Fig. 13.Huntingdon Mill.
A No. 4 Dodge stone-breaker working about 8 hours will keep a five foot Huntingdon Mill going 24 hours, and an automatic feeder is necessary. For that matter both are almost essential for an ordinary stamper battery, and will certainly increase the crushing capacity and do better work from the greater regularity of the feed.
A 10 h.-p. (nominal) engine of good type is sufficient for Huntingdon mill, rock breaker, self-feeder and steam pump. A five foot mill under favourable circumstances will crush about as much as eight head of medium weight stamps.
Fig. 14.Stone-Breaker.
Fig. 15.Stone-Breaker (Sectional View).
The Grusonwerk Ball Mills(Fig. 16), made by Krupp of Germany, as also that made by the Austral Otis Company, Melbourne, are fast and excellent triturating appliances for either wet or dry working, but are specially suitable only for ores when the gold is fine and evenly distributed in the stone. The trituration is effected by revolving the stone in a large cylinder together with a number of steel balls of various sizes, the attrition of which with the rock quickly grinds it to powder of any required degree of fineness.
Fig. 16.Grusonwerk Ball Mill.
Probably more mines have been ruined by bad mill management than by bad mining, though every experienced man must have seen in his time many most flagrant instances of bungling in the latter respect. Shafts are often sunk on the wrong side of the lode or too near or too far away therefrom, while instances have not been wanting where the (mis)manager has, after sinking his shaft, driven in the opposite direction to that where the lode should be found.
A common error is that of erecting machinery before there issufficient ore in sight to make it certain that enough can be provided to keep the plant going. In mines at a distance from the centre of management it is almost impossible to check mistakes of this description, caused by the ignorance or over sanguineness of the mine superintendent, and they are often as disastrous as they are indefensible. Another fertile source of failure is the craze for experimenting with untried inventions, alleged to be improvements on well-known methods.
A rule in the most scientific of card games, whist, is “when in doubt lead trumps.” It might be paraphrased for mining thus: “When in doubt about machinery use that which has been proved.” Let some one else do the experimenting.
Fig. 17.Double Faulted Lode.
The success of a quartz mine depends as much on favourable working conditions as on its richness in gold. Thus it may be that a mine carrying 5 or 6 oz. of gold to the ton but badly circumstanced as to distance, mountainous roads, lack of wood and water, in some cases a plethora of the latter, or irregularlyfaulted country(Fig. 17), may be less profitable than another showing only 5 or 6 dwt., but favourably situated.
It is usually desirable to choose for the battery site, when possible, the slope of a hill which consists of rock that will give a good foundation for your battery.
The economical working depends greatly on the situation, which is generally fixed more or less, in the proximity of the water. The advantages of having ample water for battery purposes, or of using water as a motive power, are so great that it is very often desirable to construct a tramway of considerable length, when, by so doing, that power can be utilised; hence most quartz mills are placed near streams, or in valleys where catchment dams can be effectively constructed, except, of course, in districts where much water has to be pumped from the mine.
If water-power can be used, the water-motor will necessarily be placed as low as possible, in order to obtain the fullest available power. One point is essential. Special care must be taken to keep the appliances above highest flood-level. If the water in the stream is not sufficient to carry off the tailings, the battery should be placed at such a height as to leave sufficient slope for tailings’ dumps. This is more important when treating ore of such value that the tailings are worth saving for secondary treatment. In this case provision should be made for tailings, dams, or slime pits.
Whether the battery is worked by water, steam, or gas power, an ample supply of water is absolutely necessary, at least until some thoroughly effective mode of dry treatment is established. If it can be possibly arranged the water should be brought in by gravitation, and first cost is often least cost; but where this is impossible, pumps of sufficient capacity, not only to provide the absolute quantity used, but to meet any emergency, should be erected.
The purer the water the better it will be for amalgamating purposes, and in cold climates it is desirable to make provision for heating the water supplied to the battery. This can be done by means of steam from the boiler led through the feed tanks; but where the boiler power is not more than required, waste steam fromthe engine may be employed, but care must be taken that no greasy matter comes in contact with the plates. The exhaust steam from the engine may be utilised by carrying it through tubes fitted in an ordinary 400 gallon tank.
Reducing appliances have often to be placed in districts where the water supply is insufficient for the battery. When this is so every available means must be adopted for saving the precious liquid, such as condensing the exhaust steam from the engine. This may be done by conducting it through a considerable length of ordinary zinc piping, such as is used for carrying the water from house roofs. Also tailings pits should be made, in which the tailings and slimes are allowed to settle, and the cleared water is pumped back to be again used. These pits should, where practicable, be cemented. It is usual, also, to have one or two tailings dams at different levels; the tailings are run into the upper dam, and are allowed to settle; the slimes overflow from it into the lower dam, and are there deposited, while the cleared water is pumped back to the battery. Arrangements are made by which all these reservoirs can be sluiced out when they are filled with accumulated tailings. It is well not to leave the sluicing for too long a period, as when the slimes and tailings are set hard they are difficult to remove.
Where a permanent reducing plant is to be erected, whatever form of mill may be adopted, it is better for many reasons to use automatic ore feeders. Of these the best two I have met are the “Tulloch”(Fig. 18), and “Challenge”(Fig. 19)either of which can be adapted to any mill and both do good work.
By their use the reducing capacity of the mill is increased, and the feeding being regular the wear and tear is decreased, while by the regulated feeding of the “pulp” in the battery box or mortar can be maintained at any degree of consistency which may be found desirable, and thus the process of amalgamation will be greatly facilitated. The only objection which can be urged against the automatic feeder is that the steel points of picks, gads, drills, and other tools may be allowed to pass into the mortar or mill, and thus cause considerable wear and tear. A remedy for this is suggested(p. 148).
Plate IV.—Tail Race
Fig. 18.Tulloch Automatic Ore Feeder.
Fig. 19.Challenge Automatic Ore Feeder.
If local conditions are favourable and the management good, some mines have paid well with a yield of from 1½ to 4 dwts. of gold per ton of ore; but others have not paid with much higher yields in test crushings. Great judgment is required in selecting the best possible site for the class of crushing machinery decided upon. The crusher should be as high as can be managed so as to give a good clear run for the tailings and room for the concentrators and amalgamators, &c., below the copper plates and blanket strakes. Solid, strong foundations are essential; many failures are due to a neglect of this point.
In a stamp mill(Fig. 20)the foundations are usually made of hard-wood logs about 5 to 6 feet long, set on end, the bottom end resting on rock and set round with cement concrete. These are bolted together, and the “box” or mortar is bolted to them. The horizontal logs to carry the “horses” or supports for the battery frame should also be of good size, and solidly and securely bolted. The same applies to your engine bed, but whether it be of timber, or mason work, above all things provide that the whole of your work is set out square and true to save after wear and friction.Fig. 21represents a 10-head stamp mill.
My experience has been that the most effective weight for stamps and height for drop largely depends on the nature of the rock foundation. I have usually found that with medium stamps, say 7½ to 8 cwt. with fair drop and lively action, about 80 falls per minute, the best results were obtained, but the tendency of modern mill men is towards the heavier stamps, 10 cwt. and even heavier.
Great improvements have been made in stamp mills since the sixteenth century, as is evident by comparingFig. 21AwithFigs. 20and21; even in the writer’s time they have been considerable.
To find the horse power required to drive a battery, multiply the weight of one stamp by the number of stamps in the battery; the height of lift in feet by the number of lifts per minute; add one-third of the product for friction, and the result will be the number of feet-lbs. per minute; divide this by 33,000 which is the number of feet-lbs. per minute equal to 1 h.-p. and the result will be the h.-p. required. Thus if a stamp weighs800 lbs. and you have five in the box, and each stamp has a lift of 9 in. = 0·75 ft. and strikes 80 blows per minute, then ÷ 800 × 5 × 0·75 × 80 = 240,000; one third of 240,000 = 80,000, which added to 240,000 = 320,000; and 320,000 divided by 33,000 = 9·7 h.-p. or 1·9 h.-p. each stamp.
Fig. 20.Stamp Mill, showing Part of Foundations.
Fig. 21.10-Head Stamp Mill.
The total weight of a battery, including stamper box, stampers, &c., may be roughly estimated at about 1 ton per stamp. Medium weight stampers, including shank, cam, disc, head, and shoe, weigh from 600-700 lbs., and need about ¾ h.-p. to work them.
Fig. 21a.Sixteenth Century Stamp Mill.
The quantity of water required for the effective treatment of gold bearing rock in a stamper battery varies according to the composition of the material to be operated upon, but generally it is more than the inexperienced believe. For instance, “mullocky” lode stuff, containing much clayey matter or material carrying a large percentage of heavy metal, such as titanic iron or metallic sulphides, will need a larger quantity of water per stamp than clean quartz. A fair average quantity would be 750 to 1000gallons per hour for each box of five stamps. In general practice I have seldom found 1000 gallons per hour more than sufficient.
As to the most effective mesh for the screen or grating, that depends largely on the size of the gold particles contained in the gangue. The finer the particles the closer must be the mesh, so that nothing but careful experiment will enable the battery manager to decide this most important point. The American slotted screens are best; they wear better than the punched gratings and can be used of finer gauge. Woven steel wire gauze is employed with good effect in some mills where specially fine trituration is required. This class of screen requires special care as it is somewhat fragile, but with intelligent treatment does good work.
The fall or inclination of the tables, both copper and blanket strakes, is also regulated by the class of ore. If it should be heavy then the fall must be steeper. A fair average drop is ¾ inch to the foot. Be careful that your copper tables are thoroughly water-tight, for, remember, where water will percolate, mercury will penetrate.
The blanket tables are simply a continuation of the mercury tables, but covered with strips of coarse blanket, green baize, or other flocculent material, intended to arrest the heavier metallic particles which have not been amalgamated.
The blanket table is, however, a very unsatisfactory concentrator at best, and is giving place to mechanical concentrators of various descriptions.
I have a device at present unpatented which will do away with the necessity for blanket tables or copper plates, as by the expenditure of from ½ to 1 h.-p. every particle of free gold after leaving the boxes must come within the embrace of a mercury amalgam. No cleaning of plates is required.
An ancient Egyptian gold washing table is shown inFig. 22. It is a representation of an old stone table such as is referred to inp. 2, which was used by the Egyptians in treating the gold ores of Lower Egypt. The ore was first ground, it is likely by means of some description of stone arrastra, and then passed over the sloping table with water, the gold being retained in theriffles. In these the material would probably be mechanically agitated. Although for its era ingenious it will be plain to practical men that if the gold were fine the process would be very ineffective. Possibly, but of this I have no evidence, mercury was used to retain the gold in the riffles, for, as previously stated, this method was known to the ancients.
Fig 22.Washing Table of Stone with Riffles.
At a mine at which I was managing director the lode was almost entirely composed of sulphide of iron, carbonate of lime or calcspar, with a little silica. In this case it was found best to crush without mercury, then run the pulp into pans, where it was concentrated. The concentrates were calcined in a common reverberatory furnace, and afterwards amalgamated with mercury in a special pan; the results were very satisfactory; but it does not follow that this process would be the most suitable for a slightly different lode stuff.
I was lately consulted with respect to the treatment of a pyritic ore in a very promising mine, but could not recommend the above treatment, because though the pyrites in the gangue was similar, the bulk of the lode consisted of silica, consequently there would be a great waste of power in triturating the whole of the stuff to what, with regard to much of it, would be an unnecessary degree of fineness. I am of opinion that in cases such as this, where it is not intended to adopt the chlorination or cyanogen process, it will be found most economical to crush to a coarse gauge, concentrate, calcine the concentrates, and finally amalgamate in some suitable amalgamator.
Probably for this mode of treatment Krom rolls would be found more effective reducing agents than stampers, as with them the bulk of the ore can be broken to any required gauge and there would consequently be less loss in “slimes.”
The great art in effective battery work is to crush your stuff to the required fineness only, and then to provide that each particleis brought into contact with the mercury either in box, trough plate, or pan. To do this the flow of water must be carefully regulated; it should not carry the stuff off too quickly nor allow the troughs and plates to be choked. In cold weather the water may be warmed by passing the feed-pipe through a tank into which the steam from the engine exhausts, and this will be found to keep the mercury bright and lively. But be careful no engine oil or grease mingles with the water, as grease on the copper tables will absolutely prevent amalgamation.
The first point, then, is to effectively crush the gangue, the degree of fineness being regulated by the fineness of the gold itself. This being done, then comes the question of saving the gold. If the quartz be clean, and the gold unmixed with base metal, the difficulty is small. All that is required is to ensure that each particle of gold shall be taken up by the mercury. The main object is to arrest the gold at the earliest possible stage; therefore, if you are treating clean stone containing free gold, either coarse or fine, I advise the use of mercury in the boxes, for the reason that a considerable proportion of the gold will be caught thereby, and settling to the bottom, or adhering to amalgamated plates in the boxes, where such are used, will not be afterwards affected by the crushing action, which might otherwise break up, or “flour,” the mercury. On the whole, I rather favour the use of mercury in the box at any time, unless the ore is very refractory—that is, contains too great a proportion of base metals, particularly sulphides of iron, arsenic, &c., when the result will not be satisfactory, but may entail great loss by the escape of floured mercury carrying with it particles of gold. Intelligent experience will assist the battery manager to adopt the right system.
The crushed stuff—generally termed the “pulp”—passes from the boxes through the “screens” or “gratings,” and so on to the “tables”—i.e., sheets of copper amalgamated on the upper surface with mercury, and sometimes electroplated with silver and afterwards treated with mercury. Unless the quartz is very clean, and, consequently light, I am opposed to the form ofstamper box with mercury troughs cast in the “lip,” nor do I think that a trough under the lip is a good arrangement, as it usually gets so choked and covered with the heavy clinging base metals as to make it almost impossible for the gold to come in contact with the mercury. It will be found better where the gold is fine, or the gangue contains much base metal, to run the pulp from the lip of the battery into a “distributor.”
The distributor is a wooden box the full width of the stamper box or mortar, having a perforated iron bottom set some three to four inches above the first copper plate, which should come up under the lip. The effect of this arrangement is that the pulp is dashed on to the plate by the falling water, and the gold at once coming in contact with the mercury begins to accumulate and attract that which follows, till the amalgam becomes piled in little crater-shaped mounds, and thus 75 per cent. of the gold is saved on the top plate.
I have tried a further adaptation of this process when treating ores containing a large percentage of iron oxide, where the bulk of the gold is impalpably fine, and contained in the “gossan.” At the end of the blanket table, or at any point where the crushed stuff last passes before going to the “tailings heap” or “sludge pit,” a “saver” is placed. The saver is a strong box about 15 in. square by 3 ft. high, one side of which is removable, but must fit tight. Nine slots are cut inside at 4 in. apart, and into these are fitted nine square perforated copper plates, having about eighty to a hundred ¼ in. holes in each; the perforations should not come opposite each other. These plates are to be amalgamated on both sides with mercury, and the top plate “fed” from time to time with mercury, in which a very little sodium has been placed. If acid ores are being treated, zinc should be employed in place of sodium, and to prevent the plates becoming bare, if the stuff is very poor, thick zinc amalgam may be used with good effect; but in that case discontinue the sodium, and occasionally, if required, say once or twice in the day, mix an ounce of sulphuric acid in a half gallon of water and slowly pour it into the launder above the saver. Underneath the “saver” you require a few riffles, or troughs,to catch any waste mercury, but if not overfed there should be no waste. This automatic appliance sometimes arrests a considerable quantity of gold.
We now come to the subsidiary processes of battery work, the “cleaning” of plates, and “scaling” same when it is desired to get all the gold off them, the cleaning and retorting of amalgam, and of the mercury, smelting gold, &c.
Plates should be tenderly treated, kept as smooth as possible, and when cleaning up after crushing, in your own battery, the amalgam—except, say, at half-yearly intervals—should be removed with a rubber only; the rubber is simply a square of black india-rubber or soft pine wood.
Fig. 23.Scraper.
When crushing rich ore, and you want to get nearly all the gold off your plates, the scraper(Fig. 23)may be resorted to. This is usually made by the mine blacksmith from an old flat file which is cut in half, the top turned over, beaten out to a sharp blade, and kept sharp by touching it up on the grinding-stone. This, if carefully used, will remove the bulk of the amalgam without injury to the plate.
Various methods of “scaling” plates will be found among“Rules of Thumb.”
Where base metals are present in the lode stuff frequent retortings of the mercury, say not less than once a month, will be found to have a good effect in keeping it pure and active. For this purpose, and in order to prevent stoppage of the machinery, a double quantity is necessary, so that half may be used alternately. Less care is required in retorting the mercury than in treating the amalgam, as the object in the one case is more to cleanse the metal of impurities than to save gold, which will for the most part have been extracted by squeezing through the chamois leather or calico. A good strong heat may therefore at once be applied to the retort and continued, the effect being to oxidise the arsenic, antimony, lead, &c., which, in the form of oxides, will not againamalgamate with the mercury, but will either lie on its surface under the water, into which the nozzle of the retort is inserted, or will float away on the surface of the water. I have also found that covering the top of the mercury with a few inches of broken charcoal when retorting has an excellent purifying effect.
In retorting amalgam, much care and attention is required.
First, never fill the retort too full, give plenty of room for expansion; for, when the heat is applied, the amalgam will rise like dough in an oven, and may be forced into the discharge pipe, the consequence being a loss of amalgam or the possible bursting of the retort. Next, be careful in applying the heat, which should be done gradually, commencing at the top. This is essential to prevent waste and to turn out a good-looking cake of gold, which all battery managers like to do, even if they purpose smelting into bars.
Sometimes special difficulties crop up in the process of separating the gold from the amalgam. At the first “cleaning up” on the Frasers Mine at Southern Cross, West Australia, great consternation was excited by the appearance of the retorted gold, which, as an old miner graphically put it, was “as black as the hind leg of a crow,” and utterly unfit for smelting, owing to the presence of base metals. Some time after this I was largely interested in the Blackborne mine in the same district when a similar trouble arose. This I succeeded in surmounting, but a still more serious one was too much for me—i.e., the absence of payable gold in the stone. I give here an extract from theAustralian Mining Standard, of December 9th, 1893, with reference to the mode of cleaning the amalgam which I adopted.
I had submitted to me lately a sample of amalgam from a mine in West Australia which amalgam had proved a complete puzzle to the manager and amalgamator. The Mint returns showed a very large proportion of impurity, even in the smelted gold. When retorted only, the Mint authorities refused to take it after they had treated the first two cakes, one of 119 oz., which yieldedonly 35 oz. 5 dwt. standard gold, and one of 140 oz., which gave 41 oz. 10 dwt. The gold smelted on the mine was nearly as bad proportionately. Thus, 128 oz. smelted down at the Mint to 87 oz. 8 dwt. and 109 oz. to 55 oz. 10 dwt. The impurity was principally iron, a most unusual thing in my experience, and was due to two causes revealed by assay of the ore and analysis of the mine water, viz., an excess of arsenate of iron in the stone, and the presence in large proportions of mineral salts, principally chloride of calcium CaCl₂, sodium NaCl, and magnesium MgCl₂, in the mine water used in the battery. The exact analysis of the water was as follows:—
It will be seen, then, that this water is nearly four times more salt than that of the sea. The effect of using a water of this character, as I have previously found, is to cause the amalgamation of considerable quantities of iron with the gold as in this case.
I received 10 oz. of amalgam, and having found what constituted its impurities proceeded to experiment as to its treatment. When retorted on the mine it was turned out in a black cake so impure as almost to make it impossible to smelt properly. I found the same result on first retorting, and after a number of experiments which need not be recapitulated, though some were fairly effective, I hit on the following method, which proved to be most successful and will probably be so found in other localities where similarly unfavourable conditions prevail.
I took a small ball of amalgam, placed it in a double fold of new fine-grained calico, and after soaking in hot water put it under a powerful press. The weight of the ball before pressingwas 1583 gr. From this 383 gr. of mercury were expressed and five-eighths of a grain of gold was retorted from this expressed mercury. The residue, in the form of a dark, grey, and very friable cake, was powdered up between the fingers and retorted, when it became a brown powder; it was afterwards calcined on a flat sheet in the open air; result, 510 gr. of russet-coloured powder. Smelted with borax, the iron oxide readily separated with the slag; result, 311 gr. gold 871-1000 fine; a second smelting brought this up to 914-1000 fine. Proportion of smelted gold to amalgam, one-fifth.
The principal point about this mode of treatment is the squeezing out of the mercury, whereby the amalgam goes into the retort in the form of powder, thus preventing the slagging of the iron and enclosure of the gold. The second point of importance is thorough calcining before smelting.
Of course it would be practicable, if desired, to treat the powder with hydrochloric acid, and thus remove all the iron, but in a large way this would be too expensive, and my laboratory treatment, though necessarily on a small scale, was intended to be on a practical basis.
The amalgam at this mine was in this way afterwards treated with great success.
For the information of readers who do not understand the chemical symbols it may be said that
FeCO₃is carbonate of iron;CaCO₃is carbonate of calcium;CaSO₄is sulphate of calcium;CaCl₂is chloride of calcium;MgCl₂is chloride of magnesium;NaClis chloride of sodium, or common salt
Before any plan is adopted for treating the ore in a new mine the management should very seriously and carefully consider the whole circumstances of the case, taking into account the quantity and quality of the lode stuff to be operated on, and ascertain by analysis what are its component parts, for, as before stated, the treatment which will yield most satisfactory results with a certain class of gangue on one mine will sometimes, even when the material is apparently similar, prove a disastrous failure in another. Some time since I was glad to note that the manager of what has since become a very prominent Australian mine strongly discountenanced the purchase of any extracting plant until he was fully satisfied as to the character of the bulk of the ore he would have to treat. It would be well for the pockets of shareholders and the reputation of managers if more of our mine superintendents followed this prudent and sensible course.
Having treated on gold extraction with mercury by amalgamated plates and their accessories, something must be said about secondary modes of saving in connection with the amalgamation process. The operations described hitherto have been the disintegration of the gold-bearing material and the extraction therefrom of the coarser free gold. But it must be understood that most auriferous lode stuff contains a proportion of sulphides of various metals, wherein a part of the gold, usually in a very finely divided state, is enclosed, and on this gold the mercury has no influence. Also many lodes contain hard heavy ferric ores, such as titanic iron, tungstate of iron, and hematite, in which gold is held. In others,again, are found considerable quantities of soft powdery iron oxide or “gossan,” and compounds such as limonite, aluminous clay, &c., which, under the action of the crushing mill become finely divided and float off in the water as “slimes,” carrying with them atoms of gold, often microscopically small. To save the gold in such matrixes as these is an operation which even the best of our mechanical appliances have not yet fully accomplished.
Where there is not too great a proportion of base metals on which the solvent will act, and when the material is rich enough in gold to pay for the extra cost of treatment, chlorination or cyanisation are the best modes of extraction yet practically adopted.
Presuming, however, that we are working by the amalgamation process, and have crushed our stone and obtained the free gold, the next requirement is an effective concentrator to secure the heavy base metals which hold a percentage of gold. There are many, and some do excellent work, but do not act equally well in all circumstances. The first and most primitive is the blanket table(p. 79); but it can hardly be said to be very effective, and requires constant attention and frequent changing and washing of the strips of blanket.
Instead of blanket tables percussion tables are sometimes used, to which a jerking motion is given against the flow of the water and pulp, and by this means the heavier minerals are gathered towards the upper part of the table, and are removed when concentrated. I have seen this appliance doing fairly good work, but it is by no means a perfect concentrator.
Another form of “shaking table” is one with a lateral motion, and this, whether with amalgamated plates, or provided with small riffles, or covered with blanket, keeps the pulp lively and encourages the retention of the heavier particles, whether of gold or base metals containing gold. There has also been devised a rocking table, the action of which is analogous to that of the ordinary miner’s cradle. This appliance, working somewhat slowly, swings on rockers from side to side, and is usually employed in mills where, owing to the complexity of the ore, difficulties have been met with in amalgamating the gold.Riffles are provided and even very fine gold is sometimes effectively recovered by their aid.
The Frue vanner will, as a rule, act well when the pulp is sufficiently and consistently fine. It is an adaptation of the old simple apparatus used in China and India for washing gold dust from river sand. The original consisted of an endless band of strong cloth or closely woven matting, run on two horizontal rollers placed about seven feet apart, one being some inches lower than the other. The upper is caused to revolve by means of a handle. The cloth is thus dragged upwards against a small stream of water and sand fed to it by a second man, the first man not only turning the handle but giving a lateral motion to the band by means of a rope tied to one side.
Fig. 24.Frue Vanner.
Chinamen were working these forerunners of the Frue vanner fifty years ago in Australia, and getting fair returns.
The Frue vanner(Fig. 24)is an endless india-rubber band drawn over an inclined table, to which a revolving and side motion is given by ingenious mechanism, the pulp being automatically fed from the upper end, and the concentrates collected in a trough containing water in which the band is immersed in its passage under the table; the lighter particles wash over the lower end.The only faults with the vanner are—first, it is rather slow; and secondly, though so ingenious it is just a little complicated in construction for the average non-scientific operative.
Of pan concentrators there is an enormous selection, the principle in most being similar—i.e., a revolving muller, which triturates the sand, so freeing the tiny golden particles and admitting of their contact with the mercury. The mistake with respect to most of these machines is the attempt to grind and amalgamate in one operation. Even when the stone under treatment contains no deleterious compounds the simple action of grinding the hard siliceous particles has a bad effect on the quicksilver, causing it to separate into small globules, which either oxidising or becoming coated with the impurities contained in the ore will not reunite, but wash away in the slimes and take with them a percentage of the gold. As a grinder and concentrator, and in some cases as an amalgamator, when used exclusively for either purpose, the Watson and Denny pan(Fig. 25)is effective; but although successfully used at one mine I know, the mode there adopted would, for reasons previously given, be very wasteful in many other mines.
Fig. 25.Watson and Denny Pan.
There is much misconception, even among men with some practical knowledge, as to the proper function of these saving appliances; and sometimes good machines are condemned because they fail to perform work they were not intended for.
It cannot be too clearly realised that the correct order of procedure for extracting the gold held in combination with base metals is—first, reduction of the particles to a uniform gauge and careful concentration only; next, the dissipation, usually by simple calcination, of substances in the concentrates inimical to the thorough absorption of the gold by the mercury; and lastly, the amalgamation of the gold and mercury.
For general purposes, where the gangue has not been crushed too fine, I think the Duncan pan will usually be found effective in saving the concentrates. In theory it is an enlargement of the alluvial miner’s tin dish, and the motion imparted to it is similar to the eccentric motion of that simple separator.
The calcining may be effectively carried out in an ordinary reverberatory furnace, the only skill required being to prevent over “roasting” and so slagging the concentrates; or not sufficiently calcining so as to remove all deleterious constituents; the subject, however, is fully treated inChapter VIII.
For amalgamating I prefer some form of settler to any further grinding appliance, but I note also improvements in the rotary amalgamating barrel, which, though slow, is, under favourable conditions, an effective amalgamator. The introduction of steam under pressure into an iron cylinder containing a charge of concentrates with mercury is said to have produced good results, and I am quite prepared to believe such would be the case, as we have long known that the application of steam to ores in course of amalgamation facilitates the process considerably.
Some twenty-five years since I also was engaged on the construction of a dry amalgamator in which sublimated mercury was passed from a retort through the descending gangue in a vertical cylinder, the material falling therefrom through an aperture into a revolving settler, the object being to save water on mines in dry country. The model, about quarter size, was completed, when my attention was called to an American invention, in which the same result was stated to be attained more effectively by blowing the mercury spray through the triturated material by means of a steam jet. I had already encountered a difficulty, since found so obstructive by experimentalists in thesame direction, that is, the getting of the mercury back into its liquid metallic form. This difficulty can be largely obviated by my own device of using a very weak solution of sulphuric acid (it can hardly be too weak) and adding a small quantity of zinc to the mercury. It is perfectly marvellous how some samples of mercury “sickened” or “floured” by bad treatment, may be brought back to the bright, clean, easy flowing metal by a judicious use of these inexpensive materials.
Thus it will probably be found practicable to crush dry and amalgamate semi-dry by passing the material in the form of a thin pasty mass to a settler, as in the old South American arrastra, and, by slowly stirring, recover the mercury, and with it the bulk of the gold.
The following is from theAustralian Mining Standard, and was headed “Amalgamation Without Overflow”:
“Recent experiments at the Ballarat School of Mines have proved that a deliverance from difficulties is at hand from an unexpected quarter. The despised Chilian mill and Wheeler pan, discarded at many mines, will solve the problem, but the keynote of success is amalgamation without overflow. Dispense with the overflow and the gold is saved.“Two typical mines—the Great Mercury Proprietary Gold Mine, of Kuaotunu, N.Z., the other, the Pambula, N.S.W.—have lately been conducting a series of experiments with the object of saving their fine gold in an economical manner. The last and best trials made by these companies were at the Ballarat School of Mines, where amalgamation without overflow was put to a crucial test, in each case with the gratifying result that ninety-six per cent. of the precious metal was secured. What this means to the Great Mercury Mine, for instance, can easily be imagined when it is understood that notwithstanding all the latest gold-saving adjuncts during the last six months 1260 tons of ore, worth £4. 17s.10⅔d.a ton, have been put through for a saving of £1 9s.1⅔d.only; or in other words over two-thirds of the gold has gone to waste (for the time being) in the tailings, and in the tailings at the present moment lie the dividends that should have cheered shareholders’ hearts.“And now for themodus operandi, which, it must be remembered, is not hedged in by big royalties to any one, rights, patent or otherwise. The ore to be treated is first calcined, then put through a rock-breaker or stamper battery in a perfectly dry state. If the battery is used, ordinary precautions, of course, must be taken to prevent waste, or the dust becoming obnoxious to the workmen. The ore is then transferred to the Chilian mill and made to the consistency of porridge, the quicksilver being added. When the principal work of amalgamation is done (experience soon teaching the amount of grinding necessary), from the Chilian mill the paste (so to say) is passed to a Wheeler or any other good pan of a similar type, when the gold-saving operation is completed.”
“Recent experiments at the Ballarat School of Mines have proved that a deliverance from difficulties is at hand from an unexpected quarter. The despised Chilian mill and Wheeler pan, discarded at many mines, will solve the problem, but the keynote of success is amalgamation without overflow. Dispense with the overflow and the gold is saved.
“Two typical mines—the Great Mercury Proprietary Gold Mine, of Kuaotunu, N.Z., the other, the Pambula, N.S.W.—have lately been conducting a series of experiments with the object of saving their fine gold in an economical manner. The last and best trials made by these companies were at the Ballarat School of Mines, where amalgamation without overflow was put to a crucial test, in each case with the gratifying result that ninety-six per cent. of the precious metal was secured. What this means to the Great Mercury Mine, for instance, can easily be imagined when it is understood that notwithstanding all the latest gold-saving adjuncts during the last six months 1260 tons of ore, worth £4. 17s.10⅔d.a ton, have been put through for a saving of £1 9s.1⅔d.only; or in other words over two-thirds of the gold has gone to waste (for the time being) in the tailings, and in the tailings at the present moment lie the dividends that should have cheered shareholders’ hearts.
“And now for themodus operandi, which, it must be remembered, is not hedged in by big royalties to any one, rights, patent or otherwise. The ore to be treated is first calcined, then put through a rock-breaker or stamper battery in a perfectly dry state. If the battery is used, ordinary precautions, of course, must be taken to prevent waste, or the dust becoming obnoxious to the workmen. The ore is then transferred to the Chilian mill and made to the consistency of porridge, the quicksilver being added. When the principal work of amalgamation is done (experience soon teaching the amount of grinding necessary), from the Chilian mill the paste (so to say) is passed to a Wheeler or any other good pan of a similar type, when the gold-saving operation is completed.”
This being an experiment in the same direction as my own, I tried it on a small scale. I calcined some very troublesome ore till it was fairly “sweet,” triturated it, and having reduced it with water to about the consistency of invalid’s gruel, put it into a little berdan pan made from a “camp oven,” which I had used for treating small quantities of concentrates, and from time to time drove a spray of mercury, wherein a small amount of zinc had been dissolved, into the pasty mass by means of a steam jet, added about half an ounce of sulphuric acid and kept the pan revolving for several hours. The result was an unusually successful amalgamation and consequent extraction—over ninety per cent.
Steam—or to use the scientific term, hydro-thermal action—has played such an important part in the deposition of metals that I cannot but think that under educated intelligence it will prove a powerful agent in their extraction. Twenty-five years ago I obtained some rather remarkable results from simply boiling auriferous ferro-sulphides in water. There is in this alone an interesting, useful, and profitable field for investigation and experiment.
The most scientific and perfect mode of gold extraction (when the conditions are favourable) is lixiviation by means of chlorine, potassium cyanide, or other aurous solvent, for by this means as much as 98 per cent. of the gold contained in suitable ores canbe converted into its mineral salt, and being dissolved in water, re-deposited in metallic form for smelting; but lode stuff containing much lime would not be suitable for chlorination, or the presence of a considerable proportion of such a metal as copper, particularly in metallic form, would be fatal to success, while cyanide of potassium will also attack metals other than gold, and hence discount the effect of that solvent.
The earlier practical applications of chlorine to gold extraction were known as Mears’ and Plattner’s processes, and consisted in placing the material to be operated on in vats with water, and introducing chlorine gas at the bottom, the mixture being allowed to stand for a number of hours, the minimum about twelve, the maximum forty-eight. The chlorinated water was then drawn off containing the gold in solution which was deposited as a brown powder by the addition of sulphate of iron.
Great improvements on this slow and imperfect method have been made of late years, among the earlier of which was that of Messrs. Newbery and Vautin. They placed the pulp with water in a gas-tight revolving cylinder, into which the chlorine was introduced, and atmospheric air to a pressure of 60 lb. to the square inch was pumped in. The cylinder with its contents was revolved for two hours, then the charge was withdrawn and drained nearly dry by suction, the resultant liquid being slowly filtered through broken charcoal on which the chloride crystals were deposited, in appearance much like the bromo-chlorides of silver ore seen on some of the black manganic oxides of the Barrier silver mines. The charcoal, with its adhering chlorides, was conveyed to the smelting-house and the gold smelted into bars of extremely pure metal. Messrs. Newbery and Vautin claimed for their process decreased time for the operation with increased efficiency.
At Mount Morgan, when I visited that mine in the eighties, they were using what might be termed a composite adaptation process. Their chlorination works, the largest in the world, were putting through 1500 tons per week. The ore as it came from the mine was fed automatically into Krumm roller mills, and after being crushed and sifted to regulation gauge was delivered into trucksand conveyed to the roasting furnaces, and thence to cooling floors, from which it was conveyed to the chlorinating shed. Here were long rows of revolving barrels, on the Newbery-Vautin principle, but with this marked difference, that the pressure in the barrel was obtained from an excess of the gas itself, generated from a charge of chloride of lime and sulphuric acid. On leaving the barrels the pulp ran into settling vats, somewhat on the Plattner plan, and the clear liquid having been drained off was passed through a charcoal filter, as adopted by Newbery and Vautin. The manager, Mr. Wesley Hall, stated that the estimated cost per ton was not more than 30s., and he expected shortly to reduce that when he began making his own sulphuric acid. As he was obtaining over 4 oz. to the ton the process was paying very well, but it will be seen that the price would be prohibitive for poor ores unless they could be concentrated before calcination.
The Pollok process is a newer, and stated to be a cheaper mode of lixiviation by chlorine. It is the invention of Mr. J. H. Pollok, of Glasgow University, and a strong Company was formed to work it. With him the gas is produced by the admixture of bisulphate of sodium (instead of sulphuric acid, which is a very costly chemical to transport) and chloride of lime. Water is then pumped into a strong receptacle containing the material for treatment and powerful hydraulic pressure is applied. The effect is stated to be the rapid change of the metal into its salt, which is dissolved in the water and afterwards treated with sulphate of iron, and so made to resume its metallic form.
It appears, however, to me that there is no essential difference in the pressure brought to bear for the quickening of the process. In each case it is an air cushion, induced in the one process by the pumping in of air to a cylinder partly filled with water, and in the other by pumping in water to a cylinder partly filled with air.
The process of extracting gold from lode stuff and tailings by means of cyanide of potassium is now largely used and may be thus briefly described:—It is chiefly applied to tailings, that is, crushed ore that has already passed over the amalgamating and blankettables. The tailings are placed in vats, and subjected to the action of solutions of cyanide of potassium of varying strengths down to 0·2 per cent. These dissolve the gold, which is leached from the tailings, passed through boxes in which it is precipitated either by means of zinc shavings, electricity, or other precipitant. The solution is made up to its former strength and passed again through fresh tailings. When the tailings contain a quantity of decomposed pyrites, partly oxidised, the acidity caused by the freed sulphuric acid requires to be neutralised by an alkali, caustic soda being usually employed.
When “cleaning up,” the cyanide solution in the zinc precipitating boxes is replaced by clean water. After careful washing in the box, to cause all pure gold and zinc to fall to the bottom, the zinc shavings are taken out. The precipitates are then collected, and after calcination in a special furnace for the purpose of oxidising the zinc, are smelted in the usual manner.
The following description of an electrolytic method of gold deposition from a cyanide solution was given by Mr. A. L. Eltonhead before the Engineers’ Club of Philadelphia.
A description of the process is as follows:—“The ore is crushed to a certain fineness, depending on the character of the gangue. It is then placed in leaching vats, with false bottoms for filtration, similar to other leaching plants. A solution of cyanide of potassium and other chemicals of known percentage is run over the pulp and left to stand a certain number of hours, depending on the amount of metal to be extracted. It is then drained off and another charge of the same solution is used, but of less strength, which is also drained. The pulp is now washed with clean water, which leaches all the gold and silver out, and leaves the tailings ready for discharge, either in cars or sluiced away by water, if it is plentiful.
“The chemical reaction of cyanide of potassium with gold is as follows, according to Elsner:—₂Au + ₄KCy + O + H₂O = ₂KAuCy₂ + ₂KHO.That is, a double cyanide of gold and potassium is formed.“All filtered solutions and washings from the leaching vats are saved and passed through a precipitating ‘box’ of novel construction, whichmay consist either of glass, iron or wood, and be made in any shape, either oval, round, or rectangular—if the latter, it will be about 10 ft. long, 4 ft. wide and 1 ft. high—and is partitioned off lengthwise into five compartments. Under each partition, on the inside or bottom of the ‘box,’ grooves may be cut a quarter-to a half-inch deep, extending parallel with the partitions to serve as a reservoir for the amalgam, and give a rolling motion to the solution as it passes along and through the four compartments. The centre compartment is used to hold the lead or other suitable anode and electrolyte.“The anode is supported on a movable frame or bracket, so it may be moved either up or down as desired, it being worked by thumb-screws at each end.“The electrolyte may consist of saturated solutions of soluble alkaline metals and earths. The sides or partitions of each compartment dip into the mercury, which must cover the ‘box’ evenly on the bottom to the depth of about a half-inch.“Amalgamated copper strips or discs are placed in contact with the mercury and extended above it, to allow the gold and silver solution of cyanide to come in contact.“The electrodes are connected with the dynamo; the anode of lead being positive and the cathode of mercury being negative. The dynamo is started, and a current of high amperage and low voltage is generated, generally 100 to 125 amperes, and with sufficient pressure to decompose the electrolyte between the anode and the cathode.“As the gas is generated at the anode, a commotion is created in the liquid, which brings a fresh and saturated solution of electrolyte between the electrodes for electrolysis, and makes it continuous in its action.“The solution of double cyanide of gold, silver, and potassium, which has been drained from the leaching vats, is passed over the mercury in the precipitating ‘box’ when the decomposition of the electrolyte by the electric current is being accomplished, the gold and silver are set free and unite with the mercury, and are also deposited on the plates or discs of copper, forming amalgam, which is collected and made marketable by the well-known andtried methods. The above solution is regenerated with cyanide of potassium by the setting free of the metals in the passage over the ‘box.’“In using this solution again for a fresh charge of pulp, it is reinforced to the desired percentage, or strengthened with cyanide of potassium and other chemicals, and is always in good condition for continuing the operation of dissolving.“The potassium acting on the water of the solution creates nascent hydrogen and potassium hydrate; the nascent hydrogen sets free the metals (gold and silver), which are precipitated into the mercury and form amalgam, leaving hydrocyanic acid; this latter combines with the potassium hydrate of the former reaction, thus forming cyanide of potassium. There are other reactions for which I have not at present the chemical formulas.“As the solution passes over the mercury, the centre compartment of the ‘box’ is moved slowly longitudinally, which spreads the mercury, the solution is agitated and comes in perfect contact with the mercury, as well as the amalgamated plates or discs of copper, ensuring a perfect precipitation.“It is not always necessary to precipitate all the gold and silver from the solution, for it is used over and over again indefinitely; but when it is required, it can be done perfectly and cheaply in a very short time.“No solution leached from the pulp, containing cyanide of potassium, gold and silver, need be run to waste, which is in itself an enormous saving over the use of zinc shavings when handling large quantities of pulp and solution.“Some of the advantages the electro-chemical process has over other cyanide processes are: Its cleanliness, quickness of action, cheapness, and large saving of cyanide of potassium by regeneration; not wasting the solutions, larger recovery of the gold and silver from the solutions; the cost of recovery less; the loss of gold, silver, and cyanide of potassium reduced to a minimum; the use of caustic alkali in such quantity as may be desired to keep the cyanide solution from being destroyed by the acidity of the pulp, and also sometimes to give warmth, as a warm cyanide solution will dissolve gold and silver quicker than a cold one.These caustic alkalies do not interfere with or prevent the perfect precipitation of the metals. The bullion recovered in this process is very fine, while the zinc-precipitated bullion is only about 700 fine.“The gold and silver is dissolved, and then precipitated in one operation, which we know cannot be done in the ‘chlorination process’; besides, the cost of plant and treatment is much less in the above-described process.“The electro-chemical process, which I have hastily sketched will, I think, be the future cheap method of recovering fine or flour gold from our mines and waste tailings or ore dumps.“Without going into details of cost of treatment, I will state that with a plant of a capacity of handling 10,000 tons of pulp per month, the cost should not exceed 8s.per ton, but that may be cheapened by labour-saving devices. There being no expensive machinery, a plant could be very cheaply erected wherever necessary.”
“The chemical reaction of cyanide of potassium with gold is as follows, according to Elsner:—
₂Au + ₄KCy + O + H₂O = ₂KAuCy₂ + ₂KHO.
That is, a double cyanide of gold and potassium is formed.
“All filtered solutions and washings from the leaching vats are saved and passed through a precipitating ‘box’ of novel construction, whichmay consist either of glass, iron or wood, and be made in any shape, either oval, round, or rectangular—if the latter, it will be about 10 ft. long, 4 ft. wide and 1 ft. high—and is partitioned off lengthwise into five compartments. Under each partition, on the inside or bottom of the ‘box,’ grooves may be cut a quarter-to a half-inch deep, extending parallel with the partitions to serve as a reservoir for the amalgam, and give a rolling motion to the solution as it passes along and through the four compartments. The centre compartment is used to hold the lead or other suitable anode and electrolyte.
“The anode is supported on a movable frame or bracket, so it may be moved either up or down as desired, it being worked by thumb-screws at each end.
“The electrolyte may consist of saturated solutions of soluble alkaline metals and earths. The sides or partitions of each compartment dip into the mercury, which must cover the ‘box’ evenly on the bottom to the depth of about a half-inch.
“Amalgamated copper strips or discs are placed in contact with the mercury and extended above it, to allow the gold and silver solution of cyanide to come in contact.
“The electrodes are connected with the dynamo; the anode of lead being positive and the cathode of mercury being negative. The dynamo is started, and a current of high amperage and low voltage is generated, generally 100 to 125 amperes, and with sufficient pressure to decompose the electrolyte between the anode and the cathode.
“As the gas is generated at the anode, a commotion is created in the liquid, which brings a fresh and saturated solution of electrolyte between the electrodes for electrolysis, and makes it continuous in its action.
“The solution of double cyanide of gold, silver, and potassium, which has been drained from the leaching vats, is passed over the mercury in the precipitating ‘box’ when the decomposition of the electrolyte by the electric current is being accomplished, the gold and silver are set free and unite with the mercury, and are also deposited on the plates or discs of copper, forming amalgam, which is collected and made marketable by the well-known andtried methods. The above solution is regenerated with cyanide of potassium by the setting free of the metals in the passage over the ‘box.’
“In using this solution again for a fresh charge of pulp, it is reinforced to the desired percentage, or strengthened with cyanide of potassium and other chemicals, and is always in good condition for continuing the operation of dissolving.
“The potassium acting on the water of the solution creates nascent hydrogen and potassium hydrate; the nascent hydrogen sets free the metals (gold and silver), which are precipitated into the mercury and form amalgam, leaving hydrocyanic acid; this latter combines with the potassium hydrate of the former reaction, thus forming cyanide of potassium. There are other reactions for which I have not at present the chemical formulas.
“As the solution passes over the mercury, the centre compartment of the ‘box’ is moved slowly longitudinally, which spreads the mercury, the solution is agitated and comes in perfect contact with the mercury, as well as the amalgamated plates or discs of copper, ensuring a perfect precipitation.
“It is not always necessary to precipitate all the gold and silver from the solution, for it is used over and over again indefinitely; but when it is required, it can be done perfectly and cheaply in a very short time.
“No solution leached from the pulp, containing cyanide of potassium, gold and silver, need be run to waste, which is in itself an enormous saving over the use of zinc shavings when handling large quantities of pulp and solution.
“Some of the advantages the electro-chemical process has over other cyanide processes are: Its cleanliness, quickness of action, cheapness, and large saving of cyanide of potassium by regeneration; not wasting the solutions, larger recovery of the gold and silver from the solutions; the cost of recovery less; the loss of gold, silver, and cyanide of potassium reduced to a minimum; the use of caustic alkali in such quantity as may be desired to keep the cyanide solution from being destroyed by the acidity of the pulp, and also sometimes to give warmth, as a warm cyanide solution will dissolve gold and silver quicker than a cold one.These caustic alkalies do not interfere with or prevent the perfect precipitation of the metals. The bullion recovered in this process is very fine, while the zinc-precipitated bullion is only about 700 fine.
“The gold and silver is dissolved, and then precipitated in one operation, which we know cannot be done in the ‘chlorination process’; besides, the cost of plant and treatment is much less in the above-described process.
“The electro-chemical process, which I have hastily sketched will, I think, be the future cheap method of recovering fine or flour gold from our mines and waste tailings or ore dumps.
“Without going into details of cost of treatment, I will state that with a plant of a capacity of handling 10,000 tons of pulp per month, the cost should not exceed 8s.per ton, but that may be cheapened by labour-saving devices. There being no expensive machinery, a plant could be very cheaply erected wherever necessary.”
The object of calcining or roasting certain ores before treatment is to dissipate the sulphur or sulphides of arsenic, antimony, lead, &c., which are inimical to treatment, whether by ordinary mercuric amalgamation or lixiviation. The effect of the roasting is first to sublimate and drive off as fumes the sulphur and a proportion of the objectionable metals. What is left is either iron oxide, “gossan,” or the oxides of the other metals. Even lead can thus be oxidised, but requires more care as it melts nearly as readily as antimony and is much less volatile. The oxides in the thoroughly roasted ore will not amalgamate with mercury, and are not acted on by chlorine or cyanogen.
To effect the oxidation of sulphur, it is necessary not only to bring every particle of sulphur into contact with the oxygen of the air, but also to provide adequate heat to the particles sufficient to raise them to the temperature that will induce oxidation. No appreciable effect follows the mere contact of air with sulphur particles at atmospheric temperature; but if the particles be raised to a temperature of 500° Fahr., the sulphur is oxidised to the gaseous sulphur dioxide. The same action effects the elimination of the arsenic and antimony associated with gold and silver ores, as when heated to a certain constant temperature these metals readily oxidise.
The science of calcination consists of the method by which the sulphide ores, having been crushed to a proper degree of fineness, are raised to a sufficient temperature and brought into intimate contact with atmospheric air.
It will be obvious then that the most effective method ofroasting will be one that enables the particles to be thoroughly oxidised at the lowest cost in fuel and in the most rapid manner.
The roasting processes in practical use may be divided into three categories:
First or Process A.—Roasting on a horizontal and stationary hearth, the flame passing over a mass of ore resting on such hearth. In order to expose the upper surface of the ore to contact with air the material is turned over by manual labour. This furnace of the reverberatory type is provided with side openings by which the turning over of the ore can be manually effected, and the new ore can be charged and afterwards withdrawn.
Second or Process B.—Roasting in a revolving hearth placed at a slightly inclined angle from the horizontal. The furnace is of cylindrical form, and is internally lined with refractory material. It has projections that cause the powdered ore to be lifted above the flame, and, at a certain height, to fall through the flame and so be rapidly raised to the temperature required to effect the oxidation of the oxidisable minerals which it is desired to extract.
The rate, or speed, of revolution of this revolving furnace obviously depends upon the character of the ore under treatment; it may vary from two revolutions per minute down to one revolution in thirty minutes. Any kind of fuel is available, but that of a gaseous character is stated to be by far the most efficient.
Any ordinary cylinder of a length of 25 ft., and a diameter of 4 ft. 6 in., inclined 1 ft. 6 in. in its length, will calcine from 24 to 48 tons per diem.
Another form of rotating furnace is one in which the axis is horizontal. It is much shorter than the inclined type, and the feeding and removal of the ore is effected by the opening of a retort lid door provided at the side of the furnace. Openings provided at each end of the furnace permit the passage of the flame through it, and the revolution of the furnace turns over the powdered ore and brings it into more or less sustained contactis continued sufficiently long to ensure the more or less complete oxidation of the ore particles.
Third or Process C.—In this process the powdered ore is allowed to fall in a shower from a considerable height, through the centre of a vertical shaft up which a flame ascends; the powdered ore in falling through the flame is heated to an oxidising temperature, and the sulphides are thus depleted of their sulphur and become oxides.
Another modification of this direct fall or shaft furnace is that in which the fall of the ore is checked by cross-bars or inclined plates placed across the shaft; this causes a longer oxidising exposure of the ore particles.
When the sulphur contents of pyritous ores are sufficiently high, and after the ore has been initially fired with auxiliary carbonaceous fuel, it is unnecessary, in a properly designed roasting furnace, to add fuel to the ore to enable the heat for oxidation to be obtained. The oxidation or burning of the sulphur will provide all the heat necessary to maintain the continuity of the process. The temperature necessary for effecting the elimination of both sulphur and arsenic is not higher than that equivalent to dull red heat; and provided that there is a sufficient mass of ore maintained in the furnace, the potential heat resulting from the oxidation of the sulphur will alone be adequate to provide all that is necessary to effect the calcination.
The construction of this furnace has already been sufficiently described. If the roasting is performed in a muffle chamber, the arrangement employed by Messrs. Leach and Neal, Limited, of Derby, and designed by Mr. B. H. Thwaite, C.E., can be advantageously employed in this furnace, which is fired with gaseous fuel. The sensible heat of the waste gases is utilised to heat the air employed for combustion; and by a controllable arrangement ofcombustion, a flame of over 100 feet in length is obtained, with the result that the furnace from end to end is maintained at a uniform temperature. By this system, and with gaseous fuel firing, a very considerable economy in fuel and in repairs to furnace, and a superior roasting effect, have been obtained.
Where the ordinary reverberatory hearth is fired with solid coal from an end grate, the temperature is at its maximum near the firing end, and tails off at the extreme gas outlet end. The ores in this furnace should therefore be fed in at the colder end of the hearth and be gradually worked or “rabbled” forward to the firing end.
One disadvantage of the reverberatory furnace is the fact that it is impossible to avoid the incursion of air during the manual rabbling action, and this tends to cool the furnace.
The cost of roasting, to obtain the more or less complete oxidation, or what is known in mining parlance as a “sweet roast” (because a perfectly roasted ore is nearly odourless) varies considerably, the variation depending of course upon the character of the ore and the cost of labour and fuel.
There are several modifications of the reverberatory furnace in use, designed mechanically to effect the rabbling. One of the most successful is that known as the Horse-shoe furnace. In plan the hearth of the furnace resembles a horse-shoe.
The stirring of the ore over the hearth is effected by means of carriages fixed in the centre of the furnace and having laterally projecting arms, carrying stirrers, that move along the hearth and turn over the pulverised ore.
In operation, half the carriages are traversing the furnace, and half are resting in the cooling space, so that a control over the temperature of the stirrers is established.
This furnace is stated to be more economical in labour than other mechanically stirred reverberatory furnaces, and there is also said to be an economy in fuel.
Usually the mechanical stirring furnaces give trouble and should be avoided, but the horse-shoe type possesses qualifications worthy of consideration.
Of these some of the best known to me are: The Howell-White, The Brückner, The Thwaite-Denny, and the Molesworth.
The Brückner is a cylinder, turning on its horizontal axis and carried by four rollers.
The batch of ore usually charged into the two charging hoppers weighs about four tons. When the two charging doors are brought under the hopper mouth, the contents of the hopper fall directly into the cylinder.
The ends or throats of the furnace are reduced just sufficiently to allow the flame evolved from a grated furnace to pass completely through the cylinder.
A characteristic size for this Brückner furnace is one having a length of 12 feet and a diameter of 6 feet. A furnace of this capacity will have an inclusive weight (iron and brickwork) of 15 tons.
The time of operation, with the Brückner, will vary with the character of the ore under treatment and the nature of the fuel employed. Four hours is the minimum and twelve hours should be the maximum time of operation.
By the addition of common salt with the batch of ore, such of its constituents as are amenable to the action of chlorine are chlorinated as well as freed from sulphur.
Where the ore contains any considerable quantity of silver which should be saved, the addition of the salt is necessary as the silver is very liable to become so oxidised in the process of roasting as to render its after treatment almost impossible. I know a case in point where an average of nearly five ounces of silver to the ton, at that time worth 30s., was lost owing to ignorance on this subject. Had the ore been calcined with salt, NaCl, the bulk of this silver would have been amalgamated and thus saved. It was the extraordinary fineness of the gold saved by amalgamation as against my tests of the ore by fire assay that put me on the track of a most indefensible loss.
The Howell-White Furnace.—This furnace consists of a cast iron revolving cylinder, averaging 25 feet in length and 4 ft. 4 in. indiameter, which revolves on four friction rollers, resting on truck wheels, rotated by ordinary gearing.
The power required for effecting the revolution should not exceed four indicated horse-power.
The cylinder is internally lined with firebrick, projecting pieces causing the powdered ore to be raised over the flame through which it showers, and is thereby subjected to the influence of heat and to direct contact oxidation.
The inclination of the cylinder, which is variable, promotes the gradual descension of the ore from the higher to the lower end. It is fed into the upper end, by a special form of feed hopper, and is discharged into a pit at the lower end, from which the ore can be withdrawn at any time.
The gross weight of this furnace, which is, however, made in segments to be afterwards bolted together, is some ninety to one hundred tons.
The furnace is fired with coal on a grated hearth, built at the lower end; it is more economical both in fuel and in labour than an ordinary reverberatory furnace.
The Thwaite-Denny Revolving Furnace.—This new type of furnace, which is fired with gaseous fuel, is stated to combine the advantages of the Stetefeldt, the Howell-White, and the Brückner.
It is constructed as follows:—Three short cylinders, conical in shape and of graduated dimensions, are superposed one over the other, their ends terminating in two vertical shafts of brickwork, by which the three cylinders are connected. The powdered ore is fed into the uppermost cylinder and gravitates through the series. The highest cylinder is the largest in diameter, the lowest the smallest.
The gas flame, burnt in a Bunsen arrangement, enters the smallest end of the lowest cylinder and passes through it; then returns through the intermediate cylinder above it, being directed by the brickwork shaft—from one cylinder to the other—till finally the gases flow through the topmost cylinder and enter into a dust depositing chamber. The gases evolved increase as the flame flows through the series and the ore is reduced by the expulsion of its sulphur, arsenic, &c., as it descends from thetop to the bottom. The top cylinder is made larger than the one below it and the middle cylinder is made larger than the lowest one in proportion to the increased bulk of gases and ore.