In the case of many complex gold and silver ores roasting before cyanidation is essential if satisfactory extraction of the precious metals is to be obtained. In such cases no practical amount of grinding or prolonged contacting of the raw ore with cyanide solution will affect more than a certain low extraction; in other cases, while the extractions may be acceptable, the consumption of cyanide is prohibitive, and roasting or, occasionally, pyritic smelting of the material is the only alternative. Roasting, which is today almost always carried out following concentration of the values into a small bulk, must be used if it becomes necessary to decompose the minerals with which the gold is associated in order to expose it to solvent action. Many of the cyanicides in the raw ore are broken up and rendered harmless by roasting, but the calcine may contain new compounds that must be removed by water or acid washing before cyanide treatment can be successfully carried out.
There were instances, where at one time the whole of the ore was roasted before cyanidation, but this practice has been for the most part discontinued on account of the large plant and high overall costs involved. With improvements in flotation methods, the generally accepted procedure today is to concentrate the gold or silver bearing minerals into a small bulk and roast the concentrate only. The calcine is then usually treated in a small cyanide circuit, with the residues sometimes passing into the main cyanide circuit, if the flow sheet includes such a step, or sent directly to waste.
Those types of gold ores which most frequently require roasting include ores carrying arsenopyrite, stibnite, sulphotellurides, and pyrrhotite. Straight pyritic ores, where the pyrite is present in small quantities, usually yield their gold to fine grinding and cyanidation alone.
Silver ores containing the values as polybasite, stephanite, pyrargyrite (the antimony sulphide), and proustite (arsenic sulphide) usually require roasting. Tetrahedrite is often refractory even after roasting. Argentite (the silver sulphide) and cerargyrite (the chloride) can frequently be cyanided without roasting. A comprehensive description of roasting practice is to be found in a series of articles entitled Roasting Gold-Silver Sulphide Ores and Concentrates by M. W. von Bernewitz.
The roasting of straight pyritic ores involves the conversion of the iron sulphides to the oxide under oxidizing conditions with the evolution of sulphur dioxide and, to some extent by catalytic action, sulphur trioxide gas also. In this chemical change the iron mineral is rendered more or less porous, thereby permitting the dissolution of the contained gold by subsequent cyanidation.
Though simple in principle, considerable temperature and other controls are necessary to avoid undesirable side reactions which are discussed below. It is not usual, for instance, to leave more than about 0.1. to 0.15%insoluble sulphur in the roasted ore, but much larger percentages of soluble sulphates are often present. It is believed that in certain cases such salts may reduce the porosity of the iron oxide and so lower the gold extraction. In addition to higher gold losses, incomplete roasting also causes trouble due to the presence of ferrous salts and other cyanicides.
The treatment scheme at Lake Shore includesdirect cyanidation followed by flotation of the sulphides, roasting of the concentrates, and cyanidation of the calcine in a separate circuit. The following notes are taken from J. E. Williamsons paper Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant, C.I.M. and M.
The pyrite values include all gold so intimately associated with the pyrite that cyanidation for a prolonged period under ideal laboratory technique will not dissolve it. Unlike the gangue values, finer grinding has comparatively little effect on reducing the pyrite values of the cyanide tailing.
Before the cyanide residue from the main plant can be floated, it is necessary to destroy the lime alkalinity present, and a good deal of experimental work was carried out before a suitable design of spray tower for conditioning the pulp with S02 gas from the roaster was worked out. This is fully described. Following recirculation of the pulp through this tower to give a pH of about 6.0 (see page 305), it is floated in Fagergren cellsusing reagent 301, copper sulphate, and pine oil. The concentrates are thickened, filtered on a drum filter provided with a flapper (see page 120), and fed by belt conveyor at 16 to 17 per cent moisture to the Edwards roaster. On the basis of a mill feed of 1200 tons per day, about 23 tons per day of concentrate carrying 26 to 27 per cent sulphur is roasted.
The installation includes two standard 70-rabble roasters, though only one unit is currently used. A 25-hp. motor drives the complete mechanism of each roaster. Special insulation was provided on the side walls and arch, and by using air-cooled rabbles a close control of temperature is possible without the use of any outside fuel.
A cooling hearth is incorporated in each of the Lake Shore roasters. This arrangement is made by providing a 10-in. drop in the bottom of the roaster, which completely stops back mixing of the charge between the cooling hearth and the roaster proper. In the original roaster, eight pairs of rabbles out of the thirty-five are used for the cooling hearth. In the second roaster, the number was decreased to six pairs.
The wet concentrate is a sticky, puttylike substance which adheres to almost any dry surface. The back mixing action of the rabbles ensures that there is always a quantity of dry, disintegrated charge at the feed end of the hearth. The fresh feed entering the furnace falls into this bed of dry, dusty material and picks up a coating of dust. This action can be compared with the bakers use of flour to prevent bread dough sticking to his bread pans.
The dust-coated or dust-lubricated concentrate will not adhere to dry surfaces, nor will it ball up. The rapid turnover and thorough mixing given by the rabbles distribute the dust-lubricated material over the first three or four bays of the roaster, where it can be dried by the heat of the charge and the gases.
The first pair of rabbles are double armed to reduce the amount of wet feed which piles up under the chute between passes of the rabble. The moisture is driven off, and the filter cake shreds break up into more or less equidimensional lumps. On further application of heat, the trapped water inside these smaller lumps turns into steam. The pressure set up inside the lumps causes them to burst.
An over-all hearth slope of about -in. to the foot is provided. However, it is stated that at no point in the furnace can the charge be said really to flow. During the elimination of the first atom of sulphur, indicated by the blue flame, the charge appears to be slightly lighter and fluffier. Even at this point, the angle of repose is over 30 deg. Elsewhere, the angle of repose is nearer 45 deg. These angles of repose compare with 35 deg, which is the angle for a rock dump.
This absence of flow makes it possible to work with a deeper bed of charge and enables the rabbles to do a more efficient job of turning the charge each revolution. The flocculent nature of the charge undoubtedly contributes materially to reducing the dusting and dust losses to the low figure obtained.
The cooled calcine is discharged from the roaster by means of a classifier rake mechanism. The rake draws the calcine out well clear of the discharge port in the roaster before dropping it into the preliming agitator.
Operations at Lake Shore have demonstrated the vital importance of this temperature control over the initial stage of the roast. The charge must be held at a low temperature for a sufficiently long period during the blue-flame stage. (The temperature has reached 900F. at the fifth port and rises slowly to a maximum of 1150 at the eighteenth port. Exit gases at 500 to 650F. contain about 2.7 per cent S02.) The ore is roasted to hematite rather than magnetitethe latter condition resulting in a higher cyanide consumption.
The roasting of ores containing arsenopyrite presents greater difficulties than straight pyritic ores mainly because of the tendency toward the formation of insoluble arsenites and arsenates, which have a detrimental effect on gold recovery.
In consideration of roasting procedure, it is quite generally agreed that reducing conditions should be maintained in the early stages of the roast to ensure elimination of the arsenic in the arsenious state. Provided this has been satisfactorily accomplished, the finishing stages of the roast can be done under active oxidizing conditions. It is not enough that reducing conditions alone be maintained during the period of arsenic elimination, as it is also necessary to maintain movement of the charge and a good flow of gas over or through the charge to carry off the arsenious oxide as it is produced.
A second important consideration in roasting procedure is that of the time-temperature sequence. For Beattie concentrate it proved best to hold the temperature about 900F. during the period of arsenic elimination followed by a rise to 1300F. before completion of the roast.
This 2000-ton per day plant is operating on an arsenical gold ore in the Duparquet area of Quebec, Canada. The operation consists of straight flotation, followed by roasting and cyaniding of the concentrates. The roasting plant, which is one of the finest installations in Canada, has paid for itself in increased recovery over that obtained by cyanidation of the concentrate direct.
The flotation concentrate is thickened, filtered to 16 per cent moisture, and partially dried (8 to 10 per cent moisture) in a coal-fired Buggies Cole drier. The discharge is in the form of round pellets, which are broken up by rollers placed on the inclined conveyors. It is delivered to 250-ton bins placed above each of three 25-ft.-diameter by 13 hearth Wedge split- draft roasters.
The furnaces are provided with an installation of two hot Cottrells for taking dust out of the hot gases, the 10 to 13 tons per day of dust carrying 2.25 per cent arsenic being returned to the third hearth by chain drags and elevators, and two cold Cottrells, which treat the gas after it has been cooled to 250 by addition of cold air (formerly by water sprays) to precipitate the arsenic.
There is also a powdered-coal-firing arrangement for adding heat to the roasters because of the low sulphur content (16 per cent) of the concentrate, and until recently heat was added to certain hearths continuously. At the present time supplemental heat is not supplied, and a feed carrying as little as 12 per cent sulphur plus 2 per cent arsenic has been successfully handled when a high tonnage rate (up to 190 tons per day) is maintained. The roasters have excellent insulation.
The roasters are operated on the split-draft principle; i.e., they are provided with bleeder flues from hearths 7 and 11 and two uptake flues from the first roasting hearths to a balloon flue. The As2O3, which condenses out of the gases when cooled below 500F., is a product that is very corrosive and difficult to handle. The product from the Cottrells runs about 77 per cent As as AS2O3, 12 per cent SO3 and carries about 0.03 oz. Au per ton. As this is not high enough grade to meet present market demands, it is stored in concrete bins. The roaster-plant gases go to a 410 ft. high concrete stack, 18 ft. bottom and 7 ft. top diameter.
The calcine is quenched, without liming, and pumped to hydroseparators which are in closed circuit with tube mills, where the calcine is ground to about 90 per cent minus 325 mesh. This overflow is thickened, filtered, repulped in cyanide solution, and cyanided by a conventional flow sheet.
A total of 1066 tons of solution running about $5 per ton is precipitated per day. Zinc dust consumed equals 0.006 lb. per ton. The present over-all recovery is 85 per cent, consisting of a flotation recovery of 90 per cent, a roaster recovery of 99 per cent, and a cyanidation extraction of 93 per cent. Table 20 summarizes the work in 1948.
The particle-size distribution of Beattie concentrate, as given by Archibald for 1940 practice, is shown in Table 21. Conditions at various points in the roasters, in 1939 when the plant was operating full capacity are shown in Tables 22 and 23.
This roasting plant, since closed down, was equipped with eight standard 70-spindle Edwards roasters, though only four were finally in use. Fuel was supplied to four ports on each furnace in the form of pulverized lignite of 8500 B.t.u. value prepared at a central plant. About 7 tons of fuel was consumed per 24 hr. for a rated furnace capacity of 120 tons of ore, which is equivalent to 117 lb. of coal per ton of ore roasted. The feed carried 3 to 4 per cent moisture, and the roasting temperature varied between 600 and 650C. The time of contact was close to 5 hr. The last 16 spindles operated in the open, and the air-cooled calcine after being sprayed with water was discharged by means of a reciprocating drag conveyor to a Cottrell precipitator. This unit resulted in improved extraction, since the fines were known to carry higher values than the ore as a whole.
Eight Edwards duplex roasters are installed, and each is capable of handling 20 tons of concentrates per day. They are fed by a constant-weight feeder and are heat controlled by means of auxiliary off-take flues placed along the crown of the roaster. These flues draw off hot gases and reduce the temperature at any given point to the desired figure.
The air is drawn through the roasters by two fans having a combined capacity of 100,000 cu. ft. per min. The fans discharge to two six-cyclone Van Tongeren dust collectors and thence to a 9 ft. 6 in. diameter chimney stack 170 ft. high. Loss of weight in roasting = 24.6 per cent.
Figure 49 shows a generalized cross-sectional view of the FluoSolids reactor which has recently been developed and is now being applied to the roasting of sulphide ores, to the calcination of various materials, and in general to problems involving reactions between solids and gases at elevated temperature.
The term fluidization denotes the fundamental principle of carrying out a gas-solid reaction in a dense suspension of solids, maintained in a turbulent mass by the upward flow of the gases that effect the reaction. This mass assumes a fluid level and acts as a fluid. The upwardly moving gas stream imparts to this mass a turbulence resembling that of a boiling liquid.
The outstanding advantage of this principle, as embodied in the FluoSolids system, is the close degree of temperature control realized and the uniformity of the temperature condition that can readily be maintained throughout the fluidized bed. Under normal conditions there is no appreciable temperature difference between parts of the bed, nor is there any appreciable temperature difference between the gaseous phase and the solid phase.
This is believed to be highly important, because it means that the heat released is immediately distributed throughout the entire bed. Thus there is no observable local overheating at any point or points, resulting in fusion and locking-up of the gold values. Furthermore, heat losses are relatively low, leading to the belief that roasting can be properly carried out on sulfide concentrates containing as little as 12 per cent or even less sulphur and without the use of purchased fuel.
At Cochenour Willans, the FluoSolids reactor consists of a cylindrical shell built of -in. steel plate 18 ft. high with an inside shell diameter of 8 ft. 8 in., reduced to an effective inside diameter of 6 ft. 8 in. by a lining of 9-in. firebrick backed with 3 in. of insulating brick. The domed top of the reactor, the hot-air ducts, and the cyclone dust collectors are lined with a castable refractory cement. Appropriate openings are provided, piercing the steel shell and lining for pressure taps, thermocouples, and feed and discharge connections.
There are two outlets in the dome top of the reactor, one, 14 in. in diameter, being a gas discharge to the cyclone dust collectors and the other, 8 in. in diameter, being merely an auxiliary stack opening which is used only when the reactor is being preheated prior to being started up. Both are lined with refractory material. The main 14-in. gas line leads to two cyclone dust collectors arranged in series. The stack is 14 in. in diameter, built of 1/8-in. steel plate, and is 125 ft. high.
In the base of the reactor shell there is fitted a steel perforated constriction plate. This plate is lined with castable refractory cement and contains 120 cup-shaped orifices in each of which a 3-in. Korundal sphere is seated. These spheres act as distributing valves for the diffusion of the air throughout the bed and as check valves when the reactor is shut downor, in the case of power failure, to prevent calcine from passing down through the perforations into the conical air chamber below. Through the conical base and constriction plate an 8-in. pipe is fitted with an external butterfly for emptying the reactor when required. All external piping is of wrought iron.
Three quench tanks are provided to receive, respectively, the hot calcine from the reactor, the dust from the first cyclone, and the relatively finer dust from the second cyclone. All the quench tanks discharge to acommon, screened receiving tank, the contents of which are pumped by a2- in. rubber-lined sand pump to the cyanide plant.
The arrangement of equipment directly ahead of the reactors is as follows: Flotation concentrates are pumped first to a small thickener then to a small rotary vacuum filter, with filter cake, at 12 per cent moisture or less, going to the reactor and with thickener overflow and filtrate going either to waste or back to the flotation circuit. The filter cake is fed to a 3-ton steel bin through a -in. vibrating screen and is introduced into the reactor by a variable speed, Coghill-type feeder, consisting of a vibrating hopper and an 8-in.-diameter by-18-in. long ribbon-type screw conveyor, suitably designed to act also as a seal to prevent gas escaping from the reactor at this point.
The current tonnage of concentrates handled in the reactor is 8 tons per day assaying about 6 oz. gold per ton. Since the reactor has a rated capacity of 15 tons per 24 hr. (actually considerably more), it is operated on two shifts only. The reactor can be shut down and started up without difficulty and without any apparent disturbance in metallurgy or extraction.
Because the Cochenour Willans ore is self-roasting, temperatures are controlled at about 1100F. by injecting water into the bed. Thermocouple connections at various points in the furnace permit automatic recording of temperature at about 5-min. intervals. Manometers connected above and below the constriction plate and at the top of the furnace indicate the pressure drop between various parts of the reactor.
For information on the cyanidation of the calcine see the general mill description in Chap. XV. The Dorrco FluoSolids reactor is also available in multicompartment design, the top or feed compartment serving to preheat the ore and effect preliminary decompositions, such as dehydration or the elimination of arsenic, the middle compartment being the main reaction chamber, while the lower compartment serves to cool the calcine and preheat the incoming gases. The material passes down from one compartment to the next through overflow standpipes, and the countercurrent flow of gases and solids makes it possible to effect considerable heat economy where this is desirable.
During the development stages of the roasting process at Lake Shore it was found that commercial cyanide extractions from the laboratory calcines could be obtained only from roasts in which salt had been added to the charge. This information was later confirmed in both the pilot plant and the commercial plant employing Edwards roasters.
The exact nature of the chemical reaction produced by salt, which has such beneficial effects in the cyanidation of the calcine, is not clear. While it can be demonstrated consistently on the Lake Shore concentrates, roasting tests carried out at Lake Shore on concentrates from mines in other camps failed to show any improvement when salt was added to the charge. These outside concentrates produced commercial cyanide extractions when roasted without salt.
The high degree of heat conservation in this type of furnace requires this step with ores of more than a certain critical calorific value in order to maintain the temperature at a predetermined figure.
A fourth effect was noted at the time of the laboratory roasting work, but the investigation of sizing technique has rendered this effect difficult to interpret. The calcine from salt roasts, when given the normal infrasizing treatment, appeared to be finer than the calcines produced with no salt.
The use of salt in roasting a gold-bearing concentrate requires a certain amount of caution. A process which depends on the use of excess NaCl and lime has been described for the extraction of gold from a concentrate by volatilization in a roaster. To avoid loss of gold, the amount of salt added must be held at such a figure that this volatilization does not occur.
After heating arsenical ores or concentrates mixed with soda ash equal to 5 to 10 per cent of their weight in the absence of air for a period of 20 to 60 min. at temperatures between 950 and 1200F. and then quenching in water, residues very low in gold values could be obtained by cyanidation of the calcine.
It is known that gold forms soluble sulpharsenate and polysulphide compounds, and it is believed that the gold was taken into the quench solution through such agency. It could be precipitated from the solution by gentle aeration, particularly in the presence of a trace of manganese salt. Such precipitates assayed from 30 to 40 oz. per ton in gold, and the gold so precipitated was not readily soluble in cyanide solution.
There appeared at the time to be a number of difficulties attending the commercial adaptation of such a process, but as new roasting techniquesare developed, it is probable that further work along these lines will be carried out.
In general, serious gold loss by volatilization is not experienced in roasting gold ores or concentrates provided that the chloride content of the feed is below a certain critical figure (at Lake Shore this was 30 lb. per ton approximately).
Considerable gold loss may be experienced in the case of arsenopyrite ores, however, if the temperature is allowed to rise too rapidly during the fuming off of the arsenic. N. S. Spence of the Department of Metallurgy,Queens University, Kingston, Ontario, describes certain tests carried out in a laboratory muffle furnace as follows.
Samples for each run weighed 150 grams. A gas-fired muffle was used, which was brought up to the desired temperature before charging the roasting dishes. Every run was done in duplicate. The temperature was recorded every 5 min., and the average calculated. The temperature used was that of the atmosphere of the muffle directly above ( in.) the roasting dishes and was read from a thermocouple pyrometer. Roasting was continued in every test until no trace of SO2 could be detected by smell in the air above the dishes. This period of time averaged 95 min. During roasting, the charges were rabbled carefully every 5 min., great care being taken to avoid dust loss. Duplicate assays were run on the roasted product, and knowing the weight and the assay of each dish, the loss was calculated.
The usual method of process and treatment of the calcines (by calcination) from the roasting of gold ores is to cool in air, quench the moderately cooled material in water, and subject it to a light grind with pebbles or balls to break up agglomerated particles, sintered prills, etc. The ground pulp is then passed over corduroy or other types of blankets to trap any free gold released in the roasting operation, and then cyanided, by either continuous or batch system.
The above procedure is frequently modified to suit local conditions, both blanket, concentration and grinding being omitted in certain cases, while in others water and even acid washes followed by filtration before repulping in cyanide solution are resorted to.
As it leaves the roaster the calcine contains less than 0.1 per cent in soluble sulphur but as much as 2.6 to 3.0 per cent soluble sulphur (as sulphate) and a variable amount of chloride, which probably never exceeds 0.15 per cent.
The calcine has an acid reaction, and before it can be handled in contact with iron and steel, it must be neutralized. This is carried out in a preliming step in a 7- by 7-ft. turbo-agitator, which oxidizes the ferrous salts and precipitates the corrosive sulphates and chlorides present as hydrates. About 60 lb. lime per ton of calcine is used to bring the solution strength to 1.0 lb. CaO per ton.
There are two stages of aeration in the treatment of the calcine pulp. The first stage is in the preliming agitator, where the pulp is given a thorough aeration. This aeration oxidizes the ferrous iron in the pulp to the ferric condition, in which form iron has a negligible cyanide consumption and no oxygen consumption. It also serves to saturate the pulp with oxygen before any cyanide is added.
The second stage of aeration is in the calcine agitation. At this point it is necessary only to maintain the oxygen content of the pulp. Aeration in the calcine agitators is obtained by blowing compressed air into the pulp through perforated pipes wrapped with six layers of canvas. These aerators are made of 8-in. lengths of 1-in. pipe, four per agitator. The aerators are located as near the bottom of the tank as possible to take full advantage of the increased solubility of oxygen in water due to pressure and to provide a maximum time of contact between the pulp and the rising column of bubbles.
It will be recalled that the flotation concentrate contains certain free gold values which resisted dissolution during the 60 hr. of main plant cyanidation. These values, after roasting, dissolve rapidly and completely.
The same rapid dissolution also applies to the values originally completely occluded in pyrite. The roast changes the pyrite grains into porous friablehematite which allows the cyanide solution to penetrate to and dissolve most of the gold values.
Test work has indicated quite clearly that cyanidation of the Lake Shore calcine is virtually complete after as short a treatment as 6 hr. After 9 hr. treatment, further time of contact will not lower the cyanide residue assay.
In plant-scale operations, the handling of a small tonnage of calcine presented certain problems. At a dilution of 2 to 1, a single 24- by 24-ft. agitator would, theoretically, furnish over 24 hr. agitation. However, short-circuiting would be so serious as to render any such treatment useless. It had been found in the main plant circuit that even six agitators in series did not completely overcome short-circuiting. However, a compromise was made in designing the calcine treatment plant. Four 12- by 8-ft. agitators were used in series. This provided some 24 to 36 hr. treatment. Thus, by more than trebling the time of contact known to be necessary for dissolution of the values, it was believed that the short-circuiting due to the use of only four tanks would be overcome.
The continuous calcine treatment required the use of a small Oliver filter. In view of the small tonnage to be handled, filtering costs per ton of calcine were high. The operation of the small filter was a constant source of inconvenience.
The high sulphate content of the calcine pulp gave rise to another problem. As the calcine pulp cooled during the agitation, CaSO precipitated. This formed a cement to bond together a heavy build-up of calcine around the rim of the agitators. The removal of this build-up necessitated shutting down the agitator and chipping off the accretions, which at times were well over a foot thick.
The same agitators are used as in the continuous process, but for batch treatment they are arranged as two pairs of tanks. The pulp from the calcine mill is pumped to one of the pair of tanks until they are filled. The flow is then diverted to the other tanks.
Cyanide is added during the filling period so that, when the batch is filled, the cyanide is approximately at full strength. Additional cyanide is added as required to maintain a cyanide strength equivalent to 0.5 to 0.7 lb. KCN per ton of solution during the treatment.
The time of treatment, i.e., from the time the tanks are filled until the pulp is run off to the filters, varies from 12 to 20 hr. but is never less than 12 hr. This allows a sufficient margin of safety at all times.
One of the regular main plant filters is used to filter the calcine pulp. The mainplant pulp is cut off shortly before the calcine batch is finished, and the filter is allowed to work out any solids remaining in its tank. Then the calcine pulp is fed to it. Once batch agitation was instituted, most of the difficulty of a mechanical nature disappeared from the calcine cyanidation.
The precautions taken in cleaning out the residual main plant pulp before using a filter for calcine are necessary to prevent the calcine filter discharge from being diluted by the main plant pulp. It is desirable, for control purposes, to have a reliable assay of each batch of calcine residue.
Calcine pulps are considerably more difficult to filter than the normal main plant pulp. The spongy porous nature of the calcine adds to the difficulty of washing the cake. However, with careful operation it is possible to recover over 99.5 per cent of the dissolved values in a single stage of filtration.
For satisfactory filtration, the calcine cake must not exceed in. in thickness. Wash water must be used in sufficient quantity to cover the cake completely at all times. Failure to keep the cake covered results in the formation of cracks. Cracks in the filter cake provide a path of negligible resistance and so decrease the amount of wash water passing through the cake.
The filters are fitted with five spray pipes. The first three carry barren solution and the last pair fresh water. The calcine treatment is carried out in a solution 0.5 to 0.7 lb. per ton of cyanide as KCN. Little improvement in gold extraction can be gained by the use of higher solution strengths, which do, however, result in higher cyanide consumption.
The roasted calcine withdrawn from the furnaces is collected in specially designed trucks. These act as cooling and storage bins, as well as a means of conveyance, and are fitted with a special bottom-discharge arrangement. By the use of these trucks the problems of cooling, storing, and dusting were eliminated, as no intermediate transfer of the calcine from truck to bin or bin to agitator was necessary. As each truck is filled with calcine, it is sampled and tested for unroasted calcine.
About 330 gal. water and 200 lb. of 90 per cent sulphuric acid are added to the agitator, the paddle gear of which is then set in motion. Seven trucks of calcine, equivalent to 2.45 tons, are fed to the agitator, and agitation proceeds for 1 hr., after which the acidity of the solution is tested. The acid-treated pulp is fed by hose from the agitator to two 5- by 2-ft. box vacuum filters, each box taking about 22.5 lb. per sq. ft. of filter area.
A dry vacuum pump connected to the filtrate receiver is put into operation. Calcine pulp is fed to the filters to within 1 in. of the top of the boxes, and the vacuum valves are then opened to allow filtration to commence. When the original copper sulphate solution has been filtered from the calcine and the cake is surface dry, a water wash of 2 in. is run into the boxes, followed by five washes of 1 in. When full, the vacuum receiver is emptied via a 1-in. acid-resisting pump, which delivers the copper-bearing solution to two sand clarifier storage tanks, each 6 by 4 by 5 ft. 3 in. The clarified solution from these tanks is fed at controlled rates to two copper- precipitation boxes, wherein the copper is precipitated from solution on steel scrap. The barren solution is allowed to ran to waste.
The precipitated or cement copper is cleaned up twice a month and is one of the materials still dispatched overseas for realization, since all attempts to find a local market were unavailing, owing to the association of arsenic with the copper.
The cyaniding of the calcine is effected in three mechanical agitators, while the gold-bearing cyanide is separated from the calcine by means of a 3- by 2-ft. Denver rotary filter. The filtrate drawn from the Denver filter is pumped to two sand clarifiers, whence it gravitates to five filiform zinc-extractor boxes arranged in series. The extractor-box effluent gravitates to a storage sump, from which solution is drawn to supply the sprays on theDenver filter and for make-up solution required in the agitators.
The zinc-gold slime cleaned up from the extractor boxes is treated with sulphuric acid for the removal of the zinc and subsequently with nitro- sulphuric acid to remove any copper which has found its way into the cyanide section of the plant from the acid leaching section. The solution siphoned off after this operation is delivered to a cyanide cement precipitation box of three compartments, steel scrap again being the precipitant. The gold slime is then transferred from the acid vat to a filter box in which it is dewatered. The dewatered gold slime is calcined and subsequently fluxed as follows: gold slime (by weight) 58 per cent, borax 20 per cent, sand 12 per cent. The manganese dioxide is 10 per cent. The fluxed gold slime is smelted in a reverberatory furnace and poured into button molds. The buttons are then remelted in a carborundum crucible, granulated in water, and fluxed with 42 per cent (by weight) borax, 33 per cent sand, and 25 per cent sodium nitrate. This flux serves to remove bismuth which is present in the gold.
An interesting problem confronting metallurgists is how to reduce the gold loss in the residues from the cyanidation of roasted calcines. An extensive investigation into this matter is reported in two recent papers published in Australia: The Condition of Refractory Gold in Lake View and Star (Kalgoolie) Ore by N. I. Haszard and Roasting and Treatment of Auriferous Flotation Concentrates by A. F. B. Norwood. The conclusion reached, as reported in the first of the above papers, was that of the three forms in which the gold occurs in Lake View and Star concentrates:
Apart from the evidence of microscopic study, roasting separately various sized fractions of the concentrate gave practically identical results, which again points to the extremely fine state of subdivision in which this refractory gold occurs. A series of tests also showed that the presence of calcium salts in the calcine had an adverse effect on cyanide extraction, presumably through the formation of compounds that reduce the porosity of the iron oxides resulting from the decomposition of the pyrite during roasting.
In the second paper, which reports a continuation of the same investigation, a rather detailed study was made of the chemical reactions taking place at various stages in the plant roaster by taking samples at various points along the hearth. An analytical method is described for determining elementary sulphur, sulphur as pyrite, and sulphur as pyrrhotite, also the proportions of magnetitie and hematite present in the calcine.
Tests were next carried out in which pyrite and pyrrhotite were roasted in sulphur dioxide. Since the reaction is endothermic, the temperature of the grains cannot rise above that of the furnace and the harmful effects of flash roasting is avoided. While the cyanide results were little better than the best roasting in air, it is worth noting that the magnetic or black roast is quite asamenable to cyanidation as the conventional red roast.
As a result of tests carried out at various temperatures, it was concluded that the high cyanidation tailings obtained following the higher temperature roasts are due to recrystallization of the iron oxide which destroys the porous structure induced by roasting and consequently locks up the submicroscopic particles of gold in dense crystals of hematite. Various methods for reducing the rate of combustion in plant furnaces are discussed.
Residual values of about 1.8 dwt. per ton appear to be the lower limit with present practice. While laboratory attempts to sulphate roast have not been very successful, the conversion of the calcine to ferric sulphate by strong sulphuric acid and subsequent ignition back to ferric oxide provides a material which will give almost 100 per cent extraction to cyanide. It is the opinion of the author that future improvements in recovery would appear to lie in the direction of devising a method of speeding up the oxidation of pyrites at temperatures below 500C.
The gold can be removed from calcines by volatilization with salt in an oxidizing atmosphere, and also by smelting with lead, but there are economic and technical problems involved in both of these schemes.
Tests carried out by the staff of Nepheline Products, Ltd., in Canada in cooperation with F. P. Archibald to determine the solubility of gold-silver-arsenic alloys in the cyanide solution led to the following conclusion:
Arsenic, alloyed with gold-silver alloys, aids rather than hinders dissolution of gold in cyanide solution, and such could be ruled out as a cause of refractory behavior of gold-bearing arsenical ores. The presence of particles of precious metal in cyanide solution residues remains unexplained.
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Grades piano scales and chord performances. Requires usb connection from keyboard piano. Not a teacher substitute though, but perhaps I should try and make stand on its own that way?DM soI can send you a free download link: reddit u/happypennygames or instagram @happypennygames or twitter @happypennygames
https://arltb.itch.io/5-levels-of-lockdownHello! This is a game that COVID-19 and the lockdown inspired me to make. I thought it would be funny. Now that I made it, I think it's quite fun to play. If you choose to play this game, you must try to survive all 5 levels of lockdown without dying of hunger, thirst, or, of course, disease. Also try not to run out of power or things can get difficult.If you're still alive when the timer runs out, you've made it through the level.How the game works is you have to get money to buy food at the grocery store, water at the water store and power at the power store, all while avoiding coming into contact with other people and getting infected. You also need to watch your surroundings because if you touch someone else's private building, you'll also get infected. And if the police catches you going off the yellow path, you'll be fined. And you must remember to stop at you're house to recover. You have to keep track of everything or you'll die one way or the other.Each level last's 100 seconds, and one last thing is if you happen to run out of power, you still have a chance, but you'll have to find your way without any light. In other words, you'll need to rely on memory of the path.I'm not sure of it's user friendly enough, or good enough in general. I made some updates.So any opinions?Thanks for reading
For thousands of years the word gold has connoted something of beauty or value. These images are derived from two properties of gold, its colour and its chemical stability. The colour of gold is due to the electronic structure of the gold atom, which absorbs electromagnetic radiation with wavelengths less than 5600 angstroms but reflects wavelengths greater than 5600 angstromsthe wavelength of yellow light. Golds chemical stability is based on the relative instability of the compounds that it forms with oxygen and watera characteristic that allows gold to be refined from less noble metals by oxidizing the other metals and then separating them from the molten gold as a dross. However, gold is readily dissolved in a number of solvents, including oxidizing solutions of hydrochloric acid and dilute solutions of sodium cyanide. Gold readily dissolves in these solvents because of the formation of complex ions that are very stable.
Gold (Au) melts at a temperature of 1,064 C (1,947 F). Its relatively high density (19.3 grams per cubic centimetre) has made it amenable to recovery by placer mining and gravity concentration techniques. With a face-centred cubic crystal structure, it is characterized by a softness or malleability that lends itself to being shaped into intricate structures without sophisticated metalworking equipment. This in turn has led to its application, from earliest times, to the fabrication of jewelry and decorative items.
The history of gold extends back at least 6,000 years, the earliest identifiable, realistically dated finds having been made in Egypt and Mesopotamia c. 4000 bc. The earliest major find was located on the Bulgarian shores of the Black Sea near the present city of Varna. By 3000 bc gold rings were used as a method of payment. Until the time of Christ, Egypt remained the centre of gold production. Gold was, however, also found in India, Ireland, Gaul, and the Iberian Peninsula. With the exception of coinage, virtually all uses of the metal were decorativee.g., for weapons, goblets, jewelry, and statuary.
Egyptian wall reliefs from 2300 bc show gold in various stages of refining and mechanical working. During these ancient times, gold was mined from alluvial placersthat is, particles of elemental gold found in river sands. The gold was concentrated by washing away the lighter river sands with water, leaving behind the dense gold particles, which could then be further concentrated by melting. By 2000 bc the process of purifying gold-silver alloys with salt to remove the silver was developed. The mining of alluvial deposits and, later, lode or vein deposits required crushing prior to gold extraction, and this consumed immense amounts of manpower. By ad 100, up to 40,000 slaves were employed in gold mining in Spain. The advent of Christianity somewhat tempered the demand for gold until about the 10th century. The technique of amalgamation, alloying with mercury to improve the recovery of gold, was discovered at about this time.
The colonization of South and Central America that began during the 16th century resulted in the mining and refining of gold in the New World before its transferal to Europe; however, the American mines were a greater source of silver than gold. During the early to mid-18th century, large gold deposits were discovered in Brazil and on the eastern slopes of the Ural Mountains in Russia. Major alluvial deposits were found in Siberia in 1840, and gold was discovered in California in 1848. The largest gold find in history is in the Witwatersrand of South Africa. Discovered in 1886, it produced 25 percent of the worlds gold by 1899 and 40 percent by 1985. The discovery of the Witwatersrand deposit coincided with the discovery of the cyanidation process, which made it possible to recover gold values that had escaped both gravity concentration and amalgamation. With E.B. Millers process of refining impure gold with chlorine gas (patented in Britain in 1867) and Emil Wohlwills electrorefining process (introduced in Hamburg, Ger., in 1878), it became possible routinely to achieve higher purities than had been allowed by fire refining.
The major ores of gold contain gold in its native form and are both exogenetic (formed at the Earths surface) and endogenetic (formed within the Earth). The best-known of the exogenetic ores is alluvial gold. Alluvial gold refers to gold found in riverbeds, streambeds, and floodplains. It is invariably elemental gold and usually made up of very fine particles. Alluvial gold deposits are formed through the weathering actions of wind, rain, and temperature change on rocks containing gold. They were the type most commonly mined in antiquity. Exogenetic gold can also exist as oxidized ore bodies that have formed under a process called secondary enrichment, in which other metallic elements and sulfides are gradually leached away, leaving behind gold and insoluble oxide minerals as surface deposits.
Endogenetic gold ores include vein and lode deposits of elemental gold in quartzite or mixtures of quartzite and various iron sulfide minerals, particularly pyrite (FeS2) and pyrrhotite (Fe1-xS). When present in sulfide ore bodies, the gold, although still elemental in form, is so finely disseminated that concentration by methods such as those applied to alluvial gold is impossible.
Native gold is the most common mineral of gold, accounting for about 80 percent of the metal in the Earths crust. It occasionally is found as nuggets as large as 12 millimetres (0.5 inch) in diameter, and on rare occasions nuggets of native gold weighing up to 50 kilograms are foundthe largest having weighed 92 kilograms. Native gold invariably contains about 0.1 to 4 percent silver. Electrum is a gold-silver alloy containing 20 to 45 percent silver. It varies from pale yellow to silver white in colour and is usually associated with silver sulfide mineral deposits.
Gold also forms minerals with the element tellurium; the most common of these are calaverite (AuTe2) and sylvanite (AuAgTe4). Other minerals of gold are sufficiently rare as to have little economic significance.
Of the worlds known mineral reserves of gold ore, 50 percent is found in South Africa, and most of the rest is divided among Russia, Canada, Australia, Brazil, and the United States. The largest single gold ore body in the world is in the Witwatersrand of South Africa.
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1848 - 1852 was unquestionably one of the most influential periods of westward movement in America. The discovery of gold at Sutter's Mill in California in 1848 lead to the 1849 Gold Rush. The dream of striking it rich and starting a new life inspired hundreds of thousands of adventurous souls from around the world to uproot and resettle in the west. This 1849 gold rush ultimately sped up the process of California moving from a U.S. territory to gaining statehood faster than neighboring areas.
Almost fifty years later gold was discovered in the Klondike River in the Alaskan Yukon leading to yet another massive rush of people seeking their fortune in a vast, unknown wilderness. This Alaskan rush has been labeled several names, but we like to think of it as The Last Great Gold Rush. Truly brave and unwavering frontiersmen forging their way through the mountains to make their own destinies.
All mine locations were obtained from the USGS Mineral Resources Data System. The locations and other information in this database have not been verified for accuracy. It should be assumed that all mines are on private property.
Primary: Gold Primary: Silver Secondary: Arsenic Tertiary: Fluorine-Fluorite Tertiary: Lead Tertiary: Zinc Tertiary: Copper Tertiary: Barium-Barite Tertiary: Tungsten Tertiary: Antimony Tertiary: Mercury Tertiary: Thallium Tertiary: Tellurium Tertiary: Bismuth Tertiary: Tin Tertiary: Molybdenum
Record Type: Site Operation Category: Producer Deposit Type: Replacement Operation Type: Surface Year First Production: 1938 Year Last Production: 1999 Discovery Year: 1933 Years of Production: Organization: Significant: Y Deposit Size: M
Type: L Description: The main Getchell deposit within the fault has been drilled to a depth of 600 m down dip from the original surface, and remains open down dip. There is a 'Main Vein' which is a dominant structure with a distinct footwall, complexed by several conjugate veins to the west. Sub-parallel, mineralised structures have also been found up to 200 m into the footwall of this main structure, while alteration, fault gouge and mineralisation occur up to 500 m to the east into its hangingwall (FirstMiss Gold Inc., 1993). Movement on the Getchell Fault has been both normal and dextral strike-slip (McCollum & McCollum 1990). On the basis of the relative displacement of the Palaeozoic sediments and the Cretaceous granodiorite of the Osgood Mountain Stock it is believed that the Getchell Fault is a reactivated older structure (D Bond, Pers. comm., 1993). The most recent displacement has taken place during the Miocene to present Basin and Range movement, representing further reactivation of an older structure. The fault cuts all three main stratigraphic units found within the pit, as well as the Osgood Mountain Stock. Altered blocks of granodiorite, rimmed by the skarn assemblage, are faulted downwards along the footwall structure into the Getchell Fault Zone and subsequently mineralised with gold (FirstMiss Gold Inc., 1993).
Type: L Description: Gold mineralization is generally found at the intersection of a number of high-angle and low-angle fault sets. The low-angle faults and associated folds are the result of Devonian and Permian-age compressional events and the higher angle faults and fracture sets formed during Tertiary extension. Mineralization is both structurally and stratigraphically controlled. The Getchell fault is a zone of overlapping fractures which have an overall strike of N10W. Hotz and Willden (1964) offer evidence for up to 3500 feet of left lateral strike slip displacement and only a relatively small amount of dip slip movement along the Getchell fault. McCollum and McCollum (1991) indicate that the sense of movement on the Getchell fault is right lateral. The Getchell Fault Zone is a complex system of sub-parallel, high angle faults which is at least 500 m wide. The zone is made up of a number of fault planes, separated by brecciated gouge and characterised by intense clay alteration, and by brecciation in the hangingwall.
Alteration Type: L Alteration Text: Alteration comments: there is a metamorphic aureole around the Osgood Mountains granodiorite which has produced in the surrounding shaly rocks a mineral assemblage consisting of cordierite-, biotite-, and andalusite-hornfels. Locally limy beds are recrystallized and calc-silicate minerals are developed. Hydrothermal alteration consists chiefly of decarbonatization accompanied by silicification in the limestone beds. Cordierite, andalusite, and biotite of the metamorphic aureole are altered to sericite and/or chlorite. Igneous dikes and portions of the main stock are altered such that plagioclase is altered to sericite and kaolinite and biotite is altered to sericite, chlorite, and pyrite.
Ore: Epidote Ore: Weilite Ore: Pharmacolite Ore: Haidingerite Ore: Guerinite Ore: Arsenolite Ore: Laffittite Ore: Galkhaite Ore: Getchellite Ore: Coloradoite Ore: Bismuthinite Ore: Garnet Ore: Chalcocite Ore: Covellite Ore: Sphalerite Ore: Galena Ore: Ferrimolybdite Ore: Molybdenite Ore: Cassiterite Ore: Fluorite Ore: Gypsum Ore: Barite Ore: Silver Ore: Electrum Ore: Orpiment Ore: Pyrite Ore: Scheelite Ore: Pyrrhotite Ore: Arsenopyrite Ore: Marcasite Ore: Magnetite Ore: Stibnite Ore: Chlorite Ore: Sericite Ore: Chabazite Ore: Calcite Ore: Hubnerite Ore: Cinnabar Ore: Ilsemannite Ore: Gold Gangue: Realgar
Comment (Development): Prospectors Edward Knight and Emmet Chase discovered gold in 1933 and located the first claims in 1934. With the financial backing of Noble Getchell and George Wingfield, the Getchell Mine, Inc. was organized in 1936 and was brought into production in 1938. In 1938, the mining rate was about 500 tpd of oxide ore and 150 tpd of sulfide ore. Sulfide ore was roasted at 1500 degrees Fahrenheit for one hour and fifteen minutes preparatory to cyanidization. In 1941, a Cottrell electric precipitating unit was installed to save the arsenic that was liberated by roasting the sulfide ore, and in 1943-45, when government wartime restrictions forced the shutdown of many gold producers, Getchell mine was permitted to continue operations as a producer of "strategic" arsenic. In 1943, arsenious oxide was being produced at the rate of 10-25 tpd from furnace fume. Also in 1942, a 227 tonne scheelite flotation plant was built to recover tungsten from Getchell ore. A slack labor supply, and high supply costs forced the gold operations to cease at the end of World War II. The US Bureau of Mines developed a carbon recovery process on site and the mine reopened in 1948 with expanded mill capacity and more underground development, but closed again in mid-1950 when known oxide reserves were exhausted. Gold production was suspended in 1951. From 1951-56, the mill processed tungsten ores mined from throughout the district. Tungsten production ceased in 1957. in 1960, Goldfield Consolidated Mines Co. purchased the interests in Getchell Mine, Inc. from the estates of Wingfield and Getchell. Gold production resumed in June 1962 and continued to December, 1967, when the mine was closed and the mill dismantled. Cyprus Mines formed a joint venture with Goldfield in 1970, with Cyprus as operator. Cyprus dropped the property at the end of 1971. Conoco leased the property from Goldfield in 1972 and completed exploration including over 300 drill holes. Metallurgically difficult sulfide reserves were identified during this program. Conoco subleased the property from 1975 to 1978 to General Electric Co. who conducted tungsten exploration along the margins of the Osgood Stock. In 1981, Conoco purchased the property from Goldfield Corp., but by 1983 had sold the property to First Mississippi for $5 million. At that time the property consisted of 14,100 acres of fee land and almost 5000 acres of unpatented claims, and reserves at the time of purchase were in excess of 750,000 ounces of gold. Mining feasibility and metallurgical studies were initiated in 1984. Heap leaching of waste rock dumps from previous mining operations commenced at the end of fiscal 1985, producing 91 ounces of gold in that fiscal year. By mid-1985, the Getchell property had increased the area of unpatented claims to 13,900 acres. In May, 1987, the board of First Mississippi Corp. authorized open pit mine development and construction of a new mill utilizing autoclave technology to process 3000 tons of ore per day. The mill was completed and production resumed in 1989 combining a traditional cyanide leach circuit with pressure oxidation. The mill started up on oxide ore in February, 1989. Sulfide ore was run through the first pressure oxidation autoclave in April, 1989 followed by the start up of the other two autoclaves in May and June, 1989. By the end of fiscal year 1989, project capital costs stood at $90.3 million, 14% over the June 1987 feasibility study estimate. In fiscal year 1989, overall gold recovery for combined oxide and sulfide mill ores was 89.8%. Heap leaching of waste rock from previous mining operations was completed in fiscal year 1989. Heap leaching continued beyond this date using oxide reserves from the Summer Camp orebody discovered in 1985.
Comment (Development): Production of oxide open pit ore commenced at the nearby Turquoise Ridge mine in 1991 and in the same year, an underground orebody adjacent to the pit area. This ore was to be mined when the pit level was deep enough to provide lateral access. In 1995, FirstMiss Gold changed its name to Getchell Gold. Underground production commenced at Turquoise Ridge Mine in May 1998. On May 27, 1999 Placer Dome completed a merger with Getchell Gold Corporation, resulting in Placer Dome owning 100% of the Getchell gold property. Gold production has been suspended since July 1999 and the property is on care and maintenance. Production from approximately 58% of the property is subject to a 2% net smelter return royalty payable to Franco Nevada Mining Corporation Ltd. Placer Dome wrote off the carrying value of the property in 2001. On October 25, 2001, Newmont Mining Corporation and Getchell Gold Corporation signed a letter of intent under which Newmont would buy ore from the Getchell mine for processing at Newmont's adjacent Twin Creeks mine.
Comment (Identification): This is the same location as old record M055410. The Getchell property consists of the Getchell, Turquoise Ridge and N Zone deposits. The Main Pit has now encompassed the earlier Central and South Pits.
Comment (Deposit): The known gold deposits within the Getchell Trend are Carlin- type, sediment-hosted, replacement deposits containing micron gold. Gold mineralization at Getchell is associated with a curvilinear fault system that strikes NNW and dips 40? to 75? east, on the eastern flank of the Cretaceous Osgood granodiorite stock. The mineralized fault zone and the Cretaceous granodiorite both cut Palaeozoic sediments of the Cambrian Preble and upper Cambrian to lower Ordovician Comus Formations which both belong to the Transition Assemblage, and the Ordovician Valmy Formation of the Western Assemblage. Thermal metamorphism along the intrusive contact formed tungsten bearing garnet-diopside skarns, passing outwards into wollastonite calc-silicates and marble. In the southern parts of the Getchell Mine area the skarn is about 30 m wide adjacent to the granodiorite contact, passing out into marble. Pelitic shales of the Preble and Comus Formations are thermally metamorphosed to cordierite-andalusite bearing hornfels nearest the contact, grading outwards into a biotite-cordierite-andalusite interval, to an outer biotite zone. The Osgood Stock and associated hornfels and skarns are found in both the footwall and hangingwall of the mineralized fault zones. Gold mineralization is found in a number of different rock types generally at the intersection of a number of high-angle and low-angle fault sets. The low-angle faults and associated folds are the result of Devonian and Permian-age compressional events and the higher angle faults and fracture sets formed during Tertiary extension. Mineralization is both structurally and stratigraphically controlled. Gold is associated with arsenic, mercury, and to a lesser extent antimony, and commonly with pervasive decalcification, silicification and carbonaceous alteration. Gold is micron-scale generally intergrown with arsenical pyrite, which in turn, is encrusted in barren, diagenetic pyrite. Late stage realgar and orpiment are commonly associated with high-grade ores. The main deposit is confined to a zone nearly 7000 ft. long at the northern end of the Getchell fault zone. Deep exploration shows that the mineralization persists at least 1 km down-dip on the Getchell fault system and also occurs along the parallel Village fault. Maximum width of ore is 200 ft., with an average width of 40 ft. Within ore zones, gold occurs as native grains that range in size from <1 micron to nearly 1 mm, with smaller grains more abundant than larger grains. Most of the gold is intimately associated with the fine grained quartz-carbon matrix of the altered rock termed "gumbo" by Joralemon (1951). Of the sulfides, pyrite and marcasite are principal hosts to gold. As of 1951, the gold:silver ratio in bullion ranged from 2:1 to 134:1 and averaged 10:1 for the entire bullion production to that date. Joralemon (1951) observed microscopic metallic grains in the Getchell ore that he concluded were native silver, although the particles were so small that conclusive chemical tests were not possible. No other silver minerals have been recognized except for very rare grains of electrum. Geochemical work at the Getchell mine and vicinity has demonstrated that As-W-Hg anomalies occur in rocks and soils over the arsenic-gold deposits and that these anomalies are not broad haloes but are restricted to the mineralized area. The highest metal contents are found in oxidized iron-rich material along fractures and bedding planes in barren bedrock, lesser values in caliche coatings on exposed bedrock, and lowest but still anomalous values in soil.
Comment (Economic Factors): From 1938 to1996 the Getchell property produced 66.8 kilotonnes of gold and more than 1.2 kilotonnes of silver from 18361 kilotonnes of ore. In 1997, the remaining Getchell resource was estimated at 14,400 kilotonnes of ore containing 153 kilotonnes of gold and an unknown amount of silver and arsenic. This resource includes Getchell underground, stockpiles, unmineable resource in Main Pit, and North Getchell underground resource.
Comment (Geology): Geology comments: Bagby and Cline (1991) offer preliminary results from research which indicate that confining pressures on the Getchell ore system varied from approximately 370-430 bars either during, or at some time subsequent to mineralization. These fluid pressures are greater than those which are normally accepted as epithermal.
Comment (Workings): The mine has been developed by both underground and surface workings at various times during its production history. The Getchell deposit was developed by the North, Center, South, and Hansen Creek Pits. The Getchell underground is fully developed and is accessed from the Getchell open pit via two portals. The Getchell underground has a relatively short remaining mine life based on current proven reserves, although its life may be extended with the lower processing costs, additional exploration drilling, and engineering analysis. Mining methods for the Getchell underground is currently 100% drift-and-fill, as the last of the longhole ore was produced in 2005.
Comment (Commodity): Gangue Materials: realgar, orpiment, pyrite, scheelite, pyrrhotite, arsenopyrite, marcasite, magnetite, stibnite, ilsemmanite, cinnabar, hubnerite, calcite, chabazite, sericite, chlorite, barite, gypsum, fluorite, getchellite, galkhaite, laffittite, arsenolite, guerinite, haidingerite, pharmacolite, weilite, coloradoite, bismuthinite, cassiterite, molybdenite, ferrimolybdite, galena, sphalerite, covellite, chalcocite, garnet, epidote
Reference (Deposit): Long, K.R., DeYoung, J.H., Jr., and Ludington, S.D., 1998, Database of significant deposits of gold, silver, copper, lead, and zinc in the United States; Part A, Database description and analysis; part B, Digital database: U.S. Geological Survey Open-File Report 98-206, 33 p., one 3.5 inch diskette.
Reference (Deposit): Madden-McGuire, D. J., 1991, Stratigraphy of the limestone-bearing part of the lower Cambrian to lower Ordovician Preble Formation near its type locality, Humboldt County, North Central Nevada, in Raines, G. L., et al, eds., Geology and Ore Deposits of the Great Basin, The Geological Society of Nevada, Reno, p. 875-893.
Reference (Deposit): Berger, B. R. and Tingley, J. V., 1985, History of discovery, mining, exploration of the Getchell mine, Humboldt County, Nevada, in Hollister, V. F., ed., Discoveries of epithermal precious metal deposits, case histories of mineral discoveries vol. 1, Society of Mining Engineers, New York, P. 49-51.
Reference (Deposit): Berger, B. R., 1985 Geological and geochemical relationships at the Getchell Mine and vicinity, Humboldt County, Nevada, in Hollister, V. F., ed., Discoveries of epithermal precious metal deposits, case histories of mineral discoveries vol. 1, Society of Mining Engineers, New York, p. 51-59.
Reference (Deposit): McCollum, L. B. and McCollum, M., 1991, Paleozoic rocks of the Osgood Mountains, Nevada, in Raines, G. L., et al, eds., Geology and Ore Deposits of the Great Basin, The Geological Society of Nevada, Reno, p. 735-738.
Reference (Deposit): Bagby, W. C. and Cline, J. S., 1991, Constraints on the pressure of formation of the Getchell gold deposit, Humboldt County, Nevada, as interpreted from secondary-fluid-inclusion data, in Raines, G. L., et al, eds., Geology and Ore Deposits of the Great Basin, The Geological Society of Nevada, Reno, p. 793-804.
Artisanal small-scale gold miners (ASM) occasionally employ whole ore amalgamation by adding mercury into ball mills to recover gold. In this process, 2530% of the mercury added is lost to the environment. It is also inefficient less than 30% of gold is recovered. Amalgamation, followed by cyanidation, has been observed at many artisanal mining sites. This combination poses additional environmental and health consequences. Tests with ore samples from Talawaan, North Sulawesi, Indonesia indicate the possibility of replacing mercury by cyanidation in the ball mill, reaching gold extraction of 93% in 6h of leaching. The gold in the Indonesian ore sample is fine and less than 8% of gold recovery was obtained with gravity concentration of the ore ground 80% below 0.25mm, which is a reasonably fine grain size for artisanal gold operations. Replacing mercury addition with cyanidation in ball mills was implemented in one artisanal gold mining operation in Portovelo, Ecuador, achieving 95% of gold extraction in 8h of mill leaching. This technique demonstrated a drastic improvement in gold recovery. It was found to be a simple, inexpensive technique well accepted by local miners. The results from laboratory and field tests are promising; however a thorough investigation into the socio-economic and environmental aspects of this presented alternative must be conducted prior to introduction.