Chapter XIII: Part 13

Undoubtedly the greatest of the gold discoveries made in the latter half of the 19th century was that of the Witwatersrand district in the Transvaal. By reason of its unusual geological character and great economic importance this district deserves a more extended description. The gold occurs in conglomerate beds, locally known as "banket." There are several series of parallel beds, interstratified with quartzite and schist, the most important being the "main reef" series. The gold in this conglomerate reef is partly of detrital origin and partly of the genetic character of ordinary vein-gold. The formation is noted for its regularity as regards both the thickness and the gold-tenor of the ore-bearing reefs, in which respect it is unparalleled in the geology of the auriferous formations. The gold carries, on an average, L2 per ton, and is worked by ordinary methods of gold-mining, stamp-milling and cyaniding. In 1899, 5762 stamps were in operation, crushing 7,331,446 tons of ore, and yielding L15,134,000, equivalent to 25.5% of the world's production. Of this, 80% came from within 12 m. of Johannesburg. After September 1899 operations were suspended, almost entirely owing to the Boer War, but on the 2nd of May 1901 they were started again. In 1905 the yield was valued at L20,802,074, and in 1909 at L30,925,788. So certain is the ore-bearing formation that engineers in estimating its auriferous contents feel justified in assuming, as a factor in their calculations, a vertical extension limited only by the lowest depths at which mining is feasible. On such a basis they arrived at more than L600,000,000 as the available gold contained in the Witwatersrand conglomerates. This was a conservative estimate, and was made before the full extent of the reefs was known; in 1904 Lionel Phillips stated that the main reef series had been proved for 61 m., and he estimated the gold remaining to be mined to be worth L2,500,000,000. Deposits similar to the Witwatersrand banket occur in Zululand, and also on the Gold Coast of Africa. In Rhodesia, the country lying north of the Transvaal, where gold occurs in well-defined quartz-veins, there is unquestionable evidence of extensive ancient workings. The economic importance of the region generally has been fully proved. Rhodesia produced L386,148 in 1900 and L722,656 in 1901, in spite of the South African War; the product for 1905 was valued at L1,480,449, and for 1908 at L2,526,000.

The gold production of Russia has been remarkably constant, averaging L4,899,262 per annum; the gold is derived chiefly from placer workings in Siberia.

The gold production of China was estimated for 1899 at L1,328,238 and for 1900 at L860,000; it increased in 1901 to about L1,700,000, to fall to L340,000 in 1905; in 1906 and 1907 it recovered to about L1,000,000.

TABLE II.--_Gold Production of Certain Countries, 1881-1908 (in oz.)._

+------+-----------+-----------+----------+---------+----------+-----------+-----------+------------+
| | Austral- | | | | | | United | |
| Year.| asia. | Africa. | Canada. | India. | Mexico. | Russia. | States. | Totals. |
+------+-----------+-----------+----------+---------+----------+-----------+-----------+------------+
| 1881 | 1,475,161 | .. | 52,483 | .. | 41,545 | 1,181,853 | 1,678,612 | 4,976,980 |
| 1882 | 1,438,067 | .. | 52,000 | .. | 45,289 | 1,154,613 | 1,572,187 | 4,825,794 |
| 1883 | 1,333,849 | .. | 46,150 | .. | 46,229 | 1,132,219 | 1,451,250 | 4,614,588 |
| 1884 | 1,352,761 | .. | 46,000 | .. | 57,227 | 1,055,642 | 1,489,950 | 4,902,889 |
| 1885 | 1,309,804 | .. | 53,987 | .. | 46,941 | 1,225,738 | 1,538,325 | 5,002,584 |
| 1886 | 1,257,670 | .. | 66,061 | .. | 29,702 | 922,226 | 1,693,125 | 5,044,363 |
| 1887 | 1,290,202 | 28,754 | 59,884 | 15,403 | 39,861 | 971,656 | 1,596,375 | 5,061,490 |
| 1888 | 1,344,002 | 240,266 | 53,150 | 35,034 | 47,117 | 1,030,151 | 1,604,841 | 5,175,623 |
| 1889 | 1,540,607 | 366,023 | 62,658 | 78,649 | 33,862 | 1,154,076 | 1,587,000 | 5,611,245 |
| 1890 | 1,453,172 | 497,817 | 55,625 | 107,273 | 37,104 | 1,134,590 | 1,588,880 | 5,726,966 |
| 1891 | 1,518,690 | 729,268 | 45,022 | 131,776 | 48,375 | 1,168,764 | 1,604,840 | 6,287,591 |
| 1892 | 1,638,238 | 1,210,869 | 43,905 | 164,141 | 54,625 | 1,199,809 | 1,597,098 | 7,102,172 |
| 1893 | 1,711,892 | 1,478,477 | 44,853 | 207,152 | 63,144 | 1,345,224 | 1,739,323 | 7,772,585 |
| 1894 | 2,020,180 | 2,024,164 | 50,411 | 210,412 | 217,688 | 1,167,455 | 1,910,813 | 8,813,848 |
| 1895 | 2,170,505 | 2,277,640 | 92,440 | 257,830 | 290,250 | 1,397,767 | 2,254,760 | 9,814,505 |
| 1896 | 2,185,872 | 2,280,892 | 136,274 | 323,501 | 314,437 | 1,041,794 | 2,568,132 | 9,950,861 |
| 1897 | 2,547,704 | 2,832,776 | 294,582 | 350,585 | 362,812 | 1,124,511 | 2,774,935 | 11,420,068 |
| 1898 | 3,137,644 | 3,876,216 | 669,445 | 376,431 | 411,187 | 1,231,791 | 3,118,398 | 13,877,806 |
| 1899 | 3,837,181 | 3,532,488 |1,031,563 | 418,869 | 411,187 | 1,072,333 | 3,437,210 | 14,837,775 |
| 1900 | 3,555,506 | 419,503 |1,348,720 | 456,444 | 435,375 | 974,537 | 3,829,897 | 12,315,135 |
| 1901 | 3,719,080 | 439,704 |1,167,216 | 454,527 | 497,527 | 1,105,412 | 3,805,500 | 12,698,089 |
| 1902 | 3,946,374 | 1,887,773 |1,003,355 | 463,824 | 491,156 | 1,090,053 | 3,870,000 | 14,313,660 |
| 1903 | 4,315,538 | 3,289,409 | 911,118 | 552,873 | 516,524 | 1,191,582 | 3,560,000 | 15,852,620 |
| 1904 | 4,245,744 | 4,156,084 | 793,350 | 556,097 | 609,781 | 1,199,857 | 3,892,480 | 16,790,351 |
| 1905 | 4,159,220 | 5,477,841 | 700,863 | 576,889 | 779,181 | 1,063,883 | 4,265,742 | 18,360,945 |
| 1906 | 3,984,538 | 6,449,749 | 581,709 | 525,527 | 896,615 | 1,087,056 | 4,565,333 | 19,620,272 |
| 1907 | 3,659,693 | 7,270,464 | 399,844 | 495,965 | 903,672 | 1,282,635 | 4,374,827 | 19,988,144 |
| 1908 | 3,557,705 | 7,983,348 | 462,467 | 504,309 |1,182,445 | 1,497,076 | 4,659,360 | 21,529,300 |
+------+-----------+-----------+----------+---------+----------+-----------+-----------+------------+

_Alloys._--Gold forms alloys with most metals, and of these many are
of great importance in the arts. The alloy with mercury--gold
amalgam--is so readily formed that mercury is one of the most powerful
agents for extracting the precious metal. With 10% of gold present the
amalgam is fluid, and with 12.5% pasty, while with 13% it consists of
yellowish-white crystals. Gold readily alloys with silver and copper
to form substances in use from remote times for money, jewelry and
plate. Other metals which find application in the metallurgy of gold
by virtue of their property of extracting the gold as an alloy are
lead, which combines very readily when molten, and which can
afterwards be separated by cupellation, and copper, which is separated
from the gold by solution in acids or by electrolysis; molten lead
also extracts gold from the copper-gold alloys. The relative amount of
gold in an alloy is expressed in two ways: (1) as "fineness," i.e. the
amount of gold in 1000 parts of alloy; (2) as "carats," i.e. the
amount of gold in 24 parts of alloy. Thus, pure gold is 1000 "fine" or
24 carat. In England the following standards are used for plate and
jewelry: 375, 500, 625, 750 and 916.6, corresponding to 9, 12, 15, 18
and 22 carats, the alloying metals being silver and copper in varying
proportions. In France three alloys of the following standards are
used for jewelry, 920, 840 and 750. A greenish alloy used by
goldsmiths contains 70% of silver and 30% of gold. "Blue gold" is
stated to contain 75% of gold and 25% of iron. The Japanese use for
ornament an alloy of gold and silver, the standard of which varies
from 350 to 500, the colour of the precious metal being developed by
"pickling" in a mixture of plum-juice, vinegar and copper sulphate.
They may be said to possess a series of bronzes, in which gold and
silver replace tin and zinc, all these alloys being characterized by
patina having a wonderful range of tint. The common alloy,
Shi-ya-ku-Do, contains 70% of copper and 30% of gold; when exposed to
air it becomes coated with a fine black patina, and is much used in
Japan for sword ornaments. Gold wire may be drawn of any quality, but
it is usual to add 5 to 9 dwts. of copper to the pound. The "solders"
used for red gold contain 1 part of copper and 5 of gold; for light
gold, 1 part of copper, 1 of silver and 4 of gold.

_Gold and Silver._--Electrum is a natural alloy of gold and silver.
Matthiessen observed that the density of alloys, the composition of
which varies from AuAg6 to Au6Ag, is greater than that calculated from
the densities of the constituent metals. These alloys are harder, more
fusible and more sonorous than pure gold. The alloys of the formulae
AuAg, AuAg2, AuAg4 and AuAg20 are perfectly homogeneous, and have been
studied by Levol. Molten alloys containing more than 80% of silver
deposit on cooling the alloy AuAg9, little gold remaining in the
mother liquor.

_Gold and Zinc._--When present in small quantities zinc renders gold
brittle, but it may be added to gold in larger quantities without
destroying the ductility of the precious metal; Peligot proved that a
triple alloy of gold, copper and zinc, which contains 5.8% of the
last-named, is perfectly ductile. The alloy of 11 parts gold and 1
part of zinc is, however, stated to be brittle.

_Gold and Tin._--Alchorne showed that gold alloyed with 1/37th part of
tin is sufficiently ductile to be rolled and stamped into coin,
provided the metal is not annealed at a high temperature. The alloys
of tin and gold are hard and brittle, and the combination of the
metals is attended with contraction; thus the alloy SnAu has a density
14.243, instead of 14.828 indicated by calculation. Matthiessen and
Bose obtained large crystals of the alloy Au2Sn5, having the colour of
tin, which changed to a bronze tint by oxidation.

_Gold and Iron._--Hatchett found that the alloy of 11 parts gold and 1
part of iron is easily rolled without annealing. In these proportions
the density of the alloy is less than the mean of its constituent
metals.

_Gold and Palladium._--These metals are stated to alloy in all
proportions. According to Chenevix, the alloy composed of equal parts
of the two metals is grey, is less ductile than its constituent metals
and has the specific gravity 11.08. The alloy of 4 parts of gold and 1
part of palladium is white, hard and ductile. Graham showed that a
wire of palladium alloyed with from 24 to 25 parts of gold does not
exhibit the remarkable retraction which, in pure palladium, attends
its loss of occluded hydrogen.

_Gold and Platinum._--Clarke states that the alloy of equal parts of
the two metals is ductile, and has almost the colour of gold.

_Gold and Rhodium._--Gold alloyed with 1/4th or 1/5th of rhodium is,
according to Wollaston, very ductile, infusible and of the colour of
gold.

_Gold and Iridium._--Small quantities of iridium do not destroy the
ductility of gold, but this is probably because the metal is only
disseminated through the mass, and not alloyed, as it falls to the
bottom of the crucible in which the gold is fused.

_Gold and Nickel._--Eleven parts of gold and 1 of nickel yield an
alloy resembling brass.

_Gold and Cobalt._--Eleven parts of gold and 1 of cobalt form a
brittle alloy of a dull yellow colour.

_Compounds._--Aurous oxide, Au2O, is obtained by cautiously adding
potash to a solution of aurous bromide, or by boiling mixed solutions
of auric chloride and mercurous nitrate. It forms a dark-violet
precipitate which dries to a greyish-violet powder. When freshly
prepared it dissolves in cold water to form an indigo-coloured
solution with a brownish fluorescence of colloidal aurous oxide; it is
insoluble in hot water. This oxide is slightly basic. Auric oxide,
Au2O3, is a brown powder, decomposed into its elements when heated to
about 250 deg. or on exposure to light. When a concentrated solution
of auric chloride is treated with caustic potash, a brown precipitate
of auric hydrate, Au(OH)3, is obtained, which, on heating, loses water
to form auryl hydrate, AuO(OH), and auric oxide, Au2O3. It functions
chiefly as an acidic oxide, being less basic than aluminium oxide, and
forming no stable oxy-salts. It dissolves in alkalis to form
well-defined crystalline salts; potassium aurate, KAuO2.3H2O, is very
soluble in water, and is used in electro-gilding. With concentrated
ammonia auric oxide forms a black, highly explosive compound of the
composition AuN2H3.3H2O, named "fulminating gold"; this substance is
generally considered to be Au(NH2)NH.3H2O, but it may be an ammine of
the formula [Au(NH3)2(OH)2]OH. Other oxides, e.g. Au2O2, have been
described.

Aurous chloride, AuCl, is obtained as a lemon-yellow, amorphous
powder, insoluble in water, by heating auric chloride to 185 deg. It
begins to decompose into gold and chlorine at 185 deg., the
decomposition being complete at 230 deg.; water decomposes it into
gold and auric chloride. Auric chloride, or gold trichloride, AuCl3,
is a dark ruby-red or reddish-brown, crystalline, deliquescent powder
obtained by dissolving the metal in aqua regia. It is also obtained by
carefully evaporating a solution of the metal in chlorine water. The
gold chloride of commerce, which is used in photography, is really a
hydrochloride, chlorauric or aurichloric acid, HAuCl4.3H2O, and is
obtained in long yellow needles by crystallizing the acid solution.
Corresponding to this acid, a series of salts, named chloraurates or
aurichlorides, are known. The potassium salt is obtained by
crystallizing equivalent quantities of potassium and auric chlorides.
Light-yellow monoclinic needles of 2KAuCl4.H2O are deposited from
warm, strongly acid solutions, and transparent rhombic tables of
KAuCl4.2H2O from neutral solutions. By crystallizing an aqueous
solution, red crystals of AuCl3.2H2O are obtained. Auric chloride
combines with the hydrochlorides of many organic bases--amines,
alkaloids, &c.--to form characteristic compounds. Gold dichloride,
probably Au2Cl4, = Au.AuCl4, aurous chloraurate, is said to be
obtained as a dark-red mass by heating finely divided gold to 140
deg.-170 deg. in chlorine. Water decomposes it into gold and auric
chloride. The bromides and iodides resemble the chlorides. Aurous
bromide, AuBr, is a yellowish-green powder obtained by heating the
tribromide to 140 deg.; auric bromide, AuBr3, forms reddish-black or
scarlet-red leafy crystals, which dissolve in water to form a
reddish-brown solution, and combines with bromides to form bromaurates
corresponding to the chloraurates. Aurous iodide, AuI, is a
light-yellow, sparingly soluble powder obtained, together with free
iodine, by adding potassium iodide to auric chloride; auric iodide,
AuI3, is formed as a dark-green powder at the same time, but it
readily decomposes to aurous iodide and iodine. Aurous iodide is also
obtained as a green solid by acting upon gold with iodine. The
iodaurates correspond to the chlor- and bromaurates; the potassium
salt, KAuI4, forms highly lustrous, intensely black, four-sided
prisms.

Aurous cyanide, AuCN, forms yellow, microscopic, hexagonal tables,
insoluble in water, and is obtained by the addition of hydrochloric
acid to a solution of potassium aurocyanide, KAu(CN)2. This salt is
prepared by precipitating a solution of gold in _aqua regia_ by
ammonia, and then introducing the well-washed precipitate into a
boiling solution of potassium cyanide. The solution is filtered and
allowed to cool, when colourless rhombic pyramids of the aurocyanide
separate. It is also obtained in the action of potassium cyanide on
gold in the presence of air, a reaction utilized in the
MacArthur-Forrest process of gold extraction (see below). Auric
cyanide, Au(CN)3, is not certainly known; its double salts, however,
have been frequently described. Potassium auricyanide, 2KAu(CN)4.3H2O,
is obtained as large, colourless, efflorescent tablets by
crystallizing concentrated solutions of auric chloride and potassium
cyanide. The acid, auricyanic acid, 2HAu(CN)4.3H2O, is obtained by
treating the silver salt (obtained by precipitating the potassium salt
with silver nitrate) with hydrochloric acid; it forms tabular
crystals, readily soluble in water, alcohol and ether.

Gold forms three sulphides corresponding to the oxides; they readily
decompose on heating. Aurous sulphide, Au2S, is a brownish-black
powder formed by passing sulphuretted hydrogen into a solution of
potassium aurocyanide and then acidifying. Sodium aurosulphide,
NaAuS.4H2O, is prepared by fusing gold with sodium sulphide and
sulphur, the melt being extracted with water, filtered in an
atmosphere of nitrogen, and evaporated in a vacuum over sulphuric
acid. It forms colourless, monoclinic prisms, which turn brown on
exposure to air. This method of bringing gold into solution is
mentioned by Stahl in his _Observationes Chymico-Physico-Medicae_; he
there remarks that Moses probably destroyed the golden calf by burning
it with sulphur and alkali (Ex. xxxii. 20). Auric sulphide, Au2S3, is
an amorphous powder formed when lithium aurichloride is treated with
dry sulphuretted hydrogen at -10 deg. It is very unstable, decomposing
into gold and sulphur at 200 deg.

Oxy-salts of gold are almost unknown, but the sulphite and
thiosulphate form double salts. Thus by adding acid sodium sulphite
to, or by passing sulphur dioxide at 50 deg. into, a solution of
sodium aurate, the salt, 3Na2SO3.Au2SO3.3H2O is obtained, which, when
precipitated from its aqueous solution by alcohol, forms a purple
powder, appearing yellow or green by reflected light. Sodium
aurothiosulphate, 3Na2S2O3.Au2S2O3.4H2O, forms colourless needles; it
is obtained in the direct action of sodium thiosulphate on gold in the
presence of an oxidizing agent, or by the addition of a dilute
solution of auric chloride to a sodium thiosulphate solution.

_Mining and Metallurgy._

The various deposits of gold may be divided into two classes--"veins" and "placers." The vein mining of gold does not greatly differ from that of similar deposits of metals (see MINERAL DEPOSITS). In the placer or alluvial deposits, the precious metal is found usually in a water-worn condition imbedded in earthy matter, and the method of working all such deposits is based on the disintegration of the earthy matter by the action of a stream of water, which washes away the lighter portions and leaves the denser gold. In alluvial deposits the richest ground is usually found in contact with the "bed rock"; and, when the overlying cover of gravel is very thick, or, as sometimes happens, when the older gravel is covered with a flow of basalt, regular mining by shafts and levels, as in what are known as tunnel-claims, may be required to reach the auriferous ground.

The extraction of gold may be effected by several methods; we may distinguish the following leading types:

1. By simple washing, i.e. dressing auriferous sands, gravels, &c.;

2. By amalgamation, i.e. forming a gold amalgam, afterwards removing the mercury by distillation;

3. By chlorination, i.e. forming the soluble gold chloride and then precipitating the metal;

4. By the cyanide process, i.e. dissolving the gold in potassium cyanide solution, and then precipitating the metal;

5. Electrolytically, generally applied to the solutions obtained in processes (3) and (4).

1. _Extraction of Gold by Washing._--In the early days of gold-washing
in California and Australia, when rich alluvial deposits were common
at the surface, the most simple appliances sufficed. The most
characteristic is the "pan," a circular dish of sheet-iron or "tin,"
with sloping sides about 13 or 14 in. in diameter. The pan, about
two-thirds filled with the "pay dirt" to be washed, is held in the
stream or in a hole filled with water. The larger stones having been
removed by hand, gyratory motion is given to the pan by a combination
of shaking and twisting movements so as to keep its contents
suspended in the stream of water, which carries away the bulk of the
lighter material, leaving the heavy minerals, together with any gold
which may have been present. The washing is repeated until enough of
the enriched sand is collected, when the gold is finally recovered by
careful washing or "panning out" in a smaller pan. In Mexico and South
America, instead of the pan, a wooden dish or trough, known as
"batea," is used.

The "cradle" is a simple appliance for treating somewhat larger
quantities, and consists essentially of a box, mounted on rockers, and
provided with a perforated bottom of sheet iron in which the "pay
dirt" is placed. Water is poured on the dirt, and the rocking motion
imparted to the cradle causes the finer particles to pass through the
perforated bottom on to a canvas screen, and thence to the base of the
cradle, where the auriferous particles accumulate on transverse bars
of wood, called "riffles."

The "tom" is a sort of cradle with an extended sluice placed on an
incline of about 1 in 12. The upper end contains a perforated riddle
plate which is placed directly over the riffle box, and under certain
circumstances mercury may be placed behind the riffles. Copper plates
amalgamated with mercury are also used when the gold is very fine, and
in some instances amalgamated silver coins have been used for the same
purpose. Sometimes the stuff is disintegrated with water in a
"puddling machine," which was used, especially in Australia, when the
earthy matters are tenacious and water scarce. The machine frequently
resembles a brickmaker's wash-mill, and is worked by horse or steam
power.

In workings on a larger scale, where the supply of water is abundant,
as in California, sluices were generally employed. They are shallow
troughs about 12 ft. long, about 16 to 20 in. wide and 1 ft. in depth.
The troughs taper slightly so that they can be joined in series, the
total length often reaching several hundred feet. The incline of the
sluice varies with the conformation of the ground and the tenacity of
the stuff to be washed, from 1 in 16 to 1 in 8. A rectangular trough
of boards, whose dimensions depend chiefly on the size of the planks
available, is set up on the higher part of the ground at one side of
the claim to be worked, upon trestles or piers of rough stone-work, at
such an inclination that the stream may carry off all but the largest
stones, which are kept back by a grating of boards about 2 in. apart.
The gravel is dug by hand and thrown in at the upper end, the stones
kept back being removed at intervals by two men with four-pronged
steel forks. The floor of the sluice is laid with riffles made of
strips of wood 2 in. square laid parallel to the direction of the
current, and at other points with boards having transverse notches
filled with mercury. These were known originally as Hungarian riffles.

In larger plant the upper ends of the sluices are often cut in rock or
lined with stone blocks, the grating stopping the larger stones being
known as a "grizzly." In order to save very fine and especially rusty
particles of gold, so-called "under-current sluices" are used; these
are shallow wooden tanks, 50 sq. yds. and upwards in area, which are
placed somewhat below the main sluice, and communicate with it above
and below, the entry being protected by a grating so that only the
finer material is admitted. These are paved with stone blocks or lined
with mercury riffles, so that from the greatly reduced velocity of
flow, due to the sudden increase of surface, the finer particles of
gold may collect. In order to save finely divided gold, amalgamated
copper plates are sometimes placed in a nearly level position, at a
considerable distance from the head of the sluice, the gold which is
retained in it being removed from time to time. Sluices are often made
double, and they are usually cleaned up--that is, the deposit rich in
gold is removed from them--once a week.

The "pan" is now only used by prospectors, while the "cradle" and
"tom" are practically confined to the Chinese; the sluice is
considered to be the best contrivance for washing gold gravels.

2. _The Amalgamation Process._--This method is employed to extract gold from both alluvial and reef deposits: in the first case it is combined with "hydraulic mining," i.e. disintegrating auriferous gravels by powerful jets of water, and the sluice system described above; in the second case the vein stuff is prepared by crushing and the amalgamation is carried out in mills.

Hydraulic mining has for the most part been confined to the country of
its invention, California, and the western territories of America,
where the conditions favourable for its use are more fully developed
than elsewhere--notably the presence of thick banks of gravel that
cannot be utilized by other methods, and abundance of water, even
though considerable work may be required at times to make it
available. The general conditions to be observed in such workings may
be briefly stated as follows: (1) The whole of the auriferous gravel,
down to the "bed rock," must be removed,--that is, no selection of
rich or poor parts is possible; (2) this must be accomplished by the
aid of water alone, or at times by water supplemented by blasting; (3)
the conglomerate must be mechanically disintegrated without
interrupting the whole system; (4) the gold must be saved without
interrupting the continuous flow of water; and (5) arrangements must
be made for disposing of the vast masses of impoverished gravel.

The water is brought from a ditch on the high ground, and through a
line of pipes to the distributing box, whence the branch pipes
supplying the jets diverge. The stream issues through a nozzle,
termed a "monitor" or "giant," which is fitted with a ball and socket
joint, so that the direction of the jet may be varied through
considerable angles by simply moving a handle. The material of the
bank being loosened by blasting and the cutting action of the water,
crumbles into holes, and the superincumbent mass, often with large
trees and stones, falls into the lower ground. The stream, laden with
stones and gravel, passes into the sluices, where the gold is
recovered in the manner already described. Under the most advantageous
conditions the loss of gold may be estimated at 15 or 20%, the amount
recovered representing a value of about two shillings per ton of
gravel treated. The loss of mercury is about the same, from 5 to 6
cwt. being in constant use per mile of sluice.

In working auriferous river-beds, dredges have been used with
considerable success in certain parts of New Zealand and on the
Pacific slope in America. The dredges used in California are almost
exclusively of the endless-chain bucket or steam-shovel pattern. Some
dredges have a capacity under favourable conditions of over 2000 cub.
yds. of gravel daily. The gravel is excavated as in the ordinary form
of endless-chain bucket dredge and dumped on to the deck of the
dredge. It then passes through screens and grizzlies to retain the
coarse gravel, the finer material passing on to sluice boxes provided
with riffles, supplied with mercury. There are belt conveyers for
discharging the gravel and tailings at the end of the vessel remote
from the buckets. The water necessary to the process is pumped from
the river; as much as 2000 gallons per minute is used on the larger
dredges.

The dressing or mechanical preparation of vein stuff containing gold
is generally similar to that of other ores (see ORE-DRESSING), except
that the precious metal should be removed from the waste substances as
quickly as possible, even although other minerals of value that are
subsequently recovered may be present. In all cases the quartz or
other vein stuff must be reduced to a very fine powder as a
preliminary to further operations. This may be done in several ways,
e.g. either (1) by the Mexican crusher or _arrastra_, in which the
grinding is effected upon a bed of stone, over which heavy blocks of
stone attached to cross arms are dragged by the rotation of the arms
about a central spindle, or (2) by the Chilean mill or _trapiche_,
also known as the edge-runner, where the grinding stones roll upon the
floor, at the same time turning about a central upright--contrivances
which are mainly used for the preparation of silver ores; but by far
the largest proportion of the gold quartz of California, Australia and
Africa is reduced by (3) the stamp mill, which is similar in principle
to that used in Europe for the preparation of tin and other ores.

The stamp mill was first used in California, and its use has since
spread over the whole world. In the mills of the Californian type the
stamp is a cylindrical iron pestle faced with a chilled cast iron
shoe, removable so that it can be renewed when necessary, attached to
a round iron rod or lifter, the whole weighing from 600 to 900 lb.;
stamps weighing 1320 lb. are in use in the Transvaal. The lift is
effected by cams acting on the under surface of tappets, and formed by
cylindrical boxes keyed on to the stems of the lifter about one-fourth
of their length from the top. As, however, the cams, unlike those of
European stamp mills, are placed to one side of the stamp, the latter
is not only lifted but turned partly round on its own axis, whereby
the shoes are worn down uniformly. The height of lift may be between 4
and 18 in., and the number of blows from 30 to over 100 per minute.
The stamps are usually arranged in batteries of five; the order of
working is usually 1, 4, 2, 5, 3, but other arrangements, e.g. 1, 3,
5, 2, 4, and 1, 5, 2, 4, 3, are common. The stuff, previously broken
to about 2-in. lumps in a rock-breaker, is fed in through an aperture
at the back of the "battery box," a constant supply of water is
admitted from above, and mercury in a finely divided state is added at
frequent intervals. The discharge of the comminuted material takes
place through an aperture, which is covered by a thin steel plate
perforated with numerous slits about 1/50th in. broad and 1/2 in.
long, a certain volume being discharged at every blow and carried
forward by the flushing water over an apron or table in front, covered
by copper plates filled with mercury. Similar plates are often used to
catch any particles of gold that may be thrown back, while the main
operation is so conducted that the bulk of the gold may be reduced to
the state of amalgam by bringing the two metals into intimate contact
under the stamp head, and remain in the battery. The tables in front
are laid at an incline of about 8 deg. and are about 13 ft. long; they
collect from 10 to 15% of the whole gold; a further quantity is
recovered by leading the sands through a gutter about 16 in. broad and
120 ft. long, also lined with amalgamated copper plates, after the
pyritic and other heavy minerals have been separated by depositing in
catch pits and other similar contrivances.

When the ore does not contain any considerable amount of free gold
mercury is not, as a rule, used during the crushing, but the
amalgamation is carried out in a separate plant. Contrivances of the
most diverse constructions have been employed. The most primitive is
the rubbing together of the concentrated crushings with mercury in
iron mortars. Barrel amalgamation, i.e. mixing the crushings with
mercury in rotating barrels, is rarely used, the process being
wasteful, since the mercury is specially apt to be "floured" (see
below).

At Schemnitz, Kerpenyes, Kreuzberg and other localities in Hungary,
quartz vein stuff containing a little gold, partly free and partly
associated with pyrites and galena, is, after stamping in mills,
similar to those described above, but without rotating stamps, passed
through the so-called "Hungarian gold mill" or "quick-mill." This
consists of a cast-iron pan having a shallow cylindrical bottom
holding mercury, in which a wooden muller, nearly of the same shape as
the inside of the pan, and armed below with several projecting blades,
is made to revolve by gearing wheels. The stuff from the stamps is
conveyed to the middle of the muller, and is distributed over the
mercury, when the gold subsides, while the quartz and lighter
materials are guided by the blades to the circumference and are
discharged, usually into a second similar mill, and subsequently pass
over blanket tables, i.e. boards covered with canvas or sacking, the
gold and heavier particles becoming entangled in the fibres. The
action of this mill is really more nearly analogous to that of a
centrifugal pump, as no grinding action takes place in it. The amalgam
is cleaned out periodically--fortnightly or monthly--and after
filtering through linen bags to remove the excess of mercury, it is
transferred to retorts for distillation (see below).

Many other forms of pan-amalgamators have been devised. The Laszlo is
an improved Hungarian mill, while the Piccard is of the same type. In
the Knox and Boss mills, which are also employed for the amalgamation
of silver ores, the grinding is effected between flat horizontal
surfaces instead of conical or curved surfaces as in the previously
described forms.

One of the greatest difficulties in the treatment of gold by
amalgamation, and more particularly in the treatment of pyrites,
arises from the so-called "sickening" or "flouring" of the mercury;
that is, the particles, losing their bright metallic surfaces, are no
longer capable of coalescing with or taking up other metals. Of the
numerous remedies proposed the most efficacious is perhaps sodium
amalgam. It appears that amalgamation is often impeded by the tarnish
found on the surface of the gold when it is associated with sulphur,
arsenic, bismuth, antimony or tellurium. Henry Wurtz in America (1864)
and Sir William Crookes in England (1865) made independently the
discovery that, by the addition of a small quantity of sodium to the
mercury, the operation is much facilitated. It is also stated that
sodium prevents both the "sickening" and the "flouring" of the mercury
which is produced by certain associated minerals. The addition of
potassium cyanide has been suggested to assist the amalgamation and to
prevent "flouring," but Skey has shown that its use is attended with
loss of gold.

_Separation of Gold from the Amalgam._--The amalgam is first pressed
in wetted canvas or buckskin in order to remove excess of mercury.
Lumps of the solid amalgam, about 2 in. in diameter, are introduced
into an iron vessel provided with an iron tube that leads into a
condenser containing water. The distillation is then effected by
heating to dull redness. The amalgam yields about 30 to 40% of gold.
Horizontal cylindrical retorts, holding from 200 to 1200 lb. of
amalgam, are used in the larger Californian mills, pot retorts being
used in the smaller mills. The bullion left in the retorts is then
melted in black-lead crucibles, with the addition of small quantities
of suitable fluxes, e.g. nitre, sodium carbonate, &c.

The extraction of gold from auriferous minerals by fusion, except as
an incident in their treatment for other metals, is very rarely
practised. It was at one time proposed to treat the concentrated black
iron obtained in the Ural gold washings, which consists chiefly of
magnetite, as an iron ore, by smelting it with charcoal for auriferous
pig-iron, the latter metal possessing the property of dissolving gold
in considerable quantity. By subsequent treatment with sulphuric acid
the gold could be recovered. Experiments on this point were made by
Anossow in 1835, but they have never been followed in practice.

Gold in galena or other lead ores is invariably recovered in the
refining or treatment of the lead and silver obtained. Pyritic ores
containing copper are treated by methods analogous to those of the
copper smelter. In Colorado the pyritic ores containing gold and
silver in association with copper are smelted in reverberatory
furnaces for regulus, which, when desilverized by Ziervogel's method,
leaves a residue containing 20 or 30 oz. of gold per ton. This is
smelted with rich gold ores, notably those containing tellurium, for
white metal or regulus; and by a following process of partial
reduction analogous to that of selecting in copper smelting, "bottoms"
of impure copper are obtained in which practically all the gold is
concentrated. By continuing the treatment of these in the ordinary way
of refining, poling and granulating, all the foreign matters other
than gold, copper and silver are removed, and, by exposing the
granulated metal to a high oxidizing heat for a considerable time the
copper may be completely oxidized while the precious metals are
unaltered. Subsequent treatment with sulphuric acid renders the copper
soluble in water as sulphate, and the final residue contains only gold
and silver, which is parted or refined in the ordinary way. This
method of separating gold from copper, by converting the latter into
oxide and sulphate, is also used at Oker in the Harz.

_Extraction by Means of Aqueous Solutions._--Many processes have been suggested in which the gold of auriferous deposits is converted into products soluble in water, from which solutions the gold may be precipitated. Of these processes, two only are of special importance, viz. the chlorination or Plattner process, in which the metal is converted into the chloride, and the cyanide or MacArthur-Forrest process, in which it is converted into potassium aurocyanide.

(3) _Chlorination or Plattner Process._--In this process moistened
gold ores are treated with chlorine gas, the resulting gold chloride
dissolved out with water, and the gold precipitated with ferrous
sulphate, charcoal, sulphuretted hydrogen or otherwise. The process
originated in 1848 with C. F. Plattner, who suggested that the
residues from certain mines at Reichenstein, in Silesia, should be
treated with chlorine after the arsenical products had been extracted
by roasting. It must be noticed, however, that Percy independently
made the same discovery, and stated his results at the meeting of the
British Association (at Swansea) in 1849, but the Report was not
published until 1852. The process was introduced in 1858 by Deetken at
Grass Valley, California, where the waste minerals, principally
pyrites from tailings, had been worked for a considerable time by
amalgamation. The process is rarely applied to ores direct;
free-milling ores are generally amalgamated, and the tailings and
slimes, after concentration, operated upon. Three stages in the
process are to be distinguished: (i) calcination, to convert all the
metals, except gold and silver, into oxides, which are unacted upon by
chlorine; (ii.) chlorinating the gold and lixiviating the product;
(iii.) precipitating the gold.

The calcination, or roasting, is conducted at a low temperature in
some form of reverberatory furnace. Salt is added in the roasting to
convert any lime, magnesia or lead which may be present, into the
corresponding chlorides. The auric chloride is, however, decomposed at
the elevated temperature into finely divided metallic gold, which is
then readily attacked by the chlorine gas. The high volatility of gold
in the presence of certain metals must also be considered. According
to Egleston the loss may be from 40 to 90% of the total gold present
in cupriferous ores according to the temperature and duration of
calcination. The roasted mineral, slightly moistened, is introduced
into a vat made of stoneware or pitched planks, and furnished with a
double bottom. Chlorine, generally prepared by the interaction of
pyrolusite, salt and sulphuric acid, is led from a suitable generator
beneath the false bottom, and rises through the moistened ore, which
rests on a bed of broken quartz; the gold is thus converted into a
soluble chloride, which is afterwards removed by washing with water.
Both fixed and rotating vats are employed, the chlorination proceeding
more rapidly in the latter case; rotating barrels are sometimes used.
There have also been introduced processes in which the chlorine is
generated in the chloridizing vat, the reagents used being dilute
solutions of bleaching powder and an acid. Munktell's process is of
this type. In the Thies process, used in many districts in the United
States, the vats are rotating barrels made, in the later forms, of
iron lined with lead, and provided with a filter formed of a finely
perforated leaden grating running from one end of the barrel to the
other, and rigidly held in place by wooden frames. Chlorine is
generated within the barrel from sulphuric acid and chloride of lime.
After charging, the barrel is rotated, and when the chlorination is
complete the contents are emptied on a filter of quartz or some
similar material, and the filtrate led to settling tanks.

After settling the solution is run into the precipitating tanks. The
precipitants in use are: ferrous sulphate, charcoal and sulphuretted
hydrogen, either alone or mixed with sulphur dioxide; the use of
copper and iron sulphides has been suggested, but apparently these
substances have achieved no success.

In the case of ferrous sulphate, prepared by dissolving iron in dilute
sulphuric acid, the reaction follows the equation AuCl3 + 3FeSO4 =
FeCl3 + Fe2(SO4)3 + Au. At the same time any lead, calcium, barium and
strontium present are precipitated as sulphates; it is therefore
advantageous to remove these metals by the preliminary addition of
sulphuric acid, which also serves to keep any basic iron salts in
solution. The precipitation is carried out in tanks or vats made with
wooden sides and a cement bottom. The solutions are well mixed by
stirring with wooden poles, and the gold allowed to settle, the time
allowed varying from 12 to 72 hours. The supernatant liquid is led
into settling tanks, where a further amount of gold is deposited, and
is then filtered through sawdust or sand, the sawdust being afterwards
burnt and the gold separated from the ashes and the sand treated in
the chloridizing vat. The precipitated gold is washed, treated with
salt and sulphuric acid to remove iron salts, roughly dried by
pressing in cloths or on filter paper, and then melted with salt,
borax and nitre in graphite crucibles. Thus prepared it has a fineness
of 800-960, the chief impurities usually being iron and lead.

Charcoal is used as the precipitant at Mount Morgan, Australia. Its
use was proposed as early as 1818 and 1819 by Hare and Henry; Percy
advocated it in 1869, and Davis adopted it on the large scale at a
works in Carolina in 1880. The action is not properly understood; it
may be due to the reducing gases (hydrogen, hydrocarbons, &c.) which
are invariably present in wood charcoal. The process consists
essentially in running the solution over layers of charcoal, the
charcoal being afterwards burned. It has been found that the reaction
proceeds faster when the solution is heated.

Precipitation with sulphur dioxide and sulphuretted hydrogen proceeds
much more rapidly, and has been adopted at many works. Sulphur
dioxide, generated by burning sulphur, is forced into the solution
under pressure, where it interacts with any free chlorine present to
form hydrochloric and sulphuric acids. Sulphuretted hydrogen, obtained
by treating iron sulphide or a coarse matte with dilute sulphuric
acid, is forced in similarly. The gold is precipitated as the
sulphide, together with any arsenic, antimony, copper, silver and lead
which may be present. The precipitate is collected in a filter-press,
and then roasted in muffle furnaces with nitre, borax and sodium
carbonate. The fineness of the gold so obtained is 900 to 950.

4. _Cyanide Process._--This process depends upon the solubility of
gold in a dilute solution of potassium cyanide in the presence of air
(or some other oxidizing agent), and the subsequent precipitation of
the gold by metallic zinc or by electrolysis. The solubility of gold
in cyanide solutions was known to K. W. Scheele in 1782; and M.
Faraday applied it to the preparation of extremely thin films of the
metal. L. Eisner recognized, in 1846, the part played by the
atmosphere, and in 1879 Dixon showed that bleaching powder, manganese
dioxide, and other oxidizing agents, facilitated the solution. S. B.
Christy (_Trans. A.I.M.E._, 1896, vol. 26) has shown that the solution
is hastened by many oxidizing agents, especially sodium and manganese
dioxides and potassium ferricyanide. According to G. Bodlander (_Zeit.
f. angew. Chem._, 1896, vol. 19) the rate of solution in potassium
cyanide depends upon the subdivision of the gold--the finer the
subdivision the quicker the solution,--and on the concentration of the
solution--the rate increasing until the solution contains 0.25% of
cyanide, and remaining fairly stationary with increasing
concentration. The action proceeds in two stages; in the first
hydrogen peroxide and potassium aurocyanide are formed, and in the
second the hydrogen peroxide oxidizes a further quantity of gold and
potassium cyanide to aurocyanide, thus (1) 2Au + 4KCN + O2 + 2H2O =
2KAu(CN)2 + 4KOH + H2O2; (2) 2Au + 4KCN + 2H2O2 = 2KAu(CN)2 + 4KOH.
The end reaction may be written 4Au + 8KCN + 2H2O + O2 = 4KAu(CN)2 +
4KOH.

The commercial process was patented in 1890 by MacArthur and Forrest,
and is now in use all over the world. It is best adapted for
free-milling ores, especially after the bulk of the gold has been
removed by amalgamation. It has been especially successful in the
Transvaal. In the Witwatersrand the ore, which contains about 9 dwts.
of gold to the metric ton (2000 lb.), is stamped and amalgamated, and
the slimes and tailings, containing about 3-1/2 dwts. per ton, are
cyanided, about 2 dwts. more being thus extracted. The total cost per
ton of ore treated is about 6s., of which the cyaniding costs from 2s.
to 4s.

The process embraces three operations: (1) Solution of the gold; (2)
precipitation of the gold; (3) treatment of the precipitate.

The ores, having been broken and ground, generally in tube mills,
until they pass a 150 to 200-mesh sieve, are transferred to the
leaching vats, which are constructed of wood, iron or masonry; steel
vats, coated inside and out with pitch, of circular section and
holding up to 1000 tons, have come into use. The diameter is generally
26 ft., but may be greater; the best depth is considered to be a
quarter of the diameter. The vats are fitted with filters made of
coco-nut matting and jute cloth supported on wooden frames. The
leaching is generally carried out with a strong, medium, and with a
weak liquor, in the order given; sometimes there is a preliminary
leaching with a weak liquor. The strengths employed depend also upon
the mode of precipitation adopted, stronger solutions (up to 0.25%
KCN) being used when zinc is the precipitant. For electrolytic
precipitation the solution may contain up to 0.1% KCN. The liquors are
run off from the vats to the electrolysing baths or precipitating
tanks, and the leached ores are removed by means of doors in the sides
of the vats into wagons. In the Transvaal the operation occupies 3-1/2
to 4 days for fine sands, and up to 14 days for coarse sands; the
quantity of cyanide per ton of tailings varies from 0.26 to 0.28 lb.,
for electrolytic precipitation, and 0.5 lb. for zinc precipitation.

The precipitation is effected by zinc in the form of bright turnings,
or coated with lead, or by electrolysis. According to Christy, the
precipitation with zinc follows equations 1 or 2 according as
potassium cyanide is present or not:

(1) 4KAu(CN)2 + 4Zn + 2H2O = 2Zn(CN)2 + K2Zn(CN)4 + Zn(OK)2 + 4H + 4Au;

(2) 2KAu(CN)2 + 3Zn + 4KCN + 2H2O = 2K2Zn(CN)4 + Zn(OK)2 + 4H + 2Au;

one part of zinc precipitating 3.1 parts of gold in the first case,
and 2.06 in the second. It may be noticed that the potassium zinc
cyanide is useless in gold extraction, for it neither dissolves gold
nor can potassium cyanide be regenerated from it.

The precipitating boxes, generally made of wood but sometimes of
steel, and set on an incline, are divided by partitions into
alternately wide and narrow compartments, so that the liquor travels
upwards in its passage through the wide divisions and downwards
through the narrow divisions. In the wider compartments are placed
sieves having sixteen holes to the square inch and bearing zinc
turnings. The gold and other metals are precipitated on the under
surfaces of the turnings and fall to the bottom of the compartment as
a black slime. The slime is cleaned out fortnightly or monthly, the
zinc turnings being cleaned by rubbing and the supernatant liquor
allowed to settle in the precipitating boxes or in separate vessels.
The slime so obtained consists of finely divided gold and silver
(5-50%), zinc (30-60%), lead (10%), carbon (10%), together with tin,
copper, antimony, arsenic and other impurities of the zinc and ores.
After well washing with water, the slimes are roughly dried in
bag-filters or filter-presses, and then treated with dilute sulphuric
acid, the solution being heated by steam. This dissolves out the zinc.
Lime is added to bring down the gold, and the sediment, after washing
and drying, is fused in graphite crucibles.

5. _Electrolytic Processes._--The electrolytic separation of the gold
from cyanide solutions was first practised in the Transvaal. The
process, as elaborated by Messrs. Siemens and Halske, essentially
consists in the electrolysis of weak solutions with iron or steel
plate anodes, and lead cathodes, the latter, when coated with gold,
being fused and cupelled. Its advantages over the zinc process are
that the deposited gold is purer and more readily extracted, and that
weaker solutions can be employed, thereby effecting an economy in
cyanide.

In the process employed at the Worcester Works in the Transvaal, the
liquors, containing about 150 grains of gold per ton and from 0.08 to
0.01% of cyanide, are treated in rectangular vats in which is placed a
series of iron and leaden plates at intervals of 1 in. The cathodes,
which are sheets of thin lead foil weighing 1-1/2 lb. to the sq. yd.,
are removed monthly, their gold content being from 0.5 to 10%, and
after folding are melted in reverberatory furnaces to ingots
containing 2 to 4% of gold. Cupellation brings up the gold to about
900 fine. Many variations of the electrolytic process as above
outlined have been suggested. S. Cowper Coles has suggested aluminium
cathodes; Andreoli has recommended cathodes of iron and anodes of lead
coated with lead peroxide, the gold being removed from the iron
cathodes by a brief immersion in molten lead; in the Pelatan-Cerici
process the gold is amalgamated at a mercury cathode (see also below).

_Refining or Parting of Gold._--Gold is almost always silver-bearing, and it may be also noticed that silver generally contains some gold. Consequently the separation of these two metals Is one of the most important metallurgical processes. In addition to the separation of the silver the operation extends to the elimination of the last traces of lead, tin, arsenic, &c. which have resisted the preceding cupellation.

The "parting" of gold and silver is of considerable antiquity. Thus
Strabo states that in his time a process was employed for refining and
purifying gold in large quantities by cementing or burning it with an
aluminous earth, which, by destroying the silver, left the gold in a
state of purity. Pliny shows that for this purpose the gold was placed
on the fire in an earthen vessel with treble its weight of salt, and
that it was afterwards again exposed to the fire with two parts of
salt and one of argillaceous rock, which, in the presence of moisture,
effected the decomposition of the salt; by this means the silver
became converted into chloride.

The methods of parting can be classified into "dry," "wet" and
electrolytic methods. In the "dry" methods the silver is converted
into sulphide or chloride, the gold remaining unaltered; in the "wet"
methods the silver is dissolved by nitric acid or boiling sulphuric
acid; and in the electrolytic processes advantage is taken of the fact
that under certain current densities and other circumstances silver
passes from an anode composed of a gold-silver alloy to the cathode
more readily than gold. Of the dry methods only F. B. Miller's
chlorine process is of any importance, this method, and the wet
process of refining by sulphuric acid, together with the electrolytic
process, being the only ones now practised.

The conversion of silver into the sulphide may be effected by heating
with antimony sulphide, litharge and sulphur, pyrites, or with sulphur
alone. The antimony, or _Guss und Fluss_, method was practised up till
1846 at the Dresden mint; it is only applicable to alloys containing
more than 50% of gold. The fusion results in the formation of a
gold-antimony alloy, from which the antimony is removed by an
oxidizing fusion with nitre. The sulphur and litharge, or
_Pfannenschmied_, process was used to concentrate the gold in an alloy
in order to make it amenable to "quartation," or parting with nitric
acid. Fusion with sulphur was used for the same purpose as the
Pfannenschmied process. It was employed in 1797 at the St Petersburg
mint.

The conversion of the silver into the chloride may be effected by
means of salt--the "cementation" process--or other chlorides, or by
free chlorine--Miller's process. The first process consists
essentially in heating the alloy with salt and brickdust; the latter
absorbs the chloride formed, while the gold is recovered by washing.
It is no longer employed. The second process depends upon the fact
that, if chlorine be led into the molten alloy, the base metals and
the silver are converted into chlorides. It was proposed in 1838 by
Lewis Thompson, but it was only applied commercially after Miller's
improvements in 1867, when it was adopted at the Sydney mint. Sir W.
C. Roberts-Austen introduced it at the London mint; and it has also
been used at Pretoria. It is especially suitable to gold containing
little silver and base metals--a character of Australian gold--but it
yields to the sulphuric acid and electrolytic methods in point of
economy.