practice on a fairly
pure solution of potassium cyanide at first, and this may conveniently
have a strength of about 1 per cent. For practice with solutions
containing zinc make a solution containing 1.1 gram of crystallised zinc
sulphate in 100 c.c. and slowly add measured quantities of from 1 to 5
c.c. of this to the 10 c.c. of cyanide liquor before diluting for the
titration.
If a cyanide solution blackens on the addition of the silver nitrate it
contains sulphide. In this case, shake up a considerable bulk of the
liquor with a few grams of lead carbonate, allow to settle and make the
assay on 10 c.c. of the clear liquor.
If the cyanide liquor be suspected to contain free prussic acid, take 10
c.c. for the assay as usual; but, before titrating, add .1 or .2 gram of
sodium carbonate. On no condition must caustic soda or ammonia be added.
The difference between the results, with and without the addition of
carbonate of soda, is supposed to measure the quantity of free prussic
acid. If this has to be reported it is best done as "prussic acid
equivalent to ... per cent. of potassium cyanide." Suppose, for example,
the difference in the two titrations equals 1 c.c. of standard silver
nitrate; the prussic acid found would be equivalent to .1 per cent. of
potassium cyanide.
~The assay with iodide.~--The standard solution of silver nitrate is
placed in a burette divided into tenths of a c.c. Take 10 c.c. of the
cyanide liquor, which should previously have been treated with white
lead for the removal of sulphides if these happened to be present.
Transfer to a small flask, add 3 or 4 drops of a solution of potassium
iodide and 2 or 3 c.c. of a solution of sodium hydrate; dilute to 60 or
70 c.c. with water. If much zinc is present the soda may be increased to
20 or 30 c.c. with advantage. The standard solution should be run in
somewhat rapidly, but a little at a time, so that the precipitate at
first formed shall be small and have only a momentary existence. The
titration is continued until there is a permanent yellowish turbidity.
The most satisfactory and exact finish is got by ignoring any faint
suspicion of a turbidity and accepting the unmistakable turbidity which
the next drop of silver nitrate is sure to produce. This finishing point
gives results which are exactly proportional to the quantity of cyanide
present; and it can be recognised with more than ordinary precision even
in solutions which are not otherwise perfectly clear.
Each c.c. of the standard silver nitrate solution corresponds to .01
gram of potassium cyanide; and if 10 c.c. of the liquor are taken for
assay this corresponds to .1 per cent. or 2 lbs. to the short ton or
2.24 lbs. to the long ton. As already explained the result should be
reported as "cyanide equivalent to so much per cent. of potassium
cyanide."
The following experimental results were obtained with a solution of
potassium cyanide made up to contain about 1.2 per cent. of the salt.
~Effect of varying cyanide.~--The bulk before titration was in each case
60 c.c.; 2 c.c. of soda and 3 drops of potassium iodide were used in
each case.
Cyanide added 40 c.c. 30 c.c. 20 c.c. 10 c.c. 5 c.c. 1 c.c.
Silver required 47.0 c.c. 35.25 c.c. 23.5 c.c. 11.7 c.c. 5.8 c.c. 1.15 c.c.
Accepting the result for 40 c.c. as correct, the others are in very
satisfactory agreement.
~Effect of varying dilution.~--The conditions were those of the 40 c.c.
experiment in the last series; but varying amounts of water were used in
diluting.
Water added none 100 c.c. 200 c.c. 400 c.c.
Silver required 47.0 c.c. 47.0 c.c. 47.0 c.c. 47.05 c.c.
Very considerable dilution therefore has no effect.
~Effect of varying soda.~--The conditions were those of the 40 c.c.
experiment in the first series, except that varying amounts of soda
solution were used.
Soda added none 10 c.c. 30 c.c.
Silver required 46.95 c.c. 47.0 c.c. 47.0 c.c.
This alkali therefore has no prejudicial effect.
~Effect of ammonia.~--Soda causes turbidity in some cyanide liquors;
with these it should be replaced by 2 or 3 c.c. of dilute ammonia with a
gram or so of ammonium chloride. The following experiments with dilute
ammonia show that larger quantities of this reagent must be avoided.
Ammonia added none 10 c.c. 30 c.c. 60 c.c.
Silver required 46.95 c.c. 47.15 c.c. 47.7 c.c. 49.5 c.c.
~Effect of sodium bicarbonate.~--In this experiment 1 gram of
bicarbonate of soda was used instead of the soda or ammonia of the other
experiments. The silver nitrate required was only 46.45 c.c. instead of
the 47.0 c.c. which is the normal result. This is probably due to the
liberation of prussic acid and shows the importance of having the
solution alkaline.
~Effect of zinc.~--In each experiment 40 c.c. of the cyanide solution
and .5 gram of zinc sulphate crystals were used and the bulk was made up
to 100 c.c. before titrating.
Soda added 1 c.c. 5 c.c. 10 c.c. 25 c.c.
Silver required 47.1 c.c. 47.0 c.c. 46.9 c.c. 46.9 c.c.
The work was easier with the more alkaline solutions. The titration in
the presence of zinc is comparatively easy, but, in learning it, it is
well to have a burette with cyanide so that if a titration be overdone
it can be brought back by the addition of 1 or 2 c.c. more cyanide and
the finish repeated; a quarter of an hour's work in this way will ensure
confidence in the method.
~Effect of other substances.~--It was found that an alkaline cyanate,
sulphocyanate, ferrocyanide, nitrite, borate, silicate or carbonate has
no effect. The ferricyanide had a small influence and, as might be
expected, hyposulphite is fatal to the assay. The addition of salts of
lead and cadmium was without effect. On the other hand, nickel produces
its full effect; and the quantity of nickel added can be calculated with
accuracy from the extent of its interference with the titration.
~Assay of commercial cyanide of potassium.~--Break off 20 or 30 grams of
the cyanide in clean fresh pieces, weigh accurately to the nearest
centigram. Dissolve in water containing a little sodium hydroxide;
transfer to a 2-litre flask: dilute to 2 litres; add a few grams of
white lead; shake up and allow to settle. Run 50 c.c. of the clear
liquor from a burette into an 8 oz. flask; add 2 or 3 c.c. of soda
solution and 3 drops of potassium iodide. Titrate with the standard
solution of silver nitrate. The percentage may be calculated by
multiplying the number of c.c. used by 40 (50 c.c. is one fortieth of
the 2 litres) and dividing by the weight of commercial cyanide
originally taken.
~Alkalinity of commercial potassium cyanide and of cyanide
solutions.~--Hydrocyanic acid like carbonic acid has no action on
methyl-orange;[43] hence the alkaline cyanides may be titrated with
"normal acid" as easily as the carbonates or hydrates. 100 c.c. of
normal acid will neutralise 6.5 grams of pure potassium cyanide.[44] A
solution of commercial cyanide prepared as for the assay last described,
but best without the addition of white lead, may be used for the test.
Take 50 c.c. of it; tint faintly yellow with methyl-orange and titrate
with normal acid till the liquor acquires a permanent reddish tint. In
the case of the purer samples of cyanide the quantity of acid used will
correspond exactly with that required to neutralise the actual quantity
of cyanide present as determined by the assay with nitrate of silver.
The less pure samples will show an excess of alkalinity because of the
presence of sodium carbonate or of potassium carbonate.
In comparing the alkalinity and cyanide strength of a solution the
simplest plan is to take 65 c.c. of the solution and titrate with
normal acid; for in this case each c.c. of normal acid corresponds to .1
per cent. of potassium cyanide. In systematic assays of this kind, the
alkalinity would no doubt be generally in excess of that required by the
cyanide present: there would be no inconvenience in recording such
excess in terms of potassium cyanide.
~Determination of the acidity of an ore.~--Most ores have the power of
destroying more or less of the alkalinity of a cyanide solution and in a
proportionate degree of damaging its efficiency. An assay is needed to
determine how much lime or soda must be added for each ton of ore in
order to counteract this. Whether this acidity should be reported in
terms of the lime or of the soda required to neutralise it will depend
on which of these reagents is to be used in the actual practice. Again,
if the ore is washed with water before treating with cyanide on the
large scale, then the assay should be made of the acidity of the ore
after a similar washing.
The _standard solutions of acid and alkali_ used for this determination
may be one-fifth normal. 200 c.c. of the normal solution should be
diluted to 1 litre in each case, 1 c.c. of the resulting solutions would
be equivalent to 8 milligrams of soda (NaHO) or 5.6 milligrams of lime,
CaO. It must be remembered this refers to the pure bases in each case.
Suppose it is desired to report as so many lbs. of lime to the short ton
(2000 lbs.) of ore. Since 1 c.c. of the standard solution is equivalent
to 5.6 milligrams of lime, if we take 2000 times this weight of ore
(_i.e._ 11,200 milligrams or 11.2 grams) for the assay, each c.c. of
standard solution will be equivalent to 1 lb. of lime to the short
ton.[45]
~Total acidity.~--Weigh out 11.2 grams of the ore, place them in a
four-inch evaporating dish and measure on to it from a burette 10 or 20
c.c. of the standard solution of soda. Stir the soda solution into the
ore and allow to stand for 15 or 20 minutes with occasional stirring.
Stir up with 30 or 40 c.c. of water, float a piece of litmus paper on
the liquid and titrate with the standard solution of acid. If the ore is
strictly neutral the quantity of "acid" required to redden the litmus
will be the same as the quantity of "soda" originally used. If the ore
is acid, less acid will be used. For example, if 10 c.c. of soda were
used and only 7 c.c. of acid were required, the ore will have done the
work of the remaining 3 c.c. of acid. And the ton of ore will require 3
lbs. of lime to neutralise its acidity.
~Acidity after washing.~--Take 11.2 grams of the ore; wash thoroughly
with water and immediately treat the residue, without drying, exactly as
just described.
~Examination of cyanide solutions for metals, &c.~--Take a measured
quantity of the solution, say 20 c.c.[46] and evaporate in a small dish
with, say, half a c.c. of strong sulphuric acid. Evaporate at first, on
a water-bath in a well ventilated place, but finish off with a naked
Bunsen flame, using a high temperature at the end in order to completely
decompose the more refractory double cyanides. Allow to cool; moisten
with strong hydrochloric acid; warm with a little water and test for the
metals in the solution by the ordinary methods. Since the quantities of
the metals likely to be present may be given in milligrams the work must
be carefully performed. It may be worth while to determine the
proportions of lime and magnesia as well as those of the metals proper.
Or the 20 c.c. of cyanide liquor may be evaporated with 5 c.c. of strong
nitric acid to dryness and gently ignited and the residue taken up with
2 or 3 c.c. of strong hydrochloric acid.
Copper, iron, and zinc can be rapidly determined in such a solution, as
follows. Dilute with water to 10 or 15 c.c., add an excess of ammonia,
and filter. The precipitate will contain the iron as ferric hydrate;
dissolve it in a little hot dilute sulphuric acid: reduce with
sulphuretted hydrogen; boil off the excess of gas, cool and titrate with
standard potassium permanganate (p. 236). Determine the copper in the
filtrate colorimetrically (p. 203); but avoid further dilution. Then add
dilute hydrochloric acid, so as to have an excess of 4 or 5 c.c. after
neutralising the ammonia; add some clean strips of lead foil, and boil
until the solution has for some time become colourless. Titrate with
standard potassium ferrocyanide (p. 263) without further dilution, and
bearing in mind that at most only one or two c.c. will be required.
~Examination of an ore for "cyanicides."~--Place 100 grams of the ore
with 200 c.c. of a cyanide solution of known strength (say .1 or .2 per
cent.) in a bottle and agitate for a definite time, such as one or two
days. Filter off some of the liquor and assay for cyanide, using say 20
c.c. Calculate how much cyanide has been destroyed in the operation.
Evaporate 20 c.c. with sulphuric or nitric acid and examine for metal.
Test another portion for sulphides, &c.
The student who has mastered the methods of assaying can greatly improve
himself by working out such problems as the above.
PLATINUM.
Platinum occurs in nature in alluvial deposits associated with gold and
some rare metals, generally in fine metallic grains, and, occasionally,
in nuggets. It is a grey metal with a high specific gravity, 21.5 when
pure and about 18.0 in native specimens. It is fusible only at the
highest temperature, and is not acted on by acids.
It is dissolved by warm aqua regia, forming a solution of "platinic
chloride," H_{2}PtCl_{6}. This substance on evaporation remains as a
brownish red deliquescent mass; on drying at 300° C. it is converted
into platinous chloride, PtCl_{2}, and becomes insoluble, and at a
higher temperature it is converted into platinum. All platinum compounds
yield the metal in this way. Platinic chloride combines with other
chlorides to form double salts, of which the ammonic and potassic
platino-chlorides are the most important.
Platinum alone is not soluble in nitric acid; but when alloyed with
other metals which dissolve in this acid it too is dissolved; so that in
gold parting, for example, if platinum was present, some, or perhaps the
whole of it would go into solution with the silver. Such alloys,
however, when treated with hot sulphuric acid leave the platinum in the
residue with the gold.
Platinum is detected when in the metallic state by its physical
characters and insolubility in acids. In alloys it may be found by
dissolving them in nitric acid or in aqua regia, evaporating with
hydrochloric acid, and treating the filtrate with ammonic chloride and
alcohol. A heavy yellow precipitate marks its presence.
The assay of bullion, or of an alloy containing platinum, may be made as
follows: Take 0.2 gram of the alloy and an equal weight of fine silver,
cupel with sheet lead, and weigh. The loss in weight, after deducting
that of the silver added, gives the weight of the base metals, copper,
lead, &c. Flatten the button and part by boiling with strong sulphuric
acid for several minutes. _When cold_, wash, anneal, and weigh. The
weight is that of the platinum and gold. The silver may be got by
difference. Re-cupel the metal thus got with 12 or 15 times its weight
of silver, flatten and part the gold with nitric acid in the usual way
(see under _Gold_), and the platinum will dissolve. The gold may contain
an alloy of osmium and iridium; if so, it should be weighed and treated
with aqua regia. The osmiridium will remain as an insoluble residue,
which can be separated and weighed. Its weight deducted from that
previously ascertained will give the weight of the gold.
When the platinum only is required, the alloy must be dissolved by
prolonged treatment with aqua regia, the solution evaporated to dryness,
and the residue extracted with water. The solution thus obtained is
treated with ammonic chloride in large excess and with some alcohol. A
sparingly soluble[47] yellow ammonic platinum chloride is thrown down,
mixed, perhaps, with the corresponding salts of other metals of the
platinum group. Gold will be in solution. The solution is allowed to
stand for some time, and then the precipitate is filtered off, washed
with alcohol, dried, and transferred (wrapped in the filter paper) to a
weighed crucible. It is ignited, gently at first, as there is danger of
volatilising some of the platinum chloride, and afterwards intensely.
With large quantities of platinum the ignition should be performed in an
atmosphere of hydrogen. Cool and weigh as metallic platinum.
IRIDIUM
Occurs in nature alloyed with osmium as osmiridium or iridosmine, which
is "rather abundant in the auriferous beach sands of Northern
California" (Dana). It occurs in bright metallic scales, which do not
alloy with lead, and are insoluble in aqua regia. Iridium also occurs in
most platinum ores, and forms as much as two per cent. of some
commercial platinum. In chemical properties it resembles platinum, but
the ammonic irido-chloride has a dark red colour, and on ignition leaves
metallic iridium, which does not dissolve in aqua regia diluted with
four or five times its volume of water and heated to a temperature of
40° or 50° C.
The other metals of the platinum group are Palladium, Rhodium, Osmium,
and Ruthenium. They differ from gold, platinum, and iridium by the
insolubility of their sulphides in a solution of sodium sulphide.
Palladium is distinguished by the insolubility of its iodide; and Osmium
by the volatility of its oxide on boiling with nitric acid.
MERCURY.
Mercury occurs native and, occasionally, alloyed with gold or silver in
natural amalgams; but its chief ore is the sulphide, cinnabar. It is
comparatively rare, being mined for only in a few districts. It is
chiefly used in the extraction of gold and silver from their ores
(amalgamation); for silvering mirrors, &c.
Mercury forms two series of salts, mercurous and mercuric, but for the
purposes of the assayer the most important property is the ease with
which it can be reduced to the metallic state from either of these.
Mercury itself is soluble in nitric acid, forming, when the acid is hot
and strong, mercuric nitrate. Cinnabar is soluble only in aqua regia.
Mercurous salts are generally insoluble, and may be converted into
mercuric salts by prolonged boiling with oxidising agents (nitric acid
or aqua regia). The salts of mercury are volatile, and, if heated with a
reducing agent or some body capable of fixing the acid, metallic mercury
is given off, which may be condensed and collected.
Mercury is separated from its solutions by zinc or copper, or it may be
thrown down by stannous chloride, which, when in excess, gives a grey
powder of metallic mercury, or, if dilute, a white crystalline
precipitate of mercurous chloride. Nitric acid solutions of mercury
yield the metal on electrolysis; and, if the pole on which the metal
comes down be made of gold or copper, or is coated with these, the
separated mercury will adhere thereto. It may then be washed and
weighed.
The best tests for mercury next to obtaining globules of the metal are:
(1) a black precipitate with sulphuretted hydrogen from acid solutions,
which is insoluble in nitric acid; and (2) a white precipitate with
stannous chloride.
DRY METHOD.
[Illustration: FIG. 47.]
Weigh up 5 grams, if the ore is rich, or 10 grams, if a poorer mineral.
Take a piece of combustion tube from 18 inches to 2 feet long, closed at
one end, and place in it some powdered magnesite, so as to fill it to a
depth of 2 or 3 inches, and on that a layer of an equal quantity of
powdered lime (not slaked). Mix the weighed sample of ore in a mortar
with 10 grams of finely powdered lime and transfer to the tube; rinse
out the mortar with a little more lime, and add the rinsings. Cover with
a layer of six or seven inches more lime and a loosely fitting plug of
asbestos. Draw out the tube before the blowpipe to the shape shown in
fig. 47, avoiding the formation of a ridge or hollow at the bend which
might collect the mercury. Tap gently, holding the tube nearly
horizontal, so as to allow sufficient space above the mixture for the
passage of the gases and vapours which are formed. Place the tube in a
"tube furnace," and, when in position, place a small beaker of water so
that it shall just close the opening of the tube. The point of the tube
should not more than touch the surface of the water. Bring the tube
gradually to a red heat, commencing by heating the lime just behind the
asbestos plug, and travelling slowly backwards. When the portion of the
tube containing the ore has been heated to redness for some time the
heat is carried back to the end of the tube. The magnesite readily gives
up carbonic acid, which fills the tube and sweeps the mercury vapour
before it. Some of the mercury will have dropped into the beaker, and
some will remain as drops adhering to the upper part of the neck. Whilst
the tube is still hot cut off the neck of the tube just in front of the
asbestos plug (a drop of water from the wash bottle will do this), and
wash the mercury from the neck into the beaker. The mercury easily
collects into a globule, which must be transferred, after decanting off
the bulk of the water, to a weighed Berlin crucible. The water is
removed from the crucible, first by the help of filter paper, and then
by exposing in a desiccator over sulphuric acid, where it should be left
until its weight remains constant. It should not be warmed.
_Example_:--5 grams of an ore treated in this way gave 4.265 grams of
mercury, equivalent to 85.3 per cent. Pure cinnabar contains 86.2 per
cent.
WET METHODS.
_Solution._--Since solutions of chloride of mercury cannot be boiled
without risk of loss,[48] nitric acid solutions should be used wherever
possible. No mercury-containing minerals are insoluble in acids; but
cinnabar requires aqua regia for solution. In dissolving this mineral
nitric acid should be used, with just as much hydrochloric acid as will
suffice to take it up.
To separate the mercury, pass sulphuretted hydrogen in considerable
excess through the somewhat dilute solution. The precipitate should be
black, although it comes down at first very light coloured. It is
filtered, washed, and transferred back to the beaker, and then digested
with warm ammonic sulphide. The residue, filtered, washed, and boiled
with dilute nitric acid, will, in the absence of much lead, be pure
mercuric sulphide. If much lead is present, a portion may be
precipitated as sulphate, but can be removed by washing with ammonic
acetate. To get the mercury into solution, cover with nitric acid and a
few drops of hydrochloric, and warm till solution is effected. Dilute
with water to 50 or 100 c.c.
GRAVIMETRIC DETERMINATION.
This may be made by _electrolysis_. The same apparatus as is used for
the electrolytic copper assay may be employed, but instead of a cylinder
of platinum one cut out of sheet copper should be taken, or the platinum
one may be coated with an evenly deposited layer of copper. Fix the
spiral and weighed copper cylinder in position, couple up the battery,
_and when this has been done_ put the nitric acid solution of the
mercury in its place.[49] The student had better refer to the
description of the _Electrolytic Copper Assay_.
The mercury comes down readily, and the precipitation is complete in a
few hours: it is better to leave it overnight to make sure of complete
reduction. Disconnect the apparatus, and wash the cylinder, first with
cold water, then with alcohol. Dry by placing in the water oven for two
or three minutes. Cool and weigh: the increase in weight gives the
amount of metallic mercury.
It must be remembered that copper will precipitate mercury without the
aid of the battery; but in this case copper will go into solution with a
consequent loss in the weight of the cylinder: this must be avoided by
connecting the battery before immersing the electrodes in the assay
solution. The electrolysed solution should be treated with an excess of
ammonia, when a blue coloration will indicate copper, in which case the
electrolysis is unsatisfactory. With a little care this need not happen.
Gold cylinders may preferably be used instead of copper; but on platinum
the deposit of mercury is grey and non-adherent, so that it cannot be
washed and weighed.
VOLUMETRIC METHODS.
Several methods have been devised: for the details of these the student
is referred to Sutton's "Handbook of Volumetric Analysis."
QUESTIONS.
1. The specific gravity of mercury is 13.596. What volume would 8 grams
occupy?
2. If 3.169 grams of cinnabar gave 2.718 grams of mercury, what would be
the percentage of the metal in the ore?
3. Pour solution of mercuric chloride on mercury and explain what
happens.
4. On dissolving 0.3 gram of mercury in hot nitric acid, and passing
sulphuretted hydrogen in excess through the diluted solution, what
weight of precipitate will be got?
FOOTNOTES:
[9] Lead may be granulated by heating it to a little above the melting
point, pouring it into a closed wooden box, and rapidly agitating it as
it solidifies.
[10] A rod of iron placed in the crucible with the assays will decompose
any regulus that may be formed.
[11] With buttons poor in silver the lowering of the temperature at this
stage is not a matter of importance.
[12] 100 grams of the lead, or of its oxide, will contain from 1.5 to
2.5 milligrams.
[13] Still the precautions of having cupels well made from bone ash in
fine powder, and of working the cupellation at as low a temperature as
possible are very proper ones, provided they are not carried to an
absurd excess.
[14] Be careful to remove the crucible before taking the bottle out of
the basin of water; if this is not done the chloride may be washed out
of it.
[15] 1 c.c. of this dilute acid will precipitate 8 or 9 milligrams of
silver.
[16] Chlorides interfere not merely by removing silver as insoluble
silver chloride, but also by making it difficult to get a good finishing
point, owing to the silver chloride removing the colour from the
reddened solution.
[17] These results were obtained when using ammonium sulphocyanate, and
cannot be explained by the presence of such impurities as chlorides, &c.
[18] Multiply the _standard_ by 1000, and dilute 100 c.c. of the
standard solution to the resulting number of c.c. Thus, with a solution
of a standard .495, dilute 100 c.c. to 495 c.c., using, of course,
distilled water.
[19] HNa_{2}AsO_{4} + 3AgNO_{3} = Ag_{3}AsO_{4} + HNO_{3} + 2NaNO_{3}.
[20] SiO_{2} + Na_{2}CO_{3} = CO_{2} + Na_{2}SiO_{3}
SiO_{2} + 2NaHCO_{3} = 2CO_{2} + Na_{2}SiO_{3} + H_{2}O.
[21] PbO + SiO_{2} = PbSiO_{3}
[22] Here and elsewhere in this article when a flux is spoken of as soda
the bicarbonate is meant.
[23] See the description of the process commencing on p. 98 and the
explanatory remarks on p. 110.
[24] Percy, _Metallurgy of Silver and Gold_, p. 258.
[25] "Limits of Accuracy attained in Gold-bullion Assay," _Trans. Chem.
Soc._, 1893.
[26] "Assaying and Hall-marking at the Chester Assay Office." W.F. Lowe.
_Journ. Soc. Chem. Industry_, Sept. 1889.
[27] Fine or pure gold is 24 carat. Nine carat gold therefore contains 9
parts of gold in 24 of the alloy; eighteen carat gold contains 18 parts
of gold in 24; and so on.
[28] The mouth of the flask must not have a rim around it.
[29] See "Assaying and Hall-marking at the Chester Assay Office," by
W.F. Lowe. _Journ. Soc. Chem. Industry_, Sept. 1889.
[30] Percy, _Metallurgy of Silver and Gold_, p. 263.
[31] See also "The Assaying of Gold Bullion," by C. Whitehead and T.
Ulke. _Eng. and Mining Journal_, New York, Feb. 12, 1898.
[32] Consult Percy's _Metallurgy of Silver and Gold_, p. 172; A.C.
Claudet, _Trans. Inst. Mining and Metallurgy_, vol. vi. p. 29; G.M.
Roberts _Trans. Amer. Inst. Mining Engineers_, Buffalo Meeting, 1898; J.
and H.S. Pattinson, _Journ. Soc. Chem. Industry_, vol. xi. p. 321.
[33] Heycock and Neville, _Journ. Chem. Soc._, 1892, p. 907.
[34] G.M. Roberts.
[35] A.C. Claudet.
[36] "The Sampling of Argentiferous and Auriferous Copper," by A.R.
Ledoux. _Journ. Canadian Mining Institute_, 1899.
[37] NaCNO + BaCl_{2} + NaHO + H_{2}O = NH_{3} + BaCO_{3} + 2 NaCl.
[38] HCy + NaHO = NaCy + H_{2}O.
[39] 2KCN + AgNO_{3} = KAg(CN)_{2} + KNO_{3}.
[40] If it be desired to make a solution so that 100 c.c. shall be
equivalent to 1 gram of sodium cyanide, then 18.085 grams of silver
nitrate should be taken for each litre.
[41] AgNO_{3} + KAgCy_{2} = 2 AgCy + KNO_{3}.
[42] AgNO_{3} + KI = AgI + KNO_{3}.
[43] See pp. 322, 323, and 324 for a description of the methods for
measuring the quantity of acid or alkali.
[44] KCN + HCl = KCl + HCN
[45] Taking 16.0 grams of ore, each c.c. = 1 lb. of soda to the short
ton. The corresponding figures for the long ton are 12.544 grams for
lime and 17.92 grams for soda.
[46] In which case each .01 gram of metal found equals 1 lb to the short
ton of solution.
[47] 100 c.c. of water dissolves 0.66 gram of the salt; it is almost
insoluble in alcohol or in solutions of ammonic chloride.
[48] According to Personne mercuric chloride is not volatilised from
boiling solutions when alkaline chlorides are present.
[49] The solution should contain about 0.25 gram of mercury, and a large
excess of nitric acid must be avoided.
CHAPTER X.
COPPER--LEAD--THALLIUM--BISMUTH--ANTIMONY.
COPPER.
Copper occurs native in large quantities, especially in the Lake
Superior district; in this state it is generally pure. More frequently
it is found in combination. The ores of copper may be classed as oxides
and sulphides. The most abundant oxidised ores are the carbonates,
malachite and chessylite; the silicates, as also the red and black
oxides, occur less abundantly. All these yield their copper in solution
on boiling with hydrochloric acid.
The sulphides are more abundant. Copper pyrites (or yellow ore),
erubescite (or purple ore), and chalcocite (or grey ore) are the most
important. Iron pyrites generally carries copper and is frequently
associated with the above-mentioned minerals. These are all attacked by
nitric acid. They nearly all contain a small quantity of organic matter,
and frequently considerable quantities of lead, zinc, silver, gold,
arsenic, bismuth, &c.
The copper ores are often concentrated on the mine before being sent
into the market, either by smelting, when the product is a regulus or
matte, or by a wet method of extraction, yielding cement copper or
precipitate. A regulus is a sulphide of copper and iron, carrying from
30 to 40 per cent. of copper. A precipitate, which is generally in the
form of powder, consists mainly of metallic copper. Either regulus or
precipitate may be readily dissolved in nitric acid.
Copper forms two classes of salts, cuprous and cupric. The former are
pale coloured and of little importance to the assayer. They are easily
and completely converted into cupric by oxidising agents. Cupric
compounds are generally green or blue, and are soluble in ammonia,
forming deep blue solutions.
DRY ASSAY.
That, for copper, next after those for gold and silver, holds a more
important position than any other dry assay. The sale of copper ores has
been regulated almost solely in the past by assays made on the Cornish
method. It is not pretended that this method gives the actual content of
copper, but it gives the purchaser an idea of the quantity and quality
of the metal that can be got by smelting. The process is itself one of
smelting on a small scale. As might be expected, however, the assay
produce and the smelting produce are not the same, there being a smaller
loss of copper in the smelting. The method has worked very well, but
when applied to the purchase of low class ores (from which the whole of
the copper is extracted by wet methods) it is unsatisfactory. The
following table, which embodies the results of several years' experience
with copper assays, shows the loss of copper on ores of varying produce.
The figures in the fourth column show how rapidly the proportion of
copper lost increases as the percentage of copper in the ore falls below
30 per cent. For material with more than 30 per cent. the proportion
lost is in inverse proportion to the copper present.
LOSS OF COPPER.
-----------------+------------+-----------+------------------
Copper present. | Dry Assay. | Margin. | Loss on 100
| | | Parts of Copper.
-----------------+------------+-----------+------------------
Per cent. | Per cent. | Per cent. |
100 | 98 | 2.0 | 2.0
95 | 92-1/2 | 2.5 | 2.6
90 | 87-3/8 | 2.6 | 2.9
85 | 82-3/8 | 2.6 | 3.0
80 | 77-3/8 | 2.6 | 3.2
75 | 72-3/8 | 2.6 | 3.5
70 | 67-1/2 | 2.5 | 3.6
65 | 62-1/2 | 2.5 | 3.8
60 | 57-5/8 | 2.4 | 4.0
55 | 52-3/4 | 2.3 | 4.2
50 | 47-3/4 | 2.2 | 4.4
45 | 43 | 2.0 | 4.5
40 | 38-1/8 | 1.8 | 4.6
35 | 33-1/4 | 1.7 | 4.8
30 | 28-1/2 | 1.50 | 5.0
25 | 23-1/2 | 1.50 | 6.0
20 | 18-1/2 | 1.56 | 7.8
18 | 16-1/2 | 1.53 | 8.5
16 | 14-1/2 | 1.48 | 9.3
14 | 12-5/8 | 1.40 | 10.0
12 | 10-5/8 | 1.37 | 11.4
10 | 8-3/4 | 1.28 | 12.8
8 | 6-7/8 | 1.14 | 14.3
6 | 5 | 1.05 | 17.5
5 | 4 | 1.00 | 20.0
4 | 3 | 1.00 | 25.0
3.75 | 2-3/4 | 0.97 | 26.0
3.50 | 2-9/16 | 0.94 | 27.0
3.25 | 2-5/16 | 0.91 | 28.0
3.00 | 2-1/8 | 0.87 | 29.0
2.75 | 1-15/16 | 0.82 | 30.0
2.50 | 1-3/4 | 0.77 | 31.0
2.25 | 1-1/2 | 0.72 | 32.0
2.00 | 1-5/16 | 0.66 | 33.0
-----------------+------------+-----------+------------------
The wet assay being known, the dry assay can be calculated with the help
of the above table by deducting the amount in the column headed "margin"
opposite the corresponding percentage. For example, if the wet assay
gives a produce of 17.12 per cent., there should be deducted 1.5; the
dry assay would then be 15.62, or, since the fractions are always
expressed in eighths, 15-5/8. With impure ores, containing from 25 to 50
per cent. of copper, the differences may be perhaps 1/4 greater.
Wet methods are gradually replacing the dry assay, and it is probable
that in the future they will supersede it; for stock-taking, and the
various determinations required in smelting works and on mines, they are
generally adopted, because they give the actual copper contents, and
since it is obvious that a knowledge of this is more valuable to the
miner and smelter. Moreover, the working of the dry method has been
monopolised by a small ring of assayers, with the double result of
exciting outside jealousy and, worse still, of retarding the development
and improvement of the process.
The principal stages of the dry assay are: (1) the concentration of the
copper in a regulus; (2) the separation of the sulphur by calcining; (3)
the reduction of the copper by fusion; and (4) the refining of the metal
obtained.
The whole of these operations are not necessary with all copper
material. Ores are worked through all the stages; with mattes, the
preliminary fusion for regulus is omitted; precipitates are simply fused
for coarse copper, and refined; and blister or bar coppers are refined,
or, if very pure, subjected merely to washing.
The quantity of ore generally taken is 400 grains, and is known as "a
full trial"; but for rich material, containing more than 50 per cent. of
copper, "a half trial," or 200 grains, is used.
~Fusion for Regulus.~--The ore (either with or without a previous
imperfect roasting to get rid of any excess of sulphur) is mixed with
borax, glass, lime, and fluor spar; and, in some cases, with nitre, or
iron pyrites, according to the quality of the ore. The mixture is placed
in a large Cornish crucible, and heated as uniformly as possible in the
wind furnace, gradually raising the temperature so as to melt down the
charge in from 15 to 20 minutes. The crucible is removed and its
contents poured into an iron mould. When the slag is solid, it is taken
up with tweezers and quenched in water. The regulus is easily detached
from the slag. It should be convex above and easily broken, have a
reddish brown colour, and contain from 40 to 60 per cent. of copper. A
regulus with more than this is "too fine," and with less "too coarse." A
regulus which is too fine is round, compact, hard, and of a dark bluish
grey on the freshly broken surface. A coarse regulus is flat and coarse
grained, and more nearly resembles sulphide of iron in fracture and
colour.
If an assay yields a regulus "too coarse," a fresh determination is made
with more nitre added, or the roasting is carried further. With low
class ores a somewhat coarse regulus is an advantage. If, on the other
hand, the regulus is too fine, less nitre or less roasting is the
remedy. With grey copper ores and the oxidised ores, iron pyrites is
added.
~Calcining the Regulus.~--It is powdered in an iron mortar and
transferred to a small Cornish crucible, or (if the roasting is to be
done in the muffle) to a roasting dish or scorifier. The calcining is
carried out at a dull red heat, which is gradually increased. The charge
requires constant stirring at first to prevent clotting, but towards the
end it becomes sandy and requires less attention. If the temperature
during calcination has been too low sulphates are formed, which are
again reduced to sulphides in the subsequent fusion. To prevent this the
roasted regulus is recalcined at a higher temperature, after being
rubbed up with a little anthracite. The roasted substance must not smell
of burning sulphur when hot. It is practically a mixture of the oxides
of copper and iron.
~Fusion for Coarse Copper.~--The calcined regulus is mixed with a flux
consisting of borax and carbonate of soda, with more or less tartar
according to its weight. Some "assayers" use both tartar and nitre, the
former of course being in excess. The charge is returned to the crucible
in which it was calcined, and is melted down at a high temperature, and,
as soon as tranquil, poured. When solid it is quenched and the button of
metal separated.
The slag is black and glassy. The small quantity of copper which it
retains is recovered by a subsequent "cleaning," together with the slags
from the next operation.
The button of "coarse copper" obtained must be free from a coating of
regulus. It will vary somewhat in appearance according to the nature and
quantity of the impurities.
~Refining the Coarse Copper.~--The same crucible is put back in the
furnace, deep down and under the crevice between the two bricks. When it
has attained the temperature of the furnace the coarse copper is dropped
into it and the furnace closed. The copper will melt almost at once with
a dull surface, which after a time clears, showing an "eye." Some
refining flux is then shot in from the scoop (fig. 48), and, when the
assay is again fluid, it is poured. When cold the button of metal is
separated.
[Illustration: FIG. 48.]
The button of "fine" copper is flat or pitted on its upper surface, and
is coated with a thin orange film; it must have the appearance of good
copper. If it is covered with a red or purple film, it is overdone or
"burnt." If, on the other hand, it has a rough, dull appearance, it is
not sufficiently refined. Assays that have been "burnt" are rejected.
Those not sufficiently fine are treated as "coarse copper," and again
put through the refining operation.
~Cleaning the Slags.~--These are roughly powdered and re-fused with
tartar, etc., as in the fusion for coarse copper. The button of metal
got is separated (if big enough refined) and weighed.
The details of the process are slightly varied by different assayers:
the following will be good practice for the student.
~Determination of Copper in Copper Pyrites.~--Powder, dry, and weigh up
20 grams of the ore. Mix with 20 grams each of powdered lime and fluor,
15 grams each of powdered glass and borax, and 5 or 10 grams of nitre.
Transfer to a large Cornish crucible and fuse under a loose cover at a
high temperature for from 15 to 20 minutes. When fluid and tranquil pour
into a mould. When the slag has solidified, but whilst still hot, quench
by dipping two or three times in cold water. Avoid leaving it in the
water so long that it does not dry after removal. When cold separate the
button, or perhaps buttons, of regulus by crumbling the slag between the
fingers. See that the slag is free from regulus. It should be light
coloured when cold and very fluid when hot. Reject the slag.
Powder the regulus in a mortar and transfer to a small crucible.
Calcine, with occasional stirring, until no odour of sulphurous oxide
can be detected. Shake back into the mortar, rub up with about 1 gram of
powdered anthracite, and re-calcine for 10 minutes longer.
Mix the calcined regulus with 10 grams of tartar, 20 grams of soda, and
3 grams of borax; and replace in the crucible used for calcining. Fuse
at a bright red heat for 10 or 15 minutes. Pour, when tranquil.
As soon as solid, quench in water, separate the button of copper, and
save the slag.
To refine the copper a very hot fire is wanted, and the fuel should not
be too low down in the furnace. Place the crucible well down in the fire
and in the middle of the furnace. The same crucible is used, or, if a
new one is taken, it must be glazed with a little borax. When the
crucible is at a good red heat, above the fusing point of copper, drop
the button of copper into it, and close the furnace. Watch through the
crevice, and, as soon as the button has melted and appears clear showing
an eye, shoot in 10 grams of refining flux, close the furnace, and, in a
few minutes, pour; then separate the button of copper. Add the slag to
that from the coarse copper fusion, and powder. Mix with 5 grams of
tartar, 0.5 gram of powdered charcoal, and 2 grams of soda. Fuse in the
same crucible, and, when tranquil, pour; quench, and pick out the prills
of metal.
If the copper thus got from the slags is coarse looking and large in
amount, it must be refined; but, if small in quantity, it may be taken
as four-fifths copper. The combined results multiplied by five give the
percentage of copper.
The refining flux is made by mixing 3 parts (by measure) of powdered
nitre, 2-1\2 of tartar, and 1 of salt. Put in a large crucible, and stir
with a red-hot iron until action has ceased. This operation should be
carried out in a well-ventilated spot.
For pure ores in which the copper is present, either as metal or oxide,
and free from sulphur, arsenic, &c., the concentration of the copper in
a regulus may be omitted, and the metal obtained in a pure state by a
single fusion.[50] It is necessary to get a fluid neutral slag with the
addition of as small an amount of flux as possible. The fusion should be
made at a high temperature, so as not to occupy more than from 20 to 25
minutes. Thirty grams of ore is taken for a charge, mixed with 20 grams
of cream of tartar, and 10 grams each of dried borax and soda. If the
gangue of the ore is basic, carrying much oxide of iron or lime, silica
is added, in quantity not exceeding 10 grams. If, on the other hand, the
gangue is mainly quartz, oxide of iron up to 7 grams must be added.
_Example._--Twenty grams of copper pyrites, known to contain 27.6 per
cent. of copper, gave by the method first described 5.22 grams of
copper, equalling 26-1/8 per cent. Another sample of 20 grams of the
same ore, calcined, fused with 40 grams of nitre, and washed to ensure
the removal of arsenic and sulphur, and treated according to the second
method, gave a button weighing 5.27 grams, equalling 26-3/8 per cent.
The ore contained a considerable quantity of lead. Lead renders the
assay more difficult, since after calcination it remains as lead
sulphate, and in the fusion for coarse copper reappears as a regulus on
the button.
~The Estimation of Moisture.~--The Cornish dry assayer very seldom makes
a moisture determination. He dries the samples by placing the papers
containing them on the iron plate of the furnace.
It is well known that by buying the copper contents of pyrites by
Cornish assay, burning off the sulphur, and converting the copper into
precipitate, a large excess is obtained.
NOTES ON THE VALUATION OF COPPER ORES.
Closely bound up with the practice of dry copper assaying is that of
valuing a parcel of copper ore. The methods by which the valuation is
made have been described by Mr. Westmoreland,[51] and are briefly as
follows:--The produce of the parcel is settled by two assayers, one
acting for the buyer, the other for the seller; with the help, in case
of non-agreement, of a third, or referee, whose decision is final. The
dry assayers who do this are in most cases helped, and sometimes,
perhaps, controlled, by wet assays made for one or both of the parties
in the transaction.
In the case of "ticketing," the parcels are purchased by the smelters by
tender, and the value of any particular parcel is calculated from the
average price paid, as follows:--The "standard," or absolute value of
each ton of fine copper in the ore, is the price the smelters have paid
for it, plus the returning charges or cost of smelting the quantity of
ore in which it is contained. The value of any particular parcel of ore
is that of the quantity of fine copper it contains, calculated on this
standard, minus the returning charges. The ton consists of 21 cwts.,
and it is assumed that the "settled" produce is the actual yield of the
ore.
If at a ticketing in Cornwall 985 tons of ore containing 63.3 tons of
fine copper (by dry assay) brought £2591 12s., the standard would be
£83 15s. This is calculated as follows:--The returning charge is fixed
at 55s. per ton of ore. This on 985 tons will amount to £2708 15s.
Add this to the actual price paid, and there is got £5300 as the value
of the fine copper present. The weight of copper in these 985 tons being
63.3 tons, the standard is £5300/63.3, or £83 15s. (nearly).
The value of a parcel of 150 tons of a 6 per cent. ore on the same
standard would be arrived at as follows:--The 150 tons at 6 per cent.
would contain 9 tons (150×6/100) of fine copper. This, at £83 15s.
per ton, would give £753 15s. From this must be deducted the returning
charges on 150 tons of ore at 55s. per ton, or £412 10s. This leaves
£341 5s. as the value of the parcel.
At Swansea the returning charge is less than in Cornwall, and varies
with the quality of the ore. This appears equitable, since in smelting
there are some costs which are dependent simply on the number of tons
treated, and others which increase with the richness. The returning
charge then is made up of two parts, one fixed at so much (12s. 2d.)
per ton of ore treated, and the other so much (3s. 9d.) per unit of
metal in the ore. In this way the returning charge on a ton of ore of
8-3/4 produce would be (12s. 2d.)+(8-3/4×(3s. 9d.)), or £2 5s.
If, for example, Chili bars, containing 96 per cent. of copper, bring
£50 per ton, the standard is £71 9s. 4d. It is got at in this way.
The returning charge on a 96 per cent. ore is (12s. 2d.)+(96×(3s.
9d.)), or £18 12s. 2d. This added to £50 gives £68 12s. 2d.,
and this multiplied by 100 and divided by 96 (100 tons of the bars will
contain 96 tons of fine copper) will give £71 9s. 4d.
The price of 100 tons of pyrites, containing 2-1/4 per cent. of copper
by dry assay, would be got on this standard as follows:--The parcel of
ore would contain 2-1/4 tons of copper. This multiplied by the standard
gives £160 16s. 0d. From this must be deducted the returning charge,
which for 1 ton of ore of this produce would be (12s. 2d.) + (2-1/4×(3s.
9d.)) or £1 0s. 7d., and on the 100 tons is £102 18s. 4d. This would
leave £57 17s. 10d. as the price of the parcel, or 11s. 7d. per ton.
This would be on the standard returning charge of 45s. (for 8-3/4 per
cent. ore); if a smaller returning charge was agreed on, say 38s., the
difference in this case, 7s., would be added to the price per ton.
WET METHODS.
The solubility of the ores of copper in acid has already been described,
but certain furnace products, such as slags, are best opened up by
fusion with fusion mixture and a little nitre.
The method of dissolving varies with the nature of the ore. With 5 grams
of pyrites, a single evaporation with 20 c.c. of nitric acid will give a
residue completely soluble in 30 c.c. of hydrochloric acid. If the ore
carries oxide of iron or similar bodies, these are first dissolved up by
boiling with 20 c.c. of hydrochloric acid, and the residue attacked by
an addition of 5 c.c. of nitric. When silicates decomposable by acid are
present, the solution is evaporated to dryness to render the silica
insoluble; the residue extracted with 30 c.c. of hydrochloric acid, and
diluted with water to 150 c.c. It is advisable to have the copper in
solution as chloride. To separate the copper, heat the solution nearly
to boiling (best in a pint flask), and pass a rapid current of
sulphuretted hydrogen for four or five minutes until the precipitate
settles readily and the liquid smells of the gas. When iron is present
it will be reduced to the ferrous state before the copper sulphide
begins to separate. The copper appears as a brown coloration or black
precipitate according to the quantity present. Filter through a coarse
filter, wash with hot water containing sulphuretted hydrogen, if
necessary. Wash the precipitate back into the flask, boil with 10 c.c.
of nitric acid, add soda till alkaline, and pass sulphuretted hydrogen
again. Warm and filter, wash and redissolve in nitric acid, neutralise
with ammonia, add ammonic carbonate, boil and filter. The copper freed
from impurities will be in the solution. Acidulate and reprecipitate
with sulphuretted hydrogen. When the nature of the impurities will allow
it, this process may be shortened to first filtering off the gangue,
then precipitating with sulphuretted hydrogen and washing the
precipitate on the filter first with water and then with ammonium
sulphide.