V. A. Pazukhin.

At the Mines of the Kuznetsk Alatau

In May of last year, 1915, I was granted, by the Council of the Tomsk Technological Institute, a travel assignment to study the organization of the cyanide process at certain mines of the Kuznetsk Alatau—a region where the development of ore gold and the application of cyanidation are rapidly growing, both in the quantity of material processed and in technical terms.

I had to visit the mines selected by me at a moment of severe crisis caused by the war. The cyanide plants were idle and could not be started up because of the absence of stocks of potassium cyanide and zinc, of the timely receipt of which there was no certainty. In addition, there was an acute shortage of labor. This circumstance deprived me of the possibility of gathering exhaustive material. Nevertheless, the accelerating pace of development of the gold industry in the Kuznetsk Alatau and its primary importance to the state give me the right to suppose that the publication of the information I have collected, though incomplete, will not be without benefit.

I. A Brief Historical Sketch.*)

Leaving aside the question of gold mining in the Kuznetsk Alatau in prehistoric and ancient times, I shall point out that the pioneer of the gold industry in this region and, in general, in Siberia was A. Ya. Popov, who in 1826 received permission to prospect gold-bearing areas. Already two years later, in 1828, he filed a claim for gold in the region of the Berikul River, and

*) See Reutovsky, Mineral Resources of Siberia.
Obruchev, Geological Survey of the Gold-Bearing Regions of Siberia.
Beresnevich. Report on Statistical, Economic, and Technical Investigations of the Tomsk Mining District, vol. I.

In the following year they had already extracted 1 pood 10 pounds 18 zolotniks 72 dolyas of gold. In 1837 the number of gold-miners reached 200, and 33 placers were being worked in the Alatau, with a total output of 57 poods 9 pounds 77 zolotniks.

In the 1870s the quantity of placer gold extracted reached its maximum (70–77 poods), and from that time, as the areas were worked out, it began steadily to decline, reaching 5–6 poods by the 1900s. Altogether, up to 1909, the region yielded 2993 poods of placer gold.

At present, work is still being carried on here and there in washing placers, but annual production does not exceed 5–6 poods.

Lode gold first began to be developed in 1877, when 5 pounds 51 zolotniks were extracted. Development soon ceased, and only from 1892 did continuous extraction of lode gold begin.

II. The Central Mine.

The Central Mine, together with the sections grouped around it, belongs to the “Gold-Mining Society of the Mariinsk Mines, formerly belonging to Ivanitsky” (Russian Gold-Mining Society), and is situated along the course of the Tata-Kozhukh brook, which belongs to the basin of the river Kiya.

The locality around the Central Mine had long been famous for its rich placers, now almost completely worked out. The figure of 1468 poods of placer gold extracted here speaks eloquently of the washing that once flourished here. It was in this very district, at the Dmitrievsky Mine, that the development of lode gold discovered here in 1878 first began. The work, however, continued only until 1881. In 1892, with the transfer of the mine into the hands of Ivanitsky, it was resumed, but not for long—in 1900 the mine was abandoned. In the same year, at the Lottery Mine, a rich vein was discovered and its development begun; afterward a whole series of vein outcrops was found, with varying contents and thicknesses. From that time on, the development of lode gold became firmly established.

In 1910–11 the Central mine, together with the adjoining areas, passed into the hands of the Gold-Industrial Company, which put the operations on a broader scale, employing improved methods for processing the ores extracted.

Like most of the mines in the Alatau, the Central mine is situated under very unfavorable conditions with respect to means of communication. The mine is 140 versts from the railway, and 95 from the village of Tisul. As late as 1912–13 there was no cart road at all from Tisul to the Central mine. Recently the Resettlement Administration undertook the construction of a road through the Mariinsk taiga. This road, although equipped quite satisfactorily, at present falls 25 versts short of the mine, and this intervening stretch still remains in its primitive state—only a taiga trail passes here. Owing to this, in summer the delivery of freight to the mine is extremely difficult: one must use either small two-wheeled carts or pack horses, and the delivery of freight from the railway to the mine costs from 80 kopecks to 1 ruble 10 kopecks per pood. During the spring and autumn muddy seasons, the mine is for a considerable time completely cut off from the outside world.

The settlement, according to local usage the stan, is situated along the course of the Taga-Kozhukh brook. The population of the settlement reaches 2,000 persons. There are a school, a club, and a hospital with medical personnel.

General Survey of Technical Installations.

As mentioned above, systematic development of ore gold in the area of the Central mine began in 1900. In 1904 the former owner of the mine built a gold-washing factory with 4 bowls. Under the name “Old Factory” it has survived to the present day, although at present only two pairs of runners are working there.

Ivanitsky, for his part, built a summer tailings plant for processing accumulated dumps, operating

and at the present time. In 1912, with the transfer of the mine into the ownership of the Gold-Mining Company, a summer silt-washing plant was built for processing the old silts, and construction was also begun on a new 8-stamp factory; somewhat later, the construction of a warm silt-washing plant was begun.

In March 1913, four stamps were started up at the new factory, and a year later, in February 1914, the factory was operating with 8 stamps. In December of the same year, the new silt-washing plant was also put into operation, and in its technical equipment it was not inferior to the Kochkar plants.

Simultaneously with the construction of the warm silt-washing plant, a central electric station, mechanical workshops, metallurgical and other buildings necessary for work on a large scale were erected.

Mining Operations.

At present, on the allotments belonging to the Company, mining operations are being carried on in many places, but the following are being worked more or less intensively:

mines:

Mine	Annual output
Pokrovskaya, of the Central mine	1,200 thousand poods
Voskresenskaya, of the Central mine	335 thousand poods
Blagoveshchenskaya, of the Lottery mine	414 thousand poods
Charkovskaya, of the Chirkovsky mine	137 thousand poods

adit

Adit	Annual output
Zakoldovannaya	1,000 thousand poods

Total annual production is approximately 3,000 thousand poods.

Alongside the Company’s own operations, work by independent prospectors is also being carried on, of course on a small scale and with fluctuating productivity.

The extracted ore is sorted after being hoisted to the surface: waste rock and solid pieces of calcedony are removed, which, as the workings deepen, in places

occur in considerable quantities. The ore from the place of extraction is brought on wheelbarrows to the new Uspenskaya mill.

The cost of hauling varies, depending on the distance, within wide limits. On average, hauling costs about 30 kopecks per 100 poods.

The ore at Tsentralnoye is rather visibly gold-bearing. For the most part it is quartz with small streaks of ocherous clay and with a varying content of sulfides, chiefly pyrite.

Crushing.

On wheelbarrows holding about 20–25 poods, the sorted ore is brought up to a trestle, where it is tipped onto a horizontal grizzly with openings of about 1″. The ore that has not fallen through is raked by two workers from the grizzly into a bin installed above the crusher; pieces that are too large are broken with hammers. From the bin the ore slides into a trough, from which the worker standing at the crusher feeds the ore, as needed, into the jaws of the crusher.

The Black-type crusher, with a capacity of 600–700 poods per hour of material, the size of whose pieces is 1″–1½″, is driven by a 30-horsepower electric motor. The crusher jaws wear out within 15–20 days.

The crushed material and the ore that has fallen through the grizzly go directly into mine cars with a capacity of about 80 poods and, along a rail track about 50 sazhen long, are taken to ore bins installed above the runners, into which it is dumped.

From each car, as it passes, a sample is taken with a shovel. In this way, over 2 days, 6–8 poods of ore accumulate; this is gradually reduced by hand to a lump size of ½–¼″, leveled out on the floor, and from it about 1 pood is taken from various places, constituting the laboratory sample.

Of course, with such a method of sampling there is no

there is no certainty that laboratory data will correspond to reality. Del-Mar points out,*) that sampling absolutely must be automatic and must not depend on the will of the worker. Moreover, in coarse crushing of the material taken, the latter is, of necessity, impoverished in the more fragile components—in the present case, gold-bearing sulfides—which become dusty and are lost.

The crusher operates 8–9 hours per day, which is sufficient to ensure the daily work of the mill.

At the crusher there are 9 workers:

2 at the ore discharge,

1 at the crusher (he is also the oiler),

5 rollers, and

1 boy taking the sample.

Grinding of the Ore.

In the new mill there have been installed 8 pairs of runners of the ordinary Kachkar type. Each pair is fed from a separate ore bin holding about 1,200 poods (16–17 tons) of ore, a quantity sufficient to feed a bowl for 24–30 hours.

The bowls are fed by hand, by raking the ore along the chute onto which the ore is dumped from the bin. One bowl is equipped with a Challenge feeder, which, according to the administration, works excellently. It is proposed to equip all the bowls with such feeders, manufactured by their own means.

A shortcoming of the Challenge feeder, as applied to the bowls, is the absence of automatic regulation of the feed rate, which here depends on the attentiveness of the worker. At the Troitsk Co. mill in Kachkar, the runners are fed by means of a belt conveyor, the speed of motion of which is varied depending—

*) Del Mar. Stamp Milling, 1912, p. 6.

the bridge according to the quantity of material in the cups. It seems to me that, by a corresponding alteration of the parts, automatic regulation of the feed could be carried out here as well.

In winter the feeding of the cups is made considerably more difficult. Snow and ice strongly cement the ore in the lar, since the premises are rather cold. It is therefore necessary to increase the number of loaders.

Of course, automatic feeding would be hindered by this circumstance to an even greater degree, and it will probably be necessary to heat the factory building.

Of the 8 cups operating at the Uspenskaya factory, 4, installed first, differ somewhat in size from the cups put into operation later.

The principal dimensions of the first series of beguns are as follows:

Item	Value
Outside lower diameter of the cup	9′7 1/2″
Inside lower diameter of the cup	6′
Upper diameter of the cup	12′9″
Height of the cup	3′1/2″
Distance between the axes of the millstones	7′10″
Diameter of the band	5′6″
Thickness of the band	7 1/2″
Width of the band	10 3/4″
Weight of the band	110 poods
Diameter of the bedstone forming the base of the begun	4′7/8″
Weight	80 poods
Total weight of the wheel	160—200 poods
Thickness of the liners	4″

The service life of the bands and liners is 10—15 months. The lower diameter of the cups of the second series is 10′6″; the remaining dimensions differ little from those of the cups of the first series.

All eight pairs have a lower drive and operate from two transmissions, which set in motion

by two electric motors of 40 HP each. Thus, one pan receives 10 HP. The runners make 10 revolutions per minute. With a screen of 8 openings per linear inch (screen No. 8), a sill height of 9″, and a ratio of liquid to solid of \(\frac{10}{1}\), the productivity of the runners is 30–40 poods per hour, which per one horsepower in 24 hours amounts to 84–96 poods, or 1.35–1.55 tons. At the same time, the slimes from the pans contain 35–40% of product—200. These figures indicate a not especially high productivity of the pans at the Uspensky mill.*)

The mill and plant are supplied with water by means of launders conducting water by gravity from the headwaters of a brook. Between the ore bins there is installed a tank with a capacity of about 1000 vedros, serving as a reservoir for water. In winter, when the water is cold and there is not enough of it, they use the drain waters from the thickener, which are lifted by a 6″ centrifugal pump, coupled to a 17-horsepower electric motor, into the same reservoir. In winter, when the low temperature of the water greatly hampers the work during washing, warm water is used, which, with the aid of an injector from the locomotive, is lifted into a small tank of 300 vedros capacity installed between the bins.

Sluices. Through the pan screen the crushed material is carried by water to the sluices. In order that the particles thrown out with force should not wear out the head plate, near the pan outlet there is a projection to which the first plate is fitted closely, cut in reverse. This projection also ensures a tight connection of the pan with the sluice.

The sluice is assembled from copper amalgamated plates of \(1 \times 2\) arshins. The head plate is laid across the sluice; the remaining ones lengthwise, in pairs. The joint between the plates, as well as their connections with the sides of the sluice, are clamped with laths. Behind the head plate and between the subsequent plates, for catching mercury and for amalgamating

*) Boyldon (Trans. Inst. of Min. a. Metall., Vol. XX, p. 141) gives for the runners of the Troitsk mill at Kochkar a productivity of about 2.4 tons in 24 hours per 1 horsepower.

panning gold, pockets are arranged. Each subsequent plate begins after the pocket 2″ lower than the preceding one. The pocket is a trough, 6″ wide and deep, measured from the edge of the upper plate, running across the entire width of the sluice. The pocket is partitioned by a board that does not reach 1—1½″ to its bottom, on which is laid an insert made from a board hollowed out to the shape of the pocket. See Fig. 1. The partition prevents the pocket from becoming clogged, while the wooden insert, by increasing the friction of the particles, promotes better catching of mercury. The pockets on one side have openings cut into the side wall of the sluice, closed with sealed hatches and serving to empty the pockets during rinsing. Each sluice ends in a trough, in the middle of which opens a pipe carrying the tailings into a common trough, along which they pass to Dorr’s classifier. The trap is a cast-iron vessel, in the bottom of which there is an opening closed by a plug, through which the accumulated mercury is drained from time to time. The pipe carrying the slimes from the sluices descends almost to the bottom of the trap, and on the side of the latter there is an overflow pipe. Thus a hydraulic seal is obtained, promoting the capture of mercury.

The total length of the sluices at a bowl is 21′—22′, the slope 32½—35½″, which amounts to a fall of 1.56″ per foot (12½%). According to data in the literature) a normal fall for plates is also considered to be 1—2½″ per foot, depending on a number of conditions*) and, among other things, on the character of the ore and the ratio of liquid to solid.

The area of the plates at one bowl at the Uspenskaya plant is equal to 21×4 = 96.6 sq. ft. If it is assumed that the capacity of the bowl is 40 poods per hour, or about 1000 poods (16.12 metric tons) in 24 hours, then the area of plates per 1 metric ton of ore crushed per day is about 6 sq. ft. (96.6 : 16.12).

Thomson (Thomson) indicates***) that in most—

* Thomson. Stamp. Milling and Cyaniding, 1915, p. 58.
* Del-Mar. Stamp. Milling, 1912, pp. 59 and 122.
** Thomson. Ibid., p. 58.

is in most cases sufficient: 3–4 sq. ft. per 1 ton in 24 hours. True, Thomson’s data refer to the practice of crushing in stamps, where the ratio of liquid to solid usually does not exceed 6–8 : 1. At the Uspenskaya factory, as was indicated, this ratio is close to 10 : 1. However, this circumstance alone does not justify such an expansion of the area of the plates, exceeding the normal by 1½–2 times, to which must also be added the “lower” sluice (49 sq. ft.) and the pulsating tables after the tube mills.

The reason for the attraction to sluices, observed at almost all Russian gold-washing factories, lies, in my opinion, in the use of internal amalgamation. Even in pans, not to mention stamps or fast-running Chilean mills, mercury, poured inside, breaks up into such fine droplets that they float over the sluices and are caught with great difficulty by the amalgamated plates, even when they come into contact with the latter; on the other hand, coarse, heavy gold does not create an unconditional necessity for internal amalgamation, since even such gold is carried out with the slimes from the stamps) or pan and is caught on the plates. Undoubtedly, in some cases the use of internal amalgamation increases the total quantity of gold caught by mercury, which is especially important when the tailings after amalgamation are not cyanided. Del Mar cites an example from Californian practice, where the use of internal amalgamation increased extraction from 40 to 80%*) and asserts that the question of the use of internal amalgamation cannot be resolved uniformly in all cases; it is necessary each time to investigate the given ore and experimentally establish whether internal amalgamation gives any additional extraction of gold for it. Depending on the results obtained, and only then, can the question of the method of amalgamation be decided, and consequently also of the type of apparatus for grinding the given ore. In those cases when the tailings of gold-washing—

) Thomson. Stamp Milling and Cyaniding, 1915, p. 57.
) Del Mar. Stamp Milling*, p. 60.

Fig. 1.

of the factory enter the cyanide plant, the gold not caught by the mercury is dissolved in potassium cyanide and precipitated in the extractors. True, in this process certain losses are unavoidable: first, because, as a general rule, only 90% of the gold that could be caught on the plates is obtained in the extractors, since part of it is lost in the solutions retained by the treated slimes or tailings; second, the enterprise loses % % on the gold that is more slowly realized under the chemical method of extraction. Nevertheless, it is doubtful that the advantages of more complete amalgamation would justify the considerable expenditure of mercury for internal amalgamation in those cases where this method yields some additional extraction.

At the Central mine, the consumption of mercury averages 10–15 zol. per 100 poods; the normal loss of mercury without internal amalgamation, according to Thomson’s data) is from 0.2 to 0.5 oz. per ton, which amounts to 2.2–5.4 zol. per 100 poods. Of course, the loss of mercury, apart from its pulverization, is also explained by a number of other causes, for example, by the presence of decomposed sulfides and the presence in significant quantities of As, Sb, Pb compounds, which produce “sickening” of the mercury)—depriving it of the ability to collect quickly into drops. In such cases the addition of lime considerably improves amalgamation. At the Central mine lime is not charged into the mortar (although in my opinion this would not be harmful), and the plates remain bright almost all the time—they do not turn green. This gives grounds to assert that the large loss of mercury is not explained by its “sickening.” That the use of internal amalgamation in the mortars is in fact the cause of the greater losses of mercury is clearly shown by the practice of the Troitsk factory at Kochkar, where, with the abolition of internal amalgamation, the loss of mercury decreased by 75%, while the quantity of gold caught on the plates did not change.**)

At gold-washing factories that employ internal amalgamation, the large loss of mercury gives rise to an effort to increase as much as possible the length

) Thomson, Stamp Milling, 1915, p. 64.
) Clennell, The Cyanide Handbook, 1910, p. 182.
**) According to private information.

sluices and to arrange an entire row of traps to catch it. However, as the figures for loss at the Uspenskaya mill show, this helps little. Increasing the area of the plates leads, first, likewise to a certain loss of mercury as a result of the friction of the slimes against the plates, and, secondly, it increases the total amount of gold on the plates and, consequently, the loss on it in %. The loss of mercury, a metal rather costly in itself, is not profitable in itself. But, in addition, the enrichment with mercury of the tailings entering the cyanide plant is the cause of difficulties in the operation of the extractors and of an excessive consumption of zinc. Mercury, similarly to gold, silver, and other metals, dissolves in potassium cyanide. Part of it is precipitated from the solutions in the form of HgS and is thus eliminated from the solution of sulphide alkalis, which, of course, is highly desirable. Another part, together with the solutions, enters the extractors and is deposited on the shavings, forming an amalgam with the zinc. As a result, the shavings become extremely brittle, break easily, and, clogging the extractor screen with fines, reduce the rate of circulation of the solutions. On the other hand, the consumption of zinc increases enormously, since, during rinsing, a large part of the shavings goes off into the precipitates.*) Perhaps this partly explains the comparatively high consumption of zinc at the Central plant for precipitating gold—0.62 English pounds per ton instead of the normal 0.125–0.5.**)

Classification. From the sluices, the slimes pass along launders to a Dorr-system classifier—duplex, of normal type and size. The classifier is driven by one of the transmissions serving the runners; power consumption is 1½ HP; the number of rake strokes is 14 per minute. In 24 hours the classifier passes 5,670 poods, or about 90 tons, of dry slimes (the average daily capacity of the mill). The maximum capacity of the classifier is 125 tons.

*) H. F. Bain, More Recent Cyanide Practice, 1910. Sec. edit., pp. 274–276.

**) Mac Farren, Text-Book of Cyanide Practice, p. 172.

The pulp entering the classifier with a liquid-to-solid ratio of 10:1 is separated into slimes and sands.

Slimes amount to about . . 50.4%
Sands „ „ . . 49.6%

The screen analysis of the slimes entering the classifier and of the resulting slimes and sands is given in Table 1.

Table 1.

Number of openings per 1″ I.M.M.	Screen analysis: slimes	Screen analysis: classifier product: slimes	Screen analysis: classifier product: sands
+60	13.09	0.57	25.78
+80	15.01	2.18	27.77
+100	10.35	4.38	16.27
+120	6.75	3.91	9.70
+150	6.95	5.75	8.17
+200	10.38	13.12	7.64
−200	36.75	69.00	4.45

The slimes from the classifier pass along launders to five Wilfley tables; the sands are carried by water to a lower sluice 21′ long and 2′4″ wide, and by launders are distributed among three Wilfley tables. At the place where the sands enter the Wilfley tables, on the latter there are arranged traps measuring 1′2″ × 2′4″, lined inside with copper sheets for catching mercury. This is the last of a series of devices serving in the mill for amalgamating gold and catching mercury.

From Table 2 it is evident what role the lower sluice and the traps at the Wilfley tables play in the overall recovery of gold. The former, with a sheet area of 49 sq. ft., catches in a month only 77 zolotniks of gold, which amounts, per 1 sq. ft. of sluice area per day, to 5.03 dolya of gold. The latter, each with 2.72 sq. ft., catch in a month 1 lb. 71 zol.,

Table 2.

Amalgam obtained	Amount of amalgam	Gold evaporated	Au content in amalgam, %	Gold extracted: gold per 100 poods	Gold extracted: %
from the bowls and from the sluice	7 p. 16 lb. 91 zol.	2 p. 28 lb. 88 zol.	36.67	6.14	79.5
from the lower sluice	— 3 „ 37 „	— — 77 „	—	—	0.55
from the Wilfley tables . .	— 5 „ 29 „	— 1 „ 71 „	24.95	0.15	1.21
from the ditches, etc.	— 2 „ 32 „	— — 20 „	—	—	0.14
from the silt plant	— 20 „ 38 „	— 4 „ 56 „	22.47	0.25	3.20
Total . .	7 „ 16 „ 91 „	2 „ 28 „ 88 „	36.67	6.49	84.6

gold, i.e., per 1 sq. ft. per day there are 65.3 dolya. From these figures it may be concluded that the benefit of the lower sluice is highly doubtful.

All 8 Wilfley tables, differing somewhat from one another in the dimensions and shape of the deck, are located considerably below the level on which the bowls and sluices are installed, and are driven from a common transmission brought from a pulley mounted on one of the transmissions serving the runners. The Wilfley tables make 230–240 pulsations per minute and a longitudinal displacement of 1″; the inclination of the tables and the water feed are regulated in such a way that part of the quartz passes into the concentrates, and not the reverse. The concentrates accumulate in boxes with which the Wilfley tables end; from there they are, from time to time, raked out onto a cart and taken under the shed by the factory. From each portion of them a sample of 1/2 pound is taken, and from the material accumulated over 2 days—5–6 poods—a laboratory sample is taken by reduction. The gold content in the concentrates is about 20–30 zol. per 100 poods. As the workings deepen, the amount of sulphides increases and at present fluctuates between 3 and 5%. As is known, gold-

that associated with sulfides is only with difficulty amenable to amalgamation and to the ordinary methods of cyanidation. Therefore, in those cases where the ores contain many concentrates rich in silver and gold, efforts are made, as far as possible, to separate them completely from the general mass of ore and subject them to special treatment.*) At Tsentralny about 10.7% of all the gold contained in the ore is bound up with the concentrates; it was therefore considered quite expedient to treat the concentrates separately. Preliminary classification of the slime into silts and tailings ensures the fullest possible separation of the concentrates. The latter are not yet being treated, since the method for extracting gold from them has not been fully developed. It is assumed that fine grinding of them in tube mills and prolonged agitation with a solution of potassium cyanide may be possible.

For an 8-hour shift at the plant the following are employed:

2 clerks, 3 feeders, 1 Wilfley-table operator,

in addition, 1 supervisor,

1 repairman (as needed), 2 oilers (12-hour shift) and 1 assistant boy.

The cost of treating 100 poods of ore at the plant, including haulage and crushing, is about 2 rubles 70 kopecks. Of this sum there falls:

Item	Share
haulage	29.85%
crushing	1.24%
grinding	20.37%
amalgamation (mercury)	7.77%
supervision and labor	26.70%
other expenses	14.07%

Table 3 presents data on the operation of the plant in April 1915 and for the first third of the year.

*) Penya, “Treatment of Concentrates,” Gold and Platinum, 1914, No. 21, p. 371.

Table

Date on which sluicing was carried out	Total number of working hours	Processed ore, poods	Mercury consumed, pounds	Amalgam obtained from trays and sluices	Evaporated gold
3	319	12700	14	19 f. 70 z.	7 f. 49 z.
5	339	13600	14	19 f. 15 z.	7 f. 02 z.
7	368	13700	15	17 f. 61 z.	6 f. 74 z.
9
10	502	17500	16	29 f. 74 z.	10 f. 86 z.
12	332½	10300	9	23 f. 05 z.	8 f. 72 z.
14	379½	11000	11	24 f. 67 z.	9 f. 44 z.
16	322	10000	12	17 f. 42 z.	6 f. 48 z.
18	329½	12000	10	16 f. 05 z.	5 f. 83 z.
20	322	10000	14	25 f. 13 z.	10 f. 04 z.
21
22	329½	10900	9	16 f. 34 z.	5 f. 82 z.
24	332	12000	10	24 f. 25 z.	8 f. 62 z.
25
26	369	13300	11	21 f. 72 z.	7 f. 13 z.
28	376	11600	10	23 f. 69 z.	8 f. 48 z.
30	371	11500	10	18 f. 29 z.	5 f. 93 z.
For April	4991½	170100	4 p. 05 f.	7 p. 16 f. 91 z.	2 p. 28 f. 83 z.
For 1/3 of a year	18150	688400	—	—	9 p. 33 f. 79 z.

Production of 1 runner in 24 hours: 816 poods.
Consumption of mercury: 9.31 zolotniks per 100 poods.

3.

Amalgam taken from other places	Gold evaporated	Laboratory assays: ores, Au	Laboratory assays: ores, Ag	Laboratory assays: concentrates, Au	Laboratory assays: concentrates, Ag	Quantity of concentrates, poods
From Wilfley 1 lb. 14 zol. From hydraulic plant 4 lb. 91 zol.	50 zol. 1 lb. 18 zol.	10.80 6.99 9.97 3.91	11.89 8.58 11.17 5.78	19.58	30.72	1720 (2.99%)
From lower sluice 92 zol. From hydraulic plant 8 lb. 76 zol.	24 zol. 2 lb. 0 zol.	11.17 10.17 5.15 4.45 15.61	12.12 6.07 6.70 8.92 12.42	26.12	34.56	2364 (4.43%)
From Wilfley 4 lb. 15 zol. From lower sluice 0 lb. 72 zol.	1 lb. 21 zol. 0 lb. 16 zol.	7.43	7.66	18.43	16.26	2484 (4.19%)
From lower sluice 0 lb. 67 zol. From hydraulic plant 3 lb. 09 zol. From canvas 1 lb. 32 zol.	0 lb. 15 zol. 0 lb. 67 zol. 0 lb. 20 zol.	11.79 6.45	7.93 8.39	18.43	16.26	2484 (4.19%)
From lower sluice 0 lb. 94 zol. From hydraulic plant 3 lb. 67 zol.	0 lb. 22 zol. 0 lb. 67 zol.	6.26 7.99	7.87 10.68	18.43	16.26	2484 (4.19%)
From Wilfley and lower sluice 11 lb. 02 zol. From hydraulic plant 20 lb. 38 zol.	2 lb. 72 zol. 4 lb. 56 zol.	7.72 —	— —			6568 (5.86%)

Rinsing. Rinsing at the factory is carried out every day. Its total duration is 2 hours, with two pairs working simultaneously. Before rinsing, feeding is stopped, 3–4 pounds of mercury are poured into the bowl (see Table 4, p. 19), and the bowl is allowed to run out, i.e., as far as possible all the material present in it is ground down. The protective screens of the plates are opened, and the plates are rubbed liberally with mercury to prevent the amalgam from being carried away during rinsing. The screen is removed, and the contents of the bowl are swept with brooms onto the sluices. So that rubbish and sands do not clog the latter, a screen is placed obliquely at the second plate, on which all the rubbish is caught. From time to time water is let in and the runners are set in motion. When the bowl is clean, and the amalgam that has accumulated in it has been washed by means of brooms and a stream of water onto the first plate, iron scrapers are used to clean off the amalgam adhering here and there to the bowl. At this time two attendants, with oars, stir up the contents of the pockets, and the greater part of the slime that has accumulated in them is carried away. Through the side hatches the remaining slime and sands are washed into a bucket and carried to the adjoining bowl. The contents of the pockets from the last sluice are brought to a small forehearth. Then the screen with the rubbish is shaken out, removed, and they begin to clean the amalgam from the plates with rubber scrapers, starting from the lower one. The amalgam from all the sluices is collected in a sealed vessel and, at the end of rinsing, is carried to the ladle, diluted with mercury, and washed from the sands with hot water; the iron is separated by a magnet. The sands and dirt are ground in a separate tub with a pestle, with water and mercury, and the amalgam obtained is added to the common mass. The washed amalgam is squeezed by means of a hand press of rather primitive construction, but one that works quite satisfactorily. The squeezed amalgam is weighed, wrapped in paper, sealed, and handed over to the cashier until roasting (evaporation). (Table 4, p. 19, shows the order in which mercury is poured into the bowls and onto the plates during the interval of time between two rinsings.)

The evaporation of the amalgam is usually carried out after rinsing (every 4 days). The amalgam is received from the cashier, weighed out into portions according to the content—

Table 4.

Date	Shift	Time of pouring	Poured mercury, lb: into bowl No. 1	Poured mercury, lb: onto sluice	Poured mercury, lb: into bowl No. 2	Poured mercury, lb: onto sluice	Poured mercury, lb: per shift
4	I	9 a.m.	1/4	—	1/4	—
4	I	12 noon	1/4	—	1/4	—	1
4	for rinsing		3 3/4	3 1/4	3 3/4	3 1/4	14
4	after rinsing		—	1 1/4	—	1 1/4
4	II	6 p.m.	1/4	—	1/4	—
4	II	9 p.m.	1/4	—	1/4	—	3 1/2
4	III	12 midnight	1/4	—	1/4	—
5	III	3 a.m.	1/4	—	1/4	—
5	III	6 a.m.	1/4	—	1/4	—	1 1/2
5	I	9 a.m.	1/4	—	1/4	—
5	I	12 noon	1/4	—	1/4	—	1
5	II	3 p.m.	1/4	2	1/4	2
5	II	6 p.m.	1/4	—	1/4	—
5	II	9 p.m.	1/4	—	1/4	—	5 1/2
5	III	12 midnight	1/4	—	1/4	—
6	III	3 a.m.	1/4	—	1/4	—
6	III	6 a.m.	1/4	—	1/4	—	1 1/2
6	I	9 a.m.	1/4	—	1/4	—
6	I	12 noon	1	—	1	—	2 1/2
6	for rinsing		4 1/2	3 1/4	4 1/2	3 1/4	15

gold in the ore and its quantity, processed from the given section of mining operations. The weighed portions of amalgam are wrapped in several layers of paper and the rolls are placed one upon another in iron crucibles, smeared inside with lime, according to the local crucible. The order of placement is carefully recorded, as shown in Table 5.

Table 5.

Crucible Nos.	Section of mining operations	Ore processed, poods	Amalgam obtained	Retorted ligature Au
I	Blagoveshch. shaft of the Lotter mine	4000	15 lb. 48 zol.	5 lb. 17 zol.
II	Voskresen. shaft, Central mine	3600	12 lb. 68 zol.	4 lb. 19 zol.
”	Pokrovsk. shaft, Central mine	4900	4 lb. 00 zol.	1 lb. 25 zol.
III	Chirkovsk. shaft of the Chirkovsk. placer	6800	5 lb. 56 zol.	1 lb. 96 zol.
”	Ore sample from the Chirkovsk. mine	1100	— 77 zol.	— 23 zol.
”	Lower sluice of the New factory		— 56 zol.	— 12 zol.
”	Lower sluice of the Old factory		1 lb. 06 zol.	— 23 zol.
IV	Prospector Petr Bushin	1200	5 lb. 20 zol.	1 lb. 76 zol.
”	Prospector Stepan Pleshkov	1000	3 lb. 18 zol.	1 lb. 23 1/2 zol.

Then the crucibles, in order, are inserted one upon another into a vertical retort, set into an ordinary furnace. The retort is closed with a lid with a helmet-shaped

by a tube that carries the mercury vapors into a vessel with water, where the mercury is collected and then poured into a jar; the quantity of it driven off from the retort is measured by volume. Usually the pressed amalgam contains about 30% fine gold (see Table 2, p. 14). The distillation lasts 3/4–1 hour. When the distillation of the mercury has ceased, the retort is opened, the crucibles are taken out, and the gold contained in them is overturned into clean, dry scoops. After burning, the paper in which the amalgam was wrapped forms a layer of ash that prevents one part of the gold from sticking to another. The weighed portions are laid out again, the evaporated gold on iron sheets; in the order of the record they are carefully cleaned of ash with a brush and weighed to an accuracy of 1/4 zol. A statement of the distillation is drawn up, and the gold is handed over to the cashier, who personally checks its weight.

When about 1 pood of gold has accumulated, it is smelted in graphite crucibles, with an addition of borax and small quantities of saltpeter, and cast into molds greased with petroleum. After cooling, the ingots are cleaned with a brush, a hammer, and, if necessary, a chisel. From each ingot two samples are taken with a drill, it is weighed, and it is marked with the initial letters of the name of the mine from which the gold was obtained, and the weight of the ingot is stamped on it. A statement is drawn up, as shown in Table 6 (p. 22), in which the weight of the gold before and after smelting is noted, as well as its loss, which usually does not exceed 1 zol. per 1 f.

When 1–2 poods of gold have accumulated, it is sent with the cashier to Tisul and delivered to the post office. Assaying of the ingot is carried out in such a way that from each of the two samples taken, two weighed portions of 0.1 gr. each are taken. The two weighed portions are cupelled without addition of silver (fineness assay). Thus from each ingot 4 separate assays are obtained, which should be sufficiently close to one another. Usually the ingots are found to be of 600–750 fineness.

Table 6.

Crucible	Gold from the placer	Weight of the evaporated gold	Weight of ingots	Mark
I	Chirkovskii . . . .	7 lb. 12 zol.
I	Kharakternyi . . .	14 ”
I	Verkhoratskii . . .	34 ” 48 dol.
I	Nezavidnyi . . .	3 ” 04 ” 48 ”
I	Sora . . . . . .	12 ”
I	Sum . .	10 ” 77 ”	10 lb. 72 zol. 72 dol.	Ch. N.
II	Lotereinyi . . . .	10 ” 24 ”
II	Goristyi . . . . .	4 ” 48 ” 72 ”
II	Sum . .	14 ” 72 ” 72 ”	14 lb. 68 zol.	L. T.
III	Lotereinyi . . . .	7 ” 06 ” 72 ”
III	Tsentralnyi . . .	6 ” 11 ”
III	Sum . .	13 ” 17 ” 72 ”	13 lb. 10 zol. 48 dol.	Ts. L.
IV	Tsentralnyi . . . .	13 ” 29 ” 48 ”	13 lb. 20 zol. 48 dol.	Ts.
IV	Total . .	1 pood 12 lb. 05 zol.	1 pood 11 lb. 75 zol. 72 dol.
IV	Loss by burning . . . . .	25 zol. 24 dol.

Slime Plant.

The slime plant is attached to the Uspenskaya gold-washing factory and is intended for the treatment of slimes by the complete slime process (all sliming process).

The slime and tailings from Wilfley tables flow by launders into a common sump, from which, by a 6-inch centrifugal pump coupled to a 25 HP electric motor, they are fed to two Dorr classifiers for secondary separation. It is difficult to say what guided the builder of the plant in introducing double classification of one and the same pulp. Classification, to be sure, is not expensive, since one classifier consumes only \(1 \tfrac{1}{2}\) HP, making 12 strokes per minute and at maximum capacity

Table 7.

From	To	Tube-mill slimes contain moisture, %	Tube-mill slimes contain slimes, %	Tube-mill slimes contain tailings, %	Classifier tailings contain moisture, %	Classifier tailings contain slimes, %	Classifier tailings contain tailings, %
8 XII	1 I	21.7	54.7	43.3	21.1	9.1	90.9
1 I	10 I	19.5	54.4	45.6	22.08	7.3	92.7
10 I	20 I	29.9	57.3	42.7	27.9	11.1	88.9
20 I	1 II	28.8	61.3	38.7	27.0	9.3	90.7
1 II	10 II	24.0	58.4	41.6	25.7	9.9	90.1
10 II	20 II	25.7	58.4	41.6	25.0	7.9	92.1
20 II	1 III	27.1	55.3	44.7	24.0	5.5	94.5
1 III	10 III	25.5	56.1	43.9	25.9	5.6	94.4
10 III	20 III	25.5	56.9	43.1	25.1	5.1	94.9

125 tons of dry slimes per day. The power consumption per ton in 24 hours does not exceed 0.01 to 0.02 horsepower.*)

The classifiers are installed on the floor above the tube mills and are driven by a common transmission from a 3-horsepower electric motor.

The lift height of the pulp is about 27′; the ratio of liquid to solid is 15 : 1; the slime content is about 40% (see Table 1, p. 13).

In Table 7 (p. 23) are presented certain figures characterizing the operation of the classifiers and tube mills.

The tailings are washed from the classifiers by water into launders and thence, through pipes, enter small sumps, from which the material is automatically directed by a Smith feeder (Shmitt) through the hollow trunnion into the tube mills.

There are two tube mills of normal dimensions, 22′ × 5′.**) They are driven by a common 150 HP motor, which operates the transmission. From the latter, power is transmitted to pulleys coupled with gears engaging a toothed wheel, which, by means of a frictional wooden ring, transmits motion to the tube mill. The number of revolutions of the latter is 27 per minute. The lining of the mill consists of armor plates, “tonopah” (tonopah), placed inside it in belts consisting of three parts. The width of such a belt is 1′. As crushing material, from 150 to 200 poods per day of coarse river pebbles are fed into the mill. An attempt was made to use vein quartz, but it proved too brittle and rounded off poorly. The pebble is fed little by little by hand into the hollow trunnion, through

*) Metall. and. Chem Engineer. 1916 Vol. XIV p. 93.

**) At the present time there is a tendency to use shorter tube mills. See Megraw, Details of Cyanide practice, 1914, pp. 105, 132, 163.

through which the pulp flows out of the mill. This trunnion is provided with a left-hand internal helical surface, along which the pebble is transported into the mill. More often, the crushing material is charged by means of the same feeder with which the pulp is scooped up.*)

To neutralize the pulp and to improve the settling of the slimes in the thickeners, 25 poods of lime per day are added to the tube mills. The lime is introduced in the form of milk of lime, which from time to time, one bucket every 15 minutes, is poured into the sumps of the feeders.

The slime plant is designed to treat 200 tons (12,400 poods) of dry slimes per day and does not correspond to the mill, whose capacity is only about 90 tons (5,500 poods). Therefore only 1 classifier and 1 tube mill operate at the plant. Since the operation of the latter is rather expensive, efforts are made, as far as possible, to relieve the tube mill by separating in the classifier, for the slime process, up to 60% of the total quantity of tailings entering the plant. With such operation the slime contains about 10–15% of product +200, which is hardly correct, since the presence of coarse particles increases the time required for the dissolution of the gold and, consequently, the duration of agitation (see below).

In all, the tube mill grinds per day about 35 tons (40% of 90), or 2,200 poods, of tailings and 3.2 tons (200 poods) of river pebble, which amounts, per 1 ton of ground product, to 184 English lb., or 9.2%, and per 1 ton of tailings treated at the plant to 72 English lb. (3.6%). Such a considerable consumption of pebble indicates its low suitability as a crushing material. On the Rand and in other places, coarse lumpy ore is used as the grinding material (pebble), and is consumed in an amount of 2½%.**) When, for this purpose, material not containing

) A Text-Book of Rand Metallurgical Practice*, 1913, Vol. I, pp. 113–116.

) Rand Practice, Vol. I, p. 116.

of gold, usually imported Greenland or Danish pebble, its consumption ranges between 1 and 4 pounds per ton. Megraw already finds the consumption of imported pebble, at 5 pounds per ton, to be excessively large.*)

A tube mill requires about 65–70 HP, and its output per 1 horsepower in 24 hours is equal to 0.54 ton, or 33.5 poods, of material containing about 90% slimes.

At the Antonov factory in Kochkar, tube mills almost identical with those installed at the Central factory grind about 100 tons per day, or 50 tons each.) With a consumption of 60 horsepower, this amounts to 0.83 ton per 1 horsepower in 24 hours. Finally, one of the tube mills, for which data on the operation are given by Hofman*) and the operating conditions of which are comparable with those of the tube mill at the Central factory, with a consumption of 47 HP grinds 97 tons, or about 2 tons per 1 horsepower in 24 hours.

Of course, in comparing the performance of grinding apparatus one must take into account a whole series of factors that are almost impossible to express by numerical ratios, and the comparison of the work of tube mills given above is more qualitative than quantitative. Nevertheless, on the basis of these data it may be asserted that the output of the tube mill at the slime plant is not large, and that its operation is disadvantageous from an economic standpoint.

One of the reasons for such low output of the tube mill, in my opinion, lies in the low moisture content of the material being ground. From Table 7 on p. 23 it is evident that the tailings from the tube mill contain on average:

25.5% moisture,
58.1% slimes,
and 42.9% sands,

) Megraw, Details of Cyanide Practice, p. 190.
) According to private information.
) Hofman, General Metallurgy, p. 600.

whereas, according to the literature,* the normal moisture content in the pulp of a tube mill is considered to be from 38 to 50%; and, although for each ore there is a certain definite percentage of moisture at which the operation of the tube mill is most advantageous, this percentage does not go beyond the indicated limits.

Upon leaving the tube mills, the pulp is fed to pulsating tables, one for each; length 5 arshins and width 2 arshins; inclination \(1/10\); number of pulsations 210 per minute. On the tables the pulp is somewhat diluted with water. Every two days the amalgam is removed and the plates are dressed with mercury, for which from \(3/4\) to 1 pound is used. In one removal, usually 56–86 zolotniks of amalgam with 20–22% ligature gold are taken from both tables (see Table 2, p. 14).

The pulsating tables are set in motion by 2 separate small transmissions, operated by motors of 3 HP each. From these same transmissions operate 2 Frenier pumps, installed for lifting the pulp from the corresponding tables back to the classifiers. The diameter of the pump wheel is 54″, width 8″, number of revolutions 12, height of pulp lift 22′.

The lower product of the classifier flows by gravity to the Dorr thickeners.

There are two thickeners. The dimensions of the tanks are \(32' \times 10'\). The speed of rotation of the agitator is 1 revolution in \(6 \tfrac{1}{2}\) minutes. Each thickener is served by a separate motor of \(1 \tfrac{1}{2}\) HP, which, through a transmission and worm gearing, sets in rotation a gear wheel mounted on the agitator shaft. If the resistance of the pulp, owing to excessive thickening of the slimes, exceeds a certain value, the agitator may break; to prevent breakage, buffers are provided, whose springs at the dangerous moment disconnect the gear wheel from the vertical shaft.

* Bernewitz, Cyanide Practice 1910 to 1913, p. 196.

The pulp enters classifiers through pipes lowered near the center of the vats, several inches below the pulp level. The ratio of liquid to solid is about 15 : 1 in the incoming pulp and 1 : 1 in the outgoing pulp. Owing to the presence of two thickeners, the thickening process proceeds very satisfactorily, and the overflow water from the thickeners contains only about 0.02% slimes. Table 8 gives figures characterizing the operation of the thickeners.

Table 8.

Period	Period	The overflow water of the thickener contains	The overflow water of the thickener contains	Pulp from the thickener contains	Pulp from the thickener contains	Pulp from the thickener contains
from	to	moisture, %	slimes, %	slimes, %	gold in 100 poods	silver in 100 poods
1—I	10—I	99.980	0.020	41.9	3.04	6.50
10—I	20—I	—	—	44.1	2.46	4.65
20—I	1—II	99.984	0.016	44.5	2.65	5.40
1—II	10—II	99.981	0.019	47.7	2.04	4.51
10—II	20—II	99.977	0.023	44.9	2.41	4.37
20—II	1—III	99.972	0.028	43.8	2.28	3.37
1—III	10—III	99.977	0.023	43.0	2.02	3.28
10—III	20—III	99.975	0.025	41.5	2.26	3.14

The pulp, thickened to a ratio of 1 : 1, continuously flows off from below through a 4″ pipe into a collector tank 30′ × 10′. In this tank an agitator is installed to prevent stratification of the pulp. The agitator makes 7 revolutions per minute and is driven by a 7.5 HP motor. Above the agitator there is a small iron dissolving tank (37′ × 3′5″), in which a solution of potassium cyanide is prepared; it is fed, as required, into the agitator and mixed with the pulp. This is the first point of intro—

tion in the process of potassium cyanide. Before adding KCN to the pulp, caustic soda is added. For the strength of the solutions see Table 9.

Table 9.

During the period	During the period	Vats treated	Charge: poods of slime in 1 vat	In the charge is contained: slime, %	In the charge is contained: tailings, %	After agitation the solutions contain: Au, in 100 p.	After agitation the solutions contain: Ag, in 100 p.	After agitation the solutions contain: KCN, %	After agitation the solutions contain: NaOH, %
from	to
8 XII	1 I	8	3900	40,0	13,0	2.00	3.00	0,029	0.010
1 I	10 I	7	4038	40,5	10,2	1.89	3.76	0,020	0.005
10 I	20 I	10	4187	41,7	10,5	1.79	3.49	0,022	0.008
20 I	1 II	10	4333	42,7	11,0	1.73	2.58	0,020	0.008
1 II	10 II	9	4198	41,6	11,1	2.05	3.09	0,019	0.010
10 II	20 II	15	4563	44,5	10,3	1.39	2.86	0,020	0.011
20 II	1 III	13	4606	44,6	9,6	1.17	2.63	0,020	0.013
1 III	10 III	16	4358	43,5	10.0	1.16	2.39	0,022	0.009
10 III	20 III	18	4170	41,5	10,3	1.07	2.10	0,019	0.007

Into the freed Dorr agitator, of which three are installed at the plant, the pulp from the thickener is raised by a 5″ centrifugal pump (capacity 5,000 cubic feet per minute; electric motor of 20 HP). The height to which the pulp is lifted is about 20′.

The Dorr agitator is a wooden vat with a capacity of about 4,100 poods of dry slime, measuring 20′ × 15′, in the center of which is suspended a 6″ iron pipe, not quite reaching the bottom of the vat. The pipe is rotated by means of gearing and a transmission common to all three agitators, driven by a 5-horsepower electric motor. At the lower end of the pipe

two titans are fastened with the vanes attached to them, which rake the settled silt toward the center of the vat, where it is entrained by air entering from below through a branch pipe opening into the central pipe. Rising upward, the pulp overflows over the edge of the pipe and is broken up against the two chutes fastened to its upper end, by which it is distributed over the entire surface of the vat. The number of revolutions of the central pipe is 4 per minute.

Compressed air is supplied under a pressure of 12–15 lb. through a 3″ pipe from the compressor and enters the vat through three 1″ branch pipes. One of them, mentioned above, opens into the central pipe; the other two serve to stir up the silt that has settled around the pipe.

The compressor of the Sullivan works, at 180 revolutions, delivers 200 cu. ft. of air per minute under a pressure of 12–15 pounds and is driven by a 30 HP electric motor. Between the compressor and the air main there is a receiver (3′×6′) with the appropriate fittings.

The duration of agitation is about 30 hours. Filling and emptying of the agitator is carried out through the same 7″ pipe, which opens near the center of the vat, and takes 2 hours each. The 7″ pipe is provided with suitable tees with quick-closing slide gates. The figures in Table 9 (p. 29) characterize the operation of the agitators.

The Dorr agitator, one of the newest devices proposed for the treatment of slimes, is, as it were, a combination of mechanical and air agitators, and combines the advantages of these two types. In the literature there are very favorable reviews of this agitator,* which are also confirmed by practice at the Central Works, where they are quite satisfied with its performance.

*) Thomson, Stamp Milling and Cyaniding, p. 164.
Rybatkin, “Gold Slime and Its Treatment,” Gold and Platinum, 1914, No. 4, p. 90.

At the end of agitation the pulp flows by gravity into a second thickener, identical with the first, which serves as an intermediate reservoir, or pulp storage tank, between the agitators and the Butters vacuum filter.

78 filtering frames are suspended in an iron rectangular riveted tank \(8' \times 38'\), the bottom of which consists of pyramids ending in pipes through which the pulp enters the tank and through which it is emptied. For lifting the pulp and the wash solutions, a 10″ centrifugal pump with a 50 HP electric motor is installed. The lift to the vacuum filters is about 20′.

The frames are connected by means of a pipeline with a vacuum pump, driven by a 20-horsepower electric motor. At 40 revolutions per minute the pump gives a vacuum of 22″ of water column.

The cakes are formed within 25–40 minutes, after which the remaining pulp is discharged back into the pulp storage tank, and the cakes are washed with solutions that have already passed through the extractors. Water is not used for washing, since even without it a large part of the solutions after the extractors has to be discharged outside. The duration of washing the cakes is \(1 \tfrac{1}{2}\) hours. The washed cakes are loosened by means of a water pressure of 10 ft., which is created by the water reservoir at the mill. Discharging the cakes takes only 10–15 minutes. The total duration of the operation is 4 hours, i.e. 5–6 operations can be performed in a day. With the plant’s present capacity of 90 tons, one operation accounts for 15–18 tons, and one frame in 24 hours—for about 1.2 tons, as against the normal 2 tons.* Table 10 (p. 32) gives the gold and silver content in the discharged cakes.

* H. F. Julian and E. Smart, Cyaniding Gold and Silver Ores, 1910, p. 249.

The solutions from the vacuum pump pass through 7″ pipes into two clarifying vats, 14′ × 15′. At the bottom of the vats there is a filter of canvas, talc, and sand, laid in successive order on a grating. From the vats the completely clear solutions are conveyed through 4″ pipes to 6 extractors.

The extractors are of the ordinary type, wooden, with 10 compartments. The total quantity of shavings charged into them is 70 poods. Table 10 gives the strength of the solutions leaving the extractors and their content of gold and silver.

Table 10.

During the time from	During the time to	KCN %	NaOH %	Au. in 100 p.	Ag. in 100 p.	Au. in 100 p.	Ag. in 100 p.
1—I	10—I	0.018	0.011	0.025	1.02	0,40	2,83
10—I	20—I	—	—	0.040	1.31	0,30	2,27
20—I	1—II	0.017	0.019	0.080	1.14	0,30	3,40
1—II	10—II	0.018	0.017	0.060	1.30	0,36	2,65
10—II	20—II	0.018	0.015	0.040	1.23	0,31	2,83
20—II	1—III	0.018	0.013	0.080	1.41	0,23	2,20
1—III	10—III	0.021	0.016	0.070	1.28	0,21	1,42
10—III	20—III	0.021	0.016	0.030	1.40	0,24	2,10

Passing through the extractors, the solutions partly enter three vats for solutions, 30′ × 10′; the excess is discharged outside.

Once every two months a washing is arranged. The flow of solution through the extractor is stopped; the zinc shavings are lifted on grates and, when the solution has drained off, they are rubbed by hand through sieves. The part remaining on the sieve is returned to the extractor. With a Brantspoit the fines and sludge that have settled in it are washed out of the extractor; this material runs along troughs into a small Sumpf, where 4 small filter frames, one-quarter of normal size, are installed. By means of a small vacuum pump the sludge and fines are filtered off and added to the precipitate washed from the shavings. For further treatment the precipitates ought to be roasted, which is expected in the near future; at present, however, the metallurgical works is not yet fully equipped, and the residues are treated in the raw state with sulfuric acid in two dissolving vats, \(7 \times 6'\), installed on a platform in the extraction room. Compressed air and steam are passed in to stir the contents.

After dissolution is complete, the precipitates are carefully decanted, washed with water, and after drying are melted in a furnace fired with oil, in graphite crucibles with fireclay inserts.

About 150 poods of raw precipitates are obtained per washing, and their quantity depends mainly on the desired completeness of extraction of gold from the extractors, the quality of the zinc shavings, the composition of the solutions, the skill of the workers, and other factors.

After treatment with acid, there remain from 12 to 20 poods of material containing about 3% gold and the same amount of silver. For obtaining and smelting 1 pood of dried residues, the following are consumed:

Item	Consumption
sulfuric acid	2 p. 36, 20 f.
borax	29, 05 f.
saltpeter	4, 17 f.
soda	9, 20 f.
oil	2 p. 31, 00 f.
graphite crucibles	19.4 marks

The treatment and smelting of 1 pood of precipitate costs about 25 rubles.

The ingots obtained contain about 400 fineness of pure gold and 350 of silver. The slags from gold smelting are remelted again, and then are sent in one of the tubs to the factory.

For an 8-hour shift at the silt plant the following are employed:

1 attendant,
1 lime charger,
1 pebble hauler,
1 vacuum operator
and 2 oilers.

At present the plant is operating at almost half of its normal capacity and treats 90 tons (5,000 poods) of dry slimes per day, with an average content of 2.7 zolotniks of gold per 100 poods. At the plant the extraction is about 85%; the total extraction, however (with amalgamation), is 96%. In unloading cakes about 0.4 zolotniks are lost (14.8%); with the solutions, about 3.7% of all the gold contained in the slimes is also lost.

For the treatment of 100 poods of dry slimes the following is consumed:

Reagent	Amount
potassium cyanide	2.50 f.
caustic soda	1.28 “
lime	6.50 “
zinc	1.82 “

The expenditure of energy per 100 poods of slimes is 56.51 K.W.-hours. Of this energy the following is used:

Use	Share
for tube mills	42.0%
for the other motors	56.3 “
for lighting	1.7 “

If the total cost of treating the slimes is taken as 100, the individual expense items are distributed as follows:

Expense item	Value
Reagents	28.76
Potassium cyanide	18.84
Sodium hydroxide	1.82
Lime	2.65
Zinc	5.45
Materials for smelting the precipitates	5.12
Energy	37.05
Tube mills	15.54
Other motors	20.99
Lighting	0.52
Heating	5.94
Workers	16.36
Supervision	5.12
Other expenses	1.65

General Layout

When considering the general layout of the mill and plant, the following weaknesses become apparent.

1. The pulp entering the plant is raised three times by pumps (see diagram II), and of the total number of horsepower engaged at the plant (327 HP), about 29% is spent on this work. This indicates poor utilization of the force of the pulp’s free fall. Meanwhile, the terrain fully permitted arranging the mill and plant in such a way as to reduce to a minimum the energy expenditure for pumping the pulp. This probably happened because the slime plant was attached to an already constructed mill.

2. A large part of the solutions, after the extractors at the Central, is discharged outside, and only a small part of them is returned to the process. Thus

about 40% of the potassium cyanide consumed at the plant is lost unproductively. This could have been avoided if crushing were carried out not in water, as is accepted at the Central plant, but in solutions. On the other hand, crushing in solutions is associated with certain difficulties) and this method takes poorly to the practice of the gold business*).

1. In choosing a method for treating a given ore, the decisive factor is the profit that the enterprise can obtain, i.e., in the words of Caldecott, “metallurgy is only a means of making money from ores” (Metallurgy is the art of making money out of ores)). By no means is the most technically perfect process always the most advantageous. In this respect the cyanide process, providing the metallurgist with great latitude in the choice of methods of operation, gives him the possibility, for each ore, of selecting the most favorable, most economical methods of treatment. In this the metallurgist’s task is greatly facilitated by the fact that the entire cyanide process, with almost all its modifications, can be carried out on a laboratory scale and, on the basis of the data from laboratory study of the ore and its properties, the question of the method of its treatment can be decided. It is not known to me from what considerations they decided at the Central plant to apply the full slime process for treating the ore. Although Bosqui does find that the slime process is cheaper and therefore more advantageous than the percolation process,) this opinion can be disputed. Caldecott, in the discussion of Bosqui’s article, points out,****) that at the present time there is a tendency to send to the percolation vats even a product consisting of the finest crystalline particles, leaving for the slime process only the amorphous, colloidal portions of the slimes. Crystalline

) Mac Farren, Text-Book of Cyanide Practice, p. 170.
) Bull. Amer. Inst. of Min. Engen. 1915, p. 2453,
) Ibid, 1915, 2439.
) Bull. Amer. Inst. of Min. Engen.* 1915, p. 1024,
**) Ibid. 1915, p. 2443.

particles, even of minimal size, do not impede good percolation, whereas their presence in slimes increases the wear of the pumps and slows the treatment of the slimes, increasing the time required for the complete dissolution of the gold. It may be for this reason that at the Central mine, where the main mass of the slimes consists of crystalline particles and a considerable quantity (10–15%) of coarse product is sent into the slime process, such prolonged agitation is required in order to ensure sufficient extraction.

Modern practice in gold-mining in other countries, where stamps are used chiefly for preliminary crushing, seeks, as far as possible, to increase the productivity of the latter by increasing the weight of the stamps, introducing coarser screens, and changing the shape of the stamp shoe itself. All this has led to obtaining material whose particles are not small enough for the gold to be extracted from them in percolators. Hence the necessity arose of introducing an additional apparatus for reducing the size of the particles to the required dimensions, at which extraction is obtained sufficiently completely. Such an apparatus is the tube mill, whose task, from this point of view, consists not at all in converting all the ore into slime, but in reducing the size of individual grains to a certain maximum value at which they readily give up the gold contained in them. In this process, of course, the conversion of part of the ore into slime is inevitable, and this slime is also treated at the slime plant. Only in certain cases is the gold contained in the ores so difficult to subject to the action of potassium cyanide that it is necessary to grind all the ore to slime and to employ the full slime process.

At the Central mine, at the summer tailings plant, old dumps are being reworked and quite satisfactory extraction is obtained. It is difficult to suppose that the character of the ore now being mined at the mine has changed substantially in comparison with former years. From prolonged exposure to the air

even if one treated the tailings poorly, could also change little. Only the sulfides, with which about 20% of all the gold contained in the ore is associated, could oxidize and increase the consumption of alkali and potassium cyanide during treatment. One may suppose that the tailings obtained anew would give quite satisfactory extraction under simple percolation, if the amorphous silts were separated from the crystalline components for the silt process. Therefore it seems to me that, in the present case, the properties of the ore and of the gold contained in it do not create an unconditional necessity for the complete silt process—a process very perfect from the technical side, but undoubtedly less advantageous, especially under the conditions of taiga placer work, than simple percolation.

1. Finally, if it had already been decided to treat the tailings by the complete silt process, then one should choose, from all the variants of this process, the least expensive; vacuum filtration cannot at present be considered such.

In his review of the practice of gold and silver milling for 1914, Fox says*) that recently an altered decantation process, the so-called counter current decantation process, has been acquiring ever greater importance; it has proved considerably more advantageous than the widely used method of vacuum filtration. One of the chief reasons that contributed to the spread of this new method was the exorbitant royalties that had to be paid to firms manufacturing various patented filtering devices. Thus, for example, one Butters filtering frame costs 300 rubles to purchase, whereas its real value is 40–50 rubles.

A number of plants have already gone over to this new method, at which vacuum filters had previously been used, in some cases left in place as equipment for dewatering the treated silts, with the aim of saving lab—

) Fox, Gold and Silver Milling in 1914, Miner. Industr.*, 1915, Vol. XXIII, p. 339.

prepare KCN. At Tsentralny, it might also be more advantageous to alter the treatment scheme currently adopted, applying continuous agitation in three Dorr agitators, followed by washing and decantation of the slimes in a series of thickeners operating in succession.

Concluding the description of the warm slime plant at Tsentralny, it should be noted that this plant is still very young and, of course, some defects in its operation are inevitable at first. In general, however, the plant stands at the level of modern gold-mining technology.

The Old Factory.

The old factory, built by Ivanitsky in 1904, at present plays a secondary role in the overall work of the mine. The factory has space for four pairs of runners, but only two have been preserved and are in operation. The runners do not differ in any essential way from those at the new Uspensky factory. They are driven by a 32 HP electric motor. Formerly, a steam engine and a Shukhov-system boiler served this purpose—an installation that has been preserved to the present time.

Ore from the bins is loaded by hand into wheelbarrows, with the larger pieces being broken up with hammers, and is brought to the bowls, where it is likewise fed by hand with shovels. The sluices, widening from 1 arshin to 3, have a total length of 11 arshins and an area of 24 square arshins.

The slimes are carried from the factory through troughs to a system of six zumps, hewn from slabs and arranged in two rows; the capacity of each zump is \(1/2\) cubic sazhen; dimensions: width \(2^{1/2}\) arshins, length \(3^{1/2}\) arshins, depth \(5/4\) arshin. One row is filled with tailings in about two days. The tailings, after a row has been filled, are unloaded onto wheelbarrows and hauled to the dump, while the slime settles in the cuts. The cleaning of the zumps is carried out on average every 70 days.

The mill at the present time processes almost exclusively prospectors’ ore. Its total capacity is about 1,000 poods per day. The amalgam is usually removed when the grinding of the ore of a given prospector has been completed. The amalgam from the lower plates goes to the benefit of the Company.

The mill is serviced by Company workers, and for a 12-hour shift the following are employed:

1 clerk,

1 ore feeder,

2 wheelbarrow men,

and 1 oiler.

The processing of 100 poods of ore at the Old Mill costs the prospectors about 36–37 kopecks; moreover, they are obliged to deliver the gold to the Company office at the established price. The loss of mercury at the mill is about 15 zolotniks per 100 poods of ore.

Summer plants.

For the reworking, chiefly, of old dumps, the mine has summer tailings and silt plants.

The tailings plant consists of 7 wooden percolation vats, 32′ × 7′. The vats are set on supports of rubble masonry, between which, beneath all the vats, runs a double rail track, serving to remove to the dump the reworked tailings discharged through four hatches in the bottom of the vat.

Above the vats is arranged a platform of planking, laid with intervals of 3″–4″ along the corresponding stringers. Along this platform there is also laid a double rail track, by which, from the old dumps, the tailings subject to reworking are delivered in mine cars. The cars are pushed up the incline by workers and tipped into the appropriate vat through the grating formed by the planking. In this process the compacted material [continues]

it is somewhat loosened, and the workers level it over the bottom of the vat without particular difficulty. During the charging of the tailings, lime is charged in layers in the amount of 60 poods per vat.

The capacity of each vat is 9,000 poods (145 tons) of dry material (120 mine cars of 75 poods each).

Charging a vat takes 10–12 hours; unloading takes the same amount of time. These operations are usually given to a contractor, who receives 20 rubles for charging and unloading one vat. When charging a vat, 10 workers and 2 horses are used; when unloading, 8 workers and 2 horses. A sample is taken from each mine car and sent to the laboratory for testing, after appropriate reduction.

The total duration of percolation is 6–7 days, i.e., the plant’s capacity is about 9,000 poods (145 tons) per day. The order in which the solutions are poured, their strength, and their quantities are given in the following table.

Table II.

Solutions	Quantity, in tons	Quantity, as % of weight of charge	Content in solution: KCN, %	Content in solution: NaOH, %
Strong	45	31	0.10	0.14
Ventilation for 6 hours
Medium	45	31	0.06	0.08
Weak	45	31	0.034	0.02

Usually at the plants they use, by weight of dry charge, \(1/3\) of strong solution, \(1/6\) of medium solution, and \(1/6\) of weak solution or water.*)

Strong and medium solutions, after a total of 3 hours of agitation, enter the settling vat and the extractors. Weak solutions and wash waters enter separate extractors. All solutions, after settling by gravity, are sent to the corresponding solution vats, where they are strengthened and from which they are delivered to the circulation vats. The 6-inch centrifugal pump, driven by a 17-horsepower electric motor, is of the usual type and size. Before the extractors, weak solutions are strengthened with potassium cyanide. Water is supplied to the tailings plant by flumes from the upper reaches of the stream and from the shafts.

During the summer campaign the plant treats about 1,000,000 poods (161,300 tons) of dry tailings with an average content of 2.5 zolotniks of gold per 100 poods.

Extraction at the plant is 70–80%. Treatment costs amount to about 3 rubles per 100 poods of tailings.

Consumption of KCN: 2–3 pounds per 100 poods.
” NaOH: 4–5 ” ” ”

The summer pulp plant, built next to the tailings plant, consists of 6 wooden vats of the “Pachuca” system, a mixing vat, solution vats, Butters vacuum filters for 18 frames, and extractors.

Above the Pachuca vats there is the same kind of platform as

) J. Park. The Cyanide Process*, 1906, p. 84.

and over the felting vats. The silt, taken directly from the old cuts, is raised by horse traction in tramcars onto a platform, where it is emptied into the vat through 2–3-inch openings between the logs. Before loading into the vat begins, 25 poods of a solution containing 0.06% KCN are pumped in; during loading this solution is continuously agitated with compressed air. In all, 32 tramcars, or about 2,460 poods (38.7 tons) of dry silts, are loaded into the vat. From time to time during loading lime is added, in an amount of 4 poods per vat. After loading is completed, agitation is continued for about 8 hours, after which the pulp is discharged into the agitator.

Compressed air is supplied by a compressor at a pressure of 25–30 pounds to the central air pipe and to 2 annular pipes arranged below along the circumference of the vat and provided with holes along their length.

The agitator is a vat with a capacity of about 120 tons of pulp; in its center rotates a crosspiece fastened to a vertical shaft, driven by a 5-horsepower electric motor.

By means of a centrifugal pump the pulp is raised to vacuum filters, onto 18 frames.

The cakes build up over the course of 15 minutes to a thickness of 1–\(1 \frac{1}{2}\) inches. Then the excess pulp is discharged back into the agitator, and the cakes are washed for 40 minutes with a potassium cyanide solution (0.02–0.03%) and for 15 minutes with water. The filtered solutions and wash waters flow by gravity into the clarifying vat and the extractors, and from there into 2 solution vats. The cakes are rinsed with milk of lime and carried off into the stream. Unloading takes 15–20 minutes. The whole operation, from the beginning of cake build-up to unloading, lasts 2–\(2 \frac{1}{2}\) hours. In one shift an average of 1,100 poods (17.8 tons) of dry silt is treated.

There are no data concerning the consumption of cyanide of potassium, since, in the author’s opinion, cyanide and lime are wasted in vain. The site contains rich silts; many silts lying in the cuts are strongly...

of cemented and poorly classified material, directly into Brown’s agitator cannot give especially good results. Usually, when reworking old silts, the pulp is first prepared in pulp-forming devices, and only then is it sent to the agitation apparatus.

III. The Bogomdarovanny Mine.

The Bogomdarovanny mine is located in the southern part of the Achinsk district of the Yenisei province, 225 versts from the Itat station of the Tomsk railway and 35 versts from the village of Chebaki. To the mine itself, through the efforts of the gold miners who once worked in this region, a cart road was laid; it is, however, in places quite difficult and inconvenient, owing to steep ascents, boggy places, and the absence of regular repair.

The mine camp and the technical structures are situated along the course of the Bolshaya Sobaka brook, which belongs to the system of the Bely Iyus River, a tributary of the Chulym.

The environs of the Bogomdarovanny mine, as well as of the Tsentralny mine, were in their time a region of lively activity by gold miners, who washed the gold-bearing areas here; the work was often carried out on a large scale, although for this reason it did not lose its generally almost wholly predatory character. At present, only old dumps stretching for whole versts, and the half-ruined remains of gold-washing installations and miners’ camps, point to the former activity.

In 1899, on the northern slopes of the neighboring Podoblachnaya and Podzvezdnaya mountains, at an elevation of more than 600 sazhens, in the headwaters of the Fedorovsky spring, a tributary of the Sobaka, outcrops of gold-bearing veins were discovered—those same bedrock formations that were the source of the rich placers in the lower-lying valleys. This deposit is being worked at the present time as well.

The gold-bearing veins occur here partly in altered argillaceous shales, partly in epidiorite, and consist of quartz with inclusions of water, liquid carbon dioxide, and sulfides, chiefly pyrite and magnetic pyrites.*) The gold is found in the vein in native form, since the sulfides are rather poor. Large visible gold grains are often encountered. It should be noted that the gold in the Bogom-Darovannyi veins is of very high fineness, 900–950.

At the present time the mine belongs to the Russian Gold-Mining Company, to which it passed from Ivanitsky.

Gold-washing mill. The ore raised to the surface is crushed at the mine in a Blake-type crusher and poured into an ore bin, whence, as required, it is loaded into the cars of the suspended railway and sent to the lower station, where it is distributed among four bins. From these latter the ore passes into carts, which take it to the mill.

Here the ore is poured into bins installed by the stamps and equipped with automatic Challenge feeders, which feed the ore to the corresponding stamps.

At the mill four five-stamp batteries have been installed, driven by a 100 HP electric motor; each battery receives 14 HP (the motor is not fully loaded). Two batteries, Nos. 1 and 4, are intended for the final grinding of the ore, and internal amalgamation is used in them. The other two batteries, Nos. 2 and 3, are used as apparatus for intermediate crushing; mercury is not poured into them, and the tailings from them go for further grinding to two pans, Nos. 1 and 2. According to the purpose of the stamps for coarse or fine crushing, the size—

*) Obruchev. Geological Survey of the Gold-Bearing Districts of Siberia.

the size of the screen openings, the height of the sill, the weight of the stamps and, of course, the capacity. It would be advisable, if this were possible, to change also the shape of the stamp battery, since for fine grinding wide batteries of the Californian type are most suitable, where the material is retained longer, whereas for coarse crushing a narrow battery is preferred, such as that used on the Rand. Table 12 presents certain data relating to the stamps and their operation.

The slimes from batteries Nos. 2 and 3 are sent for further grinding to two bowls of the kochkar type, driven by an electric motor of 30 HP. The height of the sill at the bowls is 7″; the screen has 16 openings per linear inch, No. 16. The capacity of each runner is about 1000 poods per day, which, at a power consumption of 15 HP per bowl, amounts to 66.6 poods per 1 HP in 24 hours.

Table 12.

	Stamps Nos. 4 and 1	Stamps Nos. 3 and 2
Weight of one stamp . . . poods	34½ (1098 f.)	27½ (873 f.)
Number of drops per minute . .	100	100
Height of drop, inches . . .	3—4	3—4
Width of battery, inches . . .	16	16
Height of sill, inches . . .	13	8
Screen No. . . . . . . .	30	4
Ratio of liquid to solid .	15—20 : 1	15—20 : 1
Capacity in 24 hours, poods	360	1000
” per 1 stamp, poods	72	200
” per 1 horsepower, poods	18.5	72.9

From the sieve analyses of the slimes obtained from the corresponding grinding apparatuses, presented in Table 13, it is evident that bowls and stamps Nos. 1 and 4 give material of approximately the same grain size.

Table 13.

Sieves of the I.M.M. system	Sieve analyses of slimes from: stamps No. 1 with 30-mesh screen	Sieve analyses of slimes from: stamps No. 4 with 30-mesh screen	Sieve analyses of slimes from: stamps No. 2 with 4-mesh screen	Sieve analyses of slimes from: bowl No. 1 with 16-mesh screen	Sieve analyses of slimes from: stamps No. 3 with 4-mesh screen	Sieve analyses of slimes from: bowl No. 2 with 16-mesh screen
+30	3,9	3,3	56,3	2,3	66,2	1,2
+60	14,2	21,9	16,2	18,3	12,7	15,2
+80	12,4	14,4	5,3	12,5	3,1	13,8
+100	7,8	6,5	1,2	5,6	1,3	9,1
+150	10,7	9,2	4,9	8,9	3,3	7,8
+200	12,2	10,3	5,1	10,9	4,5	11,1
−200	38,8	34,4	11,0	41,5	8,3	41,8

The output of the stamps, operating on a fine screen, per 1 HP in 24 hours is equal to 18.5 poods (see Table 12, p. 46). The combined operation of stamp and bowl, with a total expenditure of \(14+15=29\) HP and an output of 1000 poods per day, gives 34.5 poods per 1 HP in 24 hours, or 47% more.

The output per 1 HP in 24 hours of the runners at the Central mine is equal to 90 poods when operating under similar conditions. Ordinarily, however, the output of runners is still higher: thus at the Trinity factory in Kochar, according to Bayldon’s data) (Bayldon), the runners crush 149 p. per 1 HP in 24 hours; while according to Urbiter’s data*) (Urbiter), the produc-

) Bayldon. Trans. Inst. of Min. and Metall. Vol. XX, p. 141.
*) Urbiter. Engen. and Min. Journ. Vol. XCII, 1911, p. 257.

the productivity of Chilean mills, or, as they are called in the literature, arrastras, in Mexico is about 114–120 poods per 1 HP per day.

If we turn to the data available in the literature on the productivity of stamps working under conditions similar to those at Bogom-Darovannyi, we find that per 1 HP in 24 hours the stamps crush 68–99 poods;*) this exceeds several times the productivity of one horsepower at Bogom-Darovannyi, where such low productivity is probably explained by local conditions and by the somewhat obsolete type of the stamps in operation.

When comparing the practice of crushing used here in Siberia and in the Urals with practice in other countries, a substantial difference is immediately apparent: here arrastras are used almost exclusively, whereas there—stamps. The figures cited above prove that the preference given here to arrastras is explained not only by familiarity with this type of crushing apparatus, but also by their real advantages over stamps, both in terms of the utilization of energy and in constructional respects. These advantages have been clarified in detail in the literature, to which I refer interested readers.**) Here I shall only note that a simple slow-running bowl, requiring almost no repairs and not needing the attentive care of trained workers, is an indispensable apparatus under conditions of taiga work, especially at small workings that do not possess any significant funds for repair work. As far as I know, at Bogom-Darovannyi as well, the choice of stamps for crushing the ore is explained by accidental circumstances.

) Thomson, Stamp Milling and Cyaniding, p. 116.
) See Bayldon, Trans. Inst. of Min. and Metall., Vol. XX, p. 141.
Urbiter, Engen and Min Journ., Vol. XCII, 1911, p. 257.
Megraw, Engen. and Min. Journ.*, Vol. XC, 1910, p. 967.
Abstract in “Gold and Platinum,” 1913, p. 295, and
Gernet, “Gold and Platinum,” 1910, p. 155.

As was said above, internal amalgamation is not used at stamps Nos. 2 and 3, which work on coarse mesh. Into the other two stamps, Nos. 1 and 4, and into the bowls, 20 zol. of mercury are poured each hour. In addition, the threshold sill plate at these stamps is sheathed with an amalgamated copper sheet and is rubbed with mercury during washing down.

The ore entering the feeders contains, on average, about 7.96 zol. per 100 poods. The slimes, however, leaving the stamps with internal amalgamation contain only 5.71 zol. per 100 poods; i.e., the stamps recover about 30% of the gold. After the stamps, the slimes go to sluices, and from there are either directed directly into the launders leading them out of the mill (from stamps Nos. 1 and 4), or first pass through the bowls and from there are carried by similar sluices into the common discharge launder. The plates between stamps Nos. 2 and 3 and the bowls recover very little gold, since the coarse-grained slimes strip the amalgam from the plates. Pockets are arranged in the sluices, in which stirrers are installed, set in motion by the transmission. These stirrers assist in catching the floating amalgam and prevent the pockets from becoming clogged.

Through the launders the slimes are carried into a system of sumps, where they are separated into sands, which settle in them, and silts, which are carried into settling pits. Table 14 gives the screen analyses of the slimes entering the sumps, and of the silts and sands obtained after their separation.

Certain figures illustrating the operation of the mill during the first third of 1913 are given in Table 15 (p. 51).

The extraction by amalgamation, as is evident from Table 15, fluctuates around 60–65%. The evaporated gold is smelted with borax and soda in graphite crucibles, and very high-grade (over 900) ingots are obtained. For 10 pounds of gold, 3 pounds of borax and 2 pounds of soda are taken for smelting.

Thanks to the internal amalgamation in the bowls, and espe-

Table 14.

G.M.M. sieve system	Sieve analyses: slimes	Sieve analyses: sands	Sieve analyses: silts
+30	0.5	1.8	0.1
+60	9.1	19.1	1.9
+80	10.8	17.5	1.4
+100	8.4	11.4	4.1
+150	11.8	16.4	6.9
+200	4.3	4.1	1.8
−200	55.1	30.8	84.8
Au content in 100 poods		3.68	2.68

especially in the stamps, the loss of mercury at Bogom-Darovannyi is very great and amounts to about 30 zolotniks per 100 poods.

At the factory, only 21 men are employed.

The consumption of electric power per 100 poods of ore is 63.11 kW-hours. This energy is distributed as follows:

    for the stamps . . . 62.95%

    “ the Chilean mills . . . 32.43 ”

    “ lighting . . . 3.00 ”

    “ zinc point 1.62 ” (for cyanide plants)

The cost of treating 100 poods of ore fluctuates around 6 rubles, including treatment of the tailings at the plants.

Table 15.

Months of 1913	January	February	March	April	May
Ore crushed, poods	62.847	55.759	67.000	61.615	61.519
Output in 24 hours		2 0 4 4	poods
Au content in 100 poods of ore	7.37	7,2	7.51	9.36	8.87
“ “ “ sludge	2.89	2.46	3.86	2.89	3.23
% recovery	60.79	65,83	48.6	69.13	63.59
Amalgam obtained	2 p. 32 f. 8 z.	2 p. 8 f. 36 z.	2 p. 9 f. 55 z.	2 p. 33 f. 39 z.	3 p. 3 f. 80 z.
Alloyed gold steamed off	34 “ 45 “	26 “ 49 “	27 “ 72 “	35 “ 73 “	1 “ 0 “ 13 “
Gold smelted	34 “ 34 “	26 “ 47 “	27 “ 63 “	35 “ 66 “	39 “ 93 “
Pure “ obtained	32 “ 12 “	24 “ 90 “	25 “ 46 “	32 “ 67 “	36 “ 48 “
Fineness of ingots	938,1	939,25	921,0	913,0	913.1
Pure Au obtained from 100 poods	4.91	4,29	3.65	5.09	5.79

Individual expenditure items are distributed as follows:

Item	Share
Hauling in	7.44%
Crushing	29.73%
Classification	1.69%
Amalgamation	10.32%
Extraction	15.00%
Hauling away tailings	4.79%
Supervision	9.28%
General overhead expenses	22.12%

Tailings plant. The tailings separated in the sumps, as well as old dumps, are reworked at the summer tailings plant.

The latter consists of 8 leaching vats, \(28' \times 9'\), set in one row on a rubble-masonry foundation about \(7'\) high. As usual, beneath the vats there is a passage through which cars run on rails, carrying away the treated tailings by horse traction. A platform above the vats, assembled from beams with gaps of \(2''\), with a double rail track laid on it, serves for loading the tailings, which are delivered in cars from the dump or from the sumps. Parallel to the percolation vats there is a row of solution vats (4 in all), measuring \(22' \times 7'\). In order that the solutions from the extractors should enter them by gravity, the vats are sunk into the ground by \(5'\).

All the vats are built of pine wood. The staves for the percolation vats have the following dimensions: length \(7'10\frac{1}{2}''\), width \(9''\), and thickness \(4\frac{1}{2}''\). The bottom is strengthened in a rabbet cut at a distance of \(10\frac{1}{2}''\) from the lower edge of the stave, to a depth of \(1\frac{1}{2}''\). The vats are tightened with hoops of bar iron \(3'' \times \frac{1}{4}''\), 4 on each. At the bottom of the leaching vats there are 2 tight hatches, opening from below and serving for discharging the tailings into the cars placed beneath. Along the bottom of the vat, as usual, ulo-

an iron grating, divided into 4 quadrants, over which canvas is stretched, covered on top with a bast zinc mesh. There are no gratings protecting the filter from above against damage when the tailings are unloaded. In the lower part of the vat, set on a foundation with a certain slope to one side, there is a 3″ branch pipe, onto which a flexible hose is fitted; with it the flowing solutions can be directed into any of the three troughs that lead them to the extractors.

Somewhat higher than the percolation vats are installed 2 reservoirs for water, which is conveyed into them by flumes. By means of a system of 3″ pipes and troughs, the water can be directed into any of the solution or percolation vats.

A 4″ centrifugal pump, driven by an electric motor, serves to raise the solution.

The tailings are loaded into the vat in the amount of 80–83 cars of 75 poods each, or altogether about 6,000 poods (99 tons). From each car a sample is taken, and to each car are added 5 pounds of lime, or altogether about 10 poods per vat. Loading continues for 10–16 hours, with 7 workers and 2 horses employed.

Upon completion of loading, the first solution is immediately pumped in from above. Usually, according to the literature, the first solution is admitted from below.

The total quantity of solution passed through the vat is equal, by weight, to the charge, i.e., 6,000 poods. The strength of the solutions and their relative quantities are given in Table 16 (p. 54).

The tailings are not washed with water, since all the solutions before the extractors are strengthened to 0.08%, and there is always more than enough of them.

The total duration of treatment is 6 days, including the time for loading and unloading. The latter operation takes 10–12 hours and is carried out by 6 workers and 2 horses.

Table 16.

Filling sequence No.	Name	Of the total quantity, %	Poods	% KCN	% NaOH
1.	Strong	50	3000	0.12—0.14	0.08—0.1
2.	Medium	25	1500	0.09—0.1	0.04—0.08
3.	Weak	25	1500	0.06—0.09	0.05—0.07

Table 17.

	1	2	3	4
Au content in 100 poods of charge	4.18	2.92	2.61	2.72
Au content in 100 poods of tailings	0.91	1.14	0.73	0.99
% extraction	78.23	61	72.03	63.62
% moisture in the charge	10.5	12	13	13
" " " in tailings	15.5	14.4	—	17.3

The work of the vat is illustrated by the figures in Table 17 (p. 54). From this table it is evident that the extraction of gold by percolation fluctuates around 60–70%. In practice, an extraction of 85–90% is considered good. Such results of the operation of the vats at Bogom-Darovanny are explained, first, by the insufficiently fine crushing of the tailings, the grains of which are too large for the gold contained in them to be able to dissolve completely; and second, by the high content of slimes in the material being treated, which reaches 48% and on average amounts to about 30%, as is seen from Table 18.

Table 18.

Sieves, I.M.M. system	1	2	3	4
+30	1.2	0.52	1.1	1.02
+60	16.4	15.17	15.52	19.36
+80	18.8	16.78	15.16	18.25
+100	11.7	10.48	8.83	14.9
+150	11.6	15.42	15.69	17.87
+200	11.5	3.9	3.21	3.21
−200	28.5	37.73	30.21	24.79

Such a high content of slimes considerably hinders percolation. Owing to the slimes, moreover, the tailings being leached contain too much moisture.*) If

*) According to the data of the plant laboratory, about 15%, which is probably erroneous, since even in very clean tailings up to 25% moisture usually remains.

refer to Table 19, where the gradual change in the gold content in the solutions flowing out of the vat is shown, and we find that the gold content in the last fractions is 0.30–0.35 zol. per 100 poods.

Table 19.

No. in order	Au content in 100 poods
1	1.53
2	3.99
3	4.48
4	7.45
5	4.99
6	5.25
7	3.76
8	1.11
9	0.38
10	0.31
11	0.34

It is evident that the 15% of moisture contained in the discharged material includes a rather considerable quantity of dissolved gold, namely:

\[ \frac{15.6000.0,3}{100.000} = 2,7 \text{ zol.} \]

The insufficiently high recovery by means of simple percolation led to the decision to reprocess all the slimes obtained at the plant by the full silt process, which at the present time*) has already been carried out. During my stay at the mine, for the future plant only leaching was being planned. Such a solution of the question of ore treatment is, in my opinion, too radical, and perhaps it would have been sufficient merely to reduce the size of the crystalline grains by subsequent grinding in a tube mill, and to introduce a more perfect classification.

The solutions from the vats flow along launders into the extractor room and are directed, depending on the gold content, into one or another extractor.

The 2 extractors for weak and medium solutions have 11 compartments each, 32″ wide, 16″ long, and 24″ deep. The extractor for the strong solution has 12 compartments. All 3 extractors are built of wood. The frames

*) Winter of 1915–16.

for supporting the shavings, are provided with screens with 12–16 openings per linear inch. The finer screen becomes clogged with short shavings and, owing to this, the rate of circulation of the solutions is greatly reduced. The normal rate of passage through the extractors here is considered to be:

Type of solution	Throughput in 1 hour
for a strong solution	150 poods
for a medium solution	150 poods
for a weak solution	180 poods

A strong solution is considered to be one containing:

Solution	Gold content
Strong	up to 4 zolotniks of gold per 100 poods
Medium	2 zolotniks of gold per 100 poods
Weak	1 zolotnik of gold per 100 poods

After washing, the completeness of extraction is considerably reduced. The same phenomenon is observed if the strength of the solutions is lowered to 0.05% KCN or below. Therefore all solutions before the extractors are reinforced to 0.08% by adding, into the first compartment from a small reservoir, a concentrated solution of potassium cyanide, the stream of which is regulated according to the content of the incoming solution. The outgoing solutions contain from 2 to 10 parts of Au per 100 poods. The concentration of KCN in these solutions falls to 0.06–0.07%. After the extractor, the solutions are sent to the corresponding solution tanks.

Zinc shavings are placed in the extractors as needed and, as usual, from the last compartments the shavings are moved to the first, while fresh shavings are placed in the last. The shavings are preliminarily leaded by immersion for 1/2 minute in a 10% solution of lead acetate. One compartment holds about 8 pounds of shavings.

Washing is carried out once every two weeks, usually after 14–16 tanks have been treated. Before washing, the inflow of solution is stopped and the ex-

the extractor is washed with water; then the shavings are taken out and placed for washing in a mesh drum, which is rotated by hand around a vertical axis in a barrel, constantly changing direction. The shavings remaining in the drum are again loaded into one of the first compartments of the extractor, and the washed-out fine zinc and precipitate are filtered through bags of dense calico, through which the slimes obtained when cleaning the extractor are also passed. The filtrate is poured into a barrel, where it is allowed to settle for a long time. To improve settling, alum is added.

The precipitates are dried and roasted in a reverberatory furnace. The duration of roasting is 2–3 hours. An analysis of the roasted residues is given in the following table.

Table 20.

Found	% %
SiO₂	10.0
CuO	5.0
PbO	27.4
CaO	2.0
Fe₂O₃	7.4
Al₂O₃	8.31
ZnO	21.2
Au	13.61
	usually 0.34–5.9

After roasting, 9–12 poods are obtained, which are treated with sulfuric acid (1 : 5). The roasted precipitates are poured in small portions into a vat into which the acid has been poured. After the reaction is complete, over the course of 7–8 days the residues are washed by decantation, filtered through the same calico filters as were used in rinsing, and dried in a reverberatory furnace on iron sheets. Usually 3–4 poods of residues are obtained, which are fused in graphite crucibles; for 100 parts of residue the following are taken:

borax . . . 30 parts
soda . . . . 20 ”
sand . . . . 10 ”
saltpeter . . 2 ”

Gold ingots are obtained with a content of: 143 fineness of ligature, 44 fineness of silver, and 813 fineness of gold.

For dissolving 1 pud of roasted residues, 2–3 puds of sulfuric acid are consumed.

In conclusion I shall give the results of work at the Efeln plant in 1913.

Processed in all . . . . . . 80 vats.
Total amount of charges 430.457 puds.
Average gold content . . 3.27 zolotniks per 100 puds.
Total gold content . . . 14,093.19 zolotniks.
Gold content in the tailings 1.18 zolotniks per 100 puds.
Total content ” ” ” 5,086.48 zolotniks.
Extracted in all . . . . . 9,006.73 ” ”
% extraction . . . . . 63.9.
Obtained of ligature gold 120 funt. 63.35 zolotniks.
” of pure ” 92 ” 6.45 ”
Consumption of reagents: KCN 140 p. (1.3 funt. per 100 p.)
” ” NaOH 187 ” (1.72 ” ” ” ”)
” ” zinc 81 p. 33 f. (0.82 f. per 100 p.)
” ” lime 1600 p. (14.87 ” ” ” ”)

Summer silt plant. The silt plant, operating by the old decantation method, consists of 4 agitation vats of the Brown-Pachuca system, with a corresponding number of filtering and dissolving vats.

The agitation vat, 31′ high and 11′ 6″ in diameter, is built, as at the Central plant, of wooden beams 35′ × 9″ × 4½″ and is tightened with the corresponding number of hoops of strip iron. The conical bottom of the vat was likewise made of wooden beams and was the weak point of the entire structure, since a leak constantly appeared there, which could be partly eliminated only by silting up during operation of the vat. At present, the funnel made of beams, after a whole series of unsuccessful attempts to achieve a fully tight joint, has been decided to be replaced by a funnel made of bricks in cement mortar and with ce—

with the same cement plaster. Such a construction will probably at last provide the desired watertightness.

The central pipe of the vat is riveted from 1/4-inch boiler iron and has a diameter of 12 inches. The pipe is strengthened by cross braces and does not reach the bottom of the vat, while its upper edge is 12 inches below the top. For supplying the air by which agitation is produced, the following serve: 1) the middle branch pipe opening into the 12-inch pipe, 2) the lower fixed spider, fastened to it, 3) a ring with branch pipes, fastened around the circumference of the conical bottom, 4) a movable spider, lowered from above on a fire hose. Through the fixed spider and the lower ring, water may also be supplied. The central branch pipe and the branch pipes of the ring and spiders are provided with valves made of rubber tubing.

By means of the centrifugal pump that also serves the extraction plant, the vat is filled with a solution containing 0.098% KCN and 0.096% NaOH, the level of which must not come within 2 feet of the upper edge of the central pipe. In all, the vat holds about 3,400 poods, or 54.81 tonnes, of solution. Then compressed air is admitted through the central pipe, and water through the fixed spider and the ring, and the loading of the slime from angle-iron screens, delivered from the cut in small cars onto the platform above the vats, is begun. In this way 1,500 poods (24 tonnes) of slime are loaded into the vat. To improve settling and neutralize the pulp, 12–15 poods of lime are added to the vat. The slime being loaded contains on average 20–25% moisture and contains 3.2–3.72 zolotniks of gold per 100 poods. A screen analysis of the slime is given in Table 21 (p. 61).

Compressed air is delivered by a compressor at a pressure of 22–28 pounds.

The duration of agitation with the first solution is 18–24 hours, after which the blowing-in of air is stopped and the vat is left to settle for 12–15 hours. The settled solution is decanted into the filtering vat either by means of a float on a flexible hose, or by means of 2-inch taps arranged along the height

tank. Usually it is possible to decant 2,800–3,000 poods of solution containing:

0.70–0.85 zolotnik of gold per 100 poods. 0.06–0.075% potassium cyanide. 0.05–0.07 „ caustic soda.

After decantation, about 3,000 poods of solution are again poured in, the content of KCN and NaOH in which is the same as in the decanted solution.

To raise the settled slime, or, as they say, to blow out the tank, water is forced in through the lower spider, which gradually washes away the slime that has collected on the bottom of the tank, and air is admitted through the central pipe. Gradually a connection is established between the pulp in the tank and the central pipe, and proper agitation begins. Blowing out the tank lasts 1–2 hours, although it sometimes happens that the tank cannot be blown out even over the course of many hours.

Table 21.

Sieve, I.M.M. system	%
+ 30	0.8
+ 60	2.0
+ 80	1.5
+100	1.4
+150	4.2
+200	0.7
−200	89.4

Agitation and washing are continued for 6–8 hours; then 3–4 poods of lime are added and, after settling, the solution is again decanted. Usually it contains 0.3–0.4 zolotnik of gold per 100 poods.

After one washing, the slime is discharged; for this purpose the tank is filled halfway with water, air is blown for some time through the spider and the ring, and, upon opening the hatch of the discharge opening, the slime is released, being driven out by a strong jet. The entire tank is emptied in \(1/4\) hour. The total duration of discharge is \(1\frac{1}{2}\)–2 hours. The whole operation lasts from 61 to 72 hours.

The decanted solution flows by gravity into filtering vats, \(21' \times 7'\), whose filters are arranged in the same way as in the percolation vats. A 2-inch layer of tailings is loaded onto the filter; wood shavings are laid upon it, and above that another layer of tailings, 3 inches thick. The filtered, usually quite clear solutions enter the same kind of extractors as those for the tailings plant, but with 13 compartments. Precipitation, with a slow flow of the solutions and reinforcement of them with KCN, is quite satisfactory, and the outgoing solutions rarely contain more than 0.006–0.007 zol. Au per 100 poods.

The operation of the slime plant is illustrated by the following figures:

Indicator	Value
slime treated per operation	70,648 poods
with average content	3.45 zol. per 100 poods
total gold content	2,617 zol. per 100 poods
gold content in the tailings	0.99 zol. per 100 poods
total gold content in the tailings	748.91 zol.
gold extracted	1862.15 zol.
% extraction	71
duration of agitation	6,392 hours
consumption of reagents: KCN	103 poods (5.4 lb. per 100 p.)
consumption of reagents: NaOH	102 poods (5.3 lb. per 100 p.)
consumption of reagents: zinc	57 poods (2.9 lb. per 100 p.)
consumption of reagents: lime	587 poods (30 lb. per 100 p.)

The comparatively low extraction of gold (71%) is explained, first, by the method of charging the vats, into which, just as at the Central plant, settled, poorly classified slime is also dumped; as a result, stratification of the pulp is inevitable. This usually occurs even under normal operating conditions in Pachuca tanks, although to a considerably lesser degree, and is one of the essential shortcomings of these agitators; stratification of the pulp worsens the contact between the particles of slime and the solution present in the vat, which, of course, reduces the rate and completeness of gold dissolution. Secondly, there are inevitable losses of already dissolved gold in the moisture that remains in the slimes after decantation of the second wash solution. These losses

it is easy to determine from the data given above: in a charge of 1,500 poods, after decantation there remains about 500 poods (33.3%) of solution, containing on average 0.35 zolotniks per 100 poods, or in all 1.75 zolotniks of gold, which per 100 poods of charge amounts to 0.118 zolotniks, or 3.42% of all the gold contained in the charge.

In concluding this report of mine on the trip, I consider it my duty to express my profound gratitude to the managers of the mines and to their assistants for furnishing me with all materials, and also to Professors V. Ya. Mostovich and L. L. Tove, who greatly facilitated my work in compiling the report by indicating the relevant literature.

At the same time I consider it necessary to note that the benevolent attitude which I encountered on the part of the administration of the mines I visited is, among us in Russia, unfortunately, a rather rare exception. In most cases at our works visitors are treated with suspicion, and even the smallest details are made into a secret, although often everything so zealously guarded is described in detail and with care in the relevant literature. The same attitude is met by trainees who come to the works to supplement the knowledge acquired at school. But while our industry fences itself off on all sides with Chinese walls, foreign specialized publications, especially American ones, are filled with the most detailed descriptions of all sorts of factories and works, with all details of both a technical and an economic character. It is highly symptomatic that a description, for example, of the Troitsk factory at Kochkar is found in the London Transact. of the Inst. of Min. and Metall., and not in any Russian journal, where one can very rarely find at all detailed descriptions of existing enterprises. Meanwhile, the study of industry and of individual enterprises is very important, especially in such a technically backward country as Russia. For in most cases our industry is quite unable to make use of the experience of individual enterprises, and each of them entirely and in a row—

repeating the same mistakes, wasting funds and the people’s labor unproductively, in order, in the end, to work out rational methods of work that have long since been applied next door. True, sometimes the iron laws of competition, if they do not justify, at least explain the desire of individual enterprises to keep their experience for themselves. But precisely in America, where this competition is so strong, industry is most accessible to study from all sides; whereas in Russia, where various forms of commodity famine are constantly felt, and where therefore competition cannot and should not exist, individual enterprises in most cases are completely inaccessible both to the practitioner and to the researcher.

V. Pazukhin.

March 1916.

Typography of G. I. Liven, Tomsk.

Most Important Misprints

Page	Line	Printed	Should read
1	8 from bottom	historical	historical
3	19 from top	eternal	sluice
5	15 from bottom	is produced	is brought
6	13 from bottom	in 24–30	during 24–30
"	11 from bottom	trough	trough
"	"	Which	Which
8	14 from bottom	established	installed
"	footnote	gives to renew	gives for renewing
12	5 from top	as a consequence	as a consequence
"	9 from bottom	sluices	sluices
18	1 from top	Slyusok	Slyusok.
"	6 from bottom	(On table 4 p. 19)	On table 4 p. 19
"	2 from bottom	slyusok	slyusok
21	8 from bottom	Testing	Testing
24	9 from bottom	mills	mills
30	footnote	Or	or
33	11 from top	sediment	sediment
36	6 from bottom	Koldeskott	Kaldeskott
38	1 from top	sulfides	sulfides
39	15 from top	four	four
41	17 from top	quantitygiven	quantity given
49	2 from bottom	is used	is used
57	3 from bottom	Slyusok	Slyusok
60	18 from top	2,	2′
62	8 from bottom	at	at
65	2 from bottom	so that	so that.

Diagram 1.

Arrangement of the equipment at the Uspenskaya factory
and the warm silt plant.

Drawing 4. [[unclear: caption text]] (to Scheme I).

1. Driving.
2. Sampler.
3. 8 hand ore bins.
4. 2 plank platforms.
5. 6 pairs of trestles.
6. Small sluice.
7. [[unclear: item name]].
8. Lower sluice.
9. 3 false rollers.
10. 5 upper [[unclear: item name]].
11. Common sump.
12. 6-inch centrifugal pump.
13. 2 classifiers.

36 ↓

DESIGNATIONS

37 ↓

1. 2 tube mills.
2. 2 pulsating tables.
3. 2 Ferk pumps.
4. 2 Dorr thickeners.
5. 2 control pumps.
6. Slime collector.
7. Vat for concentrated solution.
8. Centrifugal pump.
9. 2 Dorr agitators.
10. Mixing vat.
11. 10-inch centrifugal pump.
12. Vacuum filters.

13. Vacuum pumps.

14. 2 clarification vats.

15. 6 electrolyzers.
16. 4 vats for solution.
17. Water vat.
18. 2 solution vats.
19. Small vacuum pump.
20. Small vacuum filters.
21. Charging centrifugal pump.
22. Discharge column from the thickener.
23. Excess solutions.
24. Discharge column.
25. Compressor.
26. End of the first section.

A. Pazukhin.

Toward the article by S. P. Yakovlev.

Scheme II.

Circulation of solutions and pulps.

• Water
• Ore crusher
• Feeder
• Edge runners
• Sluice
• Classifier
• Classifier
• Leaching vats
• Sump
• Concentrates
• Lift 22′
• Lift 22′
• Dorr thickeners
• Collector
• [[unclear: “… water”]]
• Lift 20′
• Classifier
• Tube mill
• Pulp tables
• Fréné pump
• Lift 22′
• Dorr elevator
• Collector
• a. Vacuum filters
• b. Screens
• Solution tank
• Precipitate melting
• Slimes
• Gold
• Vacuum filters
• Cake discharge
• Vacuum pump
• Clarifying tank
• Extractors
• Solution vats
• Excess solutions
• Lift 20′
• Excess solutions

[[unclear: signature]]

Diagram III

Factory and cyanide plants at Bogom-Darovannyi

Visible numbered labels in the diagram: 1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14, 15, 16, 17, 18, 19, 20, 21, 22, 23, 24, 25, 26, 27.

Designations.

Tomsk, Printing House of V. P. Yakovlev.

Legend

1. 4 ore [[unclear: workings]].
2. 4 adits.
3. 4 shafts.
4. Upper sluices.
5. Lower sluices.
6. 2 old workings.
7. Mill.
8. Butaras.

9. Sumps.

10. 3 water dams.

11. Cuts for ditches.
12. Tailings dump.
13. Rail track.
14. 2 centrifugal pumps.
15. Transmission ditch.
16. 2 locomobiles, [[unclear: Stangen oven]].
17. Tanks for solutions.

18. [[unclear: Rymduction of ore]].

19. Lighting unit.

20. Extractor for weak solution.
21. Extractor for [[unclear: rich]] solution.
22. “Dauchka” tanks, covered.
23. Filter tanks.
24. [[unclear: bins for processed sands]].
25. [[unclear: bins for unprocessed sands]].
26. Extractors.
27. Tanks for solvent.