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Platinum in Alluvial Diggings.

Pyrites and other Sulphides.—The great question at the present day, to miners in New Zealand and other parts, is the economic saving and treatment of the metallic sulphides, as although in new districts lodes consisting of quartz and free gold are occasionally found, still they are getting scarcer and scarcer every day, and even with these, after a certain depth is reached, this free milling character changes. The quantity of gold that can be saved by quicksilver is only a fraction of the assay value of the whole mass, as it is only the free gold present in the quartz that can be amalgamated with the mercury, all that is combined with pyrites and other sulphides being lost. To get the full value from the ores requires a systematic working by which "low grade ores" (ores containing less the six pennyweights of gold to the ton of 2240lbs) can be cheaply and quickly concentrated; when this is done they are ready for treatment, and can be easily converted into ingots of pure metal, the so-called base metals with which the ore is combined being at the same time brought into use and made to yield their proportion of profit.

The loss of working gold ores, even with our most modern appliances of milling and amalgamation, is still enormous, not more than one-half of the gold contained in the ore which is worked is saved. The actual average yield of ores milled and smelted, calculated from Fossett's tables of seven years' work in Colorado, showed the average value of the ore by assay to be £7 18s. per ton, the average value per ton saved by milling and smelting £3, showing a loss of £4 17s. per ton, or more than 60 per cent. The gold caught on the copper plates, under the most favourable circumstances, is only 50 per cent, of the assay value of the ore; the gold from the blankets does not amount to more than 5 per cent.; so that when treating the most tractable of these ores, battery amalgamation does not secure more than 55 per cent, of the gold.

—(M. Eissler.)

Concentrates, containing pyrites and other sulphides, have been sent from Otago to Sandhurst, and yielded from 80 to 70 ounces of gold per ton—a fair average is about 9 ounces gold per ton. Roughly speaking, it may be taken that the quantity of gold lost in the tailings is equal to that saved.

'Concentrates, Concentrators or Vanners.—After .the pulp leaves the blankets, it is in many modern mills dis- page 2 charged on to vanners or concentrating tables. Of the many kinds of vanners in use, those of Hendy and Frue appear to give the best results.

Hendy's Concentrator.—Figure 1, Has this great advantage, that it is not usually necessary all the particles or grains of ore should be of the same size, but can be treated as they come from the blankets. Hendy's concentrator consists of a shallow pan 5 or 6 feet in diameter, supported in the centre by a vertical shaft, and made to oscilate by cranks on one side; these are joined by connecting rods with the side of the pan, which turns upon a vertical axis through a short distance for every revolution of the crank shaft. The bottom of the pan is raised in the centre to nearly the height of the rim, in order to facilitate the movement of particles towards the circumference. The tailings direct from the blankets are delivered by a trough to the hopper C, from which they pass through the pipe K and distributor D into the pan near the outer edge; and the rotation of D by means of two pawls attached to it, and acted on by teeth on the rim of the pan, causes this delivery of tailings to take place evenly at all points of the circumference; rake-like arms M rotate with D in order to stir the compact mass of sand, &c., which settles at the bottom. The crank makes 210 revolutions per minute, and the accumulated sulphides are discharged through the gate E, while the amalgam and mercury collect in the depression J. The machine concentrates 5 tons of tailings in twenty-four hours.

Frue Vanner.—Figure 2. The principle of this machine is a revolving india rubber blanket with high flanges, and having a side-shaking motion, so that the motion closely resembles that given to a shovel in vanning by hand. The rubber blanket moves against the flow of water, and the sulphides are removed by dipping into a tank of water underneath the machine. The pulp as it leaves the blankets is fed with the water by means of a spreader on to the rubber belt, and thus uniformly distributed across it. To the main shaft sufficient motion is given to impart from 180 to 200 revolutions a minute, with 1 inch throw. The uphill or progressive motion varies from 2 to 12 feet a minute, and the inclination is from 8 to 6 inches, in 12 feet, varying with the ore.

The capacity of a Frue vanner for concentrating is about 0 tons per 24 hours, passing through a 40 mesh sieve (= 1600 per square inch), so that two vanners are required for every five stamps. No sizing of the material is needed, the pulp passing directly from the stamps on to the copper plates and thence on to the vanners. One quarter horse-power is required for each machine, and one man can attend to sixteen machines without difficulty. When six machines are used, the cost of treating the ore is estimated at about tenpence per ton, when ready to flow on to the machine. The side motion of the belt throws off page 3 the gangue, but never moves the very finest mineral when it has once touched the belt. The cost of one of these machines landed in New Zealand is about £130.

The concentrates after leaving the vanners will contain from 9 to 70 ounces of gold to the ton, value £33 to £250, less the cost of extraction.

Methods for assaying and proportions of fluxes required will be found at pages 12 and 13.

The next process is roasting the concentrates, and for successful working the ore requires to be free from lumps. The roasting, and further treatment, could be done better in a town, as chemicals, coals, labour, &c., would come cheaper, the cost of carriage upon the concentrates is but a small matter, and with a good supply of material the process is continuous.

Roasting Furnace.—Figure 3, plan 4, section. Single roasting furnace. Figure 3 is the plan of a roasting furnace for 1 ton of sulphides at a charge. A is the hearth bottom, about 12 feet square. It is made of the hardest bricks laid edgewise, close together, forming a bottom 4 inches thick. There are four working doors C, which enable the roaster to reach all points of the hearth conveniently with light rakes. In the middle of the length at the bottom near the doors is a square discharging hole B, which is kept shut by a slide D, during the roasting. Below the floor (fig. 4) is an arch running the whole width of the furnace, through which the hole B passes. An iron car on rails receives the roasted ore, when discharged, and wheels it to the cooling place. The bridge E is from 10 to 12 inches wide and from 8 to 10 inches high; it separates the hearth from the fireplace; if possible, it should be built of fire bricks or fire lumps. The outside wall F is often made 12 inches thick, but it is better to make it 20 or 24 inches. The roof is generally 20 inches above the bottom at its greatest distance; if it is less, although the form of the furnace is improved, it is not so likely to be durable, unless the work is perfect. The length of the bricks gives the thickness of the arch, that is 8 inches. There are three openings H in the roof, each 10 inches in diameter, communicating with the chimney by the flue I, opposite to which is a small door K, for the purpose of cleaning the flue I from time to time. For the same purpose an opening must be prepared in front. The best way to regulate the draft is by means of a cover N on the top of the chimney. The sulphides are charged through the roof by means of a hopper.

About 24 hours is required to roast a ton of sulphides. Other furnaces are in use which will put through 15 to 20 tons in 24 hours.

Test for a Perfect Roast.—The object of roasting is to get rid of all the sulphur, arsenic, antimony, &c., that can be page 4 removed by the agency of heat, and to convert the iron, copper, &c., into oxides, thus freeing a portion of the gold and fitting the remainder for chlorination.

The roasted ore when thrown up with the hoe or shovel should show but few brilliant sparks, and a portion taken out should not smell like burning sulphur (sulphurous acid); also when a portion is added to water in a glass and either filtered or allowed to settle, the clear liquid should not give a blue precipitate with ferrocyanide of potassium, showing the presence of sulphate of iron, nor a reddish brown precipitate from the presence of copper sulphate; when both are present a greenish precipitate is produced. Sulphates are very injurious to successful chlorination.

A little hydrochloric acid added to some of the roast should not give off the well-known smell of sulphuretted hydrogen, showing the presence of sulphides, as these if present, would be acted upon in the subsequent process and precipitate the gold as sulphide at the wrong time.

A practical test of roasting can easily be made in an ordinary frying-pan, first coating the inside with a mixture of chalk and water and well drying. Mix the finely powdered ore with about its own bulk of fine charcoal; this is needful only when arsenic and antimony are present; still it helps to get rid of the sulphur and can do no possible harm. If the ore contains much sulphide of lead (galena) or sulphide of antimony (stibnite, antimony glance), add some fine sand, as without this addition the mineral while roasting would soon fuse, cake together, adhere to the pan, and ruin the assay. When all is thoroughly mixed, put the pan on the fire; stir well with an iron wire till the glowing ceases and no more sparks are given off; the assay will then appear of one colour, yielding to the stirrer like dry sand. Guard against too high a heat at first. If the roasted mineral is then examined with a magnifying glass or panned off in the usual way, a quantity of gold will be found free, ready for amalgamation.

Amalgamating Pans.—Figure 5. The object of the amalgamating pan is to combine the mercury and the free gold, with the smallest possible loss of mercury. Several kinds are in use—The Wheeler, The Hydrogen Amalgam (Molloy's patent), The Knox and Attwood's, &c.; but the "Combination" pan, fig. 6, appears to be the most successful. The bottom is of cast iron, steam jacketted, with wrought iron sides; the false bottom and mullers are of cast iron, and easily replaced when worn. A pan 6 feet in diameter will treat from 1 to 2 tons pulp at a time.

The roasted ore is introduced in quantity of about 20 cwt., with water sufficient to make the pulp adhere to a stick dipped into it without dropping off, and a slow motion imparted to the machine. After three hours grinding, or longer if found neces- page 5 sary, the pulp is heated by steam under the false bottom. At the same time a small quantity of a mixture of equal parts saltpetre and salammoniac is added together with the mercury (about 140lbs.). The machine is kept in motion for a further three hours while the process of amalgamation goes on; the pulp, then very fine, is diluted with water, and a few handfuls of caustic lime added, which greatly assists the coagulation of the particles of quicksilver; the diluted pulp, which now reaches within 8 or 4 inches of the top of the pan, is gently agitated for about twenty minutes, after which it is discharged, the machine being kept in motion; first by the upper plug-hole, and afterwards through the lower one into settlers. A bucket placed in front of the discharging hole catches any quicksilver or amalgam that may escape during the discharge. Unless a final cleaning up is intended, the quicksilver and amalgam are worked by hand, the lumps broken up, and the quicksilver, after straining through a canvas filter, is ready for use with the next charge. In this process sodium amalgam, by keeping the quicksilver bright, is sometimes found very beneficial.

Settler.—The pulp from the amalgamator is run into the settler, fig. 6. This is a circular vat, in which revolves the central axis, with arms attached; the object is to keep the heavier parts of the pulp thoroughly stirred. Water is introduced during the operation, and the pulp is then drawn off into tanks by the plugged holes as required, and allowed to settle. It still contains, in many cases, a considerable quantity of gold, and is added to the coarser sulphides from the final cleaning up, and is then ready for the chlorination process.

The amalgam is put into the amalgam safe, fig. 7. It consists of a vessel with sheet iron body and cast iron top and bottom. The cover is provided with protected perforations to admit the amalgam, and is usually secured with a padlock. Inside is a canvas strainer.

The Amalgam Retort and Condenser.—In order to lose as little mercury as possible, the amalgam, which contains from 86 to 40 per cent, of gold, is put into cast iron cylinders or retorts (fig. 8). The amalgam is first moulded by hand into cakes, about the size of an egg, and placed in the retort, the lid is secured, and the fire lighted. The mercury condenses in the tube A, which is surrounded by water contained in the tube B, and drops into the tub C, placed ready to receive it. When the retort has been at a cherry-red heat for two hours, the retorting is considered complete. As the retort cools, the water in the tub C would be forced over into the retort: to prevent this, a small tap is placed at D, which, by equalising the pressure, prevents any back flow.

The Newbury-Vautin Chlorination Process.—The apparatus first required is an iron, lead-lined barrel or drum, revolving upon its axis, and capable of holding 20 or 80 cwt. page 6 of the sulphides as they come from the settler, and able to bear a pressure of about 100lbs. to the square inch. In the centre are two valves directly opposite each other—one sufficiently large to pour in a stream of the ore to be acted upon; the other one, which is smaller, serves to introduce the compressed air, Chloride of lime and sulphuric acid are used to produce the chlorine, and a sufficient quantity of the former is introduced on the top of the ore, with water, and the cover screwed down: the barrel is then turned half round, so that the small valve is uppermost, the sulphuric acid added, and the indiarubber pipe leading from the air-pump quickly attached; compressed air is forced in until a pressure of about 60lbs. to the square inch is obtained; the valve is then screwed down and the hose disconnected. The barrel is now set slowly revolving, about 9 or 10 revolutions per minute, and the chlorine gas, given off from the decomposition of the chloride of lime by the sulphuric acid, impregnates the water: the gold is thereby converted into chloride of gold, the iron, copper, lime, &c., being also converted into chlorides, and all except the sand and silver, which latter will now be in the form of insoluble chloride, are dissolved in the water. The time occupied ranges from one to four hours, according to the nature of the ore.

The chlorinator, upon being stopped, is first connected by the smaller valve with a rubber hose which leads outside the building, the compressed air and gases are discharged, and the valve closed. The cover of the larger valve is next removed, and the barrel again set revolving; at each turn, it delivers a portion of its contents into a shoot leading to the leaching vat or filter; when nearly empty, a few pails of water are thrown in, which quickly remove any remaining ore.

The Improved Filter.—Figure 9 consists of two round cylinders A and C placed one within the other, and both provided with bottoms. They are fixed together, air-tight, by a horizontal hoop or flat ring of metal; the inner cylinder A is lined with stout sheet lead; the bottom is pierced with fine pin holes slightly enlarged on the side B; at D is an opening which is attached to a pipe, so that a partial or intermittent vacuum can be produced in the space between the cylinders from an exhaust-pump or by other means, thus ensuring rapid filtering. When all the liquid chlorides have run through, the tap E is opened, and the contents allowed to flow into a vat; a slight but continuous stream of water is kept playing upon the mass in A, by which all the chlorides are washed out. The solution can be tested from time to time as it runs away at E, and when the washing is complete, the water is cut off. The stuff in the filter is removed to the tailings heap, unless a sufficient amount of silver is present to pay for leaching out, which can be readily done by a solution of calcium hyposulphite or a strong solution of common salt.

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The gold and other chlorides are now in solution in the vat, at the bottom of which is a tap; through this the liquid slowly runs on to a filter of charcoal. Gold is precipitated, in the metallic state, on the charcoal, the iron, lime, magnesia, zinc, Ac., passing through unchanged. The charcoal, after drying, is burnt, and the gold collected and melted into an ingot, which, if the various operations have been carefully conducted, will generally be over 970 fine.

The importance of obtaining the "gold contained in pyrites and other sulphides" can hardly be over-estimated. It would, if well and carefully carried out, increase the yield of gold, upon the quantity of ore at present mined, at least 70 per cent.

The value of gold exported in 1886 was £939,648; in 1887, £747,891. After making allowance for alluvial gold, 70 per cent, upon the balance means a considerable annual sum, something between three and four hundred thousand pounds (£400,000).

The following opinions of well known men are worth attention:—

Mr. A. B. Paul (California) says: "The fact is, we are not working to save gold, but to crush rock."

Professor Egliston says: "In all the methods for the extraction of the precious metals there are considerable losses. These losses are greater than is usually supposed, because, as a general rule, systematic assays of the tailings are not made."

Mr. Eissler says: "From what has been said, it will be seen that the problem of gold milling is not the easy matter which some may imagine, but that there is a wide field open for investigation. Much has no doubt been accomplished (speaking of California and Nevada), but there is plenty of room for improvement yet in the methods and appliances for securing the much-coveted metal."

Mr. H. A. Gordon, M.A., F.G.S., speaking in his report (1888) of a mine in the Skippers district, says: "This company has now all the appliances for working a mine systematically and for reducing the quartz, but is still behindhand in the principle of saving the gold. This appears to occupy a secondary place, while really it ought to be the principal."

"I drew attention in my last annual report to the fact that not only is the gold and silver left in the tailings, but in ores where there is a quantity of sulphur and arsenic; a considerable percentage of the precious metal is carried away with the water and never settles in the tailings at all. In crushing raw ores containing a large percentage of sulphur and arsenic, it is simply throwing away money by pretending to save the precious metals by amalgamation with mercury, as arsenic sickens the mercury, and so does sulphur; but the sulphur being a very light mineral, and possessing a great affinity for metals, a certain proportion of them floats on the surface of the water and is carried away. page 8 Again, antimony is found in some quartz lodes associated with gold, which sickens the mercury and creates a black scum on the surface, which prevents the gold from adhering to it."

"To carry on mining on an intelligent basis, any company of any note should have a person who is able to assay the ores; and if these assays are taken from a fair average of the stone, the company is then in a position to see whether the mode of treatment is saving a fair percentage of the gold or not; also by making assays of this description they are fully acquainted with the different metals and minerals in the stone, and in a position to know the best method of treatment to adopt from the class of ore they have to deal with."—(Reports on the Mining Industry of New Zealand, 1888.)

All the processes and apparatus described have been practically worked, generally with good results, giving 93 to 95 per cent, of the assay value of the stone; in those instances where they have failed, the failure can invariably be traced to bad management or want of the requisite knowledge. At every stage it is absolutely necessary that assays should be made. Four assays of an ore or concentrates can be made in a couple of hours, and should never be neglected. When the gangue, oxides, &c., leave the amalgamator, it becomes a question, Will they pay for chlorination? When the roasting and grinding have been well done and the gold is in a sufficiently free state (after roasting), it is very nearly all removed by amalgamation. It must not be forgotten that it is very easy "to buy gold too dear," and this should be constantly kept in mind.

Of the many books on Metallurgy, &c., the following may be consulted with advantage:—Percy's Metallurgy, Kustel, Makin, Egleston, Randall, Phillips and Bauerman, Mitchell's Assaying, Eissler, Brown, Kerl, &c.; much information has also appeared from time to time in the 'Otago Witness,' of all of which the writer has gladly availed himself.

As assays both of the ores and tailings are of frequent necessity, in place of a regular assay furnace, the following, named after the inventor "Sefstrom's furnace," will be found extremely useful, and leave nothing to be desired. The fuel used is coke, charcoal, or a mixture of both. Figure 10 consists of a double cylinder, exactly like the improved filter, without the taps; the inner cylinder is lined with a fire-resisting material about an inch thick, consisting of 1 part fire-clay and 3 to 4 parts sand, mixed with water and laid on thinly with a brush, drying between each application, till the required thickness is obtained. To hold 4 Battersea triangle S crucibles, the diameter of the inside cylinder is about 14 inches, the height 8 inches, and the distance between the two cylinders at the sides 1½ inches, at the bottom 2½ to 3. At O, about half-way up the inner cylinder, are eight holes pointing towards the centre; these are plugged with wood, when the clay is put on and can be easily removed, page 9 the air enters at D, and after having been heated in the intermediate space is carried through the holes or tuyers O. The opening D is connected with a pair of double-acting bellows, and a hood is sometimes provided to fit on the top. With this assays can be got off in from 15 to 20 minutes.

The Re-agents employed in assaying are not many, and need not be particularly pure; ordinary commercial quality will do.

Bi-carbonate of Soda.—Grind free from lumps.

Carbonate of Potash, not bi-carbonate.—Pearlash will do.

Cyanide of Potassium.—Not often required; the cakes of commerce powder in the open air, as it is very poisonous, and keep in a well-closed bottle.

Borax, for making Borax Glass.—The ordinary borax of the shops contains 30 to 47 per cent, of water of crystallization, which must be got rid of, thus: Carefully coat the interior of an S crucible with either dry chalk or chalk-wash, and dry. Place in hot fire, and drop in small pieces of borax, letting the swelling subside somewhat after each successive addition. It is well to fill the crucible only one-third full, as the borax is liable to attack the crucible and run through. When thoroughly fused pour into a frying-pan, coated with chalk; let cool, and then powder.

Black Flux Substitute,—Ten parts of bi-carb. soda (usually called soda) and three of flour.

Common Salt.—Dried and crushed, free from lumps.

Nitre or Saltpetre.—Pulverise finely and keep dry.

Wood Charcoal.—In fine powder.

Silica.—Powdered sand or quartz, if free from minerals, will do.

Litharge of Commerce.—Generally contains a trace of silver.

Lead, granulated.—Melt lead in a crucible, and when it is just hot enough to char a splinter of wood, pour into a tight square box; shake gently at first, and then vigorously from side to side. Sift through a 20-mesh sieve, and remelt what remains in the sieve. It can be assayed for silver, but this is not generally necessary.

In order to test gold or silver ores, or ores which are supposed to contain these metals, and to apply to them their proper treatment, a knowledge of the important art of fluxing is required. The nature of the gangue is important in determining the nature of the flux necessary to change it into a slag. The gangue may be acid, basic, or both acid and basic: consequently the rule for fluxing a gangue is: An acid gangue requires a basic flux; a basic gangue requires an acid flux. An acid gangue, as silica, forms salts with a basic flux, as litharge or soda, producing a slag which is composed of the silicates of lead or soda; and a basic gangue, as lime, requires an acid flux, as silica and borax, forming a slag composed of the silicate and borate of lime.

Acid Gangue.	Basic Fluxes.
Quartz, or other forms of uncombined silica; as quartz crystals, quartz rock, sandstone, sand, &c. Silicates, or silica combined with some base; as clay, clay slates, mica, &c. Rocks in which silica predominates; as granites, feldspars, prophyry, &c.
Litharge. Nitre (saltpetre). Carbonate of Soda. Carbonate of Potash. Bi-carbonate of Soda. Bi-carbonate of Potash.

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Basic Gangues.	Acid Fluxes.
Calc spar (carbonate of lime), also limestones. Heavy spar (barytes, or sulphate of barium). Fluor spar (fluoride of calcium). All so-called earths; as alumina and various combinations of lime, magnesia, baryta, Ac., without silica. Sparry iron, or carbonate of iron. Various metallic oxides; as those of iron, manganese, &c., when in sufficient quantity to be considered as gangues.
Borax. Silica. Silicates; as glass and silicate of lead, or lead glass, formed by the fusion of litharge and silica.

The first thing requisite is to determine the reducing power of the charcoal; mix well, and sample carefully.

• Soda, two ounces or 960 grains.
• Carb. potash, quarter ounce or 120 grains.
• Litharge, two ounces or 960 grains.
• Charcoal, 15 grains.
• Salt, cover.

Weigh out in the following order:—Soda, litharge, charcoal, and place on a sheet of glazed paper; mix slightly, and after adding the potash, mix quickly but thoroughly, and pour into an S crucible; put some dry salt on the paper, and dry-wash any adhering particles into the crucible. The salt should form a cover of about half-an-inch thick; it should be invariably used, as it helps, when fused, to wash down the inside of the crucible, taking with it all metals or fluxes. Put the crucible into a hot fire; remove when thoroughly fused; cool, break crucible, detach button from slag, and weigh. The weight, divided by 15, will give the quantity of lead; 1 grain charcoal (if well mixed) will reduce from litharge. It varies between 22 and 80. The trouble of making this and the next experiment will be amply repaid in the subsequent operations.

Use an S crucible, and treat exactly as in the previous operation. The difference between the weight of the lead button obtained in the first assay and in the above, divided by 75, gives the oxidizing power of 1 grain nitre. It is about 4 parts.

Example.

First assay yielded	390 grains lead.—1 grain, therefore, equals 26.
Second assay yielded	71 grains lead.—1 grain, therefore, equals 26.
	319, divided by 75, gives 4¼ grains nearly.

With the aid of these two experiments, we are enabled to control the size of the lead button. When 480 grains of ore are page 11 taken, the lead button should be about 240 grains; if it is smaller, there is the chance of some of the gold being left in the slag; if larger, it is inconvenient to work.

Preliminary charge to determine the reducing power of an ore.—This only requires to be done once, unless the character of the ore changes.

• Ore, 24 grains.
• Litharge, one and a-half ounces.
• Salt, cover.

Use W size crucible and cover, have the fire quite hot, bank round with hot coals, and heat quickly till contents are in quiet fusion, which requires fifteen to twenty minutes; pour into iron mould (figure 11; price 19s) when cold, tap away slag, and weigh. In an actual trial the button weighed 6 grains, the twentieth part of 480 grains having been taken, multiply 6 by 20=120; as the button should weigh 210 grains, we are 120 grains short. The reducing power of the charcoal used before is 26, therefore 120 divided by 26 gives say 5. Five grains charcoal will have to be used with 480 grains ore. Had the button weighed 12 grains, 12 multiplied by 20 gives 240; no charcoal will be required, sufficient reducing power being in the ore. Suppose the button weighed 40 grains, 40 multiplied by 20=800, button too large; 800 less 240, leaves 560 grains lead required to be kept oxidised. The nitre in the previous experiment had an oxidizing power of 4¼, and 560 divided by 4¼ gives 132; therefore 132 grains of nitre will have to be used.

Should the assayer decide to roast the ore, it will not be necessary to make a preliminary assay to determine its reducing power, as the roasting will drive off the reducing elements. The roasting converts the sulphides of iron, copper, manganese, lead, and their lower oxides into the higher oxides of the same metals. It will, therefore, make their ores highly oxidizing, requiring a greater amount of reducing agent in the actual assay, and perhaps necessitate a preliminary assay to determine their oxidizing power.

Use a W crucible, cover and treat in the usual manner, and weigh the resulting button. In an actual trial the button weighed 6 grains. 6 multiplied by 20=120; 240 the quantity required less 120=120; 120 divided by 26, the reducing power of the charcoal gives say 5. Five grains more charcoal, or a total of 25, will have to be used in the assay of 480 grains ore.

We have now all the data required to make the assay, and the following charge will be found very good proportions for ordinary ores.

Powder the ore very fine, and weigh all ingredients; without this good work is impossible.

To ascertain the presence or absence of sulphides, try the following very simple test:—Powder a little of the sample and heat it in an iron spoon on top of a good fire, and note its behaviour. If, shortly after heating, star-like sparks are quickly thrown off, particularly noticeable when the powder is stirred with a wire, then the mass begins to glow around the edges, while closely above small blue flames appear, and finally the entire mass becomes red hot, with fumes and an odor of a burning match is perceived; then sulphides of iron or copper, or both, are present, and so the ore will be decidedly reducing. If, in addition, an onion or garlic-like odor and whitish fumes are noticed, arsenic is also present, probably as arsenical iron pyrites. Antimony and zinc give white fumes and no odors, Zinc blende and galena are not so liable to glow and to scintillate as are the iron and copper sulphides, but the smell of a burning match should be perceptible upon heating them or any other ore containing a fairly large percentage of sulphides. It will also be noticed that after the mass has been thoroughly heated and allowed to cool, it will have lost its metallic shimmer, and become of a dull, dead color, indicating oxidation.

A simple chemical test for sulphides may be easily and quickly tried:—Place a small portion of the finely-ground ore in a test tube, pour in a little water, shake, add a few drops of hydrochloric acid, and warm gently. If the smell of rotten eggs (sulphuretted hydrogen) is perceptible, sulphides are certainly in the ore.

Iron pyrites (sulphide of iron), one of the most widely distributed of minerals, is found in rocks of every age, and is a common and abundant source of gold.

No. 1.—Use J crucible, 6¾ inches deep, 4? across outside; cover. Time, about half-an-hour.

No. 2.—Use J crucible; begin with a very gentle heat, and gradually bring up to a full red heat. Time, about 35 minutes. When in full fusion add charcoal wrapped in as small a piece of tissue paper as possible; weight of paper and charcoal, 15 grains.

No. 3.—Heat gradually; if iron matter is formed, instead of free flowing lead button, the assay must be repeated, and the proportion of nitre increased.

No. 4.—Bind with iron wire 6 four-inch wire nails, and push into the charge points down; remove them when the mass is in quiet fusion; if iron matte is formed, repeat adding a little nitre.

All these can be poured into the iron mould (fig. 11), and when cold the slag broken away.

Many assayers insist that from unroasted sulphide of iron ores no gold will be obtained by crucible fusion. This extraordinary idea, which, however, seems to spread, is entirely unfounded, or, at best, is based upon botchy experiments. The student may rest assured that he can, with careful working, extract all the gold from an unroasted ore by any of the nitre methods.— (Brown's Assaying, page 295.)

Use an S crucible. The charcoal is calculated upon a reducing power of 26. If a little sulphide has remained unoxidized during roasting, push into the charge one four-inch nail, and remove immediately after fusion.

Having obtained the lead button, the next operation is to cupel it. First, hammer to get rid of the slag, and shape to a cube, rounding the corners to prevent them injuring the cupel.

The Cupel Furnace (fig. 12).—It is hardly worth while attempting to make a substitute for this, a useful size taking muffles (fig. 18), 7 inches long, 3½ wide, and 2½ high, can be landed for about £5; the muffles two shillings each.

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The object of cupelling is to get rid of the lead by oxidation, the litharge formed by the heat being partly absorbed by the cupel and partly driven up the chimney, leaving the gold and silver together as a bead upon the surface of the cupel. Other metals that may have remained in small quantity from the previous operations, are also oxidized and got rid of.

Having heated the muffle to a good red heat, put in the cupels, and when they are well heated, by means of the tongs place in them the lead buttons, close the muffle door, and in a minute's time or less all the lead buttons will have quietly fused and on opening the door each will be seen as a little lake of molten metal, from which arise fumes of oxide of lead. The door is closed simply to melt the buttons, by the increased heat and absence of air. Do not have the heat too great, so that the cupels are whitish, or the melted lead bubbles, on the other hand for gold ores it is better to have it too hot than too cold; with silver ores too great a heat should be carefully avoided, or some of the silver will be volatilised. The buttons get gradually smaller and smaller, changing from flat liquids to convex ones, and this reduction continues until the point is reached when nearly the last of the lead leaves the bead. This is known as the brightening or flashing. As the button of gold, silver, and lead arrives near this stage, it appears to revolve with great velocity, and rainbow colors succeed each other all over its surface. Finally a film passes over the bead and then no more action is visible. Increase the heat by closing the muffle door, that the last traces of lead may be driven off, let cool gradually and remove.

The apparatus required in addition to the furnace and muffle, are tongs and cupels; the latter are better purchased. No. 5, costing two shillings per dozen, are a convenient size; the button weighing 240 grains should be divided, and two cupels used.

Parting the assay consists in removing the silver present and leaving the gold in a state fit for weighing. It is necessary that the silver should be present in the proportion of at least two and a-half parts to one part of gold, otherwise its particles will be so enveloped and protected by those of the gold that the nitric acid will be prevented from exercising its solvent action. The necessary quantity of silver can be added before the button is cupelled or after, if the quantity of silver as well as gold present is required; the button should now be weighed, the requisite quantity of silver added, the assay and silver wrapped in a little piece of sheet lead, and the whole cupelled as before. The button after cooling is removed, hammered upon any flat surface, taking care that nothing from the cupel is hammered in, and put into a small clean porcelain capsule or crucible; then fill about a quarter full with water, add four or six drops of strong nitric acid. No exact rule as to the amount of acid page 15 to add can be given, nor indeed is it necessary, but in general add drop by drop till it begins to "bite" the bead. If it is intended to make frequent assays, the acid had better be made up in quantity, 16 parts of strong nitric acid and 80 parts of water, will do for ordinary small beads; for larger ones, after having treated them with the above, add some of the following stronger solution, made up with 16 parts strong acid and 10 parts of water. Keep in well stoppered bottles. Now place the capsule on a sand bath (a tin lid filled with dry sand will do), and heat gently, not enough to cause the acid solution to boil; after a time no more action goes on; pour off and repeat the boiling with fresh acid. If there is no gold in the bead, it will not blacken on adding the acid, and nothing will remain undissolved in the capsule; it will contain only the clear solution of nitrate of silver, formed by the silver dissolving in the acid. Should a residue be left, tap the capsule lightly to cause the gold to settle at the bottom; pour the liquid off very carefully, wash three or four times with water, finally drain, remove with great care by means of blotting paper any drops of water inside the capsule, and heat gently, at first till all moisture has been driven off, then intensely for a minute or two; if the weight of the capsule has already been taken, the calculation of the result is easy.

Suppose the quantity of concentrates taken was 1 ounce (480 grains), and the gold weighed 2 grains; 32,666? ounces Troy are contained in a ton of 2240 lbs, therefore 32,666 multiplied by 2 = 65,332 grains, or 136 ounces, 2 pennyweights, 10 grains. Although the exact assaying of bullion requires great care, experience, and costly appliances, still a fair idea of the value of an ore, can with a little practice be obtained by the above methods.

Platinum.

The present high price of platinum, about 85 shillings per ounce, makes it worth looking for and saving. Platinum is generally found in small grains, of a greyish white colour approaching to that of tarnished steel. These grains are commonly flattened, and appear to have been polished by friction against other bodies. Their size usually varies from that of linseed to that of hemp seed, but fragments of much larger dimensions have occasionally been discovered.

It is frequently met with, especially in Otago, in alluvial deposits, associated with gold and being one of the heaviest of all known substances, heavier than gold, it is left behind with the gold and black sand; the latter can be partly removed with an ordinary magnet, and the residue tested for gold and platinum, thus:

The finely-powdered mineral is heated in a glass flask or porcelain dish to near boiling for some time with aqua regia, a mixture of four parts strong hydrochloric acid and one part page 16 strong nitric acid, till all but the sand is dissolved; as much as possible of the free acid is then evaporated off, and when cool the solution is diluted with water and filtered. If any silver is present, it will remain on the filter with the sand as insoluble chloride of silver, and after being well washed with water, can be tested by pouring over it a little ammonia water, collecting the liquid that runs through in a test tube and adding nitric acid in excess, that is till the smell of ammonia is removed. A white precipitate shows the presence of silver.

To the liquid from which the sand and silica have been removed, a solution of carbonate or bi-carbonate of soda is added until no more effervescence takes place; all other metals, except gold and platinum, will be precipitated, and the solution when filtered can be tested for these. If we wish to reduce the gold first, it can easily be done either by adding a solution of oxalic acid until it ceases to produce effervesence and has a sour taste, and then boiling; if any gold is present it will be precipitated as a dark powder, or by adding a solution of sulphate of iron (green vitriol) and letting stand for a few hours. Upon now filtering the liquid from which the gold has been precipitated, and testing a small portion with a solution of stannous chloride (commonly known as salts of tin), a dark brownish purple precipitate shows there is still some gold in solution, which must be got rid of by adding to the bulk of the liquid more oxalic acid or sulphate of iron, and, after boiling or standing, again filtering; if in a small portion no precipitate is produced, but a reddish brown colouring appears, platinum is present. Evaporate the remainder of the solution, add about three-fourths of its bulk of spirits of wine, and then a saturated solution of sal-ammoniac. The platinum will be thrown down as a yellow crystalline precipitate.

Or the platinum may be removed first by evaporating the aqua regia solution until it is much reduced in quantity, then adding about three-fourths of its bulk of spirits of wine, and after that a saturated solution of chloride of ammonium (sal ammoniac). By these reagents the platinum will be thrown down as a yellow crystalline precipitate, while the solution, filtered from this and treated with sulphate of iron or boiled with oxalic acid, deposits the gold.

Apothecaries' Weight.

• 20 grains = 1 scruple.
• 3 scruples = 1 dram or 60 grains.
• 8 drams = 1 ounce or 480 grains.
• 12 ounces = 1 pound or 5760 grains.

Troy Weight.

• 24 grains = 1 pennyweight.
• 20 penny weights = 1 ounce or 480 grains.
• 12 ounces= 1 pound or 5760 grains.

Avoirdupois Weight (the ordinary weights of the shops).

• 1 ounce = 437½ grains.
• 16 ounces = 1 pound or 7000 grains.